Light Metals 2011
Check out these new proceeding volumes from the TMS 2011 Annual Meeting, available from publisher John Wiley & Sons: 2nd International Symposium on High-Temperature Metallurgical Processing Energy Technology 2011 : Carbon Dioxide and Other Greenhouse Gas Reduction Metallurgy and Waste Heat Recovery EPD Congress 2011 Friction Stir Welding and Processing VI Light Metals 2011 Magnesium Technology 2011 Recycling of Electronic Waste II, Proceedings of the Second Symposium Sensors, Sampling and Simulation for Process Control Shape Casting: Fourth International Symposium 2011 Supplemental Proceedings: Volume 1 : Materials Processing and Energy Materials Supplemental Proceedings: Volume 2: Materials Fabrication, Properties, Characterization, and Modeling Supplemental Proceedings: Volume 3: General Paper Selections To purchase any of these books, please visit www.wiley.com. TMS members should visit www.tms.org to learn how to get discounts on these or other books through Wiley.
Light Metals 2011 Proceedings of the technical sessions presented by the TMS Aluminum Committee at the TMS 2011 Annual Meeting & Exhibition, San Diego, California, USA February 27-March 3, 2011
Edited by Stephen J. Lindsay
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Copyright © 2011 by The Minerals, Metals, & Materials Society. All rights reserved. Published by John Wiley & Sons, Inc., Hoboken, New Jersey. Published simultaneously in Canada. No part of this publication may be reproduced, stored in a retrieval system, or transmitted in any form or by any means, electronic, mechanical, photocopying, recording, scanning, or otherwise, except as permitted under Section 107 or 108 of the 1976 United States Copyright Act, without either the prior written permission of The Minerals, Metals, & Materials Society, or authorization through payment of the appropriate per-copy fee to the Copyright Clearance Center, Inc., 222 Rosewood Drive, Danvers, MA 01923, (978) 750-8400, fax (978) 750-4470, or on the web at www.copyright.com. Requests to the Publisher for permission should be addressed to the Permissions Department, John Wiley & Sons, Inc., I l l River Street, Hoboken, NJ 07030, (201) 748-6011, fax (201) 748-6008, or online at http:// www.wiley.com/go/permission. Limit of Liability/Disclaimer of Warranty: While the publisher and author have used their best efforts in preparing this book, they make no representations or warranties with respect to the accuracy or completeness of the contents of this book and specifically disclaim any implied warranties of merchantability or fitness for a particular purpose. No warranty may be created or extended by sales representatives or written sales materials. The advice and strategies contained herein may not be suitable for your situation. You should consult with a professional where appropriate. Neither the publisher nor author shall be liable for any loss of profit or any other commercial damages, including but not limited to special, incidental, consequential, or other damages. Wiley also publishes books in a variety of electronic formats. Some content that appears in print may not be available in electronic formats. For more information about Wiley products, visit the web site at www.wiley.com. For general information on other Wiley products and services or for technical support, please contact the Wiley Customer Care Department within the United States at (800) 762-2974, outside the United States at (317) 572-3993 or fax (317) 572-4002. Library of Congress Cataloging-in-Publication Data is available.
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TABLE OF CONTENTS Light Metals 2011 Preface About the Editor Program Organizers Aluminum Committee
xix xxi xxiii xxix
Alumina and Bauxite Bauxite Resources and Utilisation New Development Model for Bauxite Deposits P. ter Weer
5
Study on the Characterization of Marginal Bauxite from Parâ/Brazil F. Silva, J. Sampaio, M. Medeiros, andF. G arrido
13
Resource Utilization of High-sulfur Bauxite of Low-median Grade in Chongqing China J. Yin, W. Xia, andM. Han
19
Development of Bauxite and Alumina Resources in the Kingdom of Saudi Arabia A. Al-Dubaisi
23
Digestion Studies on Central Indian Bauxite P. Raghavan, N. Kshatriya, and Dasgupta
29
Effects of Roasting Pretreatment in Intense Magnetic Field on Digestion Performance of Diasporic Bauxite Z Ting-an, D. Zhihe, L Guozhi, L. Y an, D. Juan, W. Xiaoxiao, andL. Y an
33
Bayer Process I Application of Operation Integrity Management in the Alumina Industry C Suarez, D. Welshons, J. McNerney, andJ. Webb Influence of Solid Concentration, Particle Size Distribution, Ph And Temperature on Yield Stress of Bauxite Pulp C. Barbato, M. Nele, and S. Franca
41
47
A New Method for Removal Organics in the Bayer Process B. Yingwen, L. Jungi, S. Mingliang, andZ. Fei
51
Alunorte Expansion 3 - The New Lines Added to Reach 6.3 Million Tons per Year D. Khoshneviss, L. Correa, J. Ribeiro Alves Filho, H. Berntsen, andR. Carvalho
57
One Green Field Megaton Grade Large Alumina Refinery with Successful Engineering & Operation Experience...63 L. Xianqing, and Y. Xiaoping Advanced Process Control in the Evaporation Unit C. Kumar, U. Giri, R. Pradhan, T. Banerjee, R. Saha, and P. Pattnaik
v
69
Improvements in Smelter Grade Alumina Quality at Clarendon Alumina Works R. Shaw, A. Duncan, andM. Crosciale
75
Red Mud Application of Nanofiltration Technology to Improve Sea Water Neutralization of Bayer Process Residue K. Taylor, M. Mullett, L. Fergusson, H. Adams on, andJ. Wehrli
81
Caustic and Alumina Recovery from Bayer Residue S Gu
89
Investigation on Alumina Discharge into the Red Mud Pond at Nalco's Alumina Refinery, Damanjodi, Orissa, India B. Mohapatra, B. Mishra, and G Mishra Production of Ordinary Portland Cement( OPC) from NALCO Red Mud G Mishra, D. Yadav, M. Alii, and P. Sharma
93 97
Recovery of Metal Values from Red Mud P. Raghavan, N. Kshatriya, andK. Wawrynink
103
Red Mud Flocculants used in the Bayer Process S. Moffatt, F. Ballentine, andM. Lewellyn
107
Reductive Smelting of Greek Bauxite Residues for Iron Production A. Xenidis, G Zografidis, I. Kotsis, andD. Boufounos
113
Precipitation, Calcination and Properties Effect of Technological Parameters on PSD of Aluminum Tri-Hydroxide from Seed Precipitation in Seeded Sodium Aluminate Solutions 121 Y. Wu, L. Mingchun, and Q. Yanping Methods to Reduce Operating Costs in Circulating Fluidized Bed Calcination G Klett, M. Miss alla, B. Reeb, andH. Schmidt
125
Pressure Calcination Revisited F. Williams, and G Misra
131
Dynamic Simulation of Gas Suspension Calciner (GSC) for Alumina B. Raahauge, S. Wind, M. Wu, and T. Jensen
137
Physical Simulation and Numerical Simulation of Mixing Performance in the Seed Precipitation Tank with a Improved Intermig Impeller Z Ting-an, L. Y an, W. Shuchan, Z Hongliang, Z. Chao, Z Qiuyue, D. Zhihe, andL. Guozhi
145
Two Perspectives on the Evolution and Future of Alumina L. Perander, J. Metson, and G Klett
151
Significant Improvement of Energy Efficiency at Alunorte's Calcination Facility M Missalla, H. Schmidt, J. Ribeiro, andR. Wischnewski
157
Attrition of Alumina in Smelter Handling and Scrubbing Systems S. Lindsay
163
vi
Energy and Environment Perspective on Bayer Process Energy D. Donaldson
171
Optimization of Heat Recovery from the Precipitation Circuit R. Singh, S. Hial, andM. Simpson
175
Alunorte Global Energy Efficiency A. Monteiro, R. Wischnewski, C. Azevedo, and E. Moraes
179
Opportunities for Improved Environmental Control in the Alumina Industry R. Mimna, J. Kildea, E. Phillips, W, Carlson, B. Reiser, andJ. Meier
185
Alumina Refinery Wastewater Management: When Zero Discharge Just Isn't Feasible L. Martin, andS. Howard
191
High Purity Alumina Powders Extracted from Aluminum Dross by the Calcining—Leaching Process L. Qingsheng, Z. Chunming, F. Hui, andX. Jilai
197
Effect of Calcium/Aluminium Ratio on MgO Containing Calcium Aluminate Slags W. Bo, S Hui-Lan, G Dong, andB. Shi-Wen
201
Study on Extracting Aluminum Hydroxide from Reduction Slag of Magnesium Smelting by Vacuum Aluminothermic Reduction W. Yaowu, F. Naixiang, Y. Jing, H. Wenxin, P. Jianping, D. Yuezhong, and W. Zhihui
205
Application of Thermo-gravimetric Analysis for Estimation of Tri-hydrate Alumina in Central Indian Bauxites—An Alternative for Classical Techniques 211 Y. Ramana, andR. Patnaik Determination of Oxalate Ion in Bayer Liquor Using Electrochemical Method S. Turhan, B. Usta, Y. Sahin, andO. Uysal
215
Alternative Alumina Sources - Poster Session The Effect of Ultrasonic Treatment on Alumina Leaching from Calcium Aluminate Slag S. Hui-lan, W. Bo, G Dong, Z. Xue-zheng, andB. Shi-wen
221
Theory and Experiment on Cooling Strategy during Seeded Precipitation Z Liu, W. Chen, and W. Li
227
Extraction of Alumina from Red Mud by Divalent Alkaline Earth Metal Soda Ash Sinter Process S. Meher, A. Rout, andB. Padhi
231
Dissolution Kinetics of Silicon from Sintering Red Mud in Pure Water X. Li, K. Huang, and H Zhu
237
The Effect of Cooling Rate on the Leachability of Calcium Aluminate Slags W. Bo, S. Hui-lan, Z. Xue-zheng, and B. Shi-wen
241
Preparing Polymerized Aluminum-ferrum Chloride with Red Mud L. Guilin, Y. Haiyan, andB. Shiwen
245
Adsorption of Polyethylene Glycol at the Interface of Dicalcium Silicate - Sodium Aluminate Solution Y. Haiyan, X Pan, Z Lu, and T. Ding
251
vu
Production of Hematite Ore from Red Mud P. Raghavan, N. Kshatriya, and K Wawrynink
255
Aluminum Reduction Technology Enviroment- Emissions/ Anode Effect I HF Measurements Inside an Aluminium Electrolysis Cell K. Osen, T Aarhaug, A. Solheim, E. Skybakmoen, andC. Sommerseth
263
LasIRTM-R - The New Generation RoHS-Compliant Gas Analyzers Based on Tunable Diode Lasers J. Gagne, J. Pisano, A. Chanda, G. Mackay, K. Mackay, and P. Bouchard
269
Use of Spent Potlining (SPL) in Ferro Silico Manganese Smelting P. von Krüger
275
Reduction of PFC Emissions at Pot Line 70 kA of Companhia Brasileira de Aluminio H. Santos, D. Melo, J. Calixto, J. Santos, andJ. Miranda
281
Towards Redefining the Alumina Specifications Sheet - The Case of HF Emissions L Perander, M. Stam, M. Hyland, andJ. Metson
285
Design of Experiment to Minimize Fluoride and Particulate Emissions at Alumar E. Batista, P. Miotto, E. Montoro, andL. Souza
291
Innovative Distributed Multi-Pollutant Pot Gas Treatment System G. Wedde, O. Bjarno, and A. Sorhuus
295
Fluoride Emissions Management Guide (FEMG) for Aluminium Smelters N. Tjahyono, Y. Gao, D. Wong, W. Zhang andM. Taylor
301
Enviroment- Emissions/ Anode Effect II On Continuous PFC Emission Unrelated to Anode Effects X. Chen, W. Li, J. Marks, Q. Zhao, J. Yang, S. Qiu, and C. Bayliss Monitoring Air Fluoride Concentration around ALUAR Smelter in Puerto Madryn (Chubut Province, Argentina) J. Zavatti, C. Moreno, J. Lifschitz, and G. Quiroga
309
315
Reduction of Anode Effect Duration in 400kA Prebake Cells W. Zhang, D. Wong, M. Gilbert, Y. Gao, M. Dorreen, M. Taylor, A. Tabereaux, M. Soffer, X. Sun, C. Hu, X Liang, H. Qin, J. Mao, andX Lin
319
Sustainable Anode Effect Based Perfluorocarbon Emission Reduction N. Dando, L. Sylvain, J. Fleckenstein, C. Kato, V. Van Son, andL. Coleman
325
The Initiation, Propagation and Termination of Anode Effects in Hall-Héroult Cells G. Tarcy, and A. Tabereaux
329
Towards Eliminating Anode Effects A. Al Zarouni, B. Welch, M. Mohamed Al-Jallaf, and A. Kumar
333
Correlation between Moisture and HF Formation in the Aluminium Process C. Sommerseth, K Osen, T Aarhaug, E. Skybakmoen, A. Solheim, C. Rosenkilde, and A. Ratvik
339
Vili
Particulate Emissions from Electrolysis Cells H. Gaertner, A. Ratvik, and T. Aarhaug
345
Investigation of Solutions to Reduce Fluoride Emissions from Anode Butts and Crust Cover Material G. Girault, M. Faure, J. Bertolo, S. Massambi, and G. Bertran
351
PFC Survey in Some Smelters of China W. Li, X Chen, Q. Zhao, S. Qiu, andS. Zhang
357
Considerations Regarding High Draft Ventilation as an Air Emission Reduction Tool S. Broek, N. Dando, S. Lindsay, and A. Moras
361
Cells Thermal Balance Increasing the Power Modulation Window of Aluminium Smelter Pots with Shell Heat Exchanger Technology ...369 P. Lavoie, S. Namboothiri, M. Dorreen, J. Chen, D. Zeigler, andM. Taylor New Approaches to Power Modulation at TRIMET Hamburg T Reek
375
Some Aspects of Heat Transfer Between Bath and Sideledge in Aluminium Reduction Cells A. Solheim
381
Towards a Design Tool for Self-heated Cells Producing Liquid Metal by Electrolysis S. Poizeau, andD. Sadoway
387
Heat Recovery from Aluminium Reduction Cells Y. Ladam, A. Solheim, M. Segatz, and O. Lorentsen
393
Effects of Composition and Granulometry on Thermal Conductivity of Anode Cover Materials H. Wijayaratne, M. Hyland, M. Taylor, A. Grama, and T Groutso
399
Restart of 300kA Potlines after 5 Hours Power Failure X. Zhao, B. Gao, H. Han, J. Liu, J. Xiao, J. Qian, J. Yan, andD. Wang
405
Multiblock Monitoring of Aluminum Reduction Cells Performance J. Tessier, C Duchesne, andG. Tarcy
407
Cells Technology, Development and Sustainability High Amperage Operation of AP18 pots at Karmoy M Bugge, H. Haakonsen, O. Kobbeltvedt, andK. Paulsen
415
Aluminium Smelter Manufacturing Simulation - Can These Bring Real Cost Savings? M Meijer
421
Simultaneous Preheating and Fast Restart of 50 Aluminium Reduction Cells in an Idled Potline - A New Soft Restart Technique for a Pot Line 425 A. Mulder, A. Folkers, M. Stam, andM. Taylor SWOT Perspectives ofMidagePrebaked Aluminium Smelter P. Choudhury, and A. Sharma
431
Integrated Approach for Safe and Efficient Plant Layout Development R. Pires, R. Baxter, L. Tikasz, andR. McCulloch
437
IX
New Progress on Application of NEUI400kA Family High Energy Efficiency Aluminum Reduction Pot ("HEEP") Technology 443 D. Lu, J. Qin, Z. Ai, and Y. Ban Improving Current Efficiency of Aged Reduction Lines at Aluminium Bahrain (Alba) A. Ahmed, K. Raghavendra, H Hassan, andK. Ghuloom
449
Development of NEUI500kA Family High Energy Efficiency Aluminum Reduction Pot ("HEEP") Technology ..455 D. Lu, Y. Ban, X Qi, J. Mao, Q. Yang, andK Dong
Cells Process Control Current Efficiency for Aluminium Deposition from Molten Cryolite-alumina Electrolytes in a Laboratory Cell....461 G. Haarberg, J. Armoo, H Gudbrandsen, E. Skybakmoen, A. Solheim, and T Jentoftsen
Improvement in Cell Equipment and Design Retrofit of a Combined Breaker Feeder with a Chisel Bath Contact Detection System to Reduce Anode Effect Frequency in a Potroom J. Verreault, R. Gariépy, B. Desgroseilliers, C. Simard, X. Delcorde, C. Turpain, S. Simard, andS. Déry
467
Anode Dusting from a Potroom Perspective at Nordural and Correlation with Anode Properties H. Gudmundsson
471
The Application of Continuous Improvement to Aluminium Potline Design and Equipment W. Paul
477
Alcoa STARprobe™ X. Wang, B. Hosier, and G. Tarcy
483
Active Pot Control using Alcoa STARprobe™ X. Wang, G. Tarcy, E. Batista, and G. Wood
491
Technology & Equipment for Starting Up & Shutting Down Aluminium Pots under Full Amperage Y. Tao, L. Meng, C. Bin, and Y. Xiaobing
497
Study on Solution of A1203 in Low Temperature Aluminum Electrolyte H Kan, N. Zhang, andX Wang
503
Applications of New Structure Reduction Cell Technology in Chalco's Smelters F. Liu, S. Gu, J. Wang, andK Yang
509
Transport Numbers in the Molten System NaF-KF-AlF3-Al203 P. Fellner, J. Hives, andJ. Thonstad
513
Cells Process Modeling Development and Application of an ANSYS Based Thermo-electro-mechanical Collector Bar Slot Design Tool M Dupuis Impact of Amperage Creep on Potroom Busbars and Electrical Insulation: Thermal-Electrical Aspects A. Schneider, D. Richard, and O. Charette
519 525
Modern Design of Potroom Ventilation A. Vershenya, U. Shah, S. Broek, T. Plikas, J. Woloshyn, andA. Schneider
531
A Preliminary Finite Element Electrochemical Model for Modelling Ionic Species Transport in the Cathode Block ofa Hall-Héroult Cell 537 F. Gagnon, D. Ziegler, andM. Fafard CFD Modelling of Alumina Mixing in Aluminium Reduction Cells Y. Feng, M. Cooksey, and P. Schwarz
543
Bubble Transport by Electro-Magnetophoretic Foces at Anode Botttom of Aluminium Cells V. Bojarevics, andA. Roy
549
Anodic Voltage Oscillations in Hall-Héroult Cells K. Einarsrud, andE. Sandnes
555
Energy Savings by Cell Design Improvements Electrical Conductivity of the KF-NaF- A1F3 Molten System at Low Cryolite Ratio with CaF2 Additions A. Redkin, A. Dedyukhin, A. Apisarov, P. Tin'ghaev, and Y. Zaikov
563
Study of ACD Model and Energy Consumption in Aluminum Reduction Cells T. Yingfu, and W. Hang
567
Modeling of Energy Savings by Using Cathode Design and Inserts R. von Kaenel, andJ. Antille
569
Experimental Investigation of Single Bubble Characteristics in a Cold Model of a Hall-Héroult Electrolytic Cell S. Das, Y. Morsi, G Brooks, W. Yang, andJ. Chen
575
Large Gas Bubbles under the Anodes of Aluminum Electrolysis Cells A. Caboussat, L. Kiss, J. Rappaz, K. Vékony, A. Perron, S. Renaudier, and O. Martin
581
Initiatives to Reduction of Aluminum Potline Energy Consumption Alcoa Poços de Caldas/Brazil A. Abreu, M. Salles, and C. Kato
587
Overview of High-Efficiency Energy Saving for Aluminium Reduction Cell X. Canming, and Y. Xiaobing
591
Cell Voltage Noise Reduction Based on Wavelet in Aluminum Reduction Cell B. Li, J. Chen, X Zhai, S. Sun, andG. Tu
599
Poster Session Human Factors in Operational and Control Decision Making in Aluminium Smelters Y. Gao, M. Taylor, J. Chen, andM. Hautus
xi
605
Aluminum Rolling Session I An Investigation of Deformation Behavior of Bimetal Clad Sheets by Asymmetrical Rolling at Room Temperature L. Xiaobing, Z Guoyin, and D. Qiang
615
Coil Build Up Compensation during Cold Rolling to Improve Off-line Flatness L. Almeida Neto, and T. Ayhan
621
Through Process Effects on Final Al-sheet Flatness S. Neumann, andK. Karhausen
625
Cast Shop for Aluminum Production Casthouse Productivity and Safety New Casthouse Smelter Layout for the Production of Small Non-Alloyed Ingots: Three Furnaces/Two Lines J. Berlioux, A. Bourgier, andJ. Baudrenghien
635
Use of Process Simulation to Design a Billet Casthouse G. Jaouen
641
Optimizing Scrap Reuse as a Key Element in Efficient Aluminium Cast Houses T. Schmidt, J. Migchielsen, D. Ing, andK Grab
647
Implementation of an Effective Energy Management Program Supported by a Case Study R. Courchée
653
Molten Metal Safety Approach through a Network C. Pluchon, B. Hannart, L Jouet-F* astre, J. Mathieu, R. Wood, J. Riquet, F. Fehrenbach, G. Ranaud, M. Bertherat, andJ. Hennings
657
Improved Monolithic Materials for Lining Aluminum Holding & Melting Furnaces A. Wynn, J. Coppack, T. Steele, andK. Moody
663
Direct Chill Casting Cold Cracking during Direct-chill Casting D. Eskin, M. Lalpoor, andL. Katgerman
669
Surface Defects Structures on Direct Chill Cast 6xxx Aluminium Billets M. Erdegren, and T. Carlberg
675
Effect of Cooling Water Quality on Dendrite Arm Spacing of DC Cast Billets S. Mohapatra, S. Nanda, and A. Palchowdhury
681
Mould Wall Heat Flow Mechanism in a DC Casting Mould A. Prasad, andL Bainbridge
687
Productivity Improvements at Direct Chill Casting Unit in Aluminium Bahrain (ALBA) A. Noor, S. Chateeriji, and A. Ahmed
693
Xll
The Coupling of Macrosegregation with Grain Nucleation, Growth and Motion in DC Cast Aluminum Alloy Ingots M. Zaloznik, A. Kumar, H. Combeau, M. Bedel, P. Jarry, andE. Waz
699
Investment Casting of Surfaces with Microholes and Their Possible Applications T. Ivanov, A. Buehrig-Polaczek, U. Vroomen, C. Hartmann, A. Gillner, K Bobzin, J. Holtkamp, N. Bagcivan, andS. Theiss
705
Using SEM and EDX for a Simple Differentiation of?- and ?-AlFeSi-Phases in Wrought Aluminum Billets M Rosefort, C. Matthies, H. Buck, and H. Koch
711
Dross Formation, Control and Handling Oxidation of AlMg in Dry and Humid Atmospheres A. Kvithyld, D. Stevens, S. Wilson, and T. Engh
719
Study of Early Stage Interaction of Oxygen with Al; Methods, Challenges and Difficulties B. Fatela, G BrooL·, M. Rhamdhani, J. Taylor, J. Davis, andM. Lowe
725
Quality Assessment of Recycled Aluminium D. Dispinar, A. Kvithyld, and A. Nordmark
731
Melt Quality Control In-Line Salt-ACD™: A Chlorine-Free Technology for Metal Treatment P. Robichaud, C. Dupuis, A. Mathis, P. Coté, andB. Maltais
739
The Effectof TiB2 Granules on Metal Quality M. MohamedAl-Jallaf, M. Hyland, B. Welch, A. Al Zarouni, andF. Abdullah
745
Thermodynamic Analysis of Ti, Zr, V and Cr Impurities in Aluminium Melt A. Khaliq, M. Rhamdhani, G. BrooL·, andJ. Grandfield
751
Current Technologies for the Removal of Iron from Aluminum Alloys L. Zhang, J. Gao, andL. Damdah
757
Electromagnetically Enhanced Filtration of Aluminum Melts M. Kennedy, S. Akhtar, R. Aune, andJ. Bakken
763
A Review of the Development of New Filter Technologies Based on the Principle of Multi Stage Filtration With Grain Refiner Added in the Intermediate Stage 769 J. Courtenay, S. Instone, andF. Reusch Wettability of Aluminium with SiC and Graphite in Aluminium Filtration S. Bao, A. Kvithyld, T. Engh, andM. Tangstad
775
Study of Microporosity Formation under Different Pouring Conditions in A356 Aluminum Alloy Castings L. Yao, S. Cocker oft, D. Maijer, J. Zhu, and C Reilly
783
Grain Refinement Alloying, Solidification and Casting Hycast Gas Cushion (GC) Billet Casting System 7. Steen, and A. Hakonsen
793
Xlll
Studies of Fluid Flow and Meniscus Behavior during Horizontal Single Belt Casting (HSBC) of Thin Metallic Strips D. Li, J. Gill, M. Isac, andR. Guthrie
797
Development of Alba High Speed Alloy A. Ahmed, J. Hassan, G. Martin, andK Ghosh
803
Dissolution Studies of Si Metal in Liquid Al under Different Forced Convection Conditions M. Seyed Ahmadi, S. Argyropoulos, M. Bussmann, andD. Doutre
809
Modification and Grain Refinement of Eutectics to Improve Performance of Al-Si Castings M. Felberbaum, and A. Dahle
815
Production of Al-Ti-C Grain Refiners with the Addition of Elemental Carbon and K2TiF6 F. Toptan, I. Kerti, S. Daglilar, A. Sagin, O. Karadeniz, and A. Ambarkutuk
821
Effect of Mechanical Vibrations on Microstructure Refinement of Al-7mass% Si Alloys T. Tamura, T. Matsuki, andK. Miwa
827
Predicting the Response of Aluminum Casting Alloys to Heat Treatment C Wu, andM. Makhlouf
831
Electrode Technology for Aluminium Production Anode Baking Determination of Coke Calcination Level and Anode Baking Level - Application and Reproducibility of L-sub-c Based Methods 841 S. Rorvik, L. Lossius, and A. Ratvik Operation of an Open Type Anode Baking Furnace with a Temporary Crossover E. Cobo, L. Beltramino, J. Artola, J. Rey Boero, P. Roy, andJ. Bigot
847
Recent Developments in Anode Baking Furnace Design D. Severo, V. Gusberti, P. Sulger, F. Keller, andM. Meier
853
Sohar Aluminium's Anode Baking Furnace Operation S. AlHosni, J. Chandler, O. Forato, F. Morales, C. Jonville, andJ. Bigot
859
Meeting the Challenge of Increasing Anode Baking Furnace Productivity F. Ordronneau, M. Gendre, L. Pomerleau, N. Backhouse, A. Berkovich, andX. Huang
865
Wireless Communication for Secured Firing and Control Systems in Anode Baking Furnaces N. Fiot, and C. Coulaud
871
Full Control of Pitch Burn during Baking: It's Impact on Anode Quality, Operational Safety, Maintenance and Operational Costs D. Maiwald, D. Di Lisa, and P. Mnikoleiski High Performance Sealing for Anode Baking Furnaces P. Mahieu, S. Neple, N. Fiot, I. Ofico, andM. Eufrasio
xiv
875 881
Anode Raw Materials and Green Carbon Property Profile of Lab-scale Anodes Produced with 180°C Mettler Coal Tar Pitch W. Boenigk, C Boltersdorf, F. Lindner, andJ. Stiegert
889
Quality and Process Performance of Rotary Kilns and Shaft Calciners L. Edwards
895
Sub-surface Carbon Dioxide Reaction in Anodes D. Ziegler
901
Paste Quality Improvements at Alcoa Poços de Caldas Plant B. Vry, C Kato, J. Araujo, F. Ribeiro, and A. Abreu
907
Prebaked Anode from Coal Extract (2) - Effects of the Properties of Hypercoal-coke on the Preformance of Prebaked Anodes M. Hamaguchi, N. Okuyama, N. Komatsu, J. Koide, K. Kano, T. Shishido, K. Sakai, and T. Inoue The New Generation of Vertical Shaft Calciner Technology J. Zhao, Q. Zhao, and Q. Zhao
913 917
Petroleum Coke VBD Historical and Future Challenges with the Vibrated Bulk Density Test Methods for Determining Porosity of Calcined Petroleum Coke J. Panchal, M. Wyborney, andJ. Rolle
925
Prediction of Calcined Coke Bulk Density M. Dion, H. Darmstadt, N. Backhouse, F. Cannava, and M. Canada
931
Calcined Coke Particle Size and Crushing Steps Affect Its VBD Result F. Cannava, M. Canada, andB. Vitchus
937
Bulk Density - Overview of ASTM and ISO Methods with Examples of Between Laboratory Comparisons L. Lossius, B. Spencer, andH. 0ye
941
Improving the Repeatability of Coke Bulk Density Testing L. Edwards, M. Lubin, andJ. Marino
947
ASTM D7454 Vibrated Bulk Density Method - Principles and Limitations F. Laplante, andL. Duchesneau
953
Vibrated Bulk Density (VBD) of Calcined Petroleum Coke and Implications of Changes in the ASTM Method D4292 B. Spencer, L. Johnsen, D. Kirkpatrick, D. Clark, andM. Baudino
959
Anode Quality and Rodding Processes Multivariate Monitoring of the Prebaked Anode Manufacturing Process and Anode Quality J. Lauzon-Gauthier, C Duchesne, J. Tessier, K. Cantin, and I. Petit
967
Characterization of a Full Scale Prebaked Carbon Anode using X-Ray Computerized Tomography D. Picard, H. Alamdari, D. Ziegler, P. St-Arnaud, andM. Fafar d
973
xv
FEM Analysis of the Anode Connection in Aluminium Reduction Cells S. Beier, J. Chen, M. Fafard, andH. Fortin Development of Industrial Benchmark Finite Element Analysis Model to Study Energy Efficient Electrical Connections for Primary Aluminium Smelters D. Molenaar, K. Ding, and A. Kapoor
979
985
Real Time Temperature Distribution During Sealing Process and Room Temperature Air Gap Measurements of a Hall-Héroult Cell Anode 991 O. Trempe, D. Larouche, D. Ziegler, M. Guillot, andM. Fafard Effects of High Temperatures and Pressures on Cathode and Anode Interfaces in a Hall-Heroult Electrolytic Cell L. St-Georges, L. Kiss, J. Bouchard, M. Rouleau, andD. Marceau
997
New Apparatus for Characterizing Electrical Contact Resistance and Thermal Contact Conductance N. Kandev, H. Fortin, S. Chénard, G. Gauvin, M. Martin, andM. Fafard
1003
Carbon Anode Modeling for Electric Energy Savings in the Aluminium Reduction Cell D. Andersen, and Z Zhang
1009
Cathode Design and Operation Preheating Collector Bars and Cathode Blocks Prior to Rodding with Cast Iron by Passing an AC Current Through the Collector Bars 1017 E. Jensen, H Bjornstad, andJ. Hansen Development and Application of an Energy Saving Technology for Aluminum Reduction Cells P. Jianping, F. Naixiang, F. Shaofeng, L. Jun, and Q. Xiquan
1023
Study of Electromagnetic Field in 300kA Aluminium Reduction Cells with Innovation Cathode Structure B. Li, X. Zhang, S. Zhang, F. Wang, andN. Feng
1029
Evaluation of the Thermophysical Properties of Silicon Carbide, Graphitic and Graphitized Carbon Sidewall Lining Materials Used in Aluminium Reduction Cell in Function of Temperature 1035 A. Khatun, andM. Desilets Advanced Numerical Simulation of the Thermo-Electro-Mechanical Behaviour of Hall-Héroult Cells under Electrical Preheating D. Marceau, S. Pilote, M. Désilets, L. Hacini, J. Bilodeau, and Y. Caratini
1041
Influence of Technological and Constructive Parameters on the Integrity of the Bottom of Aluminum Reduction Cells during Flame Preheating 1047 A. Arkhipov, G. Arkhipov, and V. Pingin Creep Behaviors of Industrial Graphitic and Graphitized Cathodes during Modified Rapoport Tests W. Wang, J. Xue, J. Feng, Q. Liu, L. Zhan, H. He, andJ. Zhu
1053
Cathode Materials and Wear Measurement of Cathode Surface Wear Profiles by Laser Scanning E. Skybakmoen, S. Itervik, A. Solheim, K. Holm, P. Tiefenbach, and O. Ostrem
1061
Coke Selection Criteria for Abrasion Resistant Graphitized Cathodes R. Perruchoud, W. Fischer, M. Meier, and U. Mannweiler
1067
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Determination of the Effect of Pitch-Impregnation on Cathode Erosion Rate P. Patel, Y. Sato, and P. Lavoie
1073
Simplifying Protection System to Prolong Cell Life M. Mohamed Al-Jallaf, M. Hyland, B. Welch, and A. AlZarouni
1079
Aluminate Spinels as Sidewall Linings for Aluminum Smelters X. Yan, R. Mukhlis, M. Rhamdhani, and G. Brooks
1085
A New Ramming Paste with Improved Potlining Working Conditions B. Aliard, R. Paulus, and G. Billat
1091
Towards a Better Understanding of Carburation Phenomenon M. Lebeuf, M Coulombe, B. Allard, and G. Soucy
1097
Characterization of Sodium and Fluorides Penetration into Carbon Cathodes by Image Analysis and SEM-EDS Techniques 1103 Y. Gao, J. Xue, J. Zhu, K. Jiao, and G. Jiang
Inert Anodes and Wettable Cathodes Pressureless Sintering of TiB2-based Composites using Ti and Fe Additives for Development of Wettable Cathodes H. Heidari, H. Alamdari, D. Dubé, andR. Schulz Furan Resin and Pitch Blends as Binders for TiB2-C Cathodes H. Zhang, J. Hou, X. Lü, Y. Lai, andJ. Li Influence of Cobalt Additions on Electrochemical Behaviour of Ni-Fe-Based Anodes for Aluminium Electrowinning V. Singleton, B. Welch, andM. Skyllas-Kazacos Effects of the Additive Zr0 2 on Properties of Nickel Ferrite Cermet Inert Anode X. Zhang, G. Yao, Y. Liu, J. Ma, andZ. Zhang
1111 1117
1123 1129
Effect of Sintering Atmosphere on Phase Composition and Mechanical Property of 5Cu/(10NiO-NiFe2O4) Cermet Anodes for Aluminum Electrolysis 1135 Z Zou, C. Wei, Z Tian, K. Liu, H. Zhang, Y. Lai, andJ. Li
Poster Session - Electrode Influence of Ultrafine Powder on the Properties of Carbon Anode Used in Aluminum Electrolysis X. Jin, D. Songyun, L. Jie, L. Yanqing, andL. Yexiang
1143
Preparation NiFe 2 0 4 Matrix Inert Anode Used in Aluminum Electrolysis by Adding Nanopowder Z Zhang, G. Yao, Y. Liu, andX Zhang
1149
Cold Water Model Simulation of Aluminum Liquid Fluctuations Induced by Anodic Gas in New Tape of Cathode Structure Aluminum Electrolytic Cell 1155 Y. Liu, T. Zhang, Z Dou, H. Wang, G. Lv, Q. Zhao, N. Feng, andJ. He Effects of Physical Properties of Anode Raw Materials on the Paste Compaction Behavior K. Azari, H. Ammar, H. Alamdari, D. Picard, M. Fafard, andD. Ziegler
xvii
1161
Furnace Efficiency - Energy and Throughput Session I Furnaces Designed for Fuel Efficiency D. White
1169
Latest Trends in Post Consumer and Light Gauge Scrap Processing to include Problematic Materials such as UBC, Edge Trimming and Loose Swarf 1173 F. Niedermair, and G. Wimroither Investigation of Heat Transfer Conditions in a Reverberatory Melting Furnace by Numerical Modeling A. Buchholz, andJ. Rodseth
1179
Oxyfuel Optimization using CFD Modeling T. Niehoff, and S. Viyyuri
1185
Operational Efficiency Improvements Resulting from Monitoring and Trim of Industrial Combustion Systems J. Oakes, andD. Bratcher
1189
New Technology for Electromagnetic Stirring of Aluminum Reverberatory Furnaces J. Herbert, and A. Peel
1193
Evaluation of Effects of Stirring in a Melting Furnace for Aluminum K. Matsuzaki, T. Shimizu, Y. Murakoshi, andK. Takahashi
1199
Business Analysis of Total Refractory Costs C. Belt
1205
Improved Furnace Efficiency through the Use of Refractory Materials J. Hemrick, A. Rodrigues-Schroer, D. Colavito, andJ. Smith
1211
Study on the Energy-saving Technology of Chinese Shaft Calciners G. Lang, C. Bao, S. Gao, R. Logan, Y. Li, andJ. Wu
1217
Author Index
1221
Subject Index
1227
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PREFACE As editor it is my pleasure to present to you these contributed proceedings of TMS's 140th Annual Meeting and Exposition in San Diego, California. The volumes included here represent a large collective undertaking. All of it has been volunteered by the hundreds of authors, dozens of Session Chairpersons, and more than a dozen Symposium Chairpersons and Vice-Chairpersons that have created Light Metals 2011. As members we owe them all a debt of gratitude for the time and effort that they have donated. Although our industry has faced trying times in recent years contributions to this volume from industry and academia have been generous. The authors of these technical papers represent approximately fifty universities. There are almost twice that number of contributions from private industry, individual contributors, and research institutes combined. Many of these individuals or organizations have prepared technical papers on varied topics and across multiple symposia. Such support from our membership is what makes our annual meetings both productive and successful. This represents the best of TMS, a professional, diverse, and growing organization that embraces and promotes both pure and applied sciences. The volumes of Light Metals represent a large fraction of the accumulated knowledge of our industry that is in the public domain. It is often used as a primary source of reference information in the preparation of new contributions to the technical literature. Each year the wealth of information in the accumulated volumes of Light Metals grows and 2011 is no exception. Yet, it is not enough to rest on these laurels. Our future is being shaped now by forces that our industry could not have anticipated even a decade ago. We look to grow, to include academic and industrial papers from countries, universities, and enterprises that have yet to be represented in Light Metals along with those from more established contributors. We hope to allow future authors to see further, if not from standing upon the shoulders of the giants that have preceded them in our industry. I encourage our members to not only participate in annual meetings, but also to get actively involved. TMS committees are all composed of volunteers. Authors that have contributed in the past are most likely to contribute again. However, they, as I, would like to hear from members who may have been tempted to write a technical paper but never have done so. The strength of our organization is built upon new ideas and insights that come from all quarters of academia, research groups, and industry. New authors are always welcomed. On behalf of the organizers for Light Metals 2011 allow me to thank the TMS staff including Maria Boots, Chris Wood, and Christina Raabe Eck and the TMS Light Metals Aluminum Committee for their support. I would also like to recognize the contributions of John A. Johnson for his guidance and for his organization of the Plenary Session celebrating 125 Years of the Hall-Héroult Aluminum Reduction Process. I especially would like to recognize the 2011 Subject Chairpersons: Mohammed Mahmood, Abdullah Habib Ahmed Ali, Dr. Alan Tomsett, Dr. James Metson, Dr. Geoffrey Brooks, Kai Karhausen and Thomas Nieoff for their dedication and leadership in preparation for our 2011 Annual Meeting. Stephen J. Lindsay
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EDITOR'S BIOGRAPHY
STEPHEN J. LINDSAY LIGHT METALS 2011 EDITOR Stephen Joseph Lindsay holds a B.S. in Chemical Engineering from Clarkson College of Technology and an M.A. in Applied Behavioral Science from Bastyr University's Leadership Institute of Seattle program. During his time with Alcoa he has held numerous positions with responsibilities in anode, cathode, pollution control systems, and reduction technology. He has specialized in areas including emissions control, metal purity, alumina and electrolytes. In these areas he supports Alcoa's Primary Products division worldwide. His wife, Dr. Margarita Merino de Lindsay, an author, poet, and artist in her own right is his muse. She has encouraged Steve to contribute regularly to technical literature and education, plus control of pollution. It is for her sake that the colors of Light Metals 2011 are those of Spain, her home country. A member of TMS since 1985, Steve has regularly authored or co-authored in Light Metals. He has received the Best Paper Award in Reduction Technology in 2006 and again in 2009. He served as the Subject Chair for Reduction Technology in 2006, has instructed in various short courses, and has served under the direction of Dr. Halvor Kvande in the TMS Industrial Electrolysis Courses held since 2005. He has also authored or co-authored technical papers appear in the proceedings of the 8th and 9th Australasian Smelting Technology Conferences, the 7th and 8th International Alumina Quality Workshops, the International Committee for Study of Bauxite, Alumina, & Aluminium 2010, and the International Beryllium Research Conference 2007. Steve served on the Aluminum Association's Industrial Hygiene sub-committee for beryllium contributing to the understanding its mass balance in aluminum smelters. He has participated as an instructor on a regular basis in courses organized by the University of Auckland's Light Metals Research Centre, the University of New South Wales, and Alcoa's own Process Engineering Training Program. Steve is based at Alcoa's Tennessee Operations near Knoxville, Tennessee and works with Alcoa's Technology, Innovation and Center of Excellence group. His is a manager in Primary Metal's Best Practices group.
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PROGRAM ORGANIZERS ALUMINA and BAUXITE Jim Metson graduated with PhD in Chemistry from Victoria University of Wellington, New Zealand, before taking up a position at Surface Science Western, University of Western Ontario Canada. He then moved to the University of Auckland, New Zealand, where he is a Professor, the Associate Director of the Light Metals Research Centre and Head of the Department of Chemistry. He is a Director of the New Synchrotron Group Ltd, a councillor of the Australian Institute of Nuclear Science and Engineering and chairs the Research Infrastructure Advisory Group (RIAG) for the New Zealand Government. His research interests are in materials and particularly surface science, with an emphasis on applications in the aluminium industry including alumina calcination and evolution of microstructure, smelting technology and in particular the impacts of alumina properties, and the surface science of aluminium metal. He has had more than 20 years of engagement with the aluminium industry and has been a regular participant at the Annual TMS meeting. He is a past Light Metals Award winner and has co-ordinated a course "Alumina from a Smelter Perspective" held as part of the 2004 TMS meeting and was a presenter in the 2009 course "Alumina Refinery Fundamentals and Practice". Carlos Suarez has been associated with the alumina and bauxite industry for 30 years and has been a member of TMS since 1984. Carlos attended the University of Oklahoma where he obtained a degree of Science in Chemical Engineering. He also attended the University of Phoenix where he obtained a Master in Business Administration. Carlos has been involved in all aspects of alumina refining for employers such as Bauxilum, Nabalco, Vialco and Gramercy Alumina in the areas of Process Safety, Quality, Training and Development, Technical Sales, Plant Operations Research and Development, Commissioning and Start-Ups, Knowledge Management, Organizational Development, Technology Transfer and Business Development. He has been a Process Consultant for Hatch since 2004 where he has served as process lead and project manager for different alumina plant projects around the world. Carlos has been an active member of TMS. He has contributed with several technical papers and was one of the instructors for the first Alumina Refinery Fundamentals and Practice course sponsored by TMS in 2008.
xxiii
ALUMINUM REDUCTION TECHNOLOGY Mohammed Mahmood holds Master degree in Process Engineering from Strathclyde University in Scotland in 1989. He began his career with Aluminium Bahrain (ALBA) more than thirty years ago, rose through the ranks to various managerial positions, from Manager of Potlines, Manager Process & Quality Control to Manager Human Resources & Development and then to General Manager Metal Production from 2004 - 2009 and finally in 2009 to his present position as Chief Operating Officer. Among the major milestone in his career has been the retrofitting of pot lines 1 -3 that increased the production by 21 %, lead the team to further improvement and achieve 2.7% higher productivity and improve pot operation age by 16%. Being a prominent figure in Bahrain, Mohamed is very often invited to speak at International Conferences both Technical and People Development related. He is the head of the Alba Community Service Committee where his role encouraged the spirit of philanthropy amongst Alba employees and enhanced kingdom wide appreciation of Alba's corporate social responsibility initiatives. His main passion is the development of youth to become future leaders. Abdulla Habib Ahmed, Manager Research & Development in Aluminum Bahrain (Alba), joined Alba in March of 1995 after completing his degree in Chemical Engineering with first honor class as Process Engineer. Abdulla was involved in many projects and studies to maximize Aluminum Production in Alba. He gradually climbed the success ladder of Alba hierarchy to become in charge of Metal Production as Reduction Line Superintended in year 2000. On November 2004, Abdulla completed his Master degree in University of New South Wales in Australia with first honor class. In July, 2007, Abdulla embarked on doing his Ph.D. in the same University to be the first Bahraini doing the Ph.D. in Aluminium Smelting technology and among few people in Middle East specialized on the Aluminum field. In September, 2009 he has been promoted to become first R&D Manager in Alba. Abdulla is looking after the innovations; process improvements in Reduction, Carbon and Casthouse in Aluminum Bahrain (Alba). Charles "Mark" Read is Bechtel Senior Specialist - Primary Aluminium Processes. Mark is currently Bechtel's Area Manager, Reduction for the Ma'aden "Ras Az Zawr Aluminium Smelter Project, Kingdom of Saudi Arabia. Previous Bechtel roles included Engineering Manager for green-field and brown-field aluminium smelter projects, and technology and engineering oversight of studies for major Middle Eastern, North American and Russian aluminium Smelters. Mark has 33 years experience in business and technology management in the Metals Industry, over 25 years of which were in the aluminium industry including in-depth technical experience of Hall-Héroult cell design and operation, pre-baked carbon products processing and performance, and aluminium casting operations. Mark joined Bechtel's Montreal-based "Aluminium Centre of Excellence" in late 2003. Prior to joining Bechtel, Mark held various technology management positions with Elkem Metals, Kaiser Aluminium & Chemical Corporation and Alcan Inc. Mark is a graduate of Sheffield Hallam University, England. He holds a B.Sc. degree in Metallurgical Engineering and M.Sc. in Industrial Metallurgy.
xxiv
CAST SHOP for ALUMINUM PRODUCTION Geoffrey Brooks, B.Eng. (RMIT), B.A. (SUT), PhD (Melb.) F. LEng. Aust, has been a Professor in the Faculty of Engineering and Industrial Sciences at Swinburne University of Technology since 2006, where he leads the High Temperature Processing research group. He also the leader of a cluster of researchers from Australian and New Zealand Universities focussed on improving Aluminium smelting. Previously, he was a Senior Principal Research Scientist at CSIRO (2004-2006), an Associate Professor in Materials Science and Engineering at McMaster University (2000-2004) and a Senior Lecturer at the University of Wollongong (1993-2000). In the 17 years since completing his PhD at University of Melbourne, he have published over 100 papers and run many large research projects with funding from many major companies and government agencies. He is currently active in work on dross formation in aluminium processing, controlling minor elements in the casthouse, sidewall materials in aluminium cells, development of sensors for bubbling in high temperature operations, modelling of injection processes and distribution of elements in magnesium production. He has been a key reader for Metallurgical and Materials Transactions since 1998 and is a Fellow of the Institute of Engineers (Australia). Geoff has been a member of the TMS since 1990. Dr. John Grandfield is director of Grandfield Technology Pty Ltd, a consulting and technology firm. John has a Bachelor of Applied Science in Metallurgy (RMIT), a MSC in Mathematical Modelling (Monash University) and a PhD in Materials Science (University of Queensland). John has 25 years experience in light metals cast house research in industry and government laboratories (Rio Tinto Alcan, CASTcrc and CSIRO). He has developed new technology for aluminium and magnesium DC casting, and open mould conveyor ingot casting. He conducts problem solving and research projects, presents cast house technology training courses around the world, participates in in-house innovation workshops and conducts R&D program reviews. John has four patents and has published more than 50 conference and journal papers. He is chair of the Australasian Aluminium Casthouse Technology conference.
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ELECTRODE TECHNOLOGY for ALUMINUM PRODUCTION Alan Tomsett has over twenty years experience in carbon anode and cathode technology. He received his BSc and PhD in Chemical Engineering from the University of New South Wales in Sydney, Australia. He joined Rio Tinto Alcan at the R&D centre in Melbourne in 1987. His activities with R&D Group have included leadership of the global and regional Carbon R&D program, provision of technical support for the RTA Australasian smelters, carbon raw material evaluation and carbon plant technology selection for brownfield and greenfield expansions. Since 2008, Alan has been the Technical Manager - Carbon for Rio Tinto Alcan Primary Metal Pacific. Alan has been a member of TMS since 1996. He is the coauthor of several TMS papers and is a previous TMS session chair. He is also a regular contributor to the Australasian Smelting Conference. Barry Sadler has been involved in the Aluminium Industry for more than 25 years in a range of positions but always focusing on anode carbon technology. His career started in 1982 at the Comalco (Now Rio Tinto Alcan) Research Centre in Melbourne, Australia. In 1989 he moved to Comalco's New Zealand Aluminium Smelter as Carbon Plant Manager. After a stint as General Manager Organisational Effectiveness for Hamersley Iron, in 1989 Barry took up the position of Technical General Manager at Comalco Aluminium's corporate headquarters in Brisbane, Australia. Leaving Rio Tinto/Comalco in 2002 to establish Net Carbon Consulting Pty Ltd, Barry now provides consulting advice, training, and support to clients on improving plant performance, with emphasis on the practical application of statistical thinking to process management. Barry has been a regular contributor at TMS meetings for over 20 years as an author, session chairperson, and Electrodes subject organiser.
xxvi
ALUMINIUM ROLLING Kai Friedrich Karhausen is department manager for process technology at the central Rolled Products R&D of Hydro Aluminium in Bonn, Germany. Dr. Karhausen earned his doctorate at the RWTH Aachen and worked in the industrial aluminum research for 15 years both in Norway and Germany. His principal work is focused on the modeling and optimization of materials behavior in industrial production processes. Dr. Karhausen has issued 75 scientific presentations and publications. In 2003 he was awarded the Georg-Sachs-Preis of the German Materials Society (DGM) for important achievements in the field of integrated modeling of metal forming and materials behavior.
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FURNACE EFFICIENCY ENERGY and THROUGHPUT Thomas Niehoff currently Head of Non Ferrous and Mining at The Linde Group, Div. Linde Gas is based in Munich, Germany. Graduated from RWTH Aachen in Germany in mechanical engineering in 1992. Thomas has 18 years experience in combustion and metallurgical applications related to industrial gases. In his global role Thomas now overlooks the R&D activities for Linde Gas. He has in depth experience in metallurgy of aluminum, iron and steel - combustion processes and emissions from combustion. He did his PhD at RWTH Aachen on coke fired cupola process optimization with oxyfuel.
xxvni
ALUMINUM COMMITTEE 2011-2012 Chairperson John A. Johnson Johnson's Consulting Group Krasnoyarsk, Russia
Light Metals Division Chairperson John N. Hryn Argonne National Laboratory Illinois, USA
Vice Chairperson Stephen J. Lindsay Alcoa Inc. Tennessee, USA
JOM Advisor Pierre P. Homsi Rio Tinto Alcan
Past Chairperson Geoffrey Paul Bearne Rio Tinto Alcan Victoria, Australia
Secretary Charles Mark Read Bechtel Corp. Quebec, Canada
MEMBERS THROUGH 2012 Hussain H. Alali Retired, Aluminum Bahrain Manama, Bahrain
Stephen J. Lindsay Alcoa Inc. Alcoa, Tennessee, USA
Martin Iffert Trimet Aluminum AG Essen, Germany
MEMBERS THROUGH 2013 Gilles Dufour Alcoa Canada Quebec, Canada
Everett Phillips Nalco Company Illinois, USA
Pierre Le Brun Alcan Voreppe Research Center Voreppe Cedex, France
Barry Sadler Net Carbon Consulting Pty. Ltd. Kangaroo Ground, Australia
MEMBERS THROUGH 2014 John G rand field Granfield Technology Pty. Ltd. Victoria, Australia
Ketil A. Rye Alcoa Mosjoen Mosjoen, Norway
Charles Mark Read Bechtel Corp. Quebec, Canada
Carlos Suarez Hatch Associates Inc. Pennsylvania, USA
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Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
Light Metals 2011 ALUMINA and BAUXITE
ORGANIZERS
James Metson University of Auckland Auckland, New Zealand Carlos Suarez Hatch Associates Inc. Pittsburgh, Pennsylvania, USA
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
Light Metals 2011 ALUMINA and BAUXITE
Bauxite Resources and Utilisation SESSION CHAIR
Shawn Kostelak Gramercy Alumina Louisiana, USA
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
NEW DEVELOPMENT MODEL FOR BAUXITE DEPOSITS Peter-Hans ter Weer1 'TWS Services and Advice, Imkerweg 5, 1272 EB Huizen, The Netherlands;
[email protected] Keywords: Bauxite, Alumina, Project Development, Alumina Technology, Economics
1.
comprises a string of equipment which together performs the desired process step, e.g. digestion with feed tank, heat exchangers, pumps, digester vessel(s), flash vessels, etc. Such a string of equipment is often referred to as a "train", "unit" or "circuit" (e.g. digestion unit, precipitation train, mill circuit). Alumina refinery design generally takes the digestion area as plant bottleneck due to its high unit capital cost and its requirement for constant flow for optimum performance. The design / initial refinery production capacity of greenfield projects has evolved over time from about 0.5-1.0 Mt/y alumina 25-30 years ago (e.g. Worsley, Alumar, Aughinish) to 1.4-3.3 Mt/y alumina for more recently constructed and future planned projects (e.g. Lanjigarh, Yarwun, Utkal, GAC). Figure 1 illustrates this trend.
Abstract
Developing a greenfield bauxite deposit nowadays generally includes constructing an alumina refinery. Economics have resulted in ever-increasing production capacities for recently-built and future planned greenfield refineries. Rationale: economy of scale. As a result the complexity of a greenfield project has significantly increased and its capital cost has grown to several billion USD. Important consequences: • Project owners aim at risk reduction through project financing and formation of joint ventures, further complicating project implementation. • Globally only a limited number of (large) companies have the human andfinancialresources to develop greenfield bauxite & alumina projects. • Only a limited number of engineeringfirmshave the required skills and experience to successfully implement these mega projects. • Only large bauxite deposits get developed.
Greenfield Refinery Design Capacity as function of Start-up Year j
;
♦
This paper proposes an alternative development model for bauxite deposits resulting in a more efficient use of resources and a lower threshold to develop bauxite & alumina projects.
2.
Actual
♦■
;^ !
o
Bauxite Deposit Development
♦
♦ T
Planned
♦ ♦
♦
♦
*
«
i
The development of bauxite deposits is sometimes limited to the mining of bauxite for export purposes, which may or may not include drying the bauxite to a certain moisture percentage. Examples are the Boke and Kindia mines (both in Guinea), and the Bintan mine in Indonesia (now closed). In other cases the mine supplies both a local / in-country refinery, as well as exporting bauxite, e.g. the Trombetas mine (Brazil), and the Gove and Weipa mines (both in Australia). In most recent cases the projected greenfield development of a bauxite deposit includes directly or indirectly the construction of a captive alumina refinery. Examples: Utkal (India), GAC (Guinea), Aurukun (Australia), CAP (Brasil), Ma'aden (Saudi Arabia). In some cases the project may be executed in two stages: a first stage of establishing the bauxite mine with (temporary) export of bauxite, and a second stage including the construction of an alumina refinery. A recent example is the Darling Range project of Bauxite Resources Ltd in Australia as stated in press releases. How have greenfield production capacities and more specifically greenfield alumina refinery design capacities developed over time, and did this have a bearing on project implementation?
3.
♦
' 1990
' 2000
i
1
2010
Start-up Year
Figure 1 - Refinery Design Capacity vs Start-up Year Note that actual refinery production capacities increase over time as a result of de-bottlenecking, improved process efficiencies and operations performance, etc. In a paper presented at the ICSOB A 2008 conference [1] R. den Hond even suggests a doubling of design capacity by exploiting overdesign and post start-up installation of novel technology. What has been the rationale for this trend of ever-increasing design production capacities for recently built and future planned greenfield refineries and what are its consequences? 3.2 Economy of Scale The rationale offered for this trend is the economy of scale: an increased alumina production capacity improves the economics (NPV, IRR, VIR1) of a greenfield bauxite and alumina project2. In the context of alumina refinery projects, economy of scale aspects may be applied to Operating Cost and Capital Cost.
Alumina Refinery Capacity Evolution
3.1 Overview An alumina refinery consists of a number of unit operations such as grinding, digestion, evaporation, etc. A unit operation generally
1 NPV=Net Present Value; IRR=lnternal Rate of Return; VIR=Value over Investment (capital efficiency) ratio. 2 Reference [2] provides an overview of bauxite & alumina project economics.
5
3.2.1 Effect on Operating Cost3 To better assess the effects of the economy of scale on Operating Cost, we should consider its major components: • Variable costs: In $/year these costs vary with plant production, at least within certain plant production rates (typically ± 10-15%), examples: bauxite, caustic soda, coal, fuel oil, lime. The overall plant on-line time of an alumina refinery with more than one train / unit / circuit, e.g. a digestion train, is higher than a plant with one train only, as a result of moreflexibilityin equipment operation and maintenance. The effect on plant on-line time is generally limited (indie. 0.2-0.5% abs), however may vary widely and in a specific case could be significant (>1% abs). As a result the plant operates with less interruptions and operating efficiencies (e.g. bauxite, caustic soda, energy consumption) improve, albeit generally to a limited extent (indie. 0.5-3%). • Fixed costs: In $/year these costs do not vary with plant production, at least within certain plant production rates (typically ± 100,000 t/yr), examples: labour, maintenance materials, administration, other fixed costs. This is the area on which the economy of scale potentially has the largest effect, i.e. a drop in cost per tonne of alumina produced, due to the "dilution" of "fixed" annual expenses by a larger production volume. This applies particularly to labour and otherfixedcosts. If the increase in production capacity includes an increase in the number of trains, this positive effect is dampened because not just the size of the equipment involved increases, but also its number. In addition, the requirements of complex and large alumina refineries may result in disproportional increases of overhead costs. The example provided in Table 1 may illustrate the above. In this example the larger refinery capacity is based on an increase in the number of operating units in several areas, resulting in a limited improvement only of thefixedcosts per tA.
1.4
3.2
Variable Costs, $/tA
85
83
Fixed Costs, $/tA
40
34
Total Operating Cost, $/tA
125
117
* Mt/y = million tonne alumina per annum
3.2.2 Effect on Capital Cost4 Economy of scale has the following main effects5 on Capital Cost: • In general larger size equipment, particularly tanks and vessels, is more cost effective per tonne alumina (tA) produced because larger tanks have a smaller surface area over volume ratio than smaller tanks, hence are cheaper in material cost per m3 stored volume. This effect is sometimes known as the "0.6 factor rule"6, and potentially represents a significant drop in capital cost per tA (note: this factor may be different for different equipment types and unit operations). Although technological improvements have resulted over time in a general increase in equipment size available for most processing equipment (vessels, tanks, pumps, mills,filters,etc), there are physical, technical and/or economic limitations to the size of all equipment. In addition, design considerations may favor in specific cases a large number of small equipment over a small number of large equipment. • Infrastructure (both shared and non-shared) costs are diluted (e.g. piperacks, water supply, power distribution), and spare equipment may be shared in case of a larger production capacity resulting in the construction of more units. Both of these result in a lower capital cost per tA produced. As an illustration: for a refinery with two digestion trains, shared facilities represent indicatively 20-25% of its capital cost (includes raw materials handling, general facilities, shared spares, etc). Here too there are limitations: both with respect to sharing of spare equipment and because capacity increases in infrastructure are required at some stage. The overall effect is a drop in capital cost per tA produced at higher production capacities. A straightforward power factor relationship between these would look like Figure 2.
Table 1 - Effect of Economy of Scale on Opex - 1 Refinery Production Capacity, Mt/y*
The conclusion from the above is that the primary effect of economy of scale on Operating Cost is on fixed costs (expressed per tA), and particularly if a capacity increase is the result of an increase in equipment size rather than equipment number.
Power factor relationship of Greenfield Refinery Capex as function of Design Capacity
Table 2 provides an example in which the capacity increase involved an increase in equipment size rather than the number of operating units, illustrating in that case a more pronounced effect onfixedcosts per tA.
Capex * (Capacity),actor
a
Table 2 - Effect of Economy of Scale on Opex - 2
Design Capacity, kt/year
Refinery Production Capacity, Mt/y*
2.8
:::::.3.3::;::
Variable Costs, $/tA
84
84
Figure 2 - Refinery Capex vs Design Capacity - Power Factor
Fixed Costs, $/tA
50
42
Total Operating Cost, $/tA
134
126
In many cases however plant (and thus project) capacity increases are a combination of increases in equipment size and in equipment numbers (e.g. as a result of an increase in operational
* Mt/y = million tonne alumina per annum
4
Reference [4] provides an overview of Capital Cost A second-order effect is an increased plant on-line time as a result of a plant consisting of more than one train resulting in a slightly lower capex per annual tA. 6 Theoretically the factor is 0.67. 5
3
Reference [3] provides an overview of Operating Cost
6
units / trains). In addition, an increased project scope also adds (at some stage disproportionally) to its complexity. As a result, actual capital cost per tA produced may deviate from a smooth curve as shown in Figure 2. In fact Canbäck and others [5] refer to Bain who found in a study of twenty industries that at the plant level, beyond a minimum optimum scale few additional economies of scale can be exploited. Available information suggests for the alumina industry that with respect to the relationship of refinery capital cost and design capacity, a differentiation can be made in two design capacity ranges as illustrated in Figure 3: • Up to about 1.5 Mt/y: a power factor of -0.7. • Above about 1.5 Mt/y: a power factor of -0.9.
Table 3 - Effect of Capacity on Overall Project Economics Refinery Capacity
Capital Cost*, M$ Mine 115 Refinery 1,635 Infrastructure 500 (railway, port, town) Total Capital Cost*, M$ 2,250 $/AnntA 1,500
Greenfield Refinery Capex as function of Design Capacity
3 Mt/y
L5Mt/y
200 3,000 680 1,293
3,880
Operating Cost, $/tA (incl. Infrastructure opex)
137
125
Sustaining Capital, $/tA
8
8
Economics* (indie.) NPV(8%), M$ IRR, % Payback period, y
-139 7 10.5
369 9 9
* Basis W Europe, Mid 2010 US$ Alumina price at 325 $/tA
#
«ex -s- Capacity""0!7
Table 3 shows that, despite the Refinery capex per annual tA for the two options following the trend illustrated in Figure 3, the overall project economics flip from a significant negative NPV (with IRR 7% and payback period 10.5 years) to a significant positive NPV (with IRR 9% and payback period 8.5 years). A major contributor is the disproportional increase in $/tA of the Infrastructure capex. To underpin that: had the delta in capital cost between the two project options expressed in $/Annual tA remained unchanged from the delta between the two refineries, the economics of the 1.5 Mt/year project (in that case at a total capex of 1,383 $/AnntA) would have looked as follows: NPV(8%) = -12 M$; IRR = 8%; Payback period = 9.5 years. On re-considering the trend shown in Figure 1, the reasoning could be turned around: a disproportionate increase in project scale is required to result in acceptable economics. In a similar context, A. Kjar in his paper presented at the TMS 2010 Annual Meeting [6] discusses in general terms the uncompetitive capital cost of recent Western-developed greenfield alumina projects as a result of (among other reasons) large project size and increased project complexity.
Capex * Capacity 0 β
appro«. 1.5 M t / y
Design Capacity, kt/year
Figure 3 - Refinery Capex vs Design Capacity From Figure 3 it would appear that although further gains in capital cost per tA are possible at design capacities above -1.5 Mt/y, these will be limited. A design capacity of about 1.5 Mt/y for an alumina refinery might perhaps be the "minimum optimum scale" referred to by Canbäck. Note that 1.5 Mt/y is meant to be indicative only. This raises the question how this result can be reconciled with the design capacity of some future planned projects which are well above 1.5 Mt/y (refer Figure 1).
3.3 Consequences The indicated increase in the design / initial capacity of greenfield (bauxite mine and) alumina refinery projects over the past decades has had the following major consequences: • The complexity of these mega projects7 has increased significantly, especially in terms of project planning and management. Significant infrastructural works are often required, involving extensive government involvement, adding to project complexity. • Project capital cost has grown to several billion USD, and project owners reduce risk through projectfinancingand the formation of multi-party joint ventures. This is perfectly reasonable, however it complicates project implementation (e.g. with respect to decision making processes). • Due to thefinancialcommitments involved, globally only a limited number of (very) large companies have the financial and human resources to develop greenfield bauxite & alumina projects.
3.2.3 Infrastructure Costs & Overall Economics The explanation for the above result is that greenfield projects have infrastructural requirements which may include access roads and bridges, a railway line, port facilities, and employee living facilities. In case of extensive infrastructural requirements, the related capital cost is significant and has a disproportional bearing on the economics of a smaller capacity greenfield project. An example may illustrate the above for two greenfield project options at the same location: option 1 at 1.5 Mt/year alumina production design capacity, and option 2 at 3 Mt/year. Assumed infrastructural requirements for this location: • 100 km railway line. • Jetty and wharf, and ship loading/unloading facilities at the alumina export port. • Employee housing and living facilities. Table 3 provides indicative numbers for capital, operating and sustaining capital costs for the two options considered in this example and their economics.
7
7
Typically projects over 1 billion US$.
•
•
Table 4 - Greenfield 1.5 Mt/y Aa Refinery Capital Cost (typ.)
For the same reasons (project scope, complexity), only a limited number of engineering firms have the required engineering, construction and project management skills and experience to successfully implement these projects. Typically a project life of 30+ years is (implicitly) applied to justify the significant investment of a greenfield bauxite & alumina project. Reason: an alumina refinery can operate effectively for decades (refer e.g. Paranam, Gove, Kwinana, QAL). For greenfield bauxite & alumina projects with a captive refinery this means that the bauxite deposit on which a project is based should be able to sustain refining operations for such a period. Therefore only (very) large bauxite deposits are developed, indicatively 200-300 Mt and more.
Cost Item Direct Costs Equipment* Commodities* Total Direct Costs, M$ Indirect Costs Freight EPCM Temp. Construction, start-up, Commissioning, etc Owner's Engineering & Other Costs Total Indirect Costs, M$ Contingency, M$
In summary, worldwide only a small number of companies develop mostly very large greenfield bauxite and alumina projects, which often take a decade and more to develop.
Total Refinery Capital Cost*, M$
231 539
770
78 256 180 190 704 161 1,635
* Incl. steam & power generation, sub stations, residue disposal, water supply, communication & info systems # Incl. concrete, steel, mechanical bulks, piping, wire and cable, etc & Basis W Europe, Mid 2010 US$
3.4 Where from here? With an objective to lower the threshold for the development of bauxite and alumina projects, the question may be asked if the underlying trend, viz. ever-increasing alumina refinery design capacities, is inevitable, or if viable alternatives exists. The basic reason for the trend being economics (refer section 3.2), the question could be reformulated as follows: is it possible to develop smaller greenfield bauxite and alumina projects at acceptable economics? A. Kjar addresses this question and some of the issues discussed above, albeitfroma different perspective, in his earlier mentioned paper. He indicates that as a means to overcome some of these issues, attempts were made by others: 1. To gain improved control over the project execution process; and 2. To increase the level of pre-assembly to reduce total costs of on-site construction labor and low productivity - refer also a paper by R. Valenti and P. Ho [7]. A. Kjar proposes the use of replication of a modern plant design, and small increments of capacity (without quantifying a capacity), in order to quickly and more cost-effectively build a large plant / project. Although A. Kjar's paper has a different angle (viz. building a large plant at lower capital cost), there are overlaps with the subject of the current paper (investigating the possibility to lower the threshold for the development of- smaller - bauxite and alumina projects). To further explore the subject, a more in-depth look at the makeup of a greenfield project's capital cost is required.
4.
1.5 Mt/y
4.1.2 Commodities and Plant Layout Aspects Table 4 illustrates that the Commodities represent a very significant element in the refinery capital cost. Commodity amounts and their related capital costs reflect plant design including plant layout. Current alumina refinery layouts are designed to accommodate additional (future) digestion units (and all of the other required process units - e.g. precipitation, evaporation). The consequence is that plant design is not optimized for its initial production capacity. Plant layout is characterized by an "open architecture", at best compromising between on the one hand the limited layout requirements for the initial / design capacity and on the other hand the more extensive requirements to accommodate future additional process units. And in the worst case consisting of a layout of a large-capacity plant of which part is built, resulting in an inefficient plant layout for the design / initial capacity. In addition, in some cases plant design includes equipment which at some future stage might be used to its full capacity, but operates (well) below design for a considerable part of its lifetime. 4.1.3 Alternative Approach - Dedicated Plant Capacity A. Kjar's proposal to use replication means that a design is developed for a dedicated production capacity. Or putting it differently, this alternative design approach aims at designing an alumina refinery for a dedicated production capacity, i.e. without provisions for future expansions. This approach enables optimizing plant layout for the targeted production capacity, e.g. with respect to positioning similar equipment close to each other, use of common spares, etc. This more "closed" layout architecture results in a more efficient plant layout, reflected for example in the design of main plant piperacks. This is illustrated in Figure 4 which shows the main piperack layout for a typical (current design) 1.5 Mt/year capacity refinery ( i.e. in the expectation that additional production lines in the various areas will be added in the future), and the layout for a dedicated 1.5 Mt/y capacity alumina refinery (same scale). The alternative approach with its more closed layout design impacts positively on commodity volumes: for the same production capacity, commodity volumes for a greenfield plant designed along this alternative approach are similar to that of a brownfield expansion of an existing refinery. This is illustrated in Figure 5 which shows the total length of piping of greenfield and
Capital Cost Make-up
4.1 Refinery Capital Cost 4.1.1 Overview The capital cost of a greenfield alumina refinery may be split up as shown in Table 4. In this table typical numbers are shown for a low-temperature digestion alumina refinery with a 1.5 Mt/y production capacity. Note that actual numbers may deviate significantly as a result of bauxite quality, technology choices, plant location, etc.
8
reflected in lower Commodities costs, resulting in lower Direct Capital Costs, in turn lowering Indirect Capital Costs. The overall effect on the capital cost of a greenfield dedicated lowtemperature digestion alumina refinery of 1.5 Mt/y is illustrated in Table 5 (indicative numbers). As can be seen in this table, the alternative approach improves the total refinery capital cost indicatively by over 10%. In fact the capital cost expressed per annual tonne of alumina capacity is lower than that of the current-design refinery at 3 Mt/y capacity (976 vs 1,000 $/Ann tA - refer Table 3).
brownfield projects as function of plant production capacity, and the requirement of a dedicated plant of 1.5 Mt/y capacity. This approach also stimulates focusing on a "lean" design and exploit any potential overdesign right from start-up (refer the comment made in section 3.1). Layout Main Piperacks of Dedicated 1.5 Mt/y Alumina Refinery
Table 5 - Comparison of Refinery Capital Costs (indie.) 1.5 Mt/y Refinery Capacity
Cost Item
Current-design
Typical Layout Main Piperacks of 1.5 Mt/y Alumina Refinery
Direct Costs Equipment* Commodities Total Direct Costs, M$
231 539
Indirect Costs Freight EPCM Temp. Constr., start-up, Comm. Owner's Eng. & Other Costs Total Indirect Costs, M$
78 256 180 190
Contingency, M$
1
Total Capital Cost*, M$
$/AnntA 1,090
Dedicated
770
224* 459 682
704
69 227 175 168 640
161
142
1,635
976
1,464
* The more efficient plant layout enables slightly lower equipment cost as a result of a more efficient use of common spare equipment Basis W Europe, Mid 2010 US$
#
4.1.5 Compact Refinery - Simple & Limited Scope Along the lines of A. Kjar's paper (although he does not quantify "small increments of capacity"), applying the proposed dedicatedcapacity approach to a compact alumina refinery capacity of 0.4 Mt/y results in a project with a simple and much more limited scope. Available data suggest that as a result some Indirect capital cost items decrease more than proportionately, particularly costs related to temporary construction and start-up support, camp and other construction related items, and owner's costs. Table 6 illustrates the capital cost for a 0.4 Mt/y alumina refinery based on a dedicated design (indicative numbers). The table shows that the capital cost per annual tonne alumina (1,295 $/AnntA) is higher than that of the much larger 1.5 Mt/y dedicated plant (976 $/AnntA - refer Table 5), however is at a level which could result in a project with acceptable economics, provided Infrastructure capital cost is limited (compare with the 1,293 $/AnntA for the overall project capital cost of a 3 Mt/y refinery - see Table 3). Table 6 also shows that the total capital cost is at a level which would enable many more (relatively small) companies to develop such a project without necessarily requiring the formation of multi-party joint ventures, simplifying overall project management and thus enabling to lower costs (effect not included in Table 6). Note that the 0.4 Mt/y refinery production capacity used here is notfixedbut is meant to typify a capacity range of ~ 0.3-0.6 Mt/y. The higher end of this range is limited by the objective to end up with a total project capital cost well below 1 billion US$, the lower end is determined by logistical limitations (e.g. with respect to caustic soda and fuel oil shipments) and may vary for different locations.
h Figure 4 - Main Piperack Layout Comparison 400,000
Total Piping as function of Plant Capacity
Greenfield
Brownfield Plant with dedicated layout design
Plant Capacity, t/y
Figure 5 - Total Piping as function of Plant Capacity 4.1.4 Effect Alternative Approach on Commodities Cost A dedicated greenfield plant design results in lower amounts (in some cases significantly lower amounts) per annual tA produced of commodities such as steel, concrete and piping. This is
9
project with a simple and limited scope, further improving capital cost per tA produced. To ensure acceptable economics, Infrastructure capital cost should be limited. At the same time such a project has few infrastructural requirements , especially if located close to an existing port.
Table 6 - 0.4 Mt/y Refinery Capital Cost (indie.) Cost Item
Direct Costs Equipment Commodities Total Direct Costs, M$ Indirect Costs Freight EPCM Temp. Constr., start-up, Comm. Owner's Eng. & Other Costs Total Indirect Costs, M$
0.4 Mt/y Refinery Capacity (dedicated design) 95 177
28 91 37 33
Contingency, M$ Total Capital Cost*, M$ $/AnntA
* Basis W Europe, Mid 2010 US$
5.2 Main Advantages The main advantages of the new development model are: • Due to the significantly smaller project capital expenditure involved (lower risk), this approach enables the development of bauxite & alumina projects by smaller companies without a need to form multi-party joint ventures, i.e. it increases the number of companies potentially interested in developing bauxite deposits. In other words competition increases, which should result in more efficient use of resources, both in terms of capital resources and in terms of global bauxite deposits. • Due to the decreased complexity of compact alumina refining projects, the number of engineering companies potentially able to develop these projects increases, again resulting in more competition and the potential for a more efficient use of resources. • Small and simple projects carry less risks and require less time to develop, implement and start-up, all of which has a positive impact on economics. • A long term alumina refining project based on the new model requires only a relatively small bauxite deposit (a deposit of ~40 Mt could support a 0.4 Mt/y project for 30 years). This means that worldwide the number of bauxite deposits that lend themselves to development increases. • The new development model may be applied also to the development of part(s) of a large deposit. • This approach may in some cases lower the threshold to increase value creation through alumina refining rather than being limited to bauxite export sales. This is attractive both to host countries and to companies developing potential bauxite & alumina projects. • In some cases, an adapted version of this new development model may enable bauxite deposit development even in locations with little existing infrastructure, albeit at a larger than compact scale (refer e.g. to Table 5 for a dedicated 1.5 Mt/year capacity project).
272
189
!
57 1,295
518
4.2 Infrastructure Capital Cost As mentioned above, in order to realise acceptable economics for a project based on a compact dedicated production capacity, Infrastructure capital cost should be limited. Conversely a project based on a compact plant capacity has very limited infrastructural requirements and has several advantages over a large plant, particularly if the project is located close to an existing port, e.g. it may be allowed closer to residential areas (i.e. is closer to existing infrastructure); the existing infrastructure may be sufficient for a small plant, but not for a big plant; a suitable location for a small residue disposal area is easier to find than for a large one, etc. As outlined in section 5.3 several such locations exist worldwide. 4.3 Refinery Technologies Note that the alternative approach proposed above is independent of the selected refinery technologies, while at the same time stimulating to focus on improvements, e.g. positioning similar equipment close to each other, the use of common spares, etc. 4.4 Replication and Indirect Costs A. Kjar indicates in his paper that the use of replication of a modern design at small capacity increments has as one of its main advantages far lower indirect capital costs, comprising Project management; Procurement; and Technology & EPCM fees. Although no direct quantification is mentioned in the paper, this appears consistent with the results discussed above for a dedicated plant design at a compact production capacity. Some of the replication-related cost savings mentioned by A. Kjar may come on top of the cost improvements indicated in this paper.
5.
5.3 Possible Locations Following are some examples of bauxite deposits that may lend themselves to development via the proposed alternative approach (between brackets the potential alumina export port): • Haden, Queensland, Australia (Brisbane). • Bindoon, Western Australia (Fremantle). • El Palmar, Venezuela (Ciudad Guayana). • Trelawny, Jamaica (Discovery Bay). • Kibi, Ghana (Tema). The above list is not exhaustive and meant to be illustrative only. In addition some bauxite deposits which in view of their size could support the current development approach with largecapacity alumina refining projects, may also lend themselves to stage-wise development through the proposed alternative
New Bauxite Deposit Development Model
5.1 New Development Model The bauxite deposit development model proposed in this paper as detailed above is based on the development of a dedicated compact alumina refinery in the range ~ 0.3-0.6 Mt/year. The dedicated refinery design has no provisions for future expansions, enabling optimizing plant layout and resulting in lower capital cost per tonne of alumina (tA) produced compared with current plant design. The compact capacity results in a
10
approach. In this case these deposits would be able to support several (smaller) greenfield bauxite and alumina projects as outlined in the last bullet point of section 5.2 above. Example: some of the Eastern Ghats deposits in Orissa and Andhra Pradesh, India, e.g. the Kutrumali deposit (with Visakhapatnam as potential alumina export port).
6.
References
1. R. den Hond, "Technology Choices for Greenfield Alumina Plants" (paper presented at ICSOB A 2008, Bhubaneswar), pp 267-270. 2. P.J.C, ter Weer, "Greenfield Dilemma - Innovation Challenges" (paper presented at Light Metals 2005, San Francisco, California), pp 17-22. 3. P.J.C, ter Weer, "Operating Cost - Issues and Opportunities" (paper presented at Light Metals 2006, San Antonio, Texas), pp 109-114. 4. P.J.C. ter Weer, "Capital Cost: To Be or Not To Be" (paper presented at Light Metals 2007, Orlando, Florida), pp 43-48. 5. S. Canbäck, P. Samouel, and D. Price, "Do Diseconomies of Scale Impact Firm Size and Performance - A Theoretical and Empirical Overview", Journal of Managerial Economics, 2006, Vol. 4, No. l,pp 27-70. 6. Anthony Kjar, "A Case for Replication of Alumina Plants" (paper presented at Light Metals 2010, Seattle, Washington), pp 183-190. 7. R. Valenti and P. Ho, "Rio Tinto Alcan Gove G3 Experience on Pre-Assembled Modules" (paper presented at the Alumina Quality Workshop 2008, Darwin), pp 1-5. For further information, please contact P.J.C, ter Weer at
[email protected] or visit www.twsservices.eu.
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
STUDY ON THE CHARACTERIZATION OF MARGINAL BAUXITE FROM PARA/BRAZIL Fernanda A.N.G. Silva1'2, Joäo A. Sampaio2, Francisco M. S. Garrido1, Marta. E. Medeiros1 universidade Federal do Rio de Janeiro, Instituto de Quimica, Avenida Athos da Silveira Ramos, 149, Cidade Universitaria; Rio de Janeiro, RJ, 21941-909, Brasil. 2 Centro de Tecnologia Mineral / CETEM-MCT, Avenida Pedro Calmon, 900, Cidade Universitaria; Rio de Janeiro, RJ, 21941-908, Brasil. Keywords: Marginal Bauxite, Mineralogical and Chemical Characterization In this context, the aim of this work was to: ore dress, provide the chemical, structural and mineralogical characterization of crystallized amorphous bauxite (CAB) and compare its behavior and characteristic with the bauxite processed by the Bayer process, such as crystallized bauxite (CB). The techniques applied in this study were: chemical analysis by potentiometric titration and flame atomic absorption spectroscopy, X-ray fluorescence, Xray diffraction, infrared spectroscopy, scanning electron microscopy and thermal analysis.
Abstract Bauxite from Para is divided into five different layers. However, only one is processed. The crystallized-amorphous (CAB) phase is considered a marginal bauxite because it presents a high quantity of Si02reactive and its use depends on special technologies. CAB was characterized and the results were compared with the bauxite used nowadays in the alumina plant. Characterization was performed by XRD, IR, XRF, chemical analysis, TGA and SEM. XRD determined the mineral content: such bauxite is gibbsitic and has been associated with kaolinite and hematite. IR data supported the XRD results. XRF was used to determined the sample's chemical composition. The chemical content of Al203avaiiabie and Si02reactive w a s determined by potentiometric titration and FAAS. The results found for the Bayer process sample were 41.7% and 7.1%, respectively. TGA observed the bauxite decomposition and SEM supplied chemical and thermal analysis. Thus, based on stoichiometric relations of the bauxite components decomposition, it was possible to confirm the presence of the following phases: gibbsite and kaolinite.
I
I
j
j Clay layer (Laterite Soil)
Nodular bauxite
Introduction Crystallized Nodular Bauxite
Bauxites are usually considered to be of two major types: (1) lateritic (sometimes called equatorial) and (2) karst, both being weathered products from the underlying parent rock: lateritic bauxites. [1] Lateritic bauxites, formed in equatorial climates, comprise 90% of the world's exploitable bauxite reserves [1]. The weathering process has resulted in a typical profile in which the valuable aluminous material lies atop of an aluminosilicate base (often clay) and has formed from it through the leaching of silica. The main silicate mineral is kaolinite which is often associated with goethite as the iron mineral. Aluminous minerals are predominately gibbsite and to a lesser extent boehmite [1].
I
Crystallized Bauxite
I
Crystallized Amorphous Bauxite
Amorphous Bauxite
Bauxite from NE Para is divided into five layers: (1) nodular (NB), (2) crystallized nodular (CNB), (3) crystallized (CB), (4) crystallized amorphous (CAB) and (5) amorphous (AB) (Figure 1) [3]. The NB, CNB, CAB and AB layers present a high quantity of iron minerals, reactive Si0 2 and others impurities. [2-3]. These bauxites are considered marginal and their ore dressing requires special technologies arisingfromtheir content of impurities.
Figure 1. Bauxite geological profile from Northern Parâ/Brazil. Materials and Methods
Bauxite mining methods vary according to the nature of the mineralized field bodies, but in most of the cases a strip or block of bauxite is exposed and surface-mined [3]. Although the mining process is selective and only the layer of crystallized bauxite is removed, for a bauxite to be considered economically useful for the Bayer process, the available A1203 content should be between 45-55% whilst the reactive Si0 2 content should be between 4-6%. [4-6].
1 - Sample Preparation The bauxite ore was crushed in a jaw crusher and the product was classified with the use of a sieve (1.65 mm). The coarse fraction was crushed (crushing rolls) in a closed circuit. The -1.65 mm fraction was classified to remove the -37 μιη particles (sludge). The +37 urn fraction was homogenized with the product of the crushing rolls and two samples of 20 kg and 5 kg, respectively, were separated for wet granulometrie analysis. For grinding,
13
samples of 20 kg were homogenized and separated into piles of lkg. Figure 2 shows the block diagram used in the bauxite beneficiation [7].
2 - Chemical and Mineralogical Characterization The bauxite ore was submitted to chemical and mineralogical analysis with the use of: X-ray diffraction, infrared spectroscopy and X-ray fluorescence.
Samples (1 kg) of the final product of the preparation stages were wet ground in a stainless steel mill bar with 10 stainless steel bars of 20 mm diameter. The slurry in water was prepared using a bauxite solid concentration of 1000 g L"1. Grinding time varied from 0 to 30 minutes. A wet granulometrie analysis was carried out after each grinding to adjust the sample to the necessary conditions for the Bayer process [7].
During the ore dressing tests, four different samples were obtained: a work sample (WS) obtained after crushing, two work sample fractions (833 and 208 μπι) and the sludge (-37 μιη fraction obtained after desliming of the -1.65 mm fraction). These samples were submitted to the same characterization techniques of the crude sample.
The granulometrie analysis was carried out with samples of lkg, according to the damp method [8]. A vibratory sifter (684.5 rpm), equipped with a group of sieves with openings from 3.350 mm to 37 μπι, was used according to the Tyler series. All the granulometrie analysis fractions obtained in these tests were dried (100°C) and weighed.
The BP sample (the sample ground to the Bayer process requirement) was characterized with the same techniques used in the crude sample followed by potentiometric titration, flame atomic absorption spectrometry, thermal analysis and scanning electron microscopy. 2.1 - X-Rav Diffraction (XRD)
Feed Aliquot 80 kg (Archive)
«-
1 Homognization
1
Pile
1
Samples were examined by XRD in a Bruker-AXS D5005 diffractometer, with Co Ka (35 kV/40 mA) radiation, 0.02° goniometer velocity and 2Θ by path with 1 s by path counting time and data collected from 5 to 80° 2Θ.
1
2.2- Infrared Spectroscopy (IR)
Crushing (Jaw Crusher)
1
I
Fraction +1.65 mm |
Fraction -1.65 mm
2.3 - X-Rav Fluorescence (XRF)
1
*
Granulometrie -37 μπι «Analysis (Sludge) (3.32 mm to 37 μπι)
Crushing (Rolls Crusher)
1
I
Fraction f -1.65 mm |
Fraction +37 μπι
1
Infrared spectra (FTIR) was performed in a Nicolet Magna 750 Fourier transform spectrometer, from 4000 to 400 cm"1, with resolution of 4 cm'1, with sample mounting using KBr discs.
1 Homogenization
The samples were melted with lithium tetraborate at 1100°C in the proportion of 1:6 sample/fluxing agent. The melted bead was analyzed in an energy dispersive X-ray fluorescence spectrometer (Bruker-AXS model S4-Explorer), equipped with Rh tube. To obtain the semiquantitative chemical analysis, the sample spectrum was evaluated by a Spectra plus v. 1.6 software, in the standardless method mode, without a specific calibration curve.
+ r~ ON 3 |3
1
2.4 - Thermal Analysis (DTA/TGA) Thermal analysis was carried out in a Shimadzu TA-50WSI equipment (thermogravimetric analysis), DTA-50 (differential thermal analysis) in a heating grade of 10°C/min, from room temperature to 1200°C under a flow of air.
Pile
1
1 ws
1
1 (Work Sample) |
1
2.5 - Determination of Available Alumina and Reactive Silica
▼
Granulometrie Analysis (3.32 mm to 37 μπι)
[►
WS833μm WS208μm
The method to determine the amount of available alumina (the amount that will be refined to obtain A1203 in the Bayer process) and the amount of reactive silica (kaolinite), consisted of bauxite digestion in alkaline medium (NaOH) under controlled pressure and temperature, simulating the Bayer process. For the determination of available alumina, a sodium gluconate solution was added to the supernatant to form an aluminum hydroxide gluconate complex. The excess of NaOH, used in the digestion step, was neutralized with the addition of an HC1 solution. Then, a KF solution was added and back titration was carried out. Afterwards, an excess of HC1 in the standardized solution was
1
Grinding - 30 min Bayer Pulp (PB)
Figure 2. Block diagram of the stages used in the beneficiation of bauxite samples.
14
vibration [9, 12,13] and the bands near 3695 are attributed to OH stretching of kaolinite [14].
titrated with a NaOH standardized solution. The solid phase, resulting from ore digestion stage, was dissolved in an HNO3 solution [5]. The concentration of reactive silica was determined by flame atomic absorption spectrometry (FAAS).
In order to compare the mineralogical phases between CAB and CB, it is possible to observe that in CB there are more phases related to iron minerals, such as aluminum-goethite (oc(Fe,Al)OOH), goethite (FeOOH) and hematite (aFe203) [7] than in CAB. These results are associated with the distribution of the layers in the bauxite's profile since the CB layer is near the NB layer, which is composed essentially of iron minerals. For the dissolution of gibbsitic bauxite, the temperature in the Bayer process must be between 140-150°C. At this temperature the iron minerals are inert and do not react during the process.
2.6- Flame Atomic Absorption Spectrometry (FAAS). Chemical analysis was performed by flame atomic absorption spectrometry in an AA6 Varian equipment with 248.3 nm wave number, 0.5 nm slit and with air/acetylene. 2.7 - Scanning Electron Microscopy (SEM) Bauxite morphology was determined by scanning electron microscopy in high vacuum SEM (Leica/F440). The sample was embedded in resin epoxy and polished. The resin was covered again by vaporized carbon to be used as a conductor. Results and Discussion
3526
In this work, the fraction of crystallized amorphous bauxite (CAB) was characterized and the results were compared with crystallized bauxite (CB). The mineralogical phases that compose the crystallized bauxite (CB) were determined by X-ray diffraction (XRD), Figure 3. Therefore, this bauxite is essentially gibbsitic and is associated with kaolinite (Al4(SÌ4Oi0)(OH)8 and hematite (aFe203).
c
o CO
-Ω
<
9P 4000
3500
3000
— 1 — 2500
2000
1500
1000
500
Wavenumber (cm" ) Figure 4. Infrared spectra of CAB bauxite In order to determine the sample's chemical composition and how its content varied after ore dressing, XRF analysis was carried out, Table I. Table I. Chemical composition of the bauxite ore
^JuJ - 1
-
10
-r— 20
Samples 30
—r—
—1—
Bauxite Ore WS WS 833 μπι WS 208 μπι Sludge
50
40
2Θ
OGibbsite
■ Kaolinite
· Hematite
Figure 3. XRD of crystallized amorphous bauxite (CAB)
Samples
Infrared spectroscopy was used as a support to the XRD results. The infrared spectra analysis of crystallized amorphous bauxite (CAB), Figure 4, shows the bands related to the stretching vibration of Al-O-H groups observed at around 3620 cm"1. This band is related to gibbsite and kaolinite minerals [9]. The bands around 3526, 3460 and 3390 cm'1 are assigned to O-H stretching of gibbsite. The band at 914 cm'1 is due to Al-O-H group deformation vibration [10], the bands observed around 1.1001.030 cm'1 are related to the deformation vibrations of OH and are characteristic a gibbsite bauxite ore [11]. Bands appearing at around 450, 540, 660, 1100 cm'1 are assigned to the Si-O
Bauxite Ore WS WS 833 μπι WS 208 μπι Sludge
A1 2 0 3 46.06 53.10 56.54 50.70 40.94 Ti0 2 1.84 1.79 1.37 3.92 2.03
Com ponents (% in mass) MnO Si0 2 Fe 2 0 3 14.64 0.03 9.11 0.04 9.08 8.77 5.86 2.25 10.77 0.15 11.70 35.70 12.95 Com ponents (% in mass) PF Zr0 2 WO3 0.22 28.10 0.25 26.97 0.36 0.36 27.26 0.77 21.99 0.28 7.97 -
| SO3
-
|
0.13
| Total 100.00 100.00 100.00 100.00 100.00
WS: sample obtained after crushing (final product of ore dressing); Sludge: -37 μηι fraction obtained from the removal of the -1.65 mm fraction
The analysis of the results in Table 1, indicated a higher content of Si0 2 in the finer fractions than in the coarse fractions, whereas
15
the opposite behavior was observed in the content of AI2O3. The same behavior could be observed in the CB bauxite. Nevertheless, the values of the chemical content are different because CAB has a higher amount of kaolinite than CB, according to its geological profile [7]. However, in order to evaluate such behavior in relation to the contents of available alumina and reactive silica responsible for the consumption of NaOH and the formation of dessilication product (DSP) by the Bayer process, a chemical analysis by potentiometric titration and flame atomic absorption spectroscopy was carried out. These analyses were done after adjusting the CAB sample to the Bayer process.
i. The first event was related to Al(OH)3 dehydroxylation, with the formation of a AIO(OH) and χ-Α1203 mixture [7].
The grinding study showed that after a period of 20 min., the bauxite particle size distribution was similar to the optimum particle size for the Bayer process. The grinding curve, Figure 5, showed that around 90% of the particles had grain size lower than 0.208 mm and 50% under 0.043 mm.
Table II contains the data related to the samples' mass loss obtained to the Bayer process sample (BP) to both bauxites. Results analysis showed a bigger mass loss related to the first and third events. These results can be explained based on explanations i and iii [7].
ii. The second one corresponded to aFe203.nH20 dehydroxylation, since among the mineralogical phases identified through XRD, this was the only phase where it could be decomposed at such temperature [7]. iii. The third event was related to a AIO(OH) dehydroxylation combination formed in the first event and (Al4(SÌ4O10)(OH)8) dehydroxylation [7].
In order to evaluate Table II, it can be observed that the CAB bauxite had a larger mass loss related to the third event (kaolinite) when compared with the CB bauxite. An opposite behavior was observed for the mass loss related to alumina (first and third events), in which CB presented a larger mass loss than CAB. These results were in agreement with those obtained through XRD, IR and XRF, which showed that the DTA/TGA analysis could be used to observe the relationship between gibbsite and kaolinite in bauxites. 295 SC
15 20 grinding time (min)
Figure 5. Grinding curve of CAB bauxite. The time used to mill CAB was shorter than the time used to mill CB, at 20 and 30 min., respectively [5]. This result was in agreement with the results obtained from XRF, since this analysis showed that in CAB there was a larger amount of finer particles when compared with the CB. Considering that the finer fractions were composed essentially of kaolinite, according to the XRD and IR results, the time of mill was short because this mineral presented particle size smaller than 43 μπι.
365 QC —1— 200
508 QC
— i
■
400
1
600
«—
Temperature (°C)
After adjusting the CAB sample to the Bayer process requirement (20 minutes of grinding), this bauxite was characterized by thermal and chemical analysis. The chemical analysis used the following techniques: potentiometric titration and flame atomic absorption spectroscopy.
- 1 — 800
1000
1200
Figure 6. DTG/TGA of Bauxite Ore. Table II. Mass loss related to thermal events of CAB and CB samples.
Thermal analysis of the bauxite varied according to the sample's origin, impurities and crystalinity. Thus, thermal analysis (DTA/TGA) was carried out in this work as a complementary technique to XRD, IR, XRF and chemical analysis, in order to obtain a complete characterization as well as a better knowledge of the rock that composes the geological profile of NE Para [4, 5, 7].
Samples CABBP CBBP
1st
2nd
Al(OH)3 18.4 21.0
Fe 2 0 3 *
2.0 1.2
BP: sample ground to Bayer process. *To CABA aFe 2 0 3 .nH 2 0 *To CB (a(Fe,Al)OOH)
According to the results obtained through DTA/TGA for the CAB sample to the Bayer process (Figure 6) three events were observed:
16
Mass Loss (%) Events 3 rd | (AlOOH) Al4[Si4O10](OH)8 1
3.7 4.2
2.1 1.8
I
Thermal analysis results were in accordance with those obtained through the chemical analysis for available alumina and reactive silica, 41.7% and 7.1%, respectively (Table III). The ratio between A1 2 0 3 available and Si0 2 reactivem the CB bauxite was 8:1. When this ratio is below the conditions recommended by the Bayer process, 10:1, it causes a reagents and products loss during the process of bauxite dissolution. Therefore, CAB cannot be used in the Bayer process because its ratio is 6:1 and it provides a high loss to the process.
AI
7*
Table III. Chemical contents of A1 2 0 3 available and Si0 2 reactive of CAB and CB samples by potenciometric titration and atomic absorption.
Samples CABBP CBBP
A1 2 0 3
Al· -
Chemical Analysis (% in mass) available
(s<0.3%) 41.7 47.5
BP: sample ground to Bayer process.
(A) 1 1
I/7
^ 1 U 2 reactive
(s<0.1%)
7.1 5.9
>mJl·—«m
i è'
i ft a , ; ;
>!>>>··:
r—
s
■ T-
1
(B)
Scanning electron microscopy (SEM) was used to support the aforementioned analyses. Figure 7 illustrates the SEM of the Bayer process sample. This sample was divided into two different particles: pure gibbsite with smooth surface (A) and gibbsite associated with kaolinite particles with rugous surface (B).
Figure 8. Energy Dispersive Spectrum to the CABA BP sample: (A) pure gibbsite and (B) gibbsite associated with kaolinite. After the ore dressing stages, it was possible to observe that the prepared sample for the Bayer process had kaolinite particles associated with gibbsite through a physico-chemical interaction as well as in the finest fractions that comprised the sample. Conclusion The ore dressing stages aimed at the adaptation and characterization of the fractions that compose the bauxite sample of NE Para that is not currently used in the Bayer process. According to the results, the sample was adjusted to the Bayer process (BP) requirements after crushing stages, screening, removal of fine fractions and 20 minutes of grinding. XRD indicates the mineralogical composition of this bauxite. Therefore, the studied bauxite is essentially gibbsitic and is associated with the minerals kaolinite and hematite. XRF provided information related to the chemical composition of the sample, but did not inform the chemical content of A1 2 0 3 available and Si0 2 reactive making it necessary a chemical analysis by potentiometric titration and flame atomic absorption spectroscopy.
Figure 7. Scanning electron microscopy imaging of the CAB BP sample: (A) pure gibbsite and (B) gibbsite associated with kaolinite.
Thermal analysis helped us confirm the events related to dehydroxylation of the components presented in the sample: gibbsite, hematite and kaolinite.
With the use of EDS (Energy Dispersive Spectrometer), Figure 8A and 8B to gibbsite particle and gibbsite with kaolinite, respectively, it was possible to observe the semiquantitative composition of those minerals.
The contents of A1 2 0 3 available and Si0 2 reactive m the sample were 41.7 and 7.1%, respectively. The ratio between A1 2 0 3 available and Si0 2 reactive was below the quantity recommended for the Bayer process (10:1). Thus, the 6:1 rate caused a lot of loss in the process. This work has allowed us to understand why CAB bauxite is not used in the Bayer process. The reason is related to the high amount of reactive silica. This problem is likely to be associated with the presence of gibbsite particles associated with kaolinite
17
12. E. Mendelovici.2001. "Selective mechanochemical reactions on dry grinding structurally different silicates". Journal of Materials Science Letters, 20, 81-83.
particles, possibly, by a chemical surface interaction, even after the ore dressing steps. Acknowledgments
13. C.A. Azevedo, F.M.S. Garrido, M.E. Medeiros, 2006. "The effect of mechanochemical activation on the reactivity in the MgO-Al203-Si02 system", Journal of Thermal Analysis and Calorimetry, 83, 649-655.
We would like to thank CAPES and CNPq for their financial support. References
14. F.A.N.G. Silva "Estudo de Caracterizaçao e Beneficiamento do Caulim da Regiäo Borborema-Seridó (RN) (M.Sc. dissertation, PEMM/COPPE Universidade Federal do Rio de Janeiro, 2007), 51-52.
1. Smith P., 2009. The processing of high silica bauxites — Review of existing and potential processes. Hydrometallurgy, 98, 162-176. 2. B. Kotschoubey et al., "Caracterizaçao e Gênese dos Depósitos de bauxita da Provincia Bauxitifera de Paragominas, Noroeste da bacia do Grajau, Nordeste do Parâ/Oeste do Maranhäo." Caracterizaçao de Depósitos Minerais em Distritos Mineiros da Amazônia, eds. O.J. Marini, E.T. Queiroz, B.W. Ramos, (Brasilia, DF, Brasil: VGArte, 2005), Cap. 11, 691-782. 3. C. N. Barbato, S. C. A Franca, M. N Souza, "Study on the Rheological Behavior of Crystallized and CrystallizedAmorphous Bauxites", Light Metals, (2010), 87-91. 4. F.A.N.G. Silva, "Relatório de Estâgio realizado na Alunorte no periodo de 2/8/2007 A 21/9/2007" (Relatório RE2007-004-00, Centro de Tecnologia Mineral, Rio de Janeiro, RJ, 2007). 5. F.A.N.G. Silva, M. E. Medeiros, J. A. Sampaio, R. D. Santos, M. C. Carneiro, L. S. Costa, F. M. S. Garrido, "Technological Characterization of Bauxite from Parâ-Brazil", Light Metals, (2009), 139-144. 6. R. Gandi, M. Weston, M. Talavera, G.P. Brittes, E. Barbosa, "Design and Operation of the World's First Long Distance Bauxite Slurry Pipeline" Light Metals, (2008), 95-100. 7. F.A.N.G. Silva, R. D. Santos, J. A. Sampaio, F. M. S. Garrido, M. E. Medeiros, "Study on Ore Dressing and Characterization of Different Granulometrie Fractions that Compound Bauxite From Parâ/Brazil", Light Metals, (2010), 69-74. 8. J.A, Sampaio and F.A.N.G Silva, "Anâlise Granulométrica Por Peneiramento", Tratamento De Minérios Praticas Laboratoriais. ed J.A. Sampaio, S.C. Franca, and P.F.A. Braga, (Rio de Janeiro, RJ, Brasil: Centro de Tecnologia Minerai, 2007), Cap3, 55-74. 9. H. W. Marel, H. Beutelspacher, Atlas of Infrared Spectroscopy of Clay Minerals and their Admixtures, 1 ed., Amsterdam, Elsevier Scientific Publishing Company, 1976. 10. E. Mendelovici, "Selective mechanochemical reactions on dry grinding structurally different silicates", Journal of Materials Science Letters,(2001), 20, 81-83. 11. L. G. Shumskaya, "Directional Changes in the Properties of Aluminum Hydroxide-Oxides for Increase in Bauxite Reactivity in Hidrometallurgical Processing". Journal of Mining Science,(2002), 38, n° 03, 299-304.
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Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
RESOURCE UTILIZATION OF HIGH-SULFUR BAUXITE OF LOW-MEDIAN GRADE IN CHONGQING CHINA Jianguo Yin, Wentang Xia, Mingrong Han Chongqing University of Science & Technology, Chongqing 401331, PR China Keywords: High-sulfur bauxite, Resource utilization, Low-median grade, Desulfurization Abstract
It is generally thought that bauxite ores containing sulfur over 0.7% cannot be used directly to produce alumina. So high-sulfur bauxite ore has been basically abandoned, or has had a little utilization mixed into other resources until now. If a process using high-sulfur bauxite to produce a suitable quality alumina can be successfully developed, the resource life of Chinese bauxite ores will be prolonged by about 10 years. So for resource utilization of high-sulfur bauxite, it is very important not only for the development of the regional economy in Chongqing but also for the China alumina industry.
Resource utilization of high-sulfur bauxite is one of the technical problems for alumina refineries in China. There are rich highsulfur bauxite ores of low-median grade in Chongqing China, which haven't been utilized until now. Sulfur in bauxite will cause many negative effects on alumina production. There are still some disadvantages for current desulfurization technologies and it is difficult to make a breakthrough progress only from the view of application of these desulfurization technologies. So it is necessary to undertake fundamental theoretical studies on the occurrence of sulfur in high-sulfur bauxites, the reaction behavior and the impact of sulfur in the production process. Combining theoretical research with experiments, it is possible to develop a feasible desulfurization process. This will provide technical support and theoretical guidance for the utilization of high-sulfur bauxites in Chongqing, and is also advantageous for prolonging bauxite resource security for the Chinese alumina industry.
Negative effect of sulfur on the alumina production Sulfur in the bauxite ores will cause a lot negative effects on the alumina production. It will increase the soda consumption and decrease the digestion ratio of alumina. Sulfur in the form of the anions S2" and S2O32" in the liquors will erode the steel and iron of the pipes and equipment, and increase the iron content in the liquors and the products, which will cause negative effects on the quality of alumina product and in the subsequent aluminum electrolysis. When sulfur accumulates to a certain concentration, it will be precipitated in certain forms and cause scaling of the spent liquor evaporator and slurry digester, which will decrease the heat transfer coefficient of these devices or even make them not work (3). So it is necessary to desulfurize in the process of alumina production from such ores.
Introduction The alumina industry in China has developed quickly in recent years. Output of alumina in the year of 2009 was up to 24 million tons, and production capacity reached 33 million tons. It may be above 40 million tons in the following 2 or 3 years. This brings serious supply problems for bauxite resources as the production capacity of alumina expands so rapidly in China. From the view of the world, the service life of bauxite ore reserves is more than 100 years based on an annual output of 80 million tons of alumina and annual consumption of 2.5 billion tons bauxite. But from the view of China, the service life is no more than 15 years, based on an annual output of 20 million tons of alumina, even if the prospective reserves are considered (1). For Chinese bauxite ores, quality will decrease rapidly or even to exhaustion in the traditional mining areas. So it is necessary for the Chinese alumina industry to cope with this serious resource problem in order to ensure the alumina industry develops continuously, stably and healthily.
Previous desulfurization technologies Desulfurization technologies for high-sulfur bauxite ores have been studied a lot at home and abroad, and can be divided into flotation desulfurization, roasting desulfurization and wet desulfurization. Classification of desulfurization technologies is shown is Fig. 1. Desulfurization Technologies Flotation
There are some bauxite resources which are difficult to be utilized until now in the main mining areas, which include high-sulfur bauxite ores. There are 11% or so of these high-sulfur bauxite ores and about 5.6 billion tons reserves in China. With the further survey of bauxite ore under the coal seams, the portion of the ore may rise. There are rich high-sulfur reserves in Nanchuan Chongqing China, the proven reserves are more than one billion tons and the prospective reserves are about 3 billion tons. Bauxite ores of Chongqing have their own characteristics, which belong to high-sulfur diaspore bauxite of low-median grade and very different from those of Guizhou. The average quality index of the bauxite ores are as follows: A1 2 0 3 44.26%-71.39%, Si0 2 9.54%19.16%, Fe 2 0 3 0.80%-20.27%, S 0.16%-2.19%, A / S (the mass ratio of alumina to silica) 3.68-7.54 (2).
Roasting
Wet II
Direct Flotation Reverse Flotation
Muffle Furnace I
Precipitation
Rotary iTube Furnacel
Lime
Fluidized Bed
Barium Salts
Zinc Species Fig.l Classification of desulfurization technologies
19
1
Wet Oxidation
hydrophobic hematite, which improves the settling performance. But if the roasted ore is too loose, it will cause red mud fines and seriously decrease its settling performance. Compared to that roasted in a muffle furnace, the ore from the rotary tube furnace has better settling performance (18). High-sulfur bauxite ore in Henan province was investigated by Xiaolian Hu et al. with lime roasting in a muffle furnace. It shows that sulfur in the form of sulfide decreases after roasting. Lime roasting has better result for sulfur-fixation, and it decreases the content of sulfur dioxide emitted into the air. In the roasting process, pyrite is converted to hematite. After treatment, the bauxite ore decreases the sulfide content of the digestion liquor and improves its digestion performance. The optimum roasting temperature is 600. When adding 1% lime and roasting, the comparative digestion ratio is 95.35 % and sulfide in the liquor is 0.16 g / L (19). In the process of lime sintering, sulfur is removed by adding anthracite. But the desulfurizing efficiency is low (20). Improving the adding way of coal can increase desulfurizing efficiency (21).
Flotation desulfurization Through differences in the physicochemical properties of mineral surfaces, the minerals can be beneficiated, and the differences can be enlarged by adding flotation reagents and manually modified, which makes the flotation method very adaptable. Moreover the flotation method has high separating efficiency, so it is one of the most extensively adopted beneficiation methods. Sulfur in the bauxite ore mainly exist in the form of pyrite. Pyrite can be floated easily by xanthate collectors, but ores containing aluminum are hydrophilic and not easy to collect. So xanthate collectors can realize the separation pyrite from the ores easily and then decreases the sulfur content of the ores. Flotation desulfurization was firstly used by the Ural Engineering Institute of the former Soviet Union. Technologies of one-roughing, twocleaning and two-scavenging stages were adopted to treat bauxite ore containing 2% sulfur and the concentrate contains sulfur less than 0.41%. But this flotation desulfurization process is too complicated. Flotation desulfurization by potential control was reported, but it was too hard to realize on production scale (3).
It can be concluded that roasting desulfurization is relatively simpler, however, it increases the energy consumption and pollutes the environment.
Xiaomin Wang et al. studied desulfurization by reversed flotation using ethyl xanthogenate, and attainted optimum conditions as follows: pH value 12, flotation reagent dosage 0.4 kg /1, agitation time 15 minutes, the content of ore slurry 10% and particle size of ore less than 0.09 mm. Under these conditions, the sulfur content of bauxite ore can be reduced from 2.08% to 0.65%, and recovery ratio of alumina reaches 91.46%, which is suitable for alumina production (4). Using butyl xanthogenate as a reversed flotation regent, good desulfurization results can also be attained (5). Using soda as pH modifier, sodium hexametahposphate as an inhibitor, sodium sulfide and copper sulfate as activators, ethyl xanthogenates and butyl xanthogenate as collectors, Wenmi Chen treated high-sulfur bauxite in Guizhou China by reversed flotation. For the concentrate, the sulfur content is 0.44% and the recovery ratio of alumina is 96% (6). Adopting a one-stage flotation process to treat high-sulfur bauxite ore in Guizhou, the concentrate contains 0.15% sulfur and alumina losses are only 6.3% (7).
Wet desulfurization 1. Desulfurization by wet oxidation Desulfurization by wet oxidation involves pumping in air to oxidize sulfur into sodium sulfate which precipitates in the evaporation process of the spent liquor (22). But increasing the sodium sulfate in the spent liquors will lead to the double salt of burkeite precipitating in the evaporation process, which accelerates the evaporator scaling and effects its operation. So the formation of the thiosulfate anion needs to be prevented, lest it hastens the corrosion of the equipment. In addition, desulfurization by wet oxidation is a little dangerous. 2. Desulfurization by precipitation Lime is commonly used as a desulfurization reagent by precipitation. Jun Lan investigated desulfurization by precipitation for a high-sulfur bauxite ore of high grade in Maochang Guizhou China. The effects of lime dosage, alkali concentration, digestion temperature and time on the desulfurization efficiency were investigated. The results show that optimum conditions are as follows: lime dosage 10%, digestion temperature 245, desulfurization time 70 minutes and alkali concentration 240 g / L. Under these conditions, the digestion ratio of sulfur is only 31.3% and that of alumina is over 81% (23). Adopting the same technology, Runde He thought the optimum conditions are as follows: lime dosage 12%, digestion temperature 260, desulfurization time 50 minutes and alkali concentration 195 g / L. Under these conditions, the digestion ratio of sulfur is only 13% and that of alumina is over 88% (24, 25).
From the view of the technology itself flotation desulfurization is matured, but from the view of alumina production it brings some negative effects as the concentrate contains water and flotation reagents. At the same time, the process needs to treat large amount of bauxite ores, waste water and tails, and add a lot of reagents and fresh water (8). Roasting desulfurization A lot of research has focused on roasting desulfurization (9-13). Muffle furnaces, rotary tube furnaces and fluidized beds were used to pretreat high-sulfur bauxite ore. Optimum roasting conditions for muffle furnace and rotary tube furnace are a temperature of 750 centigrade and time of 30 minutes. Those for fluidized bed are temperature 800 and time 10 minutes. The digestion performance of the pretreated ore from the fluidized bed is the best, with the digestion ratio of alumina reaching 93.7% under the digestion conditions of temperature 220, molecular ratio of the proportioning 1.3 and caustic concentration 220 g / L (1417). Sulfur is removed in the form of gas, and the roasted bauxite ore has better digestion performance. It is found that roasting makes the red mud loose and porous, and converts goethite to
Barium salts are another desulfurization reagents utilizing precipitation. Barium oxide and barium hydroxide were chosen to purify industrial sodium aluminate liquors (26). Desulfurization efficiency of barium salts is satisfying, and can reach 99%. But when the contents of silicate and carbonate in the liquors are high, the consumptions of barium salts will increase. As these two salts are expensive, it will increase desulfurization expense. Adopting cheap bauxite ore of high grade, and barium carbonate to generate a desulfurization reagent of barium aluminate, the desulfurizing result is similar to those of the two barium salts, but the
20
desulfurization expense can be greatly reduced (27, 28). Desulfurization ratio reaches 94.5 % when barium aluminate is adopted to purify washing water of red mud (29). It is thought that washing water of red mud is suitable to be desulfurized and purified (30). Fanghai Lu thought that spent liquors and the first washing water of aluminum hydroxide are suitable to be desulfurized (31).
Conclusion There are rich high-sulfur bauxite ores of low-median grade in Chongqing China, which are characteristic of the region and have not been utilized till now. It is necessary to undertake fundamental theoretical studies on the occurrence of sulfur in high-sulfur bauxites of low-median grade, occurrence, reaction behavior and the impact of sulfur in the production process. Combining theoretical research with experiments, it is possible to develop a feasible desulfurization process. This will provide technical support and theoretical guidance for the utilization of high-sulfur bauxites in Chongqing, and is also advantageous for prolonging bauxite resource security for the Chinese alumina industry.
Zinc oxide has also been used as a desulfurization reagent, which makes sulfur precipitate in the form of zinc sulfide, and iron was removed at the same time. But zinc species are expensive, which will increase the cost of desulfurization. Adding zinc species will also influence the quality of the products. Wet desulfurization has the advantage of high desulfurization efficiency. The key is to choose a cheap desulfurization reagent in order to decrease the total cost of the process.
References 1. Qiuxia Wang, Keren Zhang, Junwei Zhao, et al., Status, problems and countermeasures of development and utilization of bauxite resources in China, Conservation and Utilization of Mineral Resources, 5(2008), 46-50. 2. Yang Sun, Fuchun Zhou. Study on processing industry of superior mineral resource in Chongqing, Mining Safety and Environment Protection, 32(2)(2005), 27-30,54 3. Wenmi Chen, Qiaoling Xie, Xiaolian Hu, et al., Research on reverse flotation for desulfurizing of high grade bauxite containing sulfur and producing qualified alumina products, Light Metals of China, 9(2008), 8-12. 4. Xiaomin Wang, Tingan Zhang, Guozhi LV, et al., Flotation process for desulfurization of high-sulfur bauxite, Chinese Journal of Rare Metal, 33(5)(2009), 728-732. 5. Xiaomin Wang, Tingan Zhang, Guozhi LV, et al., Flotation desulfurization of high-sulfur bauxite with butyl xanthate as collector, The Chinese Journal of Process Engineering, 9(3)(2009), 498-502. 6. Wenmi Chen, Qiaoling Xie, Xiaolian Hu, et al., Experimental study on reverse flotation technique for desulfurizing of highsulfur bauxite, Mining and Metallurgical Engineering, 28(3)(2008), 34-37. 7. Fangyin Niu, Qin Zhang, Jie Zhang, Orthogonal test research on flotation desulfurization of high-sulfur bauxite, Acta Mineralogica Sinica, 27(3/4)(2007), 393-395. 8. Runde He, Si chun Hu, Zhiying Li, et al., Discussion on the method of hydrometallurgical desulfurizing during producing alumina with high grade bauxite containing sulfur, Hydrometallurgy of China, 23(2)(2004), 66-68. 9. Chongyu Yang, Technology of alumina production (Beijing: Metallurgical Industry Press, 1993), 133-140. 10. Yiyong Wang, Tingan Zhang, Xia Chen, et al., Effects of microwave roasting on leaching behavior of diaspore ore, The Chinese Journal of Process Engineering, 7(2)(2007), 317-321. 11. Hengqin Zhao, Chongyu Yang, Hangbo Liu, Investigation on kinetics of bauxite's digestion improved by preroasting process, 6(3)(1998), 28-32. 12. A. R. Padill, D. Vega, M. C. Ruiz, Pressure leaching of sulfidized chalcopyrite in sulfuric acid-oxygen media, Hydrometallurgy, 6(l/2)(2008), 80-88. 13. Pi Liu, Zijian Lv, Zhongling Han, Analysis of effect of China bauxite ore preroasting on the Bayer process and its industrial application, Light Metals of China, 4(1997), 15-18. 14. Guozhi Lv, Tingan Zhang, Li Bao, et al., Desulfurizatin of high-sulphur bauxite by fluosolid roasting and its effect on the
Our suggestion From the above discussion, for previous desulfurization technologies for high-sulfur bauxite, there are some disadvantages such as the negative effects on alumina products, increasing material or energy consumption, polluting the environment, increasing the expense or hard to realize in the production environment etc. Under the current circumstance, it seems hard to make a breakthrough progress only from the view of the desulfurization technology itself. More importantly, previous studies focused on desulfurization only for high-sulfur bauxite ore of high grade, there is little research on bauxite ore of low-median grade. In order to deal with resource utilization of the high-sulfur bauxite of low-median grade in Chongqing China, it is necessary to have fundamental theoretical studies on the occurrence of sulfur in high-sulfur bauxites, its speciation, reaction behavior and the impact of sulfur in the production process. Combining modern instrumental analysis with chemical phase analysis, it is possible to fully investigate the occurrence of sulfur in the bauxite ore first, which will guide the following work. Then simulated study of the reaction behavior of different compounds containing sulfur in the alkali liquors can be undertaken, according to this information. Investigations of the sulfur chemical species, reaction behavior and its effect on the digestion ratio of alumina in the digestion process will follow along with a study of the digestion ratio of sulfur under the conditions of controlled digestion. On the basis of previous work, combining physicochemical properties with theoretical analysis results, and controlling the chemical reaction in the whole process, a suitable way of desulfurization for high-sulfur bauxite ore of low-median grade might be attained. It is founded that total sulfur of high-sulfur bauxite in Nanchuan Chongqing is about 1.4%. Most of this exists in the form of pyrite constituting about 1%. Other forms exist such as marcasite, melnikovite or gypsum, which are disseminated through the ore deposit and hard to separate only by flotation. Because of the diversity of sulfur species for high-sulfur bauxite ores in Nanchuan Chongqing, it is hard to get good desulfurization results by only applying a single desulfurization technology. So it is necessary to undertake some fundamental studies in order to utilize high-sulfur bauxite ores of low-median grade in Chongqing China to produce alumina of metallurgie grade.
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digestion performance, Ming and Metallurgical Engineering, 28(6)(2008), 58-61. 15. Guozhi Lv, Tingan Zhang, Li Bao, et al., Roasting pretreatment of high-sulfur bauxite, The Chinese Journal of Process Engineering, 8(5)(2008), 892-896. 16. Guozhi Lv, Tingan Zhang, Li Bao, et al., Roasting pretreatment and digestion performance of high sulfur bauxite, The Chinese Journal of Process Engineering, 9(Supplement 1)(2009), 71-75. 17. Guozhi Lv, Tingan Zhang, Li Bao, Roasting pretreatment of high-sulfur bauxite and digestion performance of roasted ore, The Chinese Journal of Nonferrous Metals, 19(9)(2009), 1684-1689. 18. Guozhi Lv, Tingan Zhang, Li Bao, et al., On the settling performance of red mud from preroasted high-sulfur bauxite, Journal of Northeastern University (Natural Science), 30(2)(2009), 242-245. 19. Xiaolian Hu, Wenmi Chen, Qiaoling Xie, Study on desulfurization of high-sulfur bauxite by calcium oxide roasting, Light Metals of China, 1(2010), 9-14. 20. Shiwen Bi, Alumina production technology (Beijing: Chemistry Industry Press, 2006), 120-122. 21. Yucheng Feng, Jinglian Zhang, Strengthened reduction and sulfur-removal with high temperature coaling for alumina clinker sintering, Light Metals of China, 1(1990), 16-19. 22. C. H. Ky3Heu;oB, Desulfurization of sodium aluminate liquors with furnace ash, Techniques of Light Metals, 1(1974), 10-14. 23. Jun Lan, Xianxi Wu, Yuancheng Xie, Research of desulfurization in the course of disposing of high grade bauxite containing sulfur, Applied Chemical Industry, 37(8)(2008): 886890. 24. Ting He, Zhiying Li, Optimal condition analysis of Guizhou high-sulfur bauxite leaching process, Nonferrous Metals, 60(2)(2008), 68-70. 25. Runde He, Zhiying Li, Nianbing Zhang, et al., Research on leaching function of Guizhou high-sulfur bauxite, Journal of Guizhou University of Technology (Natural Science Edition), 34(3)(2005), 4-7. 26. Huajun Yuan, Qinqin Li, Technological choose of Barium oxide purifying industrial sodium aluminate liquors, Light Metals of China, 4(1995), 13-17. 27. Sichun Hu, Min Guo, Hengqin Zhao, et al., Study on desulfurizing from sodium aluminate Liquors with barium aluminate, Nonferrous Metals (Extractive Metallurgy), 1(2007), 11-13. 28. Runde He, Nianbing Zhang, Sichun Hu, et al., Barium aluminate's combinative study with high grade bauxite instead of aluminum hydroxide, Light Metals of China, 4(2005), 13-17. 29. Runde He, Xifa Tan, Nianbing Zhang, et al., Barium aluminate-purifing Bayer red mud lotion desulfurization experiment, Journal of Guizhou University of Technology (Natural Science Edition), 35(3)(2006), 29-31. 30. Zhiying Li, Nianbing Zhang, Runde He, et al., Research on the new method of desulphurizing in producing alumina with exploitable sulfur-containing and high grade bauxite, Journal of Guizhou University of Technology (Natural Science Edition), 36(2)(2007), 29-31. 31. Fanghai Lu, Discussion on the desulfurizing technique of high grade bauxite containing sulfur in Maochang, Light Metals of China, 9(2008), 17-20.
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Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
DEVELOPMENT OF BAUXITE & ALUMINA RESOURCES IN THE KINGDOM OF SAUDI ARABIA AbdulGhafoor Al-Dubaisi Saudi Arabian Mining Company (Ma'aden), Al Khobar, Saudi Arabia 31952 Keywords: Saudi Arabia, Az Zabirah, Ras Az Zawr, Ma'aden, Bauxite, Alumina development of the two large phosphate and bauxite deposits discovered and utilizing the country's abundant energy resources and the developed infrastructure will generate good non-oil revenues putting Ma'aden right after Sabic in terms of revenue generation as soon as the two projects start full production.
Abstract Driven by desire to diversify its economy in an oil rich country and by the need to create jobs for the increasing number of Saudi youth, the Kingdom of Saudi Arabia established the Saudi Arabian Mining Company (Ma'aden) to develop its mineral resources. Local bauxite will be developed into an integrated mine-to-metal aluminum industry. Numerous challenges have to be overcome. Major Infrastructure has to be built; construction and operating costs have to be contained to ensure the economic viability of the project. The lack of expertise in the kingdom to run such an operation created the need to involve an international joint venture partner. The human development is no less challenging and early plans have to be in place to recruit and train a large number of Saudis to be the core of the operating organization.
The demographic changes in the kingdom are alarming. According to the Demographic Survey of 2007, the total population of the kingdom is about 24 millions including over 5 millions resident expats. With over 50% of the population under the age of 25 and a yearly growth in population over 3%, generation of new jobs becomes a high priority and warnings of social havoc due to rising unemployment rates are increasing. Mining industry- primary and secondary -offers two features that make it attractive in this regard. Firstly, this industry will contribute greatly to a balanced regional development planning—another national strategic objective—as a result of investing in remote areas where the resources are located. This will slow down and possibly reverse the migration from such areas to the already congested big cities thus helping to relax the strains, shortages, and socio-economic risks associated with such migrations.
Introduction Saudi Arabia well known for its large reserves of oil and gas contains also appreciable mineral deposits such as gold, cupper, phosphate, bauxite and others. The potential remains high as the amount of exploration for minerals done so far is limited. Most of this potential is identified to be in the Arabian Shield, the geological formation on the west of the country along the Red Sea extending inland to Najd and Qassim areas. In spite of extensive surveys conducted in Saudi Arabia in recent years, the amount of investment in exploration is very modest compared to other countries. Saudi Arabia is blessed with large reserves of phosphate, bauxite, magnesite, gold, iron, copper, and other minerals that are yet to be developed and exploited. It suffices to say that the total fund spent on exploration in the Arabian Shield in the last 30 years is less than what is spent in the Canadian Shield annually. Yet, all indications attest to the diversity of the mineral resources in the Kingdom and the tremendous potential forfindingnew resources through additional exploration.
Secondly, the industry's potential for job creation is large as the mining industry is much more labor-intensive than oil, gas and petrochemical industries. As thousands of young Saudis entering the job market each year, this industry will have the capacity to absorb a large number of job aspirants through direct and indirect hire. Location and Geology The bauxite deposit is located in a remote desert area of northern Saudi Arabia, predominantly in the province of Ha'il. Elevations range from 535 to 600 meters above mean sea level. Ma'aden's exploration license is 192 kilometers long by 35 kilometers wide, approximately centered on the town of Az Zabirah (northwest of Riyadh). Ma'aden's proposed mine targets the South Zone of the deposit and Ma'aden's mine processing facilities will be located approximately 43 kilometers northwest of the town of Qibah. The town of Az Zabirah is a similar distance to the northwest of the proposed mine processing facilities. The mine is called after a small settlement between Qibah and Az Zabirah called Al Ba'itha.
Since the discovery of oil in the late thirties and especially after the oil crisis of the seventies and the increase in oil prices, the country has enjoyed large revenues that translated in the leaps of development of both infrastructure and industry. This revenue remains however subject to fluctuating oil prices and it has seen the worst downturn cycles in the mid eighties and at the turn of the century. From the enactment of the national Five Year Development Plans in 1970, diversification of Saudi economy has remained one of the top strategic objectives of the Kingdom. Great achievement has been made in the petrochemical sector generating over $20 billion revenue in 2003 and in 2007 non-oil manufacturing contributed 10% to the GDP.
The coordinates for the approximate center of the mine processing facilities are: WGS84 latitude / longitude 27°39'52.8"N/ 44°0'38.5"E UTM Zone 38 3 060 750N / 404 600E The alumina refinery will be located on a virgin peninsula on the Arabian Gulf called Ras Az Zawr about 80 KM north of the major industrial city of Al Jubail.
Following the successful story of developing the petrochemical industry led by Sabic, the government established in 1997, the Saudi Arabian Mining Company (Ma'aden) to lead a similar development of the country's mineral resources. The
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The coordinates of the approximate center of the processing plants (Refinery, Smelter and Rolling Mill) are: WGS84 latitude / longitude 27°30'52.4"N/ 49°10'38.5"E. UTM Zone 39 3 044 753N / 319 980E The location map below shows the location of both facilities and the railway link in between. ~*-m^mmmm k
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Since 2001 Ma'aden has continued drilling in the south and central zone to improve resource and reserve estimates. The resource in the south zone of the deposit is currently estimated at about 300Mt and the reserve is estimated at about 200 Mt at 40% TAA cutoff which represent the basis of current phase of the project. The resource in the central zone is estimated by Hatch (2004) at 135 Mt with slightly higher silica and lower iron content. Ma'aden continues through yearly drilling programs to increase resource estimate and obtain better understanding of the bauxite ore in the central zone of the deposit.
Geologic evaluation of the deposit was initiated by Riofinex through drilling exploration programs performed from 1979 through 1984. The Riofinex Report identified three distinct zones (the North Zone, Central Zone and South Zone), spaced approximately 30 kilometers apart along the strike, which exhibited economic potential. Additional exploration was performed by the French BRGM from 1987 through 1993. Ma'aden's also conducted exploration drilling at the deposit in 2002 through 2003. The deposit is located discontinuously along a strike length of approximately 105 kilometers and with a width identified by drilling of at least 5 kilometers. In the Riofinex report, Riofinex interpreted that the deposit's bauxite was formed during the Early Cretaceous as a weathering profile on Late Triassic to Early Jurassic terrigenous sediments deposited in marine or littoral environments. Riofinex tentatively assigned these sediments to the Minjur Formation (North Zone), the Marrat Formation (Central Zone), and the Dhruma Formation (South Zone). Both Riofinex and BRGM referred to these collectively in their respective reports as the "Parent Rock Sequence".
The Project The Ma'aden Integrated Aluminum project, jointly developed by Ma'aden and Alcoa, includes currently the following scope:
• •
4.25 Mtpa bauxite mine in Al Ba'itha 1.8 Mtpa alumina Refinery in Ras Az Zawr 750 Ktpa Aluminum Smelter in Ras Az Zawr 460 Ktpa Rolling Mill Common Facilities and Infrastructure
The Smelter and Rolling Mill are ahead on planned completion schedule whereas Mine and Refinery are lagging with about one year behind. The first metal from smelter is scheduled for early 2012. The Smelter will start with imported alumina until Refinery is able to supply the required alumina.
The deposit's bauxite profile is several meters thick and is unconformably overlain by Late Cretaceous fluviatile and lagoonal sediments. Riofinex and BRGM refer to these sediments as the "overburden sequence (OVB)," and the Riofinex Report characterizes the sediments as part of the Wasia Formation.
On the same site and within the Ras Az Zawr Mineral Industrial City (RAZMIC), Ma'aden is developing also a large phosphate complex that produce DAP fertilizer utilizing ore from the large phosphate deposit in Al Jalamid. The phosphate complex shares with the Aluminum Project basic infrastructure such as roads, housing and port.
The bauxite outcrops in limited places on the western edge of the deposit and uniformly dips at approximately 0.5 degree to the northeast. At the eastern edge of the deposit, the BRGM Report estimated that the overburden sequence is approximately 50 meters thick.
Large Infrastructure- The first challenge Considering the remoteness of mine sites and lack of water resources in these locations, processing plants have to be located hundreds of kilometers away. Therefore, providing a means of
24
2400 MW of electrical power and just over 1 million cubic meter per day of potable water. Most of the power generated (1350 MW) will be used by the aluminum complex through a high reliability grid connection. The power and desalination plant will also provide the required water for domestic, process and cooling duties.
transportation of the ore to potential processing locations is of utmost importance. Including development of such large infrastructure in a mining project will be fatal to its viability. Early studies of the feasibility of the Az Zabirah bauxite deposit highlighted transportation of bauxite to location of process facility as the major obstacle to develop the deposit. Saudi Arabian Sabic concluded in a study conducted in 1985 that processing of Az Zabirah bauxite is not feasible due to lack of water at the mine site and the lack of means of transport. Since inception of Ma'aden in 1997, it highlighted this problem to government agencies and requested allocation of funds to building the railway to support the feasibility of the bauxite and the phosphate projects.
The mineral railway and the deepwater port are in the final construction stages and will be operational early in 2011 to support the phosphate plant which is being commissioned. The power plant was awarded for execution and is expected to be operational late 2012 in time to support the startup of the aluminum smelter.
Another major infrastructure required for the process facilities on the Gulf is a deep water port that allow import of raw materials ( caustic, liquid pitch, calcined coke ect) and export of products ( aluminum metal, alumina, DAP, etc) from and to the world markets.
The Integration Challenge The bauxite and alumina is part of a larger development for a mine-to-metal integrated project. As outlined above the level of interfaces to develop the required government-sponsored infrastructure is by itself a challenge. It required working through different governmental and semi-governmental organizations to move things and make decisions. Coupled with that is a high level of interface management required internally to get the different elements of the project aligned. These elements (Mine, Refinery, Smelter, Rolling Mill and Integrated Infrastructure) are being designed and executed by different EPCM consultants in three continents and engineering design is spread over many countries. Yet all these elements have to coordinate activities and manage interfaces related to site, systems and procedures, corporate policies, design standards and at all execution and operation readiness levels. This presents a challenge unprecedented at this scale in the aluminum industry.
Recognizing the value of these mining projects to the development of the country, the government undertook these two major infrastructure projects to build the North South Railway (NSR) which is by itself a major railway project with a track length of over 1800 Km extending from the Al Jalamid phosphate deposit near the Jordanian borders down to Az Zabiarah bauxite deposit andfinallyto the Ras Az Zawr mineral processing facility on the Arabian Gulf. Reliable and cost-effective electric power source is vital to the success of the project. The national grid is quite stretched and will not provide the reliability and the tariff that support the project. The electric power required for the Ras Az Zawr Aluminum Complex including the three operations (Refinery, Smelter and Rolling Mill) and infrastructure, is in excess of 1300 MW. Ma'aden completed many studies for the feasibility of a power plant to support the industry but this encountered many challenges: •
Ma'aden would be the first private industry to build a major power generation plant at this scale
•
The power industry regulatory framework is at infancy stage and there are no clear regulations on issues such as power generation licensing, fuel pricing, power tariffs, power wheeling, reliability of supply etc
•
Connection to the national grid is a must not only to provide backup power but to allow sales of excess power needed by the national grid.
•
The size of investment in a large power plant became huge especially with high sulfur heavy crude oil proposed.
•
The issues above required a large level of coordination among many government organizations where decision making is very slow.
Bauxite characterization and Technology Selection Extensive sampling and assaying programs were conducted for the bauxite starting with Riofenix program. Ma'aden during a prefeasibility study in 2001-2003 carried a data acquisition and assaying program. Ma'aden undertook an intensive close-spaced drilling program in a region of the South Zone forming the initial ten-year mine plan. A number of samples from that area were then composited in a devised plan to arrive at a bulk representative sample, indicative of the first ten years mining. The drill-hole samples were assayed at the Al Amri laboratory in Saudi Arabia, and the samples were further analyzed with extensive testing by Australia's Commonwealth Scientific and Industrial Research Organization (CSIRO) Minerals laboratories in Australia. These composites were characterized in terms of chemical elemental composition, quantitative mineralogy, particle size distribution, and available alumina and reactive silica under expected refinery digestion conditions (digestion temperature and liquor caustic concentration). The predesilication behavior of the composite bauxites was investigated, particularly the extent of reactive silica to desilication product (DSP) transformation under nominated conditions, and the minimum time to achieve optimum transformation. Also On behalf Ma'aden, in 2004 Hatch engaged JKTech to determine a number of comminution parameters for Az Zabirah bauxite samples, including 'average grade - AGB', 'high grade HGB' and 'low grade - LGB'. Parameters such as the ore behavior in SAG/AG mill, the primary crushing power draw, the rod milling power draw, the ball milling power draw, the abrasion of
After much iteration, and a few years of studies, consultations and interaction with government agencies, it was decided to combine Ma'aden power requirement with water needs of the country and under special arrangement Saline Water Conversion Corporation (SWCC), a government agency, will build a combined water and electricity generation facility in Ras Az Zawr with a capacity of
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was done. The choice of technology was governed by the following underlying principles:
mill lining, the strength of the ore sample when subjected to compressive forces without lateral constraints, were investigated. Later in 2005 and 2006 additional infill drilling was carried out in the South zone and a new composite sample representing the 30 years ROM was prepared. Further chemical and physical testing programs were undertaken between 2007 and 2010 including digestion and extraction tests, bauxite rheology tests and impurity analysis (especially of iron, fluoride, chloride and calcium). The results of all these tests confirmed findings of the earlier campaigns and were incorporated in the flow sheet design. The series of tests confirmed the following basic characteristics: • • • • • •
The alumina is highly boehmitic (monohydrate) High alumina and silica content ( average 56 %, and 9% respectively) Bauxite will require a high digestion temperature (260-280C) Due to high silica and high DSP, the circuit will therefore have a high capacity to incorporate impurities Low iron content Low organic matter
Table 1. Az Zabirah Bauxite (Average Values)
1
High grade
Representative Low grade
A1203
56.2
55.56
54.14
Fe203
12.59
11.17
10.99
Si02
5.48
7.71
9.29
CaO
1.24
1.14
1.16
Ti02
3.48
3.45
3.35
P205
0.16
0.15
0.15
MgO
0.09
0.09
0.09
S03
0.65
0.65
0.68
V205
0.1
0.09
0.09
Cl
0.03
0.03
0.03
F
0.08
0.08
0.08
LOI
19.77
19.64
19.4
TOC
-0.2
-0.2
-0.2
Design has to use public-domain technology as much as possible
•
Design has to reflect the most advance developments in Bayer process, especially in environmental, health and safety.
•
The bauxite is new and has not been processed before.
•
The refinery will have a new operation team and design must be robust and forgiving
•
The design has to allow for local harsh ambient conditions
The above considerations and the bauxite characteristics formed the basis of the process design. Although conventional Bayer circuit was selected for most areas, digestion design was given special consideration. Ma'aden operation experts along with external consultants carried out a qualitative comparative risk evaluation of digestion design options that include single stream full jacketed pipe heating, hybrid indirect and direct heater design and dual stream utilizing conventional shell and tube heaters. This evaluation was undertaken as possible means to mitigate the risk of unacceptably high scaling rates in high temperature indirect slurry heaters to determine which should be taken forward into process and engineering design. Evaluation criteria included process risks, capital and operating costs, and operability and maintainability considerations. Based on the evaluation criteria, the hybrid option was recommended for heater design in digestion. This option benefits from the jacketed pipe heating design in terms of energy efficiency and ease of cleaning. However, by shifting to the direct heating beyond 200C, the design minimizes the risk of potential scaling in the tube.
Table 1 below shows the composition of Az Zabirah bauxite
Components (wt%)
•
The Ma'aden alumina refinery consists of a two-unit greenfield plant operating the Bayer process and using public domain technology to produce 1.8Mtpa of smelter grade alumina. Approximately 1.4Mtpa of the refinery product is expected to be consumed by the nearby Ma'aden aluminum smelter. The remainder will be exported. The Refinery will nominally comprise: •
Three grinding mills, at 100% digester unit capacity each
•
Two digester units
•
Two precipitator trains
•
Two fluid-bed calciners, at 75% total plant capacity each
•
Two evaporation trains.
The Figure below shows a schematic of the Ma'aden Refinery flow sheet and the basic design conditions.
The Az Zabiarh bauxite underwent extensive bench scale testing by reputable research laboratories and results were reviewed by many experts from different leading operating companies. Following 2009 Alcoa JV partnership the process design was reviewed by Alcoa process consultants. The consistency in the results of these tests and reviews gave high confidence in the selected refinery flow sheet design. Based on the bauxite chemical and physical characterization, the selection of process technology
The Maaden Bayer circuit is unique in some aspects due to the bauxite chemistry and the country ambient conditions: 1. The flow sheet does not return any supernatent liquorfromthe residue area due to high evaporation rates
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2. 3.
•
The Bauxite has high reactive silica and therefore high incorporation capacity of impurity in the DSP. The casticity of the liquor is relatively high ( about 0.94)
Joint venture partner discussions, negotiations, and due diligence were carried out with the major players in the industry. At last and In December 2009, a joint venture partnership was announced with Alcoa. This paved the way to progress the project into the next phase of execution and provided depth of resources in both execution and operation expertise.
All of the above predicts a fairly clean liquor circuit. Mafaden Alumina Process
The large size and complexity of the project and large equity funding ( over $10 billion investment)
6?i8kPsg
Controlling Construction Costs and Operating Costs The Saudi government extended its great support for the first major mineral development in the kingdom. The government, through Public Investment Fund (PIF) and the Saudi Arabian Railway (SAR) built a $4 billion railway link between the two deposits ( Al Jalamid and Al Ba'itha) and the process facilities in Ras Az Zawr on the Gulf coast. The railway crossing the sand dunes of An Nufude Desert involved record breaking earthwork volumes in a difficult terrain. The government through the Saudi Port Authority (SEAPA) built a major deepwater port in Ras Az Zawr, a project close to $1 billion. The government of Saudi Arabia invested in a large combined water and power project in Ras Az Zawr to provide electricity to the Aluminum complex at attractive prices. Yet, the amount of remaining infrastructure to support the project is sizable. The mine site facilities in Al Ba'itha are in remote area which lacks all infrastructures and the project has to include development of housing, water supply and access roads. The process facilities in Ras Az Zawr are proposed on a virgin peninsula on the Arabian Gulf north of the major industrial city of Jubail. Major housing, access roads, security fencing, seawater supply and return systems, and earth works have to be developed by the project. The integrated project sought possible synergies between the smelter, refinery and rolling mill to share these infrastructure facilities to reduce costs. Value engineering studies were conducted at many stages of the project to challenge some of the equipment choice and or count to eliminate any excess capital. Best execution and delivery systems had to undergo several reviews by international and reputable consultants as well as Alcoa internal peer reviews. Alcoa experience in delivering such large and complex project was brought to table. Given the dispersed execution centers, the cost of execution including owner and EPCM consultants costs have to be contained. In some of the hot market environment that passed in 2007 and 2008 when construction costs sky rocketed, this threatened the feasibility of the project.
Joint Venture Partner Since Ma'aden have no operating experience in the alumina and aluminum industry, the search for a joint venture partner was inevitable. Given the technical, political, and business challenges with establishing a vertically integrated aluminum industry and given the financial turmoil that hit the markets in the last few years, this search was long and demanding. It is true that Saudi Arabia is rich in energy and this constitutes a major driver for investment in the energy demanding aluminum industry. However from Ma'aden point of view, with a mandate to develop the local mineral resources, exploiting the local bauxite resources is equally important for the good reasons discussed above and as mandated by the country's strategic development plans. The following issues stood out in the process of search and selection of JV partner: • The magnitude of investment in the infrastructure (at earlier stages of the JV surveillance many of the infrastructure elements that are now shouldered by government were part of the project). • The legal framework of the kingdom is not advanced enough and that was perceived as a risk to external investors • Technical risks associated with processing of the new Az Zabiarh bauxites
Controlling the operating costs of each ton of bauxite, alumina and aluminum is critical to the success of the project. Efforts were made to optimize the mining plan of waste and bauxite. Considering the quality of bauxite in Al Ba'itha, the different grade control parameters, and the variability in the ore body, the mine plan went through several optimization and reviews to reduce the per ton cost. The options of contracting all or parts of the mine operation were carefully examined. Energy costs for the production of Alumina were compared against other similar operating refineries. Despite the high temperature digestion conditions (273 C), opportunities to reduce steam consumption were exploited using experience from other Alcoa energy efficient operations. Working closely with government agencies to
27
gain their support in building the Ras Az Zawr power plant is part of the effort Ma'aden management took to control energy costs essential to competitive operation.
plan was developed to ensure selecting the best students from surrounding areas of the mine and refinery facilities to support the local communities and to establish for higher retention. Students will be recruited in different intakes depending on their role and to ensure they join in time to attend on-job training before the startup of the facility. Ma'aden Human Resources and Operational Readiness personnel approached technical colleges in the area and discussed curriculum and timetables. Because this is a new industry in the region, a level of customization of college training curriculum has to be done. Ma'aden and its joint venture partners facilitated a cooperation program with overseas technical colleges with experience in training industry professionals, to ensure training program outputs respond to the specific needs of the industry.
The consumption of caustic for processing the Az Zabirah bauxite is very high averaging 200 KG/ ton of alumina due to the high reactive silica content. This puts the yearly demand of caustic for this refinery at more than 350 Ktpa on dry basis. Considering the volatile and unpredictable caustic prices, the risk of escalating operating costs due to caustic is very serious. Ma'aden explored options available including teaming up with local investors to build a caustic production facility in Ras A Zawr. Discussions and negotiations were held with the local caustic producer, Sabic. These efforts succeeded in establishing a joint venture partnership with Sahara, a local petrochemical company, to construct a facility that will provide the bulk of caustic requirements ( Plant was sized to produce 250 Ktpa of membrane grade caustic and 300 Ktpa of EDC, and planned for startup one year earlier than Refinery). The remaining quantity will be sourced from Sabic or through global procurement network of the JV partner. In parallel, efforts to reduce caustic consumption in the process continued looking at options such as lime addition, improving wash efficiency and others.
Selected roles in the mine and refinery operating organization were targeted for overseas training in an operating bauxite mines or alumina refineries. These are mainly in lead and supervisory positions. Saudis nominated for these roles will be experienced engineers, and operation/maintenance staff from existing operation who lack the specific industry experience. They are intended to be the core of the operation organization. They will be seconded to overseas operation for varying periods to give them the required exposure before commissioning the operation of mine and refinery. The plan requires them to finish in time to return to country to participate in developing operation, maintenance and training materials, coaching of the new college students and later participate in the pre-commissioning and commissioning activities. Management recognizes that staffing as early as planned is costly, however the payback in terms of readiness for operation cannot be under estimated.
To control labor costs, the mix of labor for the mine and refinery organization has to be optimized. In Saudi industry, human resources are categorized into three types: Saudi Nationals, Western Expatriates, and Eastern Expatriates. Major industries in kingdom have achieved over the years a large percentage of Saudization approaching 90%. The target set by management for Ma'aden aluminum project is 50% as minimum. Considering the level of expertise potentially available in each labor category at startup, each role was studied carefully to ensure the mix of labor supports the intent to control labor costs but does not compromise safety or quality.
Deliverables required from vendors to support training were highlighted and given high priority in the procurement process. The delivery of these was timed to support the overall training plan to ensure training material is available earlier not later. Training of locals was included in all contracts signed for technology, engineering and project management.
These efforts to contain operating costs and measures taken will result in a very competitive operation. Cash operating costs in the first quartile or lower second quartile are forecasted for the refinery and smelter operation in Ras Az Zawr.
Thinking of long term retention of developed skills - a challenge ahead.
Developing the people and the skills- Planning for operational readiness
Ras Az Zawr is a relatively remote location. It is planned to house employees on single status in company provided campus on site. Families have to stay in the nearby Jubail. The majority of our staff will be married Saudis according to the cultural norms. They will have to drive about 80 Km to come to work and back at the end of the work day.
The challenge of building the Operation Organization including people, business systems and manufacturing systems is not less than the challenge of building the physical assets. Saudi Arabia has well developed industries in oil, gas, and petrochemicals. The surrounding region is home for Saudi Aramco and Sabic, world leaders in oil& gas and petrochemical respectively. Over decades of their presence, they trained and employed large number of skilled Saudi workforce. They established English as the industry language and created a strong base of industry knowledge and support services. However, all of the skilled manpower is tied to these industries and a new industry such as Ma'aden has to work on training and developing its manpower resources. Building on the strong industrial base available, Ma'aden took manpower development as a very high priority.
The operation and maintenance of alumina refinery include labor intensive tasks not typical in the surrounding hydrocarbon production and processing facilities such as descaling of piping and equipment, continuous isolation of sections of the plant for caustic washing, etc. These physically demanding tasks coupled with harsh ambient conditions make the refinery operation a difficult environment. These features of the site and the operation will make retention of the developed skill a challenge ahead. Plans that address human resources aspects, safety and health aspects and community relations aspects will be prepared to make Ras Az Zawr an attractive place to work and maximize talent retention.
Plans for recruitment of young Saudi students were developed. Following the design of the operating organization, the skills of each of these targeted positions were identified. A recruitment
28
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
DIGESTION STUDIES ON CENTRAL INDIAN BAUXITE P.K.N. Raghavan, N K Kshatriya, S Dasgupta Bharat Aluminium Co. Ltd., BALCO Nagar, Korba, India Keywords: bauxite, digestion parameters, caustic loss temperature, bauxite charge, lime quantity and digestion liquor concentration is discussed in this paper.
Abstract Indian bauxite deposits are grouped into five categories namely Eastern Ghats, Central Indian, West Coast, Gujarat and Jammu & Kashmir. Each of the bauxites has its own typical digestion and settling conditions due to variations in mineralogy. Bauxite digestion depends greatly on the temperature, pressure conditions, the recycled liquor concentration, bauxite mineralogy and charge quantity. Productivity of the entire Bayer process depends to a large extent on the digestion process. A series of experiments were conducted in an attempt to optimize the digestion conditions for Central Indian bauxite in order to achieve low caustic soda loss coupled with minimum bauxite consumption. Experiments were conducted to find out the best parameters for digestion on a laboratory scale. The effect of digestion temperature, bauxite charge, lime quantity and digestion liquor concentration is discussed in this paper.
Experimental Laboratory digestion tests were carried out in programmable, electrically heated digester, with small autoclave bombs of working volume 150 ml. Bauxite sample of -100 mesh size were dispersed in plant digestion liquor. The bauxite charge to be taken for digestion in the autoclaves was calculated to meet a defined charging molar ratio, which varied in quantity depending on the chemical assay of bauxite. Digestion tests were carried at different temperatures with varying retention time applicable for boehmitic bauxite. The above conditions were selected following the parameters for alumina production at the Bharat Aluminium Company Ltd, Korba (India). The quantity of alumina extracted was estimated by analyzing the bauxite and its corresponding red mud. The major constituents of bauxite and red mud were analyzed by XRF-PW-2440 Philips, Netherlands. The mineralogical analysis were carried out using PANlytical X'Pert Cubix Pro Series diffractometer, equipped with copper target tube, X'celerator detector and operated at 40kVand 30mA. Diffraction data were analyzed using PANlytical X'Pert High Score Plus Version 2.1. The caustic and alumina content of digested liquor were analyzed by potentiometric auto titrator, Mettler Toledo, DL55, Switzerland.
Introduction Alumina extraction from bauxite by the Bayer process requires an efficient digestion. The digestion step in the Bayer process is influenced by the relative amounts of the alumina hydrates present. Atmospheric pressure digestion technology can be used for only Gibbsitic bauxite with negligible or no boehmite content. Low pressure digestion technology on the other hand is used for the Gibbsitic bauxite which contains some boehmite. High pressure digestion (with or without catalyst addition) is exclusively used for the boehmitic bauxite or diasporic bauxite. The digestion process developed by Bayer was designed to treat European bauxite having moderate boehmite and some gibbsitic but negligible diaspore. These latter bauxites need to be digested at high temperature up to 240-2500C and high pressure up to 3035 atm. Lime addition facilitates the extraction of boehmite.
Results A set of digestion experiments was carried out at 240 °C for 18 minutes with varying target Molar Ratios, the outcome of these tests is given in Table 1. Digestion experiments were carried out at different digestion temperatures with varying target MR keeping constant digestion time and the results of the tests are given in Table 2. The results of digestion tests at different caustic concentration in the plant digestion liquor and for different digestion times are shown in Table 3. Digestion experiments with variation in bauxite quality at two targets MRs 1.42 and 1.45 are presented in Tables 4 and 5 respectively. Table 6 shows the test results obtained after bauxite digestion with various quantities of lime.
Caustic soda consumption is controlled by the digestion conditions and the amount of the kaolinite and quartz. The performance of the digestion process is a function of various parameters such as the digestion liquor concentration & Molar Ratio, the digestion temperature, the bauxite charge, residence time and lime addition. In practice, because of the fluctuating operations either all or a few of the above parameters do deviate from the laid down norms causing continuous variations in the digestion efficiency. The productivity of the entire Bayer process depends to.a large extent on the efficiency of the digestion process. In this case high temperature digestion (235-245°C) is used for Central Indian bauxites because of the boehmite content.
Discussion The digestion studies mentioned above have been carried out in order to optimize the digestion conditions to achieve the highest digestion efficiency, with lowest soda loss.
Plant efficiency requires an optimization of the digestion conditions in order to achieve low caustic soda loss coupled with minimum bauxite consumption per tonne of output alumina. In an effort to achieve the above mentioned goals, a series of experiments were conducted to find out the best parameters for conducting digestion on a laboratory scale. The effect of digestion
Laboratory digestion tests were conducted in order to determine the critical Blow off Molar Ratio beyond which the efficiency drops to a very low value. The results (Fig.l & Table 1) clearly show that with reduction in target MR the digestion efficiency decreases in all cases. But at the same time we may observe that below the target MR value of 1.40 there is a significant change in
29
the slope of the curve indicating the lower limit of the liquor to bauxite ratio. It can be seen from Fig.l and Table 1, that the digestion efficiency decreases to some extent with decrease in the target MR. But at times efficiency may be sacrificed in order to increase the throughput. This is a clear case of optimization as a techno-economic consideration.
The disadvantages are anticipated to be as follows: (a) Higher steam consumption (b) Increased scaling of live steam heated digesters (c) Increased chances of impurities going into the aluminate liquor, thus leading to increased impurity in product. The digestion studies mentioned above have been carried out in order to optimize the digestion conditions to achieve the highest digestion efficiency with lowest soda loss. Rg.2 Effect of temperature cm digestion efficiency Target MR 1.42 <
^
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. 242
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Table 1. Digestion tests at different target MR Bauxite Quality
Plant Liquor wt %
LOI
2235
Si0 2
3.15
Fe 2 0 3
19.43
Ti0 2
9.13
lAh
Na 2 0 c A1203
Table2. Digestion test at different temperature with varying target MR
Digestion Conditions
In gpl 188
Desili cation
90.9
Digest ion Lime
Temp,
95
Time, hrs Temp,
8
°C
°C
Bauxite Quality wt %
240
18 Time, minutes 0.75% of bauxite charge
45.00
Target MR 1.28 1.30 1.32 1.34 1.36 1.38 1.40 1.42 1.45 1.50
ML 49.81 49.76 49.17 48.99 48.76 48.33 48.40 48.36 47.97 48.46
η 79.89 81.09 81.84 82.17 82.74 82.22 83.15 83.55 84.17 84.78
SL 79.27 83.45 81.98 80.47 79.61 87.11 83.61 80.80 78.76 80.03
BXT 2.78 2.74 2.71 2.69 2.70 2.67 2.66 2.64 2.62
LOI Si0 2 Fe 2 0 3
Na2Oc AIA 19.43
Ti0 2 A1203
9.13 45.00
Temp, °C
Results
Target MR
240
ML
% % Kg
242
n
SL BXT ML H
244
ML-Mud Load (%) η - Digestion Efficiency (% TA basis) SL - Soda loss (Kg NaOH/T A1203) BXT - Dry bauxite consumption (T/T A1203 )
Plant Liquor In gpi Na2Oc 185 90.9 A1203
SL BXT ML
n
SL BXT
T
% % Kg T
% % Kg T
138 48.33 82.22 87.11 2.70 48.76 82.61 82.41 2.69 48.44 83.06 83.58 2.67
Digestion Conditions Desili Temp, °C cation Time,hrs Digest Temp, UC ion Time, minutes
95 8 240 18
1 1.40 48.40 83.15 83.61 2.67 48.55 83.60 82.41 2.66 48.18 83.93 81.78 2.65
1.42 48.36 83.55 80.80 2.66 48.12 84.04 78.29 2.64 4833 84.56 78.81 2.63
1.45 47.97 84.17 78.86 2.64 48.34 84.75 7832 2.62 48.46 85.14 77.99 2.61
Soda loss shows a declining trend when digesting bauxite for 36 minutes instead of 18 minutes (Fig.3 & Table 3).
It is well known fact that digestion efficiency increases with increase in digestion temperature. The digestion efficiency increases about 1% at 244°C (Fig.2 & Table 2), at a target MR of 1.42, which is closest to the plant operating range.
Bauxite quality plays an important role in the digestion efficiency achieved in the plant, other parameters remaining constant. Of late the bauxite quality for this operation is deteriorating, mainly with respect to recoverable alumina content, whereas the silica content is at acceptable level. In general the digestion efficiency achieved in the plant level is in direct proportion to the alumina content in the bauxite.
The advantages of operating at 244°C are: (a) Higher bauxite dissolution hence lower bauxite consumption (b) Lower soda loss
30
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Fe 2 0 3 Ti02
22.42 3.27 17.85 9.02
A1 2 0 3
46.57
LOI SÌO 2
Digestion liquor Na2Oc,gp MR 1 170.5 175.9 179.8 186.0 190.0
2.91 2.95 2.92 2.93 3.02
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BaftateôsaBi?
95 8 240 1.40
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MRUS
SL 81.61 82.81 81.61 82.09 81.10
85.48 85.93 86.01 86.13 86.32
SL 77.94 79.62 79.58 79.65 80.31
Though there are some aberrations in the experimental results, in general it can be concluded that digestion efficiency is reduced drastically when the alumina content in bauxite is below 44%. The weight ratio of A1 2 0 3 to Si0 2 known as the module of the bauxite, is more pronounced with respect to bound soda loss (Fig.4, Fig.5 & Table 4, 5) as these are more related to the silica of the bauxite processed. From the experimental results it can safely be concluded that for achieving good digestion efficiency accompanied with low bound caustic losses .The bauxite quality requirement should be as follows: A1203 > 44% , Module > 15 From Fig.8 & Table 6, it can be seen that the soda loss decreases with increasing lime charge, due to the formation of Calcium Aluminium Silicate and release of sodium hydroxide (Soda in red mud).
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as those from the coastal Gujarat are hard, highly diasporic in nature and can be best suited for Refractory applications.
TafokT.Öptiinum digestion conditions for Centrât Indian Bauxite
Fip.8. Effed of ime dosage ®n soda loss
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While carrying out the digestion studies under the conditions outlined above we can expect a lower bound on caustic loss, coupled with low bauxite consumption. We can also expect significantly lower Fe203 and Si0 2 levels in the digested liquor, which gets translated in the product hydrate and hence reports in the calcined alumina. While translating these results to plant level operation, some care must be taken on the following points:
iB
16 Saisate Poésie
Fig.6. Effect of bauxite alurrina on dkgestjon Eff. Target MR-t.AS
42
43
44
45
46
47
(1) The Bauxite used in the laboratory tests was of uniform mesh size, while that in the plant operation are of varying granulometry ranging from very coarse to fine. (2) The agitation level in the laboratory is not the same as under plant conditions. 48
(3) In the plant, flashing of the digested slurry is carried out, however this is not available in laboratory scale experiments.
Alumna« Basaste %
On careful observation of the results, one can observe that the parameters given in Table 7 would be appropriate for the digestion of Central Indian Bauxite.
Fig.?. Effect of bauxite module on soda loss Target MR-1.45
Acknowledgements
» 70 J "S 60 j co
Q2
Lksedssafs
Rg.5.£ffeet o f module on soda loss Target MR-1.42
The authors acknowledge the constant encouragement from our Chief Operating Officer, Mr. Bibhu Prasad Mishra and Chief Executive Officer and Whole Time Director, Mr. Gunjan Gupta, for their constant encouragement during the progress of this work. The Authors would like to thank the Management of Vedanta resources for allowing us to publish this paper.
1 40 + ÎG
12
14
16
18
20
Basaste til ©éufe
On Careful observations, it can be concluded that the Digestion conditions given in Table - 7 will be the optimum for high temperature - High Pressure Digestion of Central Indian Bauxites, though the digestion conditions shall vary marginally for those from the central Indian plateau towards the Eastern Ghats are softer, higher Gibbsitic in nature and hence can be digested at temperatures around 230°C, where as those in the Central India and towards the western Ghats are Bohemitic in nature and need higher temperatures and pressure.
References 1. Zhang Xiaofeng, Chen Wankun, Intensive Digestion Technique for Diasporic Bauxite, Light Metals, 199l,p 33. 2. G. Wargalla, W.Brandt, Processing of Diaspore Bauxites, Light Metals, 1981, pp 83-100. 3. V.G.Hill,R.J.Roboson, The Classification of Bauxites from the Bayer Plant Stand Point, Light Metals,1981,p 15.
The Study of Bauxite reserves from Gujarat which falls under the Western Ghats, the Bauxite Quality is mixed. We get highly Gibbsitic Bauxites (95% Gibbsite) in the regions of Kutch where
32
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
EFFECTS OF ROASTING PRETREATMENT IN INTENSE MAGNETIC FIELD ON DIGESTION PERFORMANCE OF DIASPORIC BAUXITE Zhang Ting-an1, Dou Zhi-he1, Lv Guo-zhi1, Liu Yan1, Du Juan1, Wang Xiaoxiao1, Li Yan2 Shenyang, Liaoning 110004, China \ School of Materials & Metallurgy, Northeastern University;2. Logistics Division of Northeastern University); Shenyang, Liaoning 110004, China
[email protected] Keywords: Diaspore; Digestion properties; Roasting pretreatment; Intense magnetic fields
magnetic fields. The coupling effect of a magnetic field and a temperature field is able to improve the morphology of bauxite ore and chemical speciation of the main phases, leading to activation of the ore and reducing the digestion conditions needed for the bauxite. In this paper, the effect of process conditions on digestion performance of bauxite were studied during a roasting pretreatment with an intense magnetic field. This research will provide a theoretical basis of improving digestion for the alumina industry[1~51.
Abstract This paper investigates the changes of phase and apparent morphology under the combined effects of an intense magnetic field and temperature field and the effect law of different roasting conditions on the digestion performance of roasted diaspore. The results indicated that roasting pretreatment under high magnetic fields can change the microstructure and improve the digestion properties of bauxite. The reasonable roasting conditions with intense magnetic field are as follows; the roasting temperature is 550°C, roasting time is 60min and the magnetic field intensity is 6T. The digestion rate of alumina of the roasted ore is 84.17% and the digesting liquid ratio is 1.39 while the digesting temperature is 190°C with the digestion time of 60mins. The digestion rate of alumina of the roasted ore increases to higher 52.85% than that of the raw ore , but the digesting liquid ratio is decreased to lower than 0.99 of the raw ore under the above roasting conditions. The digestion temperature of roasted ore decreases by 30°C compared to the raw ore.
Experimental Diasporic bauxite from Henan Province was used. The chemical and mineralogical compositions are shown in Table I and Table II, respectively. Mineralogical analysis of the initial sample (Fig.l) was performed on a PW3040/60 X-ray diffractometer with scanning angle from 5 to 90 degree, giving the result that the bauxite primarily consists of diaspore as well as other minor minerals such as hematite, kaolinite and anatase. 18000 • ■ *
16000
Introduction Chinese reserves of bauxite include vast quantities of diaspore with a high content of aluminum and silica and low content of iron, which at present are sub-economic due to the high levels of reactive silica that require high digestion temperatures and cause expensive losses of caustic soda during Bayer processing. At present there are many intensive leaching methods for diasporic bauxite, such as roasting pretreatment - low temperature leaching technology, mining added sweetening technology and soon. In order to promote the technology of producing alumina from Chinese bauxites, Northeastern University has presented an approach to roasting pretreatment of bauxite using intense
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Chemical composition
A1203
Si0 2
Fe 2 0 3
CaO
Na 2 0
LOI
Content/%
67.57
7.14
5.54
0.90
0.31
13.85 1
33
|
100
Table II Mineralogical composition of bauxite
|
Mineralogical composition
Diaspore
Hematite
Goethite
Quartz
Anatase
Kaolinite
Content%
69.4
3.1
2.7
1.0
3.7
12.1
The solution used in these experiments is a sodium aluminate liquor. The concentrations of total alkali, caustic alkali and A1203 in the solution are about 239g/L, 221g/L and 125g/L respectively.
where Na2Ok / A1203 is the mass ratio of Na2Ok and A1203. The relative digestion rate of A1203 is expressed as:
UIS)o
Roasting pretreatment experiment The graphite crucible (Φ20 χ 100mm) containing aliquots of sample was placed into the intense magnetic field generating device which was raised to the desired magnetic flux density ranging from 1 to 12T. Afterwards, the sample was heated. The temperature ranged from 300 to 600 °C. Once the roasting temperature was reached, the sample was left in the device for a certain time ranging from 5 to 60 minutes. After the roasting time elapsed, the device was cooled down to 150 °C. When the magnetic flux density was low enough, the crucible was removed from the device. The intense magnetic field roasting device is shown in Figure 2.
where (A/S)0 and (A/S)r are the ratio of alumina to silica in raw ore and red mud, respectively. Results and discussion Effect of roasting temperature The effect of roasting temperature ranging from 300°C to 600°C on the digestion performance of treated ore was evaluated. It can be seen from Fig.3 that digestion rate of alumina increased and then decreased with increasing temperature. 550°C is considered as the optimum roasting temperature where the caustic ratio, the real digestion rate and the relative digestion rate were 1.39, 84.17% and 95.08%, respectively. It would require over 220°C for the initial bauxite to get the same level of digestion rate. When fixing the digestion temperature at 190°C, the digestion rate of the roasted ore increased by 52.38%, and the caustic ratio decreased by 0.99, in comparison to that of the raw ore which displayed a relative digestion rate of 31.32%. Consequently, the optimum roasting temperature in the intense magnetic field is 550°C.
170
Fig.2 Intense magnetic field equipment (12T) Digestion experiment The digestion experiment was carried out in a WHFS-1 autoclave. Digestion time was 60min, and the content of CaO addition was 5% by weight. After digestion, the digested slurry was separated into solution and red mud byfiltration.The concentrations of A1203 and Na 2 0 K (concentration of caustic soda) in solution and the content of A1203 and Si0 2 in the red mud were determined by chemical analysis. The calculated formula for the molar ratio of the digestion solution is: 1 *AC
Na
2°k
a, = 1.645 x- Al203 k
Roast i ng t enrper at ur el °C
Fig. 3 Effect of intense magnetic field roasting temperature on digestion performance of diaspore
(1)
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Fig. 5 Effect of magnetic field intensity on digestion performance of diaspore
Fig. 4 Effect of intense magnetic field roasting time on digestion performance of diaspore
Mechanism analysis of pretreatment of intense magnetic field Diaspore begin to dehydrate at 500°C, which leads to the breakage of the crystal chain structure and the replacement of the coordinated octahedron of Al(0,OH)6" into a suspending state. Reaction activity of this structure is clearly much stronger. The Al3+ in the centre of the octahedron is exposed because of removal of OH" and O2" through roasting pretreatment. Because the crystal structure of the diaspore is significantly degraded, digestion performance of bauxite improved measurably. The crystal lattice of alumina transforms from the Al(0,OH)6" octahedron of a Pbnm orthorhombic structure, to an A106~ octahedron of R3c trigonal structure under the condition of roasting temperature of 550°C and roasting time of 60 minutes. The digestion conditions for pure corundum are much more demanding than that of diaspore, as the activity of the former is relatively low. Since a great deal of energy is needed to support the transformation of diaspore into pure corundum, the perfect crystal structure of pure corundum could not be generated before the roasting condition achieved. We infer that the alumina phase of the roasted ore is not a pure corundum but imperfect crystallized alumina in the transition state at roasting condition of 550°C, it is extremely beneficial for improving digestion performance, because a highly activité alumina of transition state exists in the roasted ore. XRD patterns of bauxite roasted at different temperatures are shown in Fig.6. Roasting pretreatment could improve the surface form of bauxite because of the appearance of porosity and crack, which enlarges the specific surface as measured by the method of nitrogen adsorption. The results of the specific surface of the roasted ore are shown in table 3, indicating that specific surface area reaches a maximum at a roasting temperature of 550°C, and above which, the value declines gradually with increasing roasting temperature. The increase of specific surface area enlarges the contact area between ore and liquid. So the reaction dynamic conditions are improved, and the leachability of the sample is enhanced[6]
Effect of roasting time How the roasting time varying from 5 to 60 minutes impacts on the digestion rate of roasted bauxite was investigated. The diaspore was roasted in the magnetic field with intensity of 6T and roasting temperature of 550°C. The digestion condition was 190°C for digestion temperature, 60min for digestion time, 3.1 for caustic ratio of mother liquor, 220 g/1 for caustic concentration and 5% for lime addition amount, respectively. Experimental results were shown in Fig.4. The digestion rate of alumina improved rapidly with temperature increase, which reached a maximum of 95.08% at a roasting time of 60min. Meanwhile, the caustic ratio of digested liquid achieved a minimum of 1.39 under this condition. Thus, 60 minutes is the optimal roasting time. Effect of magnetic field intensity The effect of magnetic field intensities, namely IT, 3T, 6T, 9T and 12T, on the microstructure characteristics and digestion performance of diaspore at roasting temperature of 550°C and roasting time of 30min, was investigated. The digestion condition was 190°C for digestion temperature, 60min for digestion time, 3.1 for the caustic ratio of mother liquor, 220 g/1 for caustic concentration and 5% for amount of lime addition, respectively. The experimental results are shown in Fig. 5. The digestion rate of alumina achieves a maximum, while the caustic ratio of the digested liquid reaches the lowest value, when the magnetic field intensity is 6T. Above this value, digestion rate decreases with magnetic field intensity. The reason might be that the inner crystal structure of bauxite completely transformed when the magnetic field intensity is up to 6T. With the intensity going up, the orientation of the inner molecular structure is stable and the degree of induction of various compositions to the magnetic field goes to the extreme.
35
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A-rawore ; B-300°C ; C-450°C ; D-500°C ; E-550°C ; F-600°C
Table III Effect of roasting temperature on surface area of diaspore (6T) Roasting temperature/°C
Raw ore
300
450
500
550
600
Surface area/m2-g_1
1.6205
8.7339
8.4351
38.7462
49.3558
51.0172
36
1(X
84.17% and 95.08%, respectively.
Because the magnetic susceptibility and the dielectric constant vary with specific minerals, the mineral phases in the ore will be influenced in different degrees by intense magnetic field during the roasting pre-treat process. On one hand, the value of Gibbs free energy of each phase will change, resulting in an impact on the stability of the mineral phase. On the other hand, an intense magnetic field also has an effect on the kinetics of phase transformation. The morphology of mineral phases with different magnetic properties would alter, which will affect the mineral structure and properties.
Acknowledgement This research was supported by the National Natural Science Foundation of China (NO. 51004033) References 1 . CHEN Wankun, PENG Guancai, Intensified Digestion Technology ofBauxite(Beijmg: China Metallurgy Industry Press, 1998), 112-116. 2 . WANG Yiyong et al, "Effects of Microwave Roasting on Leaching Behavior of Diaspore Ore", The Chinese Journal of Process Engineering, 2(2004), 317-321. 3 .R A Hind, K S Bhargava, C Stephen, "The Surface Chemistry of Bayer Process Solids: a review", Colloids and Surfaces A: Physicochemical and Engineering Aspects, 1-3(1999), 359-374. 4 . Guo Xianjian et al, "Microwave-assisted digestion of diasporic bauxites", Nonferrous Matels, 4(1995), 55-57. 5 . Zhao Qingjie, "Investigation of new digesting process of diaspore", Light Metals, (1)2000,17-21 6 . Lv Guo-zhi et al, "Roasting pretreatment of high-sulfur bauxite and digestion performance of roasted ore", The Chinese Journal ofNonferrous Metals, 9(2009), 1684-1689. 7 . Wang Xi'ning et al, "Effects of Magnetic Field on Solid Phase Transformation", Materials Review, 2(2002), 25. 8 . C C Koch, "Experimental Evidence for Magnetic or Electric field Effects on Phase Transformations", Mater Sei Eng, 2000, 213.
For the roasting of bauxite, proper temperature increases the internal energy, accelerates the molecular thermal motion and leads to lattice distortion and phase transition. So during the process, the energy of the magnetic field will have great influence on specific magnetic species and on the direction of thermal motion of molecules. It increases the internal active energy, the degree of phase instability, the lattice distortion and makes the phase transition less complete, so the leaching ability of the ore is enhanced. When the intensity of the magnetic field reaches 6T, the temperature field and magnetic field interact with each other. The digestion rate decreases with the further increase of temperature. There are several possible reasons to explain this phenomenon. When the intensity of magnetic field is too high, the atoms and molecules will be oriented in a certain range according to the direction of the magneticfield[7~8].So this will hinder the lattice distortion and the transitions of the crystal structure. This decreases the activity of roasted ore and leads to a decrease in digestion rate. So there should be an optimum magnetic intensity, rather than an infinitely high value. Conclusions (1) The digestion performance of diasporic bauxite is enhanced effectively by a roasting pre-treat in the presence of a magnetic field. After pre-treatment under appropriate conditions, the observed leaching temperature decreases by at least 30°C compared with the raw ore. (2) The mechanism of the effect of intense magnetic field in the pre-processing is that the influence of magnetic field varies with mineral phase, due to the changes of magnetic susceptibility and dielectric constant for different minerals. It will affect the value of Gibbs free energy of different phases as well as the stability of the ore. When the intensity of magnetic field is high, the atoms, and molecules will be oriented in a certain range according to the direction of the magnetic field. In higher ranges, the activation effect is weak. (3) The optimal conditions of intense magnetic field pre-treatment are determined as temperature of 550°C, roasting time of 60 minutes, magnetic intensity of 6T. When the roasted ore was digested in the sodium aluminate solution with a mole ratio of 3.1 and alkali concentration of 220g/L at 190°C for 60 minutes, the mole ratio of digestion solution decreases to 1.39 and the absolute and relative digestion leaching rates are
37
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
Light Metals 2011 ALUMINA and BAUXITE Bayer Process I SESSION CHAIR
Peter-Hans ter Weer TWS Services and Advice Netherlands
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
APPLICATION OF OPERATION INTEGRITY MANAGEMENT IN THE ALUMINA INDUSTRY Carlos Suarez1, Daniel Welshons1, John McNerney2, Jim Webb2 ^atch Associates Consultants Ine; 1600 West Carson Street, Pittsburgh PA 15219, USA 2 Warren - Forthought Ine, 1212 North Velasco, Suite 207 Angleton, Texas 77515, USA Keywords: Process Safety, Operations Integrity Management, Process and Operations Improvement strategy and implementation of such a system in an alumina refinery environment.
Abstract In today's economic environment the Safety of our industry assets - People, Equipment and Processes - have become even more demanding. This paper provides means to apply Operation Integrity Management System (OIMS) to the Alumina Industry. It discusses what OIM is about, its implementation parameters such as cost and time, as well as an effective way to integrate such a system into refineries day to day operations. A series of Key Process Performance Indicators (KPPIs) are presented as well as an Electronic Knowledge Support System (EKSS) Mockingbird ® as a tool to support the implementation of OIMS.
What is OEM Operational Integrity Management (OIM) consists of an integrated set of theories, practices and techniques for ensuring that an industrial facility operates "with integrity", i.e., that the facility's performance is what it should be - no more, no less. A Brief History of Integrity Management The path of integrity management systems began in the United States, where in 1992 the Occupational Safety and Health Administration (OSHA) introduced its Process Safety Management Program. The goal of its 14 strands was to prevent or minimize the consequences of catastrophic accidents caused through the release of chemicals. The program requires a holistic approach that integrates technologies, procedures and practices, creating multiple barriers of protection.
Introduction Alumina refinery operators can use Operations Integrity Management Systems (OIMS) to improve their Process Safety and Operations Performance. Throughout this paper the authors will refer to Process Safety Management (PSM) as an equivalent system adopted by the Highly Hazardous Materials chemicals and petrochemicals operations.
For the oil and gas sector, the impetus behind the development of integrity management systems came from the Piper Alpha disaster in 1988, in which 167 people lost their lives. The inquiry into the disaster produced the Cullen Report, where the primary recommendation was that operating companies should be required to implement safety management systems that ensure safe design and operation of offshore installations.
Alumina refinery safety in the USA is mainly regulated by the Mining Safety Health Administration (MSHA). This branch of the Department of Labor focuses on employee's health and safety aspects related to surface and underground mining operations. The Occupational Safety and Health Administration (OSHA); another branch of the Department of Labor, regulates chemical and petrochemicals operations. It is important to mention this difference since OSHA regulates process safety activities and systems through PSM while MSHA does not. It is up to alumina refineries management to adopt PSM or OIMS elements to cover the safety of their process, their employees and their surrounding communities.
The report specified that such system should draw on quality assurance principles similar to ISO 9000. The Cullen Report's recommendations were accepted immediately by the British government and the new regime that resulted has influenced the development of integrity management systems around the world. The following have been reported as precursors to the Piper Alpha incident:
As the different elements of OIM are presented you will realize that a lot of them already form part of your refinery ISO, Health, Safety and Environmental systems. This makes it easier for the integration, implementation, cost and time reduction to deploy such a system.
• • •
Corporate Pride and Craftsmanship Complaisency rather than Competency Change to the Rules
The initial response by the industry included the following:
The success in implementing OIMS depends among other things on how quick the refinery culture embraces and supports it.
• • •
Once the elements are in place, the deployment, auditing and further improvements can be enhanced by using an Electronic Knowledge Support System. Mockingbird ® and its suite of applications have been successfully used by the chemical and petrochemical industry over the years. This paper discusses the
•
41
Mitigation of smoke hazards Installation of sub-sea pipeline isolation systems Improvements to the "Work Permits" management system Relocation of some pipelines emergency shutdown valves
that they perform in a manner that is consistent and compatible with Company's policies and business objectives. Evaluate and train contractors before starting any work.
The Elements of OEM Below is a brief description of the OIM elements shown in figure 1. 1.
Management establishes policy, provides perspective, sets expectations and provides the resources for successful operations. Assurance of Operations Integrity requires management leadership and commitment visible to the organization, and accountability at all levels. The adoption of such system is driven by the General Manager and his/her team of Functional Managers.
2.
Comprehensive risk assessments can reduce safety, health, environmental and security risks and mitigate the consequences of incidents by providing essential information for decision-making. Mitigating risks at the early stages of design and the continuation in the operation phase is a must.
3.
Inherent safety and security can be enhanced, and risk to health and the environment minimized, by using sound standards, procedures and management systems for facility design, construction and startup activities.
4.
Accurate information on the configuration and capabilities of processes and facilities, properties of products and materials handled, potential Operations Integrity hazards, and regulatory requirements is essential to assess and manage risk. Process Safety Information such as P&IDs, Materials Safety Data Sheets should reflect current operation.
5.
6.
7.
8.
9.
10. Effective management of stakeholder relationships is important to enhance the trust and confidence of the communities where the business operate. Emergency planning and preparedness are essential to ensure that, in the event of an incident, all necessary actions are taken for the protection of the public, the environment and company personnel and assets. Make it a point to run simulations and drills periodically so that everybody involved react according to plan. 11. Assessment of the degree to which expectations are met is essential to improve Operations Integrity and maintain accountability. Involve all levels of the organization in routine audits. Make it a point to discuss results and focus on corrective actions. OIMS11 Elements Driver
Operations
Evaluation
2. Risk Assessment and Management 1. Management Leadership. Commitment and Accountability
Control of operations depends upon people. Achieving Operations Integrity requires the appropriate screening, careful selection and placement, ongoing assessment and training of employees, and the implementation of appropriate Operations Integrity programs. Apply Management of Change when dealing with new personnel assignments, particularly in the plant areas.
3. Facilities Design and Construction 4. Information / Documentation 5. Personnel and Training 6. Operations and Maintenance
11. Operations Integrity Assessment and Improvement
7. Management of Change 8. Third Party Services 9. Incident Investigation and Analysis 10. Community Awareness and Emergency Preparedness
Operation of facilities within established parameters and according to regulations is essential. Doing so requires effective procedures, structured inspection and maintenance programs, reliable Operations Integrity critical equipment, and qualified personnel who consistently execute these procedures and practices. Make sure that Standard Operating Procedures for both Process Operations and Maintenance reflect current systems. Changes in facilities, or managed to arising from level.
Effective incident investigation, reporting and followup is necessary to achieve Operations Integrity. They provide the opportunity to learn from reported incidents and to use the information to take corrective action and prevent recurrence. Monitor and request reports from any incident, no matter how small or insignificant to the perception of the affected.
Ret·, e« mUoM
Figi. OIM 11 Elements Refinery Culture and How to Promote OIM Cultural Attributes 1. Culture is a feature of the entire organization, not just of some of the individuals within that organization. Therefore, if someone — even the general manager — leaves the organization, the culture of that organization should not change significantly.
operations, procedures, site standards, organizations must be evaluated and ensure that Operations Integrity risks these changes remain at an acceptable
2. Culture is on-going — it is not a one-time event. A facility in which everyone is continuously striving to identify and correct problems and to eliminate hazardous conditions has a strong operational integrity culture, whereas a facility which makes only
Third parties carrying out work on the company's behalf impact its operations and its reputation. It is essential
42
Loss of market share is reduced — After an incident, this loss continues until the company's reputation is restored. Adverse publicity and negative public image can have insurmountable effects
spasmodic and irregular efforts to improve such conditions does not. 3. In a strong OIM culture there is minimal disconnect between words and actions. All managers and workers 'walk the talk'; their words and deeds match.
Litigation costs are reduced — These are unavoidable after an incident and can total five times the cost of the regulatory fines.
4. The creation and maintenance of an organizational culture requires leadership from the top. Allowing lower level employees to "do their own thing" does not create a culture.
Incident investigation costs are reduced — Investigating an incident and implementing corrective actions can cost millions of dollars
5. It is difficult for any organization to truly assess the quality of its own culture. It takes an outsider to truly evaluate the quality of a company's culture. Therefore, an organization with a strong OIM culture will make frequent use of outside auditors, inspectors and reviewers to identify areas of weakness and to suggest corrective actions. Moreover, the auditors' reports will go directly to the facility managers
Regulatory penalties are reduced — For many incidents, a fine after litigation can total 1 million dollars or more Regulatory attention is reduced — A major incident usually results in increased regulatory audits and inspections Key Performance Indicators
6. A strong OIM culture is one in which employees and contract workers feel free to report on difficulties and problems, even if those employees and workers are potentially opening themselves up to criticism
Examples of KPIs for Operation Integrity Management are shown infigure2 below.
7. With regard to SHE (Safety, Health and Environmental) issues, the organization places excessive emphasis on the safety term, to the detriment of the health and environmental elements.
% P&IDs Conformance to Current Process Installations As Built Drawings Available and in Conformance Facilty Design and Built per Sound Standards % of Critical Operational and Maintenance Procedures Reviews Completed to Schedule % Compliance with Critical Procedures Material Safety Data Sheets (MSDSlAvailabilityto^^ Mandatory Training Completed to Schedule Personnel Trained on New Standard OperatingProcedures
Prepare and publish a Mission Statement that spells out the organization's stated commitment to operational integrity management principles.
2.
Develop guiding tenets that show how the OIM program is to be implemented.
3.
Develop a detailed program showing how the guiding tenets are to be achieved.
S of Risk Assessment Corrective Actions Completed to Schedule % of Risk Assessments Reviewed to Schedule
man—-
8. A strong operational integrity management culture adapts to new circumstances without its basic values being affected by issues such as economic downturns or the adoption of new technologies. It is suggested here that management can go about creating a strong operational integrity culture by following the three steps shown below: 1.
Leadership Participation in Incidents Investigation Participation in OIM Program Assessment OIM Contribution as Part of the Employees Performance Assessment
EEEEBEEEEEEa All Critical Controls for Process Safety Identified \ of Controls Inspected to Schedule ft of Controls Outside Tolerance % Compliance with Critical Procedures
% of MOC Documents Compliant with Procedures % of Temporary Changes Overdue % of MOC Physically Installed but Awaiting Completion of Documentation
llll Γ
^ ^ ^ ^ ^ ^ Μ ^ — ■
Assessment of Capabilities to Performed Work Deficiencies Corrected Effective Communication
Ü
OIM provides unparalleled capacity for enhanced risk reduction. A company's risk exposure is reduced in the following areas when well-founded process safety systems are in place.
% of Overdue Incident Investigations No of Repeat Incidents Occuring % of Follow Up Corrective Actions Completed to Schedule 1 Lessons Learned from Company and Industry Incidents
a ■ BE ',- m i mm jjuLlma-
No of Emergency Exercises /Desktop Exercises Completed to Schedule Emergency Plan Reviewed to Schedule
Lives are saved and injuries are reduced — Both the personal impact of human loss and cost of deaths or injuries are painful. A solid OIM program can help prevent these costs
% of Inspections or Tests Completed to Schedule % Compliance with Standards and Procedures
Figure 2 - KPIs for OIM Elements
Property damage costs are reduced — In the U.S., major industrial incidents cost an average of $80 million each
OEM Implementation Parameters
Business interruptions are reduced — These losses can amount to four times the cost of the property damage from an incident
43
To succeed in business a company must: • • •
Protect its license to operate Meet ever more demanding regulatory requirements Manage the sustainability of your business
• •
•
Raise stakeholder and public confidence Minimize and, where practicably possible, eliminate the risk of incidents
The cost areas for years 1-5 includes (1) the remaining cost to reach 100 % compliance and (2) the ongoing cost to maintain compliance (or quality) for the remaining years if the company reaches 100 percent compliance in that period.
How to Implement an OIM System • • • • • • • •
The cost for developing the program is described below:
Assign an OIM Manager (or team) Learn from the literature (check key references) Learn by training (from process safety professionals) Learn from other companies - align; network; participate in industry alliances Note strong synergies with ISO, TQM, RMP, Responsible Care Set some clear OIM goals (one tofiveyears) Track performance versus the goals on a regular basis Reassess OIM/plan & modify (every 3-5 yrs)
Developing an OEM Program. The cost, primarily in equivalent labor costs, to bring the OIM program (and individual element programs) from the concept stage through the final design (such as developing an MOC or MI written program that the facility personnel are confident will work). This category also includes the cost of training personnel to be proficient in various OIM activities, such as leading PHAs, leading incident investigations, leading compliance audits, writing procedures, and leading employee training.
The best OIM companies show the following attributes: •
OIM Champions who affiliate themselves with multiple disciplines (e.g., EHS, Engineering, Operations, Insurance) to work collaboratively
•
Functionally having two platforms to identify, analyze, select, implement, control & monitor process; i.e., worst case (top down) and more frequent events (bottom up)
•
Regularly conduct reviews of the OIM program against the defined elements, and
•
Implementing an OEM Program. The cost (again primarily in equivalent labor costs) to do implementation tasks, such as writing operating procedures, updating PSI, doing initial training of operators and maintenance personnel, and performing/documenting Process Hazard Analysis - PHAs. Responding to Recommendations. The cost, primarily capital costs and expenses, to implement improvements to address recommendations from PHAs, MOC hazard reviews and incident investigations. As a rule of thumb $22,000 - $25,000 per P&ID might be used as the cost associated for the development and implementation of an OIM system.
Especially promote risk engineering in the conceptual engineering design phase.
Mockingbird ® (EKSS)
Time
Mockingbird ® is an Electronic Knowledge Performance System widely used by the Chemical and Petrochemical industry to support OIM / PSM systems development, implementation and manage compliance.
Research has shown that the following level of compliance can be attained after the implementation of an OIM system: • • • •
Initial implementation
40% level of compliance First Year (Baseline) 50% level of compliance Second Year 100% level of compliance anticipated in Fifth Year Excellence in OIM program anticipated Seventh Year
The initial population of the system takes place during the "Miracle Month". This period is used for training on the system, designing its structure and have all involved take ownership. Mockingbird ® becomes the portal for all OIM elements and its tools.
Time is dependent on how many documented similar elements the business already have in place in the organization that can be brought into OIM or easily adapted.
.incMocKingbirdd
Cost
MocKingbird® lTts^u
The labor cost for developing and implementing an OIM element can be accounted for in one or more of the following categories: Meetings Writing Reviewing Revising Training/Orientation Pilot testing More revising
44
Conclusions •
~ΤΛ, , ·ι · ι ,Λι · r * OIM can be easily implemented m any alumina refinery for which other systems such as ISO, Safety, Health and Environmental already exists.
•
OIM can be a combination of the elements listed in this paper and those defined for PSM
•
The successful implementation and continuity in the use of OIM depends on the organization commitment to support it
•
It is important that a strong safety culture be promoted and nurtured
•
The existence of other systems such as ISO, Safety, Health and Environmental, etc will make the OIM system implementation less costly and shorter in time
•
Refineries that reach an excellence level of compliance with OIM will also show an overall improvement in financial and operating levels
•
Mockingbird ® and its application suites offer a sound and robust platform on which to manage OIM
CCPS, 'Guidelines for Auditing Process Safety Management Systems", -g ' .- β *. „ (μ« 169-0556-8 ' * '
Further Readings Exxon Management Systems h{tp;//w\vw.
Exxon Mobil, Operations Integrity Management System CCPS, "Guidelines for Technical Management of Chemical Process Safety", 1989, ISBN No, Ö-8169-0423-5. CCPS, "Plant Guidelines for Technical Management of Chemical Process Safety", 1992, ISBN No. 0-8169-0499-5. American Petroleum Institute, Recommended Practice 750 "Management of Process Hazards", 1990, reaffirmed 1995. Available via the API at www.api.org Ü.S. Occupational Safety and Health Administration, 29-ŒR-1910.119, "Process Safety Management of Highly Hazardous Chemicals". Canadian Chemical Producers' Association (CCPA), Responsible Careen codes of practice. Available on the CCPÄ website www.ccpa.ca American Institute of Chemical Engineers, "Daw's Fire and Explosion Index Hazard Classification Guide", latest edition, ISBN No. 0 4 1 6 9 4 6 2 3 - & American Iristttute öf Chemical Engineers, "Dow's Chemical Exposure index", latest edition, ISBN No. 0-81694647-5. Kietz, T.Ä., "An Engineer's View of Human Error", London; The Institution of Chemical Engineers, third edition 20Û1, ISBN No. 1 -5603-291Ö-5 (available from CCPS). CCPS, "Guidelines for Preventing Human Error in Process Safety", 1994, ISBN No. Ö4169446V8. Health and Safety Executive (UK), "Human Factors in Industrial Safety", HS(G)48, London HMS0,1989, ISBN No. 0-11-6854864.
45
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
INFLUENCE OF SOLIDS CONCENTRATION, PARTICLE SIZE DISTRIBUTION, pH AND TEMPERATURE ON YIELD STRESS OF BAUXITE PULP Carla Napoli Barbato1'2, Silvia Cristina Alves Franca2, Marcio Nele de Souza1 1-Escola de Quimica/ Universidade Federai do Rio de Janeiro- Av. Horacio Macedo, 2030. Edificio do Centro de Tecnologia, Bloco E. Cidade Universitaria. Cep: 21941-909. Rio de Janeiro - RJ 2-Centro de Tecnologia Minerai- Av. Pedro Calmon, 900. Cidade Universitaria. Cep:21941-908. Rio de Janeiro- RJ Keywords: yield stress, zeta potential, van der Waals forces Thus, the purpose of this work is to study the effect of solids concentration, particle size distribution produced by different grinding time, pH and temperature on the yield stress of bauxite pulp.
Abstract In Northern Brazil, bauxite pulp is transported through pipelines to the plant where alumina is produced. In slurry transportation through pipelines, knowing the yield stress value is essential for the pumps and pipeline design. Yield stress is the minimum shear stress and corresponds to the first evidence of flow. This rheological property is influenced by factors, such as: particle form, temperature, particle size distribution and interaction among the particles. Within the context above, the objective of this work is to verify the influence of solids concentration, particle size distribution produced by different grinding time, temperature and pH on the yield stress of bauxite pulp. It was verified that the yield stress of bauxite slurry increases as solids concentration and grinding time go up and decreases with temperature and pH.
$&mpê~Açti Cap*ft*maO
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Introduction
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Brazil is a country with large territory and its mining activity usually takes place far from the production or consumption centers. Ore transportation by rail or road is expensive and lacks adequate infrastructure, which makes its transportation in the form of pulps through pipelines an attractive alternative.
J*
*
-
The first bauxite slurry pipeline in the world is located in the North of Brazil and it aims at transporting the aluminum ore from the processing plant, located in Paragominas (PA), to the alumina production plant in the municipality of Barcarema (PA) through a 244 km long pipeline [1]. The mine and the pipeline are located in the State of Para, Brazil, as shown in Figure 1.
φ Figure 1. Map of Brazil and pipeline showing location of Mine and alumina plant.
In slurry transportation through pipelines, it is important to get information on yield stress in order to carry out the pumps and pipeline project. Furthermore, high values of yield stress can retard solid particles from gravity settling for long periods of time [2].
Materials and Methods The bauxite used in this work comes from North of Brazil and it is basically composed of the minerals gibbsite and kaolinite. The dressing process followed the methodology described by [6] and the steps consisted of crushing and grinding.
Yield stress is the minimum value of shear stress and is the first evidence of flow, i.e., the value of shear stress when the velocity gradient tends to zero [3,4].
The particles size distributions studied were based on industrial conditions in pipeline transportation of bauxite located in Para State, in which the particle size distribution is 6% plus 0.208 mm and 40-47% minus 0.043 mm [1]. The particle size distribution of the bauxite sample ground for 30, 35 and 40 min in a rod mill was determined through wet screening with the use of a set of Tyler sieves with openings from 1.2 to 0.037 mm.
Yield stress is related to the internal structure of the suspension. For this structure to begin to flow it is necessary to overcome the minimum shear stress necessary to break the interparticle contact and the unions of aggregates or floccules. This structure is formed by particles whose electric charges are of opposite signs. The interparticle force of attraction of this structure is of the van der Waals type. The variables which influence yield stress are: solids concentration, particles size, shape, pH and the nature of the materials [5].
The zeta potential was determined in a DT 1200 equipment manufactured by Dispersion Technology. The suspensions were prepared with 10% bauxite and 0.01 M KC1. The zeta potential
47
measurements were carried out in a pH range of 2.0 to 12.5. The pH was adjusted with diluted solutions of KOH and HC1
Figure 2 illustrates the size distribution of bauxite ground for 30, 35 and 40 min, obtained through wet screening with a set of Tyler sieves with openings from 1.2 to 0.037 mm. It was verified that the bauxite ground for 30 min presented a size distribution similar to that of the bauxite transported in the pipeline located in the North of Brazil. The bauxite ground for 40 min had a higher amount of fine particles (<0.037 mm) than the bauxite ground for 30 and 35 min, at approximately 13 and 6%, respectively.
In order to study the influence of solids concentration, particles size distribution produced by different grinding times, temperature and pH on the yield stress of bauxite slurry, a matrix of experiments 24"1 around mean values was performed (Table 1). The solids concentration studied was 50 and 60% (w/w), the grinding time of bauxite was 30 and 40 min, the pH studied was 7 and 12 and the temperature was 25 and 45°C.
io(H
Table 1. Experimental conditions following matrix experiments 24"1 around mean values.
90
T GT SC pH (min) (°Q (% w/w) 1 25 30 7 50 2 45 30 60 7 3 45 40 50 7 4 25 40 60 7 12 5 45 30 50 12 6 25 30 60 7 25 40 50 12 12 8 40 60 45 9 35 35 55 9.5 10 35 35 55 9.5 11 35 55 9.5 35 T- Temperature; GT- Grinding Time; SC- Solids Concentration Run
80
60
0.1
K=
*D,
Figure 2. Size distribution of bauxite ground for 30, 35 and 40 min obtained through wet screening. The zeta potential is useful to evaluate the electrical double-layer repulsive forces among particles in suspensions, as charge density and the potential of the particles depend on pH and ionic strength. Figure 3 illustrates the zeta potential of bauxite in relation to pH. It can be observed that the point of zero charge occurs at pH = 10.5. At this pH the degree of particle flocculation is maximum. In the pH range between 7 and 11, the zeta potential is small. It is expected that in this range particle flocculation is high because the density of charge on the particles surface is small, which results in high attraction due to van der Waals forces. At pH 12 the zeta potential is higher, which means that the density of charge on the particles surface is large and the repulsive forces among particles are greater than the attraction forces of van der Waals and, consequently, the degree of flocculation is lower.
(1)
H J_ A,
+
(2)
3
A suspension will be stable when the repulsive forces dominate while the presence of strong attractive forces will cause particle aggregation. Thus, the suspensions prepared at pH 12 are more stable than the suspensions prepared at pH 7 and 9.5.
Tm - measured torque Dv - vane diameter xy = shear stress H= vane height Equation 03 was used to adjust the yield stress values of the bauxite pulps, having as its variables: temperature, grinding time of bauxite, solids concentration and pH. Y=
a0+£aiXi+'£aijXiXj
1
Particle Size (mm)
Measurements of yield stress were obtained from an ARES rheometer, manufactured by TA instruments. A geometry Vane was used. The tests were performed at 1 rpm for 200 s. Torque values were converted to shear stress by means of Equations 1 and 2. Yield stress is the maximum shear stress obtained at low speeds.
-TJK
Bauxite ground for 30 min Bauxite ground for 35 min Bauxite ground for 40 min Bauxite transported along the pipeline
(03)
With Y as the dependent variable (yield stress), Xi the independent variables (temperature, grinding time of bauxite, solids concentration, and pH) and ^ and a^ as the parameters. Results and Discussion
48
H
Figure 3. Zeta potential of bauxite particles.
"
■
The model parameters for yield stress used normalized independent variables [+1,-1] so that the values of the parameter could be associated with the variable effect. Estimates for model parameters were obtained with standard linear regression procedures [7]. Standard statistical tests of significance (t-test of student) were used to allow the evaluation of parameter significance. Whenever parameter significance was smaller than 5%, the parameter and respective effect would be removed from Equation 3. Regression results obtained for the yield stress determination are shown in Table 2.
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1
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7
8
1
9
1 10
1 11
1
12
1
13
1
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Experimental Run
1.964 + 0.064 1.232 + 0.082 0.752 ±0.082 -0.038 + 0.082 0.073 +0.082 0.448 +0.082 0.188 ±0.082 -0.023 ±0.082 0.946
Figure 4. Comparison between (■) observed and (O) predicted yield stress.
aoH aT asc*GT asc*oH asc*T R' SC-Solids Concentration; GT - Grinding Time and T-Temperature
Figure 5 illustrates the values of yield stress from the bauxite slurry obtained under different conditions of solids concentration, grinding time, temperature and pH. It was verified that the experimental conditions that favored a higher value of yield stress of the bauxite pulp was 60% (w/w) solids, bauxite ground for 40 min, 25°C and pH 7.
** Significant effects in bold 5-
It is possible to observe from the correlation coefficient (Table 2) and the model fitting (Figure 4) that the model gave a good fit to the experimental parameters. The value of R2 is larger than a confidence interval of 95% indicating a very good fit. The model relating the yield stress (τ0) and the statically signification preparation variable is:
4-
3CD CO
τ0 = 1.964 +1.232 *SC + 0.752 *GT- 0.448 *SC*GT
(04)
m
Fh
2-
τ
IiI I -
It may be concluded, by observing the parameters values in Table 2, that the yield stress depends mainly on solids concentration and the grinding time of bauxite and these variables show a direct relationship to yield stress. The remaining variables, pH and temperature, are not statistically significant. This means that the effects of temperature and pH are negligible compared to the yield stress measurement error.
ΓΤ 1
m
m
T
m
m
n ..
1
2
3
4
5
6
7
Firn 10
m
11
Experimental Run
Figure 5. Values of yield stress of bauxite pulp obtained under different conditions of solids concentration, grinding time, temperature and pH. It can be stated that yield stress increases when solids concentration and grinding time of bauxite increase and when temperature and pH decrease. This behavior can be explained by the following: 1. The increase of solids concentration results from the decrease of the water layer between particles, so the inter-particle distance is smaller in a denser slurry which produces an increased attractive potential and a larger probability of collisions among particles, resulting in more particles attracting one another and a high close-packing of particles [8,9,10].
2. The increase of bauxite grinding time allows a bigger production of fine particles (< 0.037 mm) which accentuates the strong particle-particle interaction due to the van der Waals forces of attraction, forming floes or aggregates of low packing density and the formation of chain or networks of particles (Figure 6) that reduces the fluidity of the suspension. For the suspension to flow, it is necessary to apply a greater force to break the packing particles network, such as that shown in Figure 7. In this Figure the state of the network of very voluminous particles packing is similar to a "sponge".
The yield stress of the bauxite slurry increases with increase of solids concentration, grinding time of the bauxite and with decreasing temperature and pH. Acknowledgements The authors would like to thank CAPES for their financial support, the School of Chemistry / Federal University of Rio de Janeiro and the Center for Mineral Technology. References
3. When the pH is between 7 and 11, the zeta potential is low which means a high charge surface density and strong van der Waals attraction forces among the particles and the formation of floes. At pH 12, the zeta potential is higher and the yield stress is low because the net attractive force decreases due to the increase of the surface charge density. When this repulsive force exceeds the van der Waals attraction at a pH far from the IEP, dispersed suspensions are usually generated (Figure 7).
1. Gandhi, R. Weston, M. Talavera M. Brittes G. P. & Barbosa E."Design and Operation of the World's First Long Distance Bauxite Slurry Pipeline". Light Metals, 2008, 95-100. 2. Dzuy, N.Q. e Boger, D. V., "Yield Stress Measurement for Concentrated Suspensions", Journal of Rheology, 27 (1983), 321349.
4. The yield stress increases as temperature decreases because the viscosity of dispersant phase (water) increases.
3. Liddell, P. V.e Boger D. V., "Yield Stress Measurements with the Vane", Journal of Non-Newtonian Fluid Mechanic, 63 (1996), 235-261. 4. Stokes, J. R. e Telford, J. H., "Measuring the Yield Behaviour of Structrured Fluids", Journal of Non-Newtonian Fluid Mechanics, 124 (2004), 137-146. 5. Alejo, B. and Barrientos, A., "Model for yield stress of quartz pulps and copper tailing", International Journal of Mineral Processing, 93 (2009), 213-219. 6. Silva F. A. N. G. Medeiros M. E. Sampaio, J. A. Santos R. D. Carneiro M. C. Costa L. S. Garrido F. M. S. "Technological Characterization of Bauxite from Parâ-Brazil". Light Metals, 2009,139-144.
Figure 6. Configuration of maximum packing fraction, particles forming networks of particle-particle contact (floes or aggregates) due to the van der Waals forces of attraction.
7. Box, G. E. P., Hunter, W. G., Hunter, J. S., Statistics for Experiments (New York, NY: Wiley, 1978). 8. Nuntiya, A., Prasanphan, S., "The Rheological behavior of kaolin suspensions", Chiang Mai Journal Sience, 33 (2006), 271281.
OOOOO O OOOOOO
9. He M. Wang Y. Forssberg E. "Parameter Studies on the Rheology of Limestone Slurries" International Journal of Mineral Processing, 78 (2006), 63-77.
OOOOO O OOOOO O
10. He, M., Wang, Y. e Forssberg, E., "Slurry Rheology in Wet Ultrafine Grinding of Industrial minerals: a Review", Powder Technology, 147 (2004), 94-112.
Figure 7.Configuration of minimum packing fraction; disperse particles due to repulsive forces.
11. Krester, R. G., Scales, P. J. "The effect of temperature on the yield stress of mineral suspensions". J. Colloid and Interface Sci., 328 (2008), 187-193.
Conclusion A bauxite slurry prepared at pH 12 is more stable than that prepared at pH 7 because the bauxite particles have a higher surface charge at pH 12. Thus, the repulsive forces between particles are higher at this pH. The parameter that has the greatest effect on the yield stress of the bauxite pulp is solids concentration, followed by grinding time, pH and temperature.
50
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
A NEW METHOD FOR REMOVAL OF ORGANICS IN THE BAYER PROCESS Bai Yingwei1, Shen Mingliang1, Li Junqi2, Zhao Fei2 1. Guiyang Aluminum Magnesium Design & Research Institute, Guiyang, Guizhou, 550081; 2. Materials Science and Metallurgy school,Guizhou University, Guiyang. Guizhou, 550003 Keywords: Bayer process, Organics, Super-concentration, Carbonate, causticization refineries with the Bayer process. The impurities, especially Na20 content in the product will be increased, and this is not good for aluminum hydrate settlement, filtration and classification, which results in generating foam in the settlers, reducing the alumina whiteness, and accelerating the aluminum hydrate scaling in the seed precipitation tanks.
Abstract This article introduces briefly the main organics sources in Bayer process alumina production and the harm due to accumulation during process flows, specifies the organics removal methods applied generally in world alumina refineries, and explains briefly advantages & disadvantages of each method. The article stresses the super-concentrated organics removal method developed successfully by GAMI with Guizhou University together, which not only has good organics removal effects (total removal percent of 57-66%), but also effectively removes the carbonates in the process flow.
(3) It also has effects on the evaporation of mother liquor. The organic impurities can make Na2Oc concentration in the mother liquor increase (make Na2C03 in the liquor supersaturated and concentration increased), the organics content is higher and this phenomenon is more obvious. The organics makes the liquor viscosity increase and the size of the precipitated monohydrate sodium carbonate fine, resulting in difficulties with settlement and filtration & separation. The production practice of Guangxi Branch Company of China Aluminum Company estimates that: the oxalate is easy to be precipitated from the evaporator and flash evaporator when the sodium oxalate concentration in the liquor is high, having the effects on the evaporation and increasing the scaling cleaning workload.
1. Introduction The organics in the Bayer liquor arise mainly from the bauxite. A few of the organics are also brought in from the flocculants, defoaming agents and dewatering agents, but with lesser amounts and smaller effects, according to the literature reports. The organic carbon content in the bauxite is generally 0.1-0.3%, but also can be lower to 0.03% or higher to 0.6% (some surface mines). The organic carbon content in tropical bauxite is higher, generally 0.20.4%; however, that in diasporic bauxite is lower, generally 0.1%. The organics content in the bauxite in Poncho, Africa and Australia is higher and that in Europe, Russia and China is lower [1].
(4) The pitch organics often attach to the surface of filter cloths and heat exchangers, which has effects on the filtration and heat exchange. Because the organics have some serious effects on the alumina production process flow mentioned above, alumina refineries in the world adopt varied methods for the removal of organics from the process flow. Two kinds of methods are applied in the actual production: one is to remove the organics from the liquor, mainly by spent liquor calcination, adsorption, deposition etc; the other is to totally or partly damage the organics in the liquor, mainly by the oxidization which is to partly or totally oxidise the organics to Na2C03. At present the main application methods are as follows [2" 61 : adsorption, oxidization, calcination, aluminum hydrate washing liquor settlement, crystallization etc.
There are three main kinds of organics in the Bayer liquor: natural matter (such as roots), humic acids and pitch [1] formed by precipitation of the organics (plants) and microorganisms with the chemical reaction. These organics undergo some changes during the alumina production flow, and finally become humic acids, benzenecarboxylic acids, oxalic acid etc. The organics content in the production process accumulates step by step along with the progress of production, and harms the Bayer alumina production process when it reaches a certain value. The main effects are as follows:
2. Test principle and test method
(1) It has effects on the digestion and red mud separation & washing [1]. Along with the increase of the organics content, the liquor viscosity will increase, the red mud settlement speed will be reduced and the overflow suspended matters in the settler will increase.
2.1 Test principle The super-concentrated causticization method this article puts forward is different from the traditional organic removal methods. The principle is as follows: according to the characteristic that the solubility of oxalate in the alkaline liquor is reduced, along with the increase of the alkali concentration in a certain alkali concentration range, the liquor required for organics removal is concentrated to a higher concentration (super-concentrated) so as to precipitate the most oxalate from the liquor by crystallization,
(2) It has effects on the crystal seed precipitation[l]. The precipitation speed and the alumina output rate shall be reduced, which results in the alumina size becoming fine and fragile, and especially fragile during calcination. This problem becomes one of the difficulties with producing sandy alumina in the alumina
51
then the precipitated oxalate is translated into the insoluble compound by a causticization reaction so that the compound is removed from the process flow by washing and filtration. Therefore, the effect of factors on the solubility of oxalate in the liquor and the optimum conditions of causticization reaction is required to be found by testing, moreover, the authors have also studied the solution condition of the causticization slag in the liquor so that this method can be better applied to the industry production.
Effect of crystal seed addition on organics removal percent The optimum super-concentrated conditions of organics removal are adopted for concentration and slag making. The concentrated slag is dried and milled to be fine for 1-section causticization test, using the orthogonal method. The chemical compounds in the liquor and solid after causticization and separation, is analyzed and the organics removal percent is calculated so as to determine the optimum 1-section causticization conditions.
In the test: Nk expresses the caustic sodium oxide concentration in the liquor (NaAJ g/1
•
The 2-section causticization treatment with different causticization conditions is done based on the 1-section causticization. The organics removal condition after 2section causticization separation is analyzed so as to determine the optimum 2-section causticization conditions.
•
The anti-solution test of the slag obtained after causticization is done so as to study the anti-solution condition of the organics in the caustic slag during the washing.
Nc expresses the carbonate sodium oxide concentration in the liquor (Na2Oc) g/1 NT expresses the total soda concentration in the liquor (Na2Or) g/1 AO expresses the alumina concentration in the liquor (A1203) g/1 Corganks expresses the organics content in the sodium oxalate (Na2C204) in the liquor
3. Test result and analysis
AC organics expresses the precipitation (settlement) percent (%) of C
3.1 Effect of initial organic carbon concentration on organics removal percent
organics
Crystal seed coefficient of sodium oxalate
When the mother liquor (super-concentrated green liquor) (Nk=153g/1; Corganics =2.46-8.0g/l) is super-concentrated to about Nk 300g/l, table 1 shows that the crystal precipitation percent of sodium oxalate is directly proportional to the concentration of Corganics in the green liquor.
Added Na2C204 Na2C204 in the liquor
(mass ratio)
Crystal seed coefficient of sodium carbonate =
Added Na2C204
Table 1 Relation of initial organics concentration and organics removal percent
(mass ratio)
Na2C204 in the liquor
BSC= Before super concentration ASC= After super concentration
The organics removal percent formula can be expressed as the followings: C% =
Item
cox^o-Qxv/ X l o o % c 0 xv 0
No.
Ντ
Conditions
C0 and V0 respectively express the concentration and the volume of organic carbon (Corganics) in the test liquor.
1
Cj and Vj respectively express the concentration and the volume of organic carbon (Corganics) in the concentrated liquor.
2
2.2 Test method
3
•
Result
Effect of initial organic carbon concentration on organics removal percent: initial organic carbon concentration = 2.468.0 g/1
Ao
Nk (g/i)
Ne (g/i)
m
Q
Ne/Ντ
(%)
BSC
163.8
91.35
153.0
15.3
2.46
9.1
ASC
335.4
191.38
314.0
21.4
1.81
6.4
BSC
168.3
91.35
153.0
15.3
4.0
9.1
ASC
331.2
213.35
309.0
22.2
2.51
6.7
52
(%)
66.89
71.39 BSC
168.3
91.35
153.0
15.3
6.0
9.1
ASC
321.2
181.84
303.0
18.2
3.6
5.7
BSC
168.3
91.35
153.0
15.3
8.0
9.1
ASC
312.6
179.27
298.0
14.6
4.31
4.7
73.36
4
Effect of concentrated Nk on organics removal percent: concentrated Nk= 300 g/1. 320 g/1. 340 g/1. 360 g/1. 380g/l
ACcxgnics
75.97
? 76
I
!.. i
£ 74 J Έ
8
72
i n i t i a l c o n c e n t r a t i o n of o r g a n i c C ( g / I )
concentrated N K v a l u e ( g / l )
Figure 1: Relation of initial organic carbon concentration and organics removal percent
Figure 2: Relation of concentrated Nk value and organics removal percent (1)
The organics removal percent has a direct relationship to the initial organic carbon content, and will increase along with the increase of the organic carbon content in the green liquor. There are main two reasons for this: one is that the more organics shall be precipitated by the crystallization after the initial organics content increases, the first precipitated organics itself can be the crystal seed for the later precipitated organics so as to promote the crystallization precipitation of the organics during the concentration and improve the organics removal percent. The other one is that the solubility of the organics in the sodium aluminate liquor has no big change under certain conditions, the organics removal percent must be increased if the initial organic carbon concentration is increased.
concentrated NKvalue(g/l)
3.2 Effect of concentrated Nk on organics removal percent
Figure 3: Relation of concentrated Nk value and solubility percent of carbonate in the liquor
The mother liquor (super-concentrated green liquor) (Nk=136.8/1; Corgnics=l-44g/l) is super concentrated. The precipitation percent of sodium oxalate is increased along with the increase of the superconcentrated liquor Nk value. The precipitation percent reaches the peak value (refer to table 2 and figure 2) when the super concentrated liquor Nk value reaches about 360g/l, moreover, the solubility of the carbonate in the liquor is continuously reduced along with the increase of the concentrated liquor Nk value, the lowest value is here when the Nk is 360g/l, and men it will be little increased along with the increase of the Nk value (refer to table 2 and figure 2).
The mother liquor (super-concentrated green liquor) (Nk=153g/1; Corganics =6g/l) is super concentrated, the study result shows that the precipitation percent of sodium oxalate is increased along with the increase of Nk value of super concentrated liquor under the condition of higher Corganjcs content in the super-concentrated liquor. The precipitation percent reaches a peak value (refer to Table 3 and Figure 4) when the Nk value of super-concentrated liquor reaches about 360g/l.
Table 2: Relation of concentrated liquor Nk and organics removal percent (1)
Table 3: elation of concentrated Nk value and organics removal percent (2)
\jtem
No.
Ντ
Ao
Nk
Ne
(-Organics
NC/NT
ACotgnics
(g/i)
m
m
(g/0
(g/i)
(%)
(%)
BSC
160.4
82.9
136.8
23.6
1.44
14.7
ASC
333.4
172.04
310.0
23.4
1.06
7.0
BSC
160.4
82.9
136.8
23.6
1.44
14.7
ASC
352.6
101.78
335
17.6
0.98
14.7
BSC
160.4
82.9
136.8
23.6
1.44
14.7
ASC
374.2
200.19
364.0
10.2
0.79
2.7
BSC
160.4
82.9
136.8
23.6
1.44
14.7
ASC
399.0
219.27
386.0
13.0
0.91
3.3
^ResiA Condition
1 2 3 4
No.
I
Ντ
Ao
Nk
Ne
Corgarks
Ν^τ
(g/i)
(g/i)
(g/i)
(g/i)
(g/i)
(%)
BSC
160.4
82.9
136.8
23.6
1.44
14.7
ASC
333.4
172.04
310.0
23.4
1.06
7.0
^ResultX Conditiohs^
1 70.3
2 75.3
3 81.3
4 79.2
BSC
160.4
82.9
136.8
23.6
1.44
14.7
ASC
352.6
101.78
335
17.6
0.98
14.7
BSC
160.4
82.9
136.8
23.6
1.44
14.7
ASC
374.2
200.19
364.0
10.2
0.79
2.7
BSC
160.4
82.9
136.8
23.6
1.44
14.7
ASC
399.0
219.27
386.0
13.0
0.91
3.3
BSC= Before super concentration ASC= After super concentration
BSC= Before super concentration ASC= After super concentration
53
ÂCognics 1
(%) 70.3
75.3
81.3
79.2
Table 5: Sodium carbonate causticization percent and organics removal percent of one-section causticization Factors Test No
Nc
[CaO] [Na2Oc]
(%)
40
350
360
concentrated
I I
I8 |
0.5
ACorganics
ACorganics
%
%
%
1
ACorganics
%
55
2
86.73
62.1
|
10
1.3
75
3
91.21
61.8
I
10
1.4
95
4
94.81
67.8
|
12
1.2
75
4
94.52
69.3
I
12
1.3
95
2
96.12
64.7
I
12
1.4
55
3
93.67
62.9
|
14
1.2
95
3
93.98
76.45
I
14
1.3
55
4
90.64
75.36
|
14
1.4
75
2
93.90
72.68
|
[NC12C2O4]
I|
Table 6: Two-section causticization organics removal percent and total organics removal percent (%)
ACorganics I
%
81.8
82.1
81.5
No.
in table 2 No.4
2
1.2
^s^^organics 81.3
79.2
81.9
82.3
(%)
10
causticization, then we can get the results shown in table 6.
No.3
1
I
(%)
The causticization slurry No.7 to No.9 in table 5 is cooled to 5070°C and added the lime — \ £ α θ ] — = 0 5 ^ 1 0 for two-section
I
1.2
ACorganics
3.4.2 Two-section causticization
(added Na 2 C 2 0 4 liquor / Na 2 C 2 04 in the liquor)
0
9
Time
According to the test results, the maximum difference in organics removal is shown after concentrated slag causticization is carried out. This shows the order of the effect of factors on the organics removal percent are: Nc> time > temperature > [CaO/Na2Oc]; the optimum caustic conditions of the organics removal are: N c -14%, time-3 hours, temperature-95 °C, [CaO/Na2Oc]-1.2. These conditions are also the optimum conditions for carbonate removal.
Effect of crystal seed addition on organic carbon Crystal seed coefficient
5 6
I 7
The crystal seed addition test is carried out under the conditions No.3 and No.4 specified in table 2. The results indicates that the crystal seed addition can increase the settlement percent (ACorganics) of caustic sodium oxalate a littlebit and also is good for improving the settlement and filtration performance. The effect of crystal seed addition on the sodium oxalate precipitation refers to the following table (table 4).
Original conditions
I
2
I
3.3 Effect of crystal seed addition on organic carbon removal percent
No
1
I 34
370
NK value(g/l)
Figure 4: Relation of concentrated Nk value and organics removal percent (3)
Table 4: percent
|
Temperature
Sodium carbonate causticization percent
\. 1
82.1
(%)
N.
Super concentratio n
Onesection
Twosection
Anorganic
ACorganics
Δ^-Organics \ IO)
.(%)
75.3
76.12
(%)
87.6 3
Total ACorganics
(%)
75.3x87.63
=66.0% 1
in table2
The results show that the two-section causticization treatment based on the one-section causticization can increase the organic carbon removal percent in the concentrated slag by about 10%. The total removal percent of the organics in the liquor after super concentration, one-section causticization and two-section causticization is about 66.0%.
3.4 Causticization test 3.4.1 One-section causticization The No. 3 liquor (main components: NT160.4g/l; A1 2 0 3 82.9g/l; NK136.8 g/1; N c 23.6 g/1; Corganics1.44 g/1) in table 2 is concentrated to 364g/l, and then the slurry after concentration is treated by the one-section causticization and the control conditions (such as causticization green liquor Nc, molar ratio [CaO/Na 2 O c ], temperature, reaction time etc.) of the causticization reaction are adjusted. The chemical components of the precipitated liquor and solid are respectively analyzed, and the sodium carbonate casuticization percent and the organics removal percent are shown as follows:
3.5 Anti-solution test of causticization slag The resolution test shall be respectively done for the causticization slag from the test No.7 in table 4 and the causticization slag of this test slurry after two-section causticization. The 2 kinds of test results are almost consistent with the resolution law: the alumina concentration during washing has only a small effect on the resolution percent of Corganics, and the temperature has a big effect
54
on it. The resolution test result of organic carbon in causticization slag is shown in table 7. Table 7: Resolution of organic carbon in causticization
Temperature (D
carbon
Resolution
No.
Nk(g/L)
Ao(q/L)
1
4.0
4.8
70
2.9
4.0
4.8
70
3.1
4.0
4.8
80
5.1
4.0
4.8
95
7.5
5
6.0
7.2
60
3.4
6
6.0
7.2
70
4.2
7
6.0
7.2
80
7.6
8
6.0
7.2
95
10.8
9
8.0
9.6
60
3.6
10
8.0
9.6
70
7.1
11
8.0
9.6
80
11.5
I 12
8.0
9.6
95
15.2
I
2
I
4
3
)
the salt settlers for concentration and to the downstream separators for separation. After separation, the solid after solution is sent to one-section causticization tank with lime milk addition, the onesection causticization slurry is obtained by heating the reaction and is sent to a two-section causticization tank after it is cooled. T h e two-section causticization slurry is obtained and sent to separation washing machines, and then the washing liquor is sent to the alumina production system, the slag is sent to the alumina red m u d washing system and is discharged after washing along with the red mud.
Corganics( /o) 1
|
Crystal seed
1 -
I Liquor *
:T
I
'
Lime nilk* JLJL
Super concentration*·'
Test liquor proportion
4.
Mother liquor
Separation ♦>
Causticization*·'
I Slag.
> Causiticization liquor
,
t Related systems of alumina* production*'
? Solid ^
Conclusion
f
The organic carbon removal percent has the direct relation with the initial organic carbon content; it is increased along with the increase of the initial organic carbon content.
Discharge**
When the initial organic carbon content reaches a certain level, the organics removal percent is increased first by the linearity along with the increase of the concentrated caustic soda concentration; however, the organic carbon removal percent is reduced along with the increase of the high concentrated caustic soda concentration after the concentrated caustic soda concentration reaches the maximum value 360g/l.
References 1. Center-south University, Guiyang Aluminum Magnesium Design & Research Institute. "Study report on organics removal of alumina production with bauxite by Bayer process". 2009.
The crystal seed addition can slightly improve the settlement percent of the causticization sodium oxalate product.
2.
The reaction percent of the oxalate crystallization after onesection causticization is 62.1%-72.68%.
Chen Wengu, Tang Jiaming, Zhang Li. "Study on sodium oxalate in the Bayer Process". Light Metal 2005, 5: 11-15.
3.
B Gnyra and G Lever. "Review of Byer Organics-Oxalate Control Processes". Light Metals 1979,151-161.
The two-section causticization treatment based on one-section causticization can increase the total organic carbon removal percent in the concentrated slag by about 10%.
4. W Arnswald et al. "Removal of Organic Carbon from Liquor by Wet Oxidation in Tjbe Digesters". Light Metals 1991, 2327.
The alumina concentration has a small effect on the resolution percent of CorganiCS, and the temperature has a big effect on it. This method can not only effectively remove the oxalate, but also has a relatively good carbonate removal effect. 5. Process flow of super concentrated causticization The mother liquor in the alumina system is sent to the general evaporators for preliminary concentration, and the concentrated mother liquor is sent to the super-concentrated evaporators for super concentration, then the super-concentrated slurry is sent to
55
5.
G M Bell. "Osidation of Organic Substances in the Bayer Process". Light Metals 1981,117-128.
6.
P Atkins and S C Grocott. "The Liquid Anion Exchange Process for Organics RemovaF. Light Metals 1993, 151-157.
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
ALUNORTE EXPANSION 3 - THE NEW LINES ADDED TO REACH 6.3 MILLION TONS PER YEAR Daryush Albuquerque Khoshneviss1, Luiz Gustavo Correa2, Joaquim Ribeiro Alves Filho1, Hans Marius Berntsen3, Ricardo Rodrigues de Carvalho2 alunorte - Alumina do Norte do Brasil S.A., Rodovia PA 481, km 12. Distrito de Murucupi, CEP: 68.447-000, Barcarena-PA, Brasil 2 Vale S.A., Av. Graça Aranha, 26 - 10° andar, CEP: 20.030-900, Rio de Janeiro-RJ, Brasil 3 Hydro Aluminium AS, N-0240 Oslo, Norway Keywords: Expansion Project, Alunorte Performance, Plant Management Abstract Alunorte started operation in 1995 with a design production capacity of 1.1 Mtpy. Since then the plant was expanded three times and consists today of seven production lines with a total production capacity of close to 6.3 Mtpy. Expansion 3 included the process lines 6 and 7. The lines were commissioned in 2008 and reached their nominal production capacity after a short startup phase. In Expansion 3 a number of new technology developments were consolidated, such as the bauxite transport by pipeline, an improved precipitation concept and others. The impact of the process modifications applied in lines 6 and 7 are discussed in comparison to the previous lines with regard to effects on production volume, productivity, alumina quality and availability.
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However, a number of challenges are associated with the operation of a plant of the size of Alunorte which has grown very fast: the management of seven process lines, the operation of the plant with a very young organization and strong demand for training.
ceeded well its design production capacity. From Expansion 2 onwards the actual production remains slightly behind the design production capacity. However, all lines individually have demonstrated their design capacity. But the plant has become more and more complex with time. A number of issues arose which are linked to the size of the plant and its fast growth. On the other hand operations, maintenance, logistics can be made very efficient and a number of benefits can be achieved.
Introduction Alunorte started operation in July 1995. The plant had two production lines and a nominal production capacity of 1.1 Mtpa. This production rate was reached in 1997 and exceeded in the following years. A number of process improvements were implemented so that the production capacity could be increased to 1.6 Mtpa in 2000. In eight years only from 2000 to 2008 the plant production capacity increased three times. Work on Expansion 1 started in 2000 and in the first quarter 2003 the third process line was commissioned. The plant design production capacity was increased to 2.325 Mtpa. This production rate could be reached and even exceeded during the first year after the expansion. In 2006 process lines 4 and 5 were commissioned as part of Expansion 2. The production capacity increased further to close to 4.4 Mtpa. This number includes afirm-upof the precipitation area of lines 1-3. Finally, in the fourth quarter of 2008, lines 6 and 7 started operation as part of Expansion 3. The plant has now a nominal production capacity of close to 6.3 Mtpa.
In the following, the Expansion 3 project is described followed by a review of Alunorte's current performance. Then the current main challenges are discussed - technically and organizationally. Some of the projects started to improve the situation are presented as well as an outlook about Alunorte's expected development and performance. Scope of the Project Expansion 3 consists of new filters for bauxite dewatering, two new process lines (lines 6 and 7), two 3,300 t/d calciners, a port expansion and the construction of a new high-pressure circulating fluidized bed (CFB) boiler. This paper focuses on the implementation of the new process lines from bauxite dewatering to calcination and discusses port expansion and boiler erection briefly.
In Fig. 1 the development of production capacity and actual production are shown. The graph makes clear that historically Alunorte exceeded the expectations with regard to production volumes. For Expansion 1 and 2 the actual production before the expansion - which were higher than design - was the design basis onwards. In this way additional production volume is included in today's nameplate production capacity of close to 6.3 Mtpa when compared to the production volume of the original plant and the expansions. Before Expansion 2 was commissioned Alunorte ex-
The design of the new process lines 6 and 7 is partly based on the original design of Alunorte of Alcan International from 1995, partly it includes improvements which have been implemented in the previous expansions and some new elements were added. The most obvious difference between the expansions is their size. Each of the new lines has a nominal production capacity of 930 ktpa or 1.86 Mtpa in total for Expansion 3. The original design capacity of the plant was 1.1 Mtpa, 550 ktpa per line. Thus, the
57
design capacity of lines 6 and 7 is about 70% larger than the original design capacity of the first two lines. The dimensioning of the new lines is larger than that of the old ones but also the process performance considerably increased. In 2000, after the firm-up, lines 1 and 2 produced 1.6 Mtpa and operated at a precipitation productivity of close to 84 g/1. With lines 6 and 7 a precipitation productivity of 89 g/1 can be achieved. This is a result of the continuous improvement of the precipitation concept. Lines 1/2/3, 4/5 and 6/7 apply three different concepts. In lines 6/7 coarse seed and fine seed filtration are installed, lines 4/5 use seed cyclones and lines 1/2/3 coarse seed filtration. The newest lines are operated with the highest solids concentration and achieve the highest precipitation yield. Furthermore, the cooling capacity is well dimensioned. In addition to traditional in-tank plate heat exchangers external plate heat exchangers (Barriquand coolers) are installed. The number of tanks has been increased as well as their size in order to increase residence time for higher precipitation productivity.
The performance of the calcination facilities was increased again with Expansion 3. The energy efficiency of the new two 3,300 t/d circulating fluidized bed (CFB) calciners is higher than that of the old ones. A high degree of automatic control is installed [4]. In total seven CFB calciners of Outotec (formerly Lurgi) are installed. Alunorte received the energy efficiency award 2010 from the German Energy Agency for the two newest calciners [5]. A part of the electrical power used by Alunorte is co-generated at site while the rest is received from the national grid. As part of Expansion 3 a circulating fluidized bed boiler was installed. The utilities system for steam and electrical power generation of Alunorte has a high overall efficiency and good operational flexibility. This is important in order to be able translate the low specific consumption of steam and electrical power into a low specific consumption of boiler fuel. The environmental concept is continued with Expansion 3 in order to guarantee a generally high standard. Emissions and effluents are within the legislative limits. Of high importance is Alunorte's bauxite residue disposal concept - dry-stacking which is state of the art. The mud in filtered in drum filters and then disposed at high solids concentrations. Water from the bauxite residue area (RDA) is collected, cleaned and then send to a nearby river.
The bauxite for lines 6 and 7 is supplied from Mineraçao Bauxita Paragominas (MBP) through the world's first long-distance bauxite slurry pipeline [1]. The bauxite slurry is pumped at a solids content of 50 % and is then dewatered at Alunorte in hyperbaric filters [2,3]. A residual moisture content of about 15 % in the dewatered bauxite is achieved. This is slightly higher than that of Trombetas bauxite which arrives by ship. To be pumpable through the pipeline the Paragominas bauxite is finely ground at MBP. Its particle size distribution is sufficiently fine to be charged to the Bayer plant without additional milling. Different to the other lines no mills are installed in lines 6 and 7 and the bauxite is sent from filtration by conveyor belts directly to the bauxite slurry re-suspension tanks and heating units.
Project Execution The installation of process lines 6 and 7 as part of Expansion 3 was finished on schedule, on budget and the design performance of the two new lines could be demonstrated during the first two months after start-up. The project had a duration of 32 months. The master schedule of the project for the refinery is shown in Fig. 2.
The clarification area has undergone a number of changes compared to the original design. Initially conventional decanters and washers were installed. As part of the expansions deep thickeners were added. Alunorte made good experience with their operation so that in Expansion 3 only deep thickeners are installed and conventional decanters or washers are completely eliminated. As part of Expansion 3 Alunorte has decided to install Diastar filters for liquor filtration. The filters are operated fully automatically. The operational performance of the liquor filtration area could be improved in comparison to the previous process lines.
A number of factors contributed to the fast execution of the project. One of them was the strategy to use the latest expansion as design basis for the new lines. Improvements in specific plant areas were included. However, the concept was not fundamentally changed in most areas. The biggest differences are found in precipitation. The project was executed with an extended owners' team with no EPCM contractor. Alunorte Board of Directors acted as steering committee for the project. An advantage of the organization of the project was the ability to take decisions very fast and avoid decision-related delays of the project. Furthermore, it was possible to work with a very experienced team. Most members of the Alunorte expansion team have already been involved in the previous expansions. Finally, it was possible to perform the start-up of the lines earlier than originally scheduled. Less time than planned was required to reach the nominal production capacity of lines 6 and 7.
In digestion, evaporation and vacuum flash cooling bottom-entry flash tanks are installed. The carry-over of caustic soda to the condensate is lower than for conventional side-entry flash tanks. The regenerative condensate has better quality, the scaling in the vapor transfer lines is reduced and longer cleaning cycles are achieved. Originally side-entry flash tanks were installed in digestion lines 1 and 2. These were modified to bottom entry and in all expansions only bottom entry tanks were used.
The project completion system ProCoSys™ was used for construction and commissioning of the new lines. The system was initially developed by Hydro and mainly used for oil and gas projects. The system provides a systematic approach to ensure a fast project completion. It is ensured that all equipment is available and checked at the time it is required. The system became a well recognized tool at Alunorte and helped to achieve fast commissioning and a start-up of the new lines without major delays.
A modern fully digitalfieldbusprocess control system is installed. At Expansion 2 it was decided to complete the move from an analog system to a digital system. This system is implemented as part of Expansion 2 and Expansion 3. The degree of automatic control increased with each expansion. The high degree of automatic control of the new lines helps to achieve higher overall performance in the new lines.
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Alunorte's shareholders are very satisfied with the performance of the project organization, management and execution in general. An exception, however, is the unacceptable safety performance of Expansion 3. A fatality occurred during the construction caused by a crane fell on a passing construction worker. A TRI rate of 3.4 was reached for the overall Expansion 3 project. Although the implementation of the process lines went according to schedule and the production capacity was achieved faster than initially planned there were some delays in the Expansion 3 project. Both the new boiler and the port expansion were delayed. The delays caused some operational challenges. Solutions were found to operate Alunorte at close to full capacity without having these projects finished. Expansion 3 included a high-pressure circulating fluidized bed boiler for steam and electrical power generation. As the boiler was delayed the required steam for the new process lines was provided by already existing low pressure boilers. For more than one year Alunorte had to operate without spare boilers. Delays in the early phase of the port expansion could not be recovered so that finally the installation of a new alumina shiploader could not befinishedon time. Therefore some logistic challenges needed to be overcome at the port. The situation was challenging but Alunorte succeeded to ship alumina at a rhythm exceeding to 5 million t/yr with only one ship loader without loosing production due to logistical problems at the port.
With Expansion 3 new personnel was contracted by Alunorte. The hiring started enough time before the planned start-up of the new lines. Training for the new operators started about a year before commissioning. In this way it was possible to perform the start-up of the new lines with a team of sufficient experience. Overall, the new lines performed as expected. No major negative surprises were experienced when the lines were started up. Although a new precipitation concept was installed for which no operational experience existed no major problems turned up. Lines 6 and 7 are fully based on Paragominas bauxite and some minor differences in process behavior were observed caused by the bauxite. These differences were mostly found in the clarification area. The fine particle size distribution of the bauxite might be the cause of a somewhat slower settling mud as compared to Trombetas bauxite. Measures were taken to correct the situation in the clarification area.
Start-up and Operational Performance
The energy utilization of Alunorte is among the best in the world with about 8 GJ per ton of alumina. As there are few plants with a similarly low energy utilization this defines the world-wide benchmark for alumina production. A detailed review of Alunorte's energy efficiency is presented in [6]. Alunorte's specific caustic consumption is mainly determined by the noncontrollable losses which are driven by the content of reactive silica in the bauxite. Controllable losses are small with less than 8 kg per ton alumina. Although the content of reactive silica is somewhat high in both bauxites, Trombetas and Paragominas, their quality is generally very good. The content of available alumina is high and the content of organic material very low. In combination this leads to a low specific consumption of bauxite and very clean liquor with low concentrations of impurities. This
23 25 27 29 31 33 35 37 39 41 43 45 47 49 51 53 week(2008)
Figure 3. Ramp-up of Alunorte's lines 6 and 7.
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Table I. Typical alumina quality parameters of Alunorte. 5.0% 0.015 % Si0 2 cc-alumina LOI (300-1000 °C) 0.8% Fe 2 0 3 0.015 % 0.005 % LOI (110-300 °C) 0.7% Ti0 2 Na 2 0 0.4% BET 75 m2 3.0% ZnO 0.001 % + 149 μπι (+100#) + 74 μπι (+200#) 70.0 % CaO 0.01 % -44μm(-325#) 7.0% v 2 o 5 0.002 % MnO 0.001 % app. density 0.97 g/cm3 0.001 % attrition index 17% p2o5
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the previous alumina quality characteristics could be maintained. However, some minor differences can be observed when the product from lines 6 and 7 is compared to that of the older lines. The content of occluded soda is slightly higher and the lines are operated with slightly coarser crystal size distribution in precipitation. No difference is observed for Attrition Index. The product of the new lines is blended with the product of the other lines before shipped or sent to the nearby Albras smelter. In Table I the typical alumina quality parameters of Alunorte are shown. As result of improvements done in lines 6 and 7 compared to the older lines variations in alumina quality parameters became smaller. As the variations in all lines are generally small and the product of all lines is blended this effect is not observed in the final product, in particular not for alumina which is shipped. The storage in silos at Alunorte, transport by ship and storage at the smelters lead to additional homogenization of the product. However, the target is to operate the process as stable as possible in order to achieve the best possible performance.
Alunorte has grown as fast as no other alumina refinery before and has a size which exceeds the second-largest alumina refinery by about 50 %. In this way Alunorte stands out in comparison to all other alumina refineries in the world. In the following section a critical review is presented about Alunorte's current performance, actual challenges and expected development in the coming years. Safety There is a high focus on safety at Alunorte. The YTD TRI rate in October 2010 was 1.95 (based on one million working hour), after having achieved a TRI rate of 2.37 in 2009. Alunorte has developed and continuously improved a strategy to manage health and safety during the past 15 years of operation [7]. Alunorte is ambitious to further improve this performance in the coming years. The decision to install new technologies at Alunorte always includes some new risks for accidents which need to be carefully addressed.
Project Budget
Process & Technology
The project cost of Expansion 3 was 629 USD per annual ton of alumina. The project budget was set up in both currencies US Dollar and Brasilian Real (BRL). The project budget was met in BRL but there was a budget overrun in US Dollar. The main cause was the development of the Brasilian currency during the project. At the beginning of the project the exchange rate was 2.2 while it dropped to about 1.6 at the end of the project. In Fig. 4 the development of the Brasilian Real in comparison to the US Dollar is shown for the period from 2000 until the end of 2008. The situation of Expansion 3 is compared to the Expansions 1 and 2. A large amount of the material and equipments was purchased in Brasil so that project cost in BRL were not significantly affected by the development of the currencies.
The general technical performance of Alunorte is described above. Operation is characterized by a high productivity, low controllable caustic consumption, low energy utilization, a stateof-the art bauxite residue disposal concept and low plant emissions. For good plant performance and good product quality it is important to operate the refinery as stable as possible. Operational variations must be kept small and the degree of non-planned maintenance must be kept as small as possible. One of the challenges to achieve a good performance in these aspects is the fact that Alunorte can be considered as one huge alumina production complex which can be divided into three major units. The first unit are lines 1/2/3, the second lines 4/5 and the third 6/7. Since the equipment installed in these units is not identical and different process concepts exist, e.g. in precipitation, specific operation and maintenance procedures are required. The three units are with small exceptions process-wise fully independent
Expansion 3 was finished close to the beginning of the worldwide economic downturn end of 2008. The additional alumina volume from Expansion 3 was added to the market at a time of falling LME and alumina prices. The sudden development of the alumina market made clear how important it is to operate a plant
60
from each other. The plant liquor of one unit is usually not mixed with liquor of another unit so that individual characteristics for each of the units exist. One example was given above. The settling behavior of Paragominas bauxite is different from that of Trombetas. The advantage of the separation of the lines is that operational disturbances of the process in one line do usually not affect the other lines. Even a stop of a line is possible without affecting the production of the lines in the other operational units.
Organization & Management In early days when Alunorte consisted of two lines only 450 people were employed. Nowadays, Alunorte employs 1,600 people and about 1100 more workers as permanent contractors. Due to the very fast growth of the plant Alunorte's team is very young and many of the employees have only few years of work experience. The plant has not only grown very fast but also new technologies were employed. The dewatering of the bauxite received through the pipeline from MBP is a new technology not only for Alunorte but also for the whole alumina industry. Alunorte could not profit from the experience of other alumina refineries. The first circulating fluidized bed boilers were installed as part of Expansion 2. Although the technology is not new as such it was new for Alunorte. Furthermore, there is just a small number of circulating fluidized bed boilers installed in South America. Both, the young organization and the application of new technologies at Alunorte result in a strong demand for training for operators and engineers.
The three production units are, however, not fully independent from each other. They share utilities such as the electrical system, steam generators, bauxite supply and port. Operational problems in one of these plant areas can affect the whole process and will in this case lead to significant losses in production. Alunorte recognized that this is a risk for the operation of the whole refinery and started a program to minimize the interconnections between the units as far as possible. This task became more and more important as the plant grew. The complexity of the plant increases exponentially with its size and the possibilities for equipment failure which can affect larger parts of the plant get bigger. The availability of the plant is exposed in a similar way. The more complex the refinery got the more efforts were required to maintain a high global availability.
As explained above the plant consists of seven lines which can be divided into three units consisting of lines 1/2/3, 4/5 and 6/7. The operational issues are not the same in the different lines. In some plant areas, such as precipitation, different process concepts are chosen. Furthermore, the first lines have an age of 15 years while the newest lines are just two years old. All this has to be considered in daily operation, for maintenance, training of the employees, etc. The principal management organization of Alunorte is still very similar to that in place during the early years of operation. It can be questioned if today's organization is still ideal or if other concepts are more suitable. Alunorte works actively to evaluate the current issues and challenges to manage the plant. Alunorte's is generally open to new ideas, technologies and concepts and works also on a continuous improvement of the plant organization.
Logistics are another important issue. Some of the challenges arose from the delay of the port expansion. This issue could be handled as described above. However, the tonnages of alumina, bauxite, caustic and boiler fuels are huge and require a careful planning of port operations. Unsatisfactory performance in this part of the plant could result in an increase of transport and handling cost or in the worst case in production losses. A similar challenge exists for the transport of bauxite residue to the residue disposal area (RDA). It is transported by trucks from the filtration area to the disposal area. In case of bad weather conditions as it canfrequentlyhappen during the rainy season from December to April the disposal to the RDA can become a bottle neck. As the plant has the potential to creep in production in the coming years and the bauxite quality slowly deteriorates the amount of bauxite residue which needs to be disposed will increase in the future and the challenge will gets bigger. Alunorte studies options to improve this situation in order to improve the logistics and also eliminate a potential safety risk caused by the large number of trucks in service. In addition to logistics of materials, transport logistics for Alunorte's employees needs to be mentioned as well. Alunorte is located at about 40 km from Belém, the capital of the state Para, in straight line. The road, however, is much longer about 100 km. Many of Alunorte's employees live in Belém and travel forth and back day by day. To improve the situation Alunorte launches a ferry connection between the plant and Belém city.
Conclusion With Expansion 3 Alunorte has become the largest alumina refinery in the world with a production capacity of close to 6.3 Mtpa. Alunorte has grown as fast as no other refinery in the world. The development of Alunorte from its start-up in 1995 is presented and Expansion 3 is reviewed in detail. The design of Expansion 3 is based on that of the previous lines. However, a number of new technologies are applied, such as a new precipitation concept, different filters for pregnant liquor or the use of deep thickeners only for mud wash. The Expansion 3 project had a duration of 32 month with a short start-up phase of two month for both new lines to reach design production capacity. The project cost of Expansion 3 was USD 629 per annual ton of alumina. The installation of the two new process lines could be performed on schedule and on budget. The generally good performance of Alunorte, such as energy efficiency or alumina quality parameters, could be maintained also after the addition of the two new lines. The challenges of managing a plant of the size of Alunorte with a young team with on average few years of work experience are discussed. Alunorte has not only grown very fast but has implemented a number of new technologies, such as bauxite transport as slurry in a pipeline or the installation of CFB boilers. Training of Alunorte's operators and engineers is therefore important. Nowadays, Alunorte can be considered as a big alumina production complex with seven process lines which can be divided into three units, namely Lines 1/2/3, 4/5 and 6/7.
After many years with a high focus on growth Alunorte has entered into a phase of optimizing the operation of the plant. There is potential for some increase in production. Furthermore, Alunorte works constantly to improve energy utilization and specific consumption figures. Some counteracting factors have to be considered such as the aging of the equipment or a slow deterioration of the bauxite quality in the long-term perspective. As part of the current work on process optimization new technologies are reviewed and tested to guarantee good operational performance also in the long term.
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This offers technologically and organizationally possibilities to operate the units independently of each other with the aim to achieve a high global availability of the plant and to minimize the risk of failures in one line affecting the rest of the refinery. Alunorte has the goal to further improve the operational performance of the plant in order stay among the best plants in the world also in the long term. References 1. Ramesh Gandhi, Mike Weston, Mani Talavera, Geraldo Pereira Brittes and Eder Barbosa, "Design and Operation of the World's First Long Distance Bauxite Slurry Pipeline", In Light Metals 2008, De Young, D.H., Ed., TMS (The Minerals, Metals & Materials Society), 95-100. 2. Ayana Oliveira, Juarez Dutra and Jorge Aldi, "Alunorte Bauxite Dewatering Station - A Unique Experience", In Light Metals 2008, De Young, D.H., Ed., TMS (The Minerals, Metals & Materials Society), 85-87. 3. R. Bott, T. Langeloh and J. Hahn, "Filtration of Bauxite after Pipeline Transport: Big Challenges - Proper Solutions", In Proceedings of 8th International Alumina Quality Workshop, Armstrong, L, Ed., AQW Inc.,2008, 319-323. 4. Michael Missalla, Jan Jarzembowski, Roger Bligh, Hans Werner Schmidt, "Increased Availability and Optimization of Calciner Performance due to Automation", Light Metals 2009, Bearne, G., Ed., TMS (The Minerals, Metals & Materials Society). 5. Michael Missalla, Hans Werner Schmidt, Joaquim Ribeiro Alves Filho and Reiner Wischnewski, "Significant Improvement of Energy Efficiency at Alunorte's Calcination Facility", In Light Metals 2011, Lindsay, S., Ed., TMS (The Minerals, Metals & Materials Society). 6. Reiner Wischnewski, Cleto Maues de Azevedo Jr, Emerson L.S. Moraes, Arthur Barros Monteiro, "Alunorte Global Energy Efficiency", In Light Metals 2011, Lindsay, S., Ed., TMS (The Minerals, Metals & Materials Society). 7. Jorge Aldi Dias Lima, "Safety Performance at Alunorte", In Light Metals 2008, De Young, D.H., Ed., TMS (The Minerals, Metals & Materials Society), 5-11.
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Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
One Green Field Megaton Grade Large Alumina Refinery with Successful Engineering & Operation Experience Luo Xianqing
Yang Xiaoping
(Guiyang Aluminium Magnesium Design & Research Institute, Guiyang, Guizhou, 550081) Keywords: diaspore bauxite, alumina production, Bayer process, comprehensive energy consumption accumulated bauxite and the ancient weathering crust bauxite, in which the accumulated diasporic bauxite is 83.5% and the ancient weathering crust diasporic bauxite is 16.5%. From the ore bed scale it is known that all ore beds in 10 large-size ore beds are accumulated diasporic bauxite and 3 ores in medium-size ore beds are accumulated diasporic bauxite. The accumulated bauxite mainly is distributed in Pingguo, Debao, Jingxi etc. in the west of Guangxi. This bauxite is characterized by medium aluminum content, high ferrous content, high alumina silica ratio, low sulfur content etc. and consists of diasporic bauxite, goethite, hematite and kaolinite being 85% content ,with lower content of other mineral, is the few high-quality bauxite resource in China can be used for producing the aluminum with Bayer process. The ore grade: A1203 is 45%-65%, generally 54%-60%; Si02 is 3.4%-14% &, generally 5%-9%; alumina silica ratio is 4-11; Fe 2 0 3 is 5%-25%, generally 10%-20%.
Abstract: The Phase I alumina project (l.ómillion tons/a metallurgical grade alumina) of the Guangxi Huayin Aluminum Corporation Limited is a greenfield construction project with to-date, the largest disposable investment in the Chinese alumina industry. The project is based on a diasporic bauxite which is difficult to grind and dissolve, with a high aluminum & low silicon content. Two years of operational experience shows that the comprehensive energy consumption per ton alumina production is only 10.51GJ/t. This article introduces the process scheme & technical measures adopted in the engineering and operational stages of this project.
1. Introduction Guangxi Huayin Aluminum Corporation Limited is a partnership of the Guangxi Investment (group) Corporation Limited, China Minmetals Non-ferrous Metals Corporation Limited and China Aluminum Stock Corporation Limited. The major investment is located in Mayi Town, Debao County, Guangxi Chuang Municipality, Guangxi. The construction of its phase I alumina project (hereinafter to be referred as Huayin phase I alumina project) has been targeted to produce 1.6 million tons/a of metallurgical grade alumina and is based on a total budgetary estimated investment of RMB 9.12 billion..
The bauxite in Debao and Jingxi County is the raw material for the Huayin phase I alumina project, and there is a mining division and an ore wash division respectively built in these two counties. The parameters of the mined bauxite in the first mining area are as follows:
The project was designed by the Guiyang Aluminium Magnesium Design & Research Institute, constructed and started up by China Aluminum International Engineering Corporation Limited as the general contractor. Construction was started in June 2005; the first production line was put into production in Dec. 2007, with all the production lines in operation in June 2008 and production & operational standards achieved in June 2009.
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£7.0
£55.80
£5.94
>9.39
£7.0
|
>56.22
£5.84
£9.62
£^0
|
|
The ore processing performance test [ ] result indicates that the bauxite in Debao and Jingxi is a diaspore dominated bauxite, characterized by high aluminum & low silicon content, and its performance is as follows:
2. Bauxite resource and ore processing performance Guangxi Chuang Municipality is abundant in high-quality bauxite resource reserves. The total explored bauxite reserves are 0.569 billion ton and the proven ones are 0.375 billion ton up to the end of 2005. As far as known from the ore bed investigation degree in Guangxi, the resource reserves reaching exploration degree are 28%, reaching detailed investigation are 22% and reaching general investigation are 48%, thus the bauxite investigation degree in Guangxi is high. Guangxi bauxite contains 3 types of ore, including ancient weathering crust bauxite, accumulated bauxite and laterite bauxite. The proven resource reserves are mainly the
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•
This bauxite is difficult to grind and its work index is large. The Bond rod mill work index is 20 mesh, Wi=16.02; the Bond ball mill work index is 120 mesh, Wi=22.55.
•
Under the test conditions with a solid content of 320g/l in the slurry, test liquor caustic soda (Na20) being 245g/l and lime addition being 7%, the predesilication rate of this bauxite is close to 80% when the predesilication temperature is 90 to 100°C.
•
Under the test conditions with digestion temperature being 260°C, digestion time being 40 min and lime addition
being 7%, the actual digestion rate reaches 86% and the relative digestion rate is 96%, however, with digestion time being 30 min, the actual digestion rate is 83% and the relative digestion rate is 92%. With a diluted slurry retention of 3 hours, the ratio of alumina content and silicon oxide content in the sodium aluminate liquor is always above 230. The bauxite in Debao and Jingxi is difficult to dissolve, so it requires higher digestion temperatures, higher caustic liquor concentration and finer bauxite granularity. •
belt weigher & roller conveyor. The test liquor required for bauxite grinding is taken directly from the evaporation station, and is added in proportion respectively from the outlet & inlet of the rod mill and the outlet & inlet of the ball mill. The slurry ground by the rod mill and the ball mill together enter the slurry tank, and are then sent to the hydraulic cyclone for classification by the slurry pump. The underflow is sent back to the ball mill and the overflow enters the slurry storage tank through the rotary screen and becomes the qualified slurry, then sent to the high-pressure digestion by slurry pump.
After some flocculant is added to the slurry after digestion, the red mud settling speed can satisfy the industry production requirements, thus the red mud settlement performance is acceptable.
This grinding scheme effectively solves the problem that the bauxite from Debao and Jingxi is difficult to grind and satisfies the digestion requirements to the bauxite granularity, which reduces the power consumption for the bauxite grinding by 6.22kw.h/t compared with the general bauxite grinding process and makes the bauxite granularity more uniform.
3. Process scheme and main operation indexes The alumina refinery uses the Bayer process, based on the bauxite supply conditions. It consists of raw material plant (including bauxite storage & transportation, bauxite grinding, lime storage & transportation, lime slaking etc.), digestion shop (including high-pressure pump house, sleeve-tube preheating digestion & dilution etc.), red mud settling & washing plant (including red mud settling & washing, red mud filtration, red mud transportation, flocculant preparation, hot water station, security filtration etc.), precipitation shop (including pregnant heat exchange, seed precipitation, hydrate classification, intermediate cooling, seed filtration etc.), hydrate calcination (including product filtration, hydrate silo & hydrate conveying, calcination, alumina conveying & packing etc.), evaporation plant (including evaporation station, green liquor storage tank & water washing, test liquor storage tank, desalt, causticization etc.), circulating water shop etc.
3.2 Predesilication Considering the bauxite from Debao and Jingxi possesses good predesilication performance, the predesilication section is not here in the design and just requires more heating in the slurry storage tank. Fast slurry predesilication is achieved in the slurry storage tank, with 2 hours predesilication (the scaling in the single sleeve-tube preheater is very slow and the acid washing cycle of the single sleeve tube preheater is above 80 days during actual production, indicating that the predesilication effect is good and this measure is very successful). 3.3 Digestion An eleven-stage preheating and ten-stage flash evaporation digestion device has been adopted for this plant.
The process flow of the whole plant is as follows: the bauxite is made into slurry after it is ground with the lime and qualified test liquor by the ball mill. The slurry after predesilication enters the digestion unit consisting of the sleeve-tube preheater and the autoclave, and the digested slurry after multi-stage self-evaporation enters the dilution tank. The diluted slurry enters the red mud settling & washing system consisting of large flat-bottomed settlers and drum filters. The red mud after washing is sent to the red mud stockyard and the settler overflow after security filtration and heat-exchange cooling, enters the seed precipitation system. The aluminum hydrate generated from the seed precipitation in converted to alumina through gas suspension calcination and the spent liquor after precipitation is fed back to bauxite grinding through the evaporation.
The slurry from the raw material grinding section is sent to the high-pressure pump house, and to six-stage sleeve-tube preheater through a high-pressure diaphragm pump, and heated to 170 °C, with the secondary vapour generated by the self -evaporator. It then enters 4 autoclaves heated to 220 °C with secondary vapour, and is heated to the digestion temperature being 260 °C in a follow-up reaction autoclave with fresh 6.1MPa steam. It then enters an insulated autoclave for insulated retention for 40 min. The slurry after high temperature digestion, is sent to the dilution tank after the temperature reduces to below 126 °C through a ten-stage self-evaporation, and sent to the after tank by the pump, after it is diluted with the primary washing liquor from the red mud settling & washing and remains for some time. It is then sent to the red mud settler by pump. The feeding pump of this digestion device is a triple cylinder single action diaphragm pump.
The technical scheme of each process section is as follows: 3.1 Bauxite grinding
There are many improvements in the utilization of digestion fresh steam condensate secondary vapour, the pore plate structure of the flash evaporator, heating pipes of the autoclave, continuous emission of non-condensing gas from autoclave, utilization of high-pressure digestion secondary condensate etc. in this engineering design compared with before. These make the steam consumption of the digestion system more reasonable.
A two-step bauxite grinding process is used, with the first step a rod mill open circuit and the second step a ball mill & hydraulic cyclone closed circuit. The bauxite, being no more than 15mm from the homogenization stackpile, is sent to the bauxite silo of the bauxite grinding plant, and the lime from the lime warehouse is sent to the lime silo of the plant.
3.4 Red mud settling & washing
The bauxite and the lime are sent to the rod mill in proportion respectively by plant-type feeder & belt conveyor and electronic
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A reverse washing system is adopted, consisting of red mud settlement by settler, triple settling & washing and primary washing by drum filter.
precipitation output rate. So intermediate cooling unit is added for higher precipitation output rate (the output of pregnant liquor can reach more than 88kg/m3 ).
The solid-liquor separation of digestion slurry (concentration of caustic soda Na20 being 172g/l ) after dilution in the settler, the overflow (coarse liquor) is sent to security filtration and the underflow is sent to red mud washing system by pump. In this washing system, the red mud is treated using hot water addition through triple settling & washing. The washing water flows along the reverse routing of red mud and its concentration is gradually increased, finally being used for diluting the slurry after digestion.
Moreover, some crystallizing agent is added so as to intensify the precipitation of aluminum hydrate and also increase its granularity. 3.8 Seed filtration This section has the function of seed supply for seed precipitation. The aluminum hydrate in the slurry from the last precipitator of the precipitation system is subjected to liquor-solid separation so as to obtain the aluminum hydrate seed; the spent liquor (caustic alkali Na20 concentration being 181g/l) generated after filtration is sent to evaporation section for carrying out mother liquor evaporation, so as to obtain the test liquor
The settler is a large flat-bottomed settler of 40m diameter. Some flocculant is added during operation so as to ensure the solid content of the settlement overflow is controlled within the specified limit. The red mud slurry after triple washing is sent to drum vacuum filter by the pump so it is washed and filtered with water injection for the last time. The red mud after washing is sent to mixing reaction tank; the filtrate is back to red mud washing settler through the filtrate tank by pump.
3.9 Recycle of aluminum hydrate in spent liquor In the spent liquor obtained after seed filtration and product filtration, the aluminum hydrate content is about 2g/l. The existence of aluminum hydrate in the spent liquor increases the pipe blocking risk index of the evaporator heating pipe and reduces the circulating efficiency of the caustic alkali. So this project is equipped with an aluminum hydrate recycling section and a vertical leaf filter is the main device for this section.
The filtration area of the drum filter is 100m2, being equipped with water injection device. It can simultaneously carry out red mud filtration and washing. The alkali content in the residual liquor of the red mud after filtration, is reduced to 0.32% (dry basis) and the water content is adjusted to about 40% (wet basis). Finally the red mud is transported by a diaphragm pump to the red mud stockyard being 2 km away for dry storage.
The underflow from the spent liquor conical tank in the seed filtration section is sent to the vertical leaf filter for filtering out the aluminum hydrate in the liquor. The filtrate is sent back to the spent liquor tank in the seed filtration section, and the filter cake (aluminum hydrate) is sent to the overflow tank in seed filtration and is finally sent to the precipitation section as seed.
3.5 Security filtration The large double-star vertical leaf filter of the filtration area being 308m2 is adopted to filter the coarse liquor from the red mud settling & washing. It is characterized by light weight, less occupation area, higher automatization, being convenient for maintenance and high capacity. The solid content of filtrate is no more than 15mg/l.
3.10 Evaporation and desalt The super-concentrated desalting technology with a 6-effect falling film evaporator and forced circulating evaporator has been adopted. It is characterized with maturity & reliability, low energy consumption (steam ratio <0.27) and high efficiency desalting.
The refined liquor(pregnant) from security filtration is sent to the pregnant heat exchange section by pump.
The spent liquor being 80 °C from the heat exchange is respectively sent to the 6th and 4th evaporator unit.
3.6 Pregnant heat exchange
The liquor entering the 6th evaporator unit reaches the requirements after it is concentrated through the 6th effect evaporator and the 5th effect evaporator.
This section is for simultaneously satisfying the seed precipitation temperature requirements and recycling the heat resource. The two mediums of heat exchange are respectively the pregnant liquor (inlet temperature being 100-105 °C and outlet temperature being 60-62 °C) from the security filtration and the spent liquor (inlet temperature being 45-48 °C and outlet temperature being 80-82 °C) from the product filtration.
The liquor entering the 4th evaporator unit is concentrated through the 4th evaporator, the 3th evaporator, the 2nd evaporator and the 1st evaporator, and enters 2-stage flash evaporation, then respectively enters the 3th stage flash evaporation and super-concentrated evaporator in proportion. The concentration of the evaporation liquor after 3-stage flash evaporation reaches the plant requirements; the concentration of caustic alkali Na20 in the liquor after concentration through super-concentrated evaporator reaches 320g/l. The sodium carbonate crystals are removed from the slurry with salt grain discharged from the super-concentrated evaporator through the settler and filter, the overflow from the settler and the filtrate from the filter as well as the liquor from 5-effect evaporator and 3-stage flash evaporation converge to become the mother liquor.
The equipment for this section is a plate-type heat exchanger. 3.7 Seed precipitation The seed precipitation system consists of large mechanical agitation precipitators of diameter 14m, intermediate cooling unit and classifier (hydraulic cyclone). The sodium aluminate pregnant liquor after cooling is sent to the precipitation system where hydrate seed is added, and then the hydrate is generated by the precipitation over 45-50 hours. Guangxi is in semi-tropical area in south of China, with a hot climate and high humidity. This limits the natural drop in temperature during precipitation and the increase in
Mother liquor, liquor after causticization, less evaporation green liquor and liquid caustic soda with alkali addition are mixed to be the test liquor with caustic alkali concentration being 245g/l,
65
and then it is sent to the bauxite grinding by pump after circulation through the liquor storage tank.
The four production lines ran under full load in the last month of 2009 along with the recovery of the alumina market. The technical & economic indexes were obviously improved and the production cost is reduced, being below RMB 1100, thus the profit was achieved in July to December 2009.
3.11 Productfiltration& washing The underflow from the hydraulic cyclone classifier flows to an aluminum hydrate slurry tank and the secondary condensate from the evaporation section enters a hot water tank, and the hydrate slurry and the hot water respectively are sent to a horizontal pan filter for hydrate separation & washing.
Production statistical indexes of Guangxi Huayin Aluminum Corporation Limited in July to December 2009 are indicated below.
The spent liquor obtained through filtration is sent to the spent tank in seed filtration by pump. The washing liquor is sent to the red mud washing settler by pump and the filer cake is sent to the calciner or hydrate silo through a belt conveyor.
No
This section is mainly equipped with a large horizontal pan filter. Separation and 2-stage reverse washing of the aluminum hydrate is carried out in this filter. Some dehydrating agent is added so as the water content in aluminum hydrate is below 3%. 3.12 Calcination
Item
Unit
1
Bauxite component
2
Output
3
Product One-class percent
kt/a
product
Comprehensive energy consumption
The aluminum hydrate calcination uses a gas suspension calciner. The aluminum hydrate from the product filtration & washing or the aluminum hydrate silo, is sent to a small silo before the calciner, through the belt conveyor.
4.1 4.2 4.3
The aluminum hydrate in the small silo is sent to a gas suspension calciner through the feeding weigher and screw feeder, being burned to the required alumina specification.
6 7 8 9 10 11
Steam Electrical power Coal gas Alkali (converted into Na2C03) Bauxite Lime Fresh water Compressed air Relative digestion rate Pregnant liquor output Alumina recycling rate
There are two heat exchange stages in calcination: heat exchange between the aluminum hydrate and the exhausted gas generated through coal gas burning and between alumina and air required for coal gas burning.
I 12
The exhausted gas, with dust generated during calcinations, is discharged after passage through electrostatic filters to meet discharge standards.
827.3 ( converted into 1654.7 for whole year)
|
Sandy AI2O3
4
5
Production statistical I indexes in July to December 2009 AI2O3 53.86% A/S 8.504
%
99.78
GJ/t(kg standard coal/t)
10.51 (358.8)
t/t kWh/t Nma/t
2.438 246.3 617.1
kg/t
118.7
t/t
2.26 243.8 2.97 39.93 93.4 87.63 81Ό1
kg/t
t/t ma
%
kg/ma
%
|
5 Conclusions Huayin phase I alumina project is a green field project with the largest disposable investment and construction scale in the Chinese alumina industry. It achieved the target of full commissioning in three years, and production & standards achieved after 1 year's commission, and thus establishes a new record in China alumina construction history.
The utilization of calciner waste heat has been greatly improved in later stage of this project, so RMB 10 million economic benefits per year can be gained. 4 Actual production conditions
Guangxi Huayin Aluminum Corporation Limited as a new plant can get away from the difficult position in the international finance crisis, because this plant possesses the right technical guidelines, excellent engineering design, better construction quality and strong production management, thus various indexes are at international advanced levels, especially comprehensive energy consumption per ton alumina production, being only 10.51GJ reaches the international leading standard.
The first production line of this project was put into production in December 2007; it took only 33 hours to go through the whole alumina production process flow. The second and third production line were successively finished in March and April 2008, and the commissioning of the fourth production line in May 2008 indicated that Huayin phase I alumina project was fully finished. However, due to the international finance crisis coming, the China aluminum product was overstocked heavily and the alumina price declined significantly and continuously. So Guangxi Huayin Aluminum Corporation Limited had to reduce the output to limit loss, and ran at only 50% production before May 2009.
References 1. Tian Fengming Fan Na, Potentional and availability analysis on bauxite resources in Guangxi, China Mining, 19th Volume, 5th Period, May 2010.
Guangxi Huayin Aluminum Corporation Limited checked the performance of Huayin phase I alumina project with the technical guidance of China Aluminum International Engineering Corporation Limited in June 2009. The checking results indicate that various indexes reached or exceeded the designed value, thus the performance checking was qualified.
2. Deng Feiyue, Huang Manxiang, Processing technology performance study of bauxite ore in Debao Guangxi, Mining study and development, 27th volume, 1st period, Feb. 2007. 3. Tang Shitai, Development on grinding process technology
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with two-section grinding and hydro-cyclone classification in alumina production. 11th volume, 5th period, Oct. 2001.
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Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
ADVANCED PROCESS CONTROL IN THE EVAPORATION UNIT C.Satish Kumar1, Tonmoy Banerjee1, Uttam Giri1, Rosalin Pradhan1, Ramu Saha1, Pratichi Pattnaik1 Vedanta Aluminium Limited; Lanjigarh; Orissa - 766027; India Keywords: Advanced Process Control, RMPCT, steam economy Abstract
Process Description
Energy consumption is a significant constituent of production cost in an Alumina refinery. Indian refineries are striving to reach global benchmarks and have made conscious efforts to minimize energy consumption. Evaporation plays a significant role in conserving energy and is responsible for 30% of the thermal energy consumption. In the Evaporation unit, due to a large number of interacting processes, frequent disturbances and the presence of single line equipment, it is necessary to maximize the utilization of assets to ensure the production volumes at targeted efficiencies. The need for reduction in process variability resulting in optimizing operations, challenged by plant constraints, was derived by adopting Advanced Process Control (APC) application in the Evaporation unit.
The evaporation unit removes water that has entered to the liquor circuit from bauxite moisture, mud washing, hydrate washing and miscellaneous dilution streams. The evaporation section consists of two identical parallel batteries of six backward feed, falling film evaporator units in each battery. The weak Spent liquor is collected and received in evaporator feed tanks. The weak liquor is evaporated in the six backward feed falling film evaporators and strong liquor is stored in the evaporator test tanks. In each evaporator train six nos. online calandrias are provided as tubular heat exchangers according to the falling film principle. Each calandria is equipped with a detachable cover and liquor distribution system (on top) for even distribution of spent liquor into the tubes.
APC incorporates a matrix of modeled dynamic process responses with an optimizer to maintain the unit at the most profitable vertex of the allowable operating envelope. This paper will discuss the area selection criteria, activities carried out during the project tenure and post benefit analysis.
Vapor separators and ducts are provided next to the calandrias for efficient separation of the vapor from the product. Low pressure (LP) steam provides the energy required for evaporation. LP steam is fed to the 1st effect from main header through flow control loop. A desuper heater is provided in the steam main header in order to extract superheat from the steam coming from main header and thereby preventing high temperature steam entering the evaporator body. Spent liquor is pumped from the feed tanks to feed flash vessel connected to 5th effect unit where it gets flashed and evaporated and then transferred successively to feed flash vessels connected to the 6th unit and mixing condenser for further concentration. After flashing the feed in three flash vessels, the liquor is pumped to the 6th unit. From the 6th unit, the liquor is pumped successively to the 5th, 4th, 3rd, 2nd and 1st units through transfer pumps. Finally the concentrate coming out of the 1st unit is flashed successively in product flash vessels connected to 3rd unit followed by 4th unit and thus final concentration is achieved. Each body is provided with a recirculation pump and corresponding piping. The recirculation pump circulates the required flow of liquor ensuring the proper wetting rates (to avoid any dry spots on the evaporating surface) to the top of the calandria on liquid distributor. The steam condensate from the 1st effect and 2nd effect is extracted withoutflashing.Balance process condensate is flashed three times before being extracted at 60 °C. Non-condensable gases from 1st unit to 6th unit are connected to a common header and in turn are connected to a condenser. The vaporsfromthe 6th unit & last feed flash vessel are condensed in a direct mixing type condenser (Barometric Condenser) with cooling water from an alkaline cooling tower entering condenser at 35°C and returning at 50°C. The vacuum in the evaporator is achieved & maintained through a two stage steam ejector system. LP Steam is being used in the ejectors to maintain the desired vacuum in the system.
Introduction The challenges for any alumina refinery are to minimize the cost of production per tonne of alumina, while meeting safety and environmental considerations, along with maximum production volumes. In view of the recent step increase in the cost of energy generated from fossil fuels, Alumina refinery operators are constantly looking for various ways and means to reduce specific energy consumption per tonne of alumina production. Utilization of new control technologies using existing infrastructure with a reduced support team has emerged as a new dimension for significant economic saving in alumina production. Multivariable predictive control technology becomes one of the main tools to optimize the capital investment. M/s Vedanta Aluminium Limited is located at Lanjigarh, Orissa state, India, having a capacity of 1MMTPA smelter grade alumina production, employing the Bayer process with low temperature and pressure digestion. Evaporation consumes about 30% of the thermal energy and is an ideal unit in Alumina refinery for APC implementation to generate significant potential savings. Apart from this, other benefits accrue, such as uniform test liquor caustic concentration, which is a key parameter in the Digestion unit for improving productivity. In classic process control, a controller brings a single value close to a set point using a single manipulated variable, or it can involve a reactive process based on currently measured conditions. Classic process control can do things automatically, but it can't respond to other kinds of variables, such as cost constraints. Process variability reduction resulting in optimizing operations challenged by plant constraints was derived by deploying an effective APC application in the Evaporation unit.
Area Selection criteria Digestion, Calcination and Evaporation are the three major areas for APC implementation to generate profits in terms of yield improvement, increase in throughput, minimization of standard
69
deviations of process parameters and reduced energy consumption. The digestion process has the most potential for APC benefits because it is a key production unit, critical for sustainability of the process and a good state of digestion control. Due to frequent changes in the quality of the bauxite feed to the refinery, it is difficult to control the process parameters in the digestion unit. So the option of APC implementation in the digestion unit was kept on hold tempora-rily. However, auto control of alumina to caustic ratio with test liquor and bauxite slurry has been implemented.
future response with respect to control. The prediction model will send communication to optimizer for taking moves on targets of Controlled Variables (CVs) and Manipulated Variables (MVs) and final control element will take action accordingly. The basic idea of the control strategy is to take action before a disturbance reaches the CVs. Disturbances which enter the process are detected using the predictive model and appropriate changes are made in the manipulated variables such that the controlled variables are held constant. Figure 2. shows the behavior of variables with and without control process.
The existing Calcination unit has been planned for an upgrade in throughput by 20%. After completion of the upgrade and achieving the rated throughput, APC implementation will be done in Calcination. The two Evaporation units are designed for 300 Tonnes/hr each. Due to lower bauxite quality than design (-37% Trihydrate alumina (THA) compared with a design of 41% THA), mud generation is on the higher side which results in more wash water demand per tonne of red mud to minimize residual caustic losses. So both the Evaporation units are operated with 65%-70% of their capacity. To maintain consistent product quality, minimizing the effect of disturbance and enhancing performance of single line equipment, raised a challenge to improve the steam economy, so as to reduce the specific energy consumption per tonne of alumina production.
MV Past Move* DV P»*t Vahtei CV Unforced CV Current Velue CV Unforced +Bi»s SP Funnel Error MV Move PUn CV ConlroDeil
Figure 2. RMPCT controlling a CV inside limits APC brings several values close to optimal targets simultaneously using several manipulated variables simultaneously. Also, APC includes a predictive process based on a process model that foresees how variables will behave in the future, and the controllers in an APC system find optimal trade-offs in case of conflicts among the goals. Because several actuators act simultaneously to achieve the best trade-offs, the application can achieve its goals in the fastest or most efficient fashion. In addition, APC's predictive character allows early recognition of potential violations and timely implementation of remedies, which means improved stability and reduced defects. It also lets applications work nearer to process constraints, which further reduces costs. A major objective of APC is to minimize the dotted area (Error) shown in Fig.2
Robust Multivariable Predictive Control Technology (RMPCT) Control designs in the process industry are almost exclusively based on PID controllers these days. Even though they are simple to implement and easily integrated into the control system, they quickly reach their limits when more complexity is involved. The development of advanced control theory has received considerable attention and applications in the process industry are gradually increasing. APC opens new opportunities, as even complex situations can be mathematically described with process parameters or variables and used for automatic, flexible plant operation. APC employs robust control techniques and can handle several constraints simultaneously by handling multiple inputs and outputs, and maximizes the profits derived by pushing the operations closer to limits.
APC Advantages ■ Standard deviations are minimized and faster adjustment can be done ■ Improvement in throughput, yield and consistent product quality ■ Manual intervention will become significantly less, which leads to a reduction in operator stress ■ Lifetime of the plant with the existing automation increases with less maintenance work ■ Significant reduction in energy consumption
Multivariable controller Predicted Steady Optimizer state values
H
Prediction
' Targets, WV targets
Process
Predicted! Controller values
Project Implementation Steps
•"H Model
The scope of the project was the implementation of APC in both evaporation trains of the alumina refinery. A detailed project planning schedule was prepared in collaboration with M/s Honeywell. The following activities were carried out over five months.
Prediction error
Figure 1. Advanced process control components Advanced process control consists of optimizer and prediction model. The function of the prediction model is to predict the
70
Data Collection and Analysis Base case data was collected to understand the operation of evaporation unit and evaluate the variability in process parameters. Six minute average data was considered for variability analysis. Minimum, maximum, average and standard deviation of all process parameters were collected for 15 days to understand the existing operation, and major process upsets during this period were discarded.
Functional Design The functional design specification was based on preliminary plant test and earlier data analysis, discussion with plant operating personnel and engineers and information gathered on operating constraints and future operating strategy. Since the operating variables in the evaporation unit are interacting with each other, RMPCT was selected for advanced control. Two identical RMPCT were selected for evaporation train land 2 to provide reliable operational control and optimization objectives, while maintaining the constraints within specifications. To achieve this, RMPCT manipulates a set of independent variables MVs to maintain a set of dependent variables CVs at target or within constraints. The controller also uses disturbance variables (DVs) as feed forward control to reject the disturbances. A linear programmed optimizer and dynamic control algorithm coordinate the movement of MVs. The optimizer functions were decided to maximize steam economy, improve product density control and to maintain process condensate conductivity within acceptable limit.
Preliminary Plant Test The main objective of Preliminary plant test was as follows: Estimate step size, dead time and settling time of the variables, Evaluate the tuning of the DCS controllers and changes accordingly wherever is required, Check the existing instrumentation. Preliminary plant test was performed for one week. During preliminary plant test, the operation of a few controllers was improved from earlier Manual mode to Auto mode by making the adjustment in tuning factors of the controllers. Fig.3 shows the controllability of steam flow in AUTO mode. This valve was operated manually earlier. After taking the controller in AUTO the fluctuation in steam flow has come down significantly.
^
7
^
Step Test A step test plan was prepared based on preliminary plant tests, functional design and earlier operational experience. The step test was required to build the process model, which is the basis of predictive control. The step test plan was designed to build up a good model without interrupting normal plant operation. A test was carried out for 15 days in both evaporation trains one after another. Eleven numbers of manipulated variables were considered for the step test plan. Those variables were Steam flow controller, Feed low controller, Bypass feed flow controller, Barometric condenser pressure controller, Barometric condenser cooling water flow controller and each effect's level controller. During the step test, one variable was moved at a time and all other variables were kept in auto mode. The test was carried out round the clock and a normal range of operation of the evaporation unit was ensured. The results of each step were monitored and analyzed continuously. Based on this analysis step size and numbers of steps were increased and or decreased in a few cases.
^
Figure 3. Steam Flow control before and after tuning
A few snapshots of the correlation between the MVs and DVs with CVs obtained after the step test are shown below. FIR Order 200 Lap Order 2 Settle T = 2 0 0 TfSettie = 1Q3 FIR Form = Pos Rank = 3Tria! 7 G[sl= .000171
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V^yT^s^yC^V 150 2£ 100 -20$
193**.2 + 27 8s * 1
Figure 5. shows Correlation between Product Density and Feed Flow
»p •x*
.*,'.*<
Figure 4. shows Product Flash Vessel#3 level before and after tuning
71
FIR Order 1 2 0 L a p Order 1 Settle T - 1 2 0 TfSettle « 115 FIR Form « Pos Rank = 2 Trial 5 Q 1 Gis) - 0 6 6 26.7* + 1
After the detail validation of the model, the APC Server configuration was done in the existing Experion Process Knowledge System (EPKS) network. Ole Process Control (OPC) client was installed in the APC server for the communication to the OPC server. Redirection manager was installed in APC server machine to communicate to the redundant server in case the primary server failed. A Distributed Control System (DCS) resident watchdog timer was maintained to allow mode shedding of MVs in case of any communication loss between DCS & the profit controller. The APC controller plan moves for all the MVs, and those moves, will be written to DCS MVs. Finally, the APC model was successfully commissioned in both evaporation units, 15 weeks after the start of the project.
120
Figure 6. shows Correlation between Vacuum and 1st effect temperature difference Training Two types of training were given: 1). Engineer Training before Step test: 25 engineers attended engineer training. The objective of the training was to understand the basic function, design and implementation of APC. Also, identification of control models and tuning of the controller were covered in the training. 2). Operator Training after Commissioning: All concerned operators were trained under the operators training in which procedures for taking APC online, offline, and emergency control were covered.
Benefit Analysis Reduction of Variability in CVs Significant reduction of the variability of all the CVs has been achieved as measured by standard deviation and shown in normal distribution graphs. For example, Product density is the important controlled variable in the evaporation circuit for product quality control. Before the implementation of APC the product density variability had a range from 1.321 to 1.350 whereas, after APC implementation the range narrowed to between 1.302 and 1.310. Product density standard deviation has reduced from 0.006 to 0.001. The following figure shows the variability reduction before and after APC implementation.
Detail Design and Commissioning The APC detailed design was developed based on the information generated during the Plant step test. To maximize the steam economy, the optimizer was designed to minimize the steam flow, maximize the vacuum, minimize the product density and minimize the temperature difference across each effect. In order to achieve the minimum temperature difference across each effect, optimizer was designed to maximize the feed flow through the evaporator and minimize the bypass feed. During the step test it was observed that the step change in level controller of each effect didn't give any correlation with CVs. So the basic controller was designed by taking five MVs which are Feed Flow Controller , Bypass feed flow controller, Steam Flow controller, Barometric condenser pressure controller, Barometric condenser cooling water flow controller and fifteen numbers of CVs out of which the important variables were Product Density, Process Condensate Conductivity, Temperature difference across each effect, first effect jacket temperature, Feed density, feed temperature, inlet water temperature to barometric condenser, deviation of feed flow and barometric condenser pressure from set point has been taken as DVs in the detail design. Preference has been given for MVs control in the following order: Feed flow Controller, Direct condenser pressure Controller, Steam Flow Controller, Bypass feed flow controller, Direct condenser cooling water flow controller.
Figure 7. shows Normal distribution curve for the Product density before implementation of APC
Two types of limits have been set for all MVs and CVs, one is the engineering limit and other one is the operating limit. The engineering limit has been set to ensure safe operating range of equipment. Operating limits are set to serve the process control and optimization. The operating limit is accessible to the control room engineer. The high limit and low limit of operating range are set based on forward unit's requirement and overall condition of the plant. After the detail design, the model was run in offline mode at M/s Honeywell, Pune. The validations with respect to achieving the objective were carried out at different input conditions. The direction of movement of each CV with respect to the step change in each of the MVs was checked and validated.
Figure 8. shows Normal distribution curve for the Product density after implementation of APC.
72
Maximization of Steam Economy: Steam Economy has increased by an average of 2% after implementation of APC. This was achieved mainly by minimizing the direct condenser pressure and optimizing process parameters continuously. Before APC implementation, the average direct condenser pressure was -648 rnmHg(g), which improved to -655 mmHg(g) after implementation of APC. Previously, maintaining pressure below -655 mmHg (g) was a constraint for operation as it was leading to an increase in process condensate conductivity intermittently, because of disturbances created by other process parameters. But under present conditions, due to continuous optimization by APC, lower pressures up to -662 mmHg were achieved without deterioration in the process condensate, as well as the direct condenser cooling water quality. Barometric Cond. Pr.Actual kmujiv2}rjrtwNe>HMirtN»TH«ujr>.cnH«inr«.o>t -645 - — — B i—,.y,i
-650
-655 -660 *0*¥4Φ*Ε*Τ*
m
; -665
I -670 Figure 9. shows the direction of movement of condenser pressure towards high limit after giving a step change in the limits. Conclusion RMPCT implementation in the Evaporation unit at Lanjigarh was completed within the planned project tenure and has proven successful benefits. RMPTC provided an opportunity to operate the plant closer to the operational constraints which helped plant debottlenecking without adding assets. Project outcomes such as minimizing standard deviation of process parameters, improvement in throughput, plant uptime with less breakdown maintenance and finally improvement in steam economy were successfully delivered. The way forward involves Digestion and Calcination as the next key units for APC implementation to maximize the production volume and throughout.
References 1. Ayana Oliveira, Jefferson Batista, Jedson Santos, Marcia Ribero, Jorge Charr, Rafael Lopes "Advanced process control in alumina digestion unit" Page no: 91-96, Light Metals 2009 2. Robert K.Jonas, "Application and benefits of advanced control to alumina refining", page no: 43-49, Light Metals 2004 3. Freeman N, Pandya M and De Roos R, "Simulation applications for alumina refining", page no: 52-55, Alumina quality workshop 2008 4. Oliveria A, Batista J, Ribero M, Charr J and Lopes R, "Robust multivariable predictive control technology implementation in an alumina digestion unit", page no: 47-51, Alumina quality workshop 2008
Acknowledgements The authors wish to gratefully acknowledge Mr. Jaydeep Bhattacharya and Mr. Vinoth Ramachandran from M/s Honeywell, Pune for providing assistance during the project.
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
IMPROVEMENTS IN SMELTER GRADE ALUMINA QUALITY AT CLARENDON ALUMINA WORKS Ruth Shaw1, Ajamu Duncan2, Marlon Crosdale3 Clarendon Alumina Works; Halse Hall; Maypen, Clarendon, Jamaica Keywords: Diastar, Process Management, Hydro Cyclone Abstract Alcoa's Clarendon Alumina Works recorded a significant improvement in its smelter grade alumina, SGA, quality. This marked improvement came about as a result of changes in green liquor filtration technology and process management in precipitation. This paper discusses the impact that these changes had on iron, calcia, titanium, silica and soda in SGA. Introduction Alcoa's Clarendon Alumina Works, CAW, located in South Central Jamaica, is a low temperature digestion plant with an annual production capacity of approximately 1.4 million tonnes SGA. The refinery is ISO 9001:2008 certified.
Parameter
2006
2007
2008
2009
%Fe 2 0 3
0.019
0.010
0.008
0.009
%Si0 2
0.017
0.017
0.016
0.014
%NazO
0.45
0.45
0.40
0.43
%CaO
0.057
0.054
0.051
0.050
%Ti02
0.002
0.002
0.001
0.001
Table 1: CAW Alumina Quality 2006 - 2009
Since 2007 the quality of the alumina produced at CAW has made progressive improvement as measured by the Customer Satisfaction Index (CSI). The CSI is the metric that measures the refinery's ability to meet critical parameters required by the smelter customers. For Clarendon these parameters include iron oxide (Fe203), silica (Si02), sodium oxide (Na20) and calcia (CaO). These parameters are affected by the quality of the liquor sent to precipitation and operating practices. At the beginning of 2007 the CSI averaged 55%. Following the commissioning of improved technologies and improvement in process management systems, the CSI increased to an average 92% in 2009. See Figure 1. Table 1 show the trend of the alumina quality parameters in SGA for the period 2006 - 2009.
Iron oxide (Fe203), titanium (Ti02) and calcia (CaO) all showed a step reduction which was directly related to the commissioning of the high pressure diastar filters. This resulted in a significant drop in the particulates in the liquor sent to the precipitation circuit. Further, improvement in thefiltrationprocess control and monitoring minimized deviations and sustained gains. In precipitation, strategic changes in process management and hydrate classification using hydro-cyclones were the main contributors to the improvement in silica and soda levels in SGA. Improvement in Fe203, CaO and Ti0 2
Customer Satifaction index
The Effects of High Pressure Diastar Filters grioo
Since the construction of the refinery in the early 1970s the plant has employed the use of sand filtration technology to remove suspended particles from the liquor being fed to the precipitation circuit. In this process the liquor is filtered by gravity through a bed of suitably sized sand and the filtrate fed to the precipitation stage.
E 80 c
t
40
ä
20
There were several issues associated with sand filters that resulted in high particulates in filtrate. These included:
m** 1
y**1
10*
„fi*
y**
<**
* *
Frequent screen ruptures and sand leakages Changes in the particle size distribution of the sand bed over the life cycle of the filter
0ec^
Figure 1: Customer Satisfaction Index 2007 - 2009
75
No efficient way of flushing out the mud entrained in the sand bed during the filtration cycle.
Operations Management: Re-clothing practices Process Control to prevent Cloth Degradation Inventory Management of Filter parts Cloth Design
As a consequence of aforementioned, the operations experienced high variability in the control of iron oxide, titanium oxide and calcia in SGA.
Process Management: Changes to dosing of flocculants to thickeners A visual management system Spigot sample to reduce time taken to identify defective filter elements.
Iron in SGA and Filtrate Solids — -Iron m SGA
——Filtrate Solids
0.04 0.035 0.03 g
0.025
S c
0<02
|
"I 4"AJ
ITP Calcia and Calcia in SGA — -Calcia to Precip
io I
0.015
»10
i^>->^.r^·.
0.01
90
o 1
/V
œ
70
16°
-20
(*>«r°1
^ * &
a#*
i9itt&
g 0ec"^
40
,5
Figure 2: Iron in SGA and Filtrate solids 2006 - 2009 Titanium in SGA and Filtrate Solids — »Titanium in SGA
/^^
·§ 50
-30 i^·01
0.09 0.08 0.07
<:■■■: M - k k
Λ -0M
a!
0.05 * | \
/
, V v - ^
\
■■ ' \
•
t
χ
X
^
•
* ' Λ * * ' * ,
0.04 O
30 20
0.03 0.02
10
0.01
yvtfv06
— F i l t r a t e Soiids
k f\
; :
80
0.005 0 y,***
— C a l c i a in SGA
oe*06
itf**1
oec-0 1
ypSfi
o e
c ^
&*&
o e
c-^
Figure 4: Calcia in SGA and LTP Calcia solids 2007 - 2009 Improvement in Na 2 0 and Si0 2
y****
v***
ju*°1
ç**
1
jtfv**
o e c ^ „g***
In 2006 CAW s precipitation department advanced its classification system by moving from gravity classifiers to hydrocyclone classifiers. These changes required modification of the hydrate management and first precipitator control strategies so as to assure that the soda and silica levels in product meets the customer specification. Figures 5 and 6 illustrate the improvement in both the variation and absolute value of the Soda and Silica in product.
oe c-«
Soda in SGA ofi
Figure 3: Titanium in SGA and Filtrate solids 2007 - 2009
0.55
During the 2007 plant expansion, the refinery replaced the sand filters with Diastar filters which are widely regarded as the most advanced fine filtration technology.
0.5
The change to Diastar filters resulted in the step reduction in iron oxide and titanium oxide in SGA. Refer to Figures 2 and 3.
* 1
cu
* 0.35 0.3
Process Improvements - Diastar Filters and Thickener Operations
0.25 -
Following the commissioning, the refinery was faced with the challenge of adapting to the Diastar operations. Significant improvements were made in the following areas with the gains in calcia control being realized during 2008. See Figure 4.
**<*
#**
tfH*1
yfi*
ç**
***
„**
*#*
Figure 5: Soda in SGA 2006 - 2009
16
o*&
«*'*
CAW experienced high variability in its first precipitator soda control immediately following the introduction of the hydrocyclone classifiers in 2007. Changes in head tank control with the improved classification system positively influenced the fixed soda in SGA. These changes allowed the sustained reduction in the average value of soda in product from 0.45% in the second half of 2008 to 0.41% in 2009 and a reduction in variability from 0.09% in 2008 to 0.03% in 2009. Silica in product is CAW's highest-weighted component capturing over 35 % of the CSI. Silica in SGA is affected by inter alia hydrate seed in contact with liquor for extended periods of time. Vessel failures in precipitation are one of the leading causes that present this condition. Silica in SGA - Box and Whisker
T
s
-E 3-
H2 - 08
Π.
H1 - 09
H2- 09
Figure 6: Silica - Box and Whisker plot Introduction of a new hydrate management system and the revision to the procedure for returning a failed vessel to service has eliminated occurrences of silica excursions. In 2007 silica averaged 0.17 + 0.02 while in 2009 the silica averaged 0.14 ±0.02 Conclusion CAW has significantly reduced iron oxide, calcia, titanium oxide, soda and silica in SGA through: Switching from gravity sand filter to short cycle Diastar pressure filters Improved process management and control of the Diastar and precipitation operations CAW has seen marked improvement in its product quality and continues to drive for improvement through the process of best practice transfer. References 1 Diastars filters, manufactured by Filters Gaudfrin (France), are cylindrical shaped pressure vessels with a cone shaped bottom and a domed shape removable top. They utilize a filter cloth as the filter medium, similar to a Kelly filter.
77
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
Light Metals 2011 ALUMINA and BAUXITE
Red Mud SESSION CHAIR
Linda-Therese Kristiansen Hydro Aluminium Oslo, Norway
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
APPLICATION OF NANOFILTRATION TECHNOLOGY TO IMPROVE SEA WATER NEUTRALIZATION OF BAYER PROCESS RESIDUE Kelvin Taylor1}, Mark Mullett^, Lee Fergusson2), Helen Adamson1* & Juerg Wehrli1} υ Hatch Associates Pty Ltd, 144 Stirling Street, Perth, WA 6000, Australia 2) Virotec Global Solutions Pty Ltd, Coomera Waters, QLD 4209, Australia Keywords: Nanofiltration, Membranes, Red Mud, Bayer Process, Sea Water Neutralization plants. Typical sea water NF plant operating pressures are around 18 bar compared to around 60 bar for SWRO plants [2]. The capacity for concentrating calcium (Ca) and magnesium (Mg) ions is therefore likely to be limited by precipitation rather than osmotic pressure.
Abstract Sea water neutralization of alkaline Bayer process solid waste residue is a sustainable solution to turn a hazardous waste material into a benign, non-hazardous material that can be re-used in some applications. The concentrations of the active neutralizing agents in sea water, calcium and magnesium, are low and large volumes of sea water are required to neutralize the alkaline residue.
The production of alumina by the Bayer process results in the generation of large quantities of "red mud". Between one and two tonnes of residue are typically generated for every one tonne of alumina produced and, due to its high alkalinity and pH, the red mud generated from alumina refineries is classified internationally under the Basel Convention (Article 1, paragraph la) as a "hazardous waste" (classification #B2110) and its trans-boundary movement is tightly controlled [3].
Nanofiltration of sea water can produce a concentrate containing up to four times the calcium and magnesium levels in sea water, while the mono-valent ions, like sodium and chloride, are only marginally concentrated. This significantly reduces the volume of sea water required for the neutralization process and hence the size of neutralization equipment. Other advantages include improved reaction kinetics and reduced salinity of the neutralized residue, making it more suitable for certain applications. The nanofiltration permeate, with lower scaling potential, is also an improved feed for sea water reverse osmosis plants to produce potable water.
The alumina industry has for many years sought ways to minimize and reduce the environmental and social impacts of red mud and alumina refineries face two basic choices in managing their red mud residue, either:
Introduction
i)
store the residue indefinitely at considerable cost; or
ii)
attempt to reuse it in some form [4],[5].
Long-term storage costs and environmental liabilities are high, and it is estimated that management of residue in Australia alone costs more than AUD80 million per year [6].
Nanofiltration (NF) is a membrane-based separation method that is similar to reverse osmosis (RO) but the NF membranes are more selective in retaining chemical species than RO membranes. RO membranes retain >99.0% of all chemical species in solution, except for some small, polar organic molecules such as acetone and acetaldehyde. NF membranes on the other hand will allow mono-valent ions such as sodium and chloride to pass through to permeate but will retain divalent species, such as calcium and magnesium, in the retentate.
Refinery waste reuse requires some form of conversion from its highly caustic state to a more environmentally benign and "userfriendly" state for both transport and reuse applications. Thus, it is desirable to neutralize refinery residues in a way that they are no longer highly caustic, irrespective of their potential or actual reusability. Most neutralization options that attempt to address red mud residue storage and management problems do not make full use of the inherent acid neutralizing, metals and phosphate binding, and odour destroying properties of the neutralized red mud residue. One method for neutralizing alumina refinery residues, called Basecon Technology can be used to economically convert the residue and thereby imbues it with environmental and commercial value [7]-[ll]. Such a technology can reduce the long-term environmental, financial and social liabilities of storing caustic residues, and has the added advantage of producing a reusable solid material, called Bauxsol Raw Material (BRM), which has favorable environmental and industrial reuse characteristics. Different reactants, including sea water, brines, concentrated brines and metal salts can be used to neutralize red mud residue.
In general, the species that are retained by an NF membrane are multi-valent inorganic ions and large organic molecules. Different NF membranes have different rejection characteristics depending on their surface chemistry and surface charge density and are normally characterized by measuring the rejection of magnesium sulfate, typically in a range between 95% and 99%. Typical applications for NF membranes are concentration of dissolved metals in acid solutions, concentration of dyes and sugars, removal of chloride from sulfate solution and removal of calcium from solution to stop scaling. Nanofiltration technology has also been applied to pre-treat sea water for desalination plants, where the NF membranes have been used to remove calcium, magnesium and sulfate prior to reverse osmosis treatment. It has been shown that the RO membranes, or thermal equipment in the case of distillation plants, operate more efficiently if NF pretreatment is applied [1].
Sea water neutralization is the most widely implemented neutralization process, due to its cost-efficiencies and readily available feedstock [12] for alumina refineries located close to the sea, and having environmental licenses that provide access to sea water. When sea water is used to neutralize red mud, the volume of sea water added is between three and ten times the volume of
The concentration of species retained by the NF membranes is low and hence the operating pressure is also low when compared to the operating pressures of sea water reverse osmosis (SWRO)
81
red mud slurry being treated (also refer Table 4). Nanofiltration now offers an opportunity to significantly reduce the volumes required. The expected advantages of using NF concentrate to neutralize red mud are: •
Reduction in reacting volume and therefore tanks, pumps, piping and the capital and operating costs associated with this equipment.
•
Reduction in salinity of the neutralized product, thus making it more useable for soil remediation and other similar applications.
•
MEMBRANE FILTRATION
m
Production of an NF permeate stream that will perform better in SWRO plants than untreated sea water.
PERMEATE
Figure 1: Schematic Diagram of Test Arrangement
An R&D program was initiated to determine the extent to which sea water could be concentrated using NF membranes before scaling commenced, or before the osmotic pressure became excessive. In addition the comparative reactivity of the sea water NF concentrate and untreated sea water with red mud was investigated in order to gain an understanding of the equipment capacities required and capital reduction benefits that could reasonably be anticipated by alumina refiners. Experimental Membrane tests were conducted in two phases: laboratory bench tests (2L start volume), and pilot scale (lkL start volume). Cross flow filtration over a membrane surface under pressure was used in all tests to effect the chemical separations, the aim being to concentrate magnesium and calcium by rejecting it to the retentate, while passing sodium chloride to the permeate. All tests were run using a sample of sea water collected from the Western Australian coast near Perth, and pre-filtered using 5 μπι cartridge filters.
Figure 2: Pilot Test Rig
Two types of tests were conducted. Perturbation tests (P-tests) determined the relationship between rejection, pressure and flux for a given membrane and solution type, while recovery tests (Rtests) provided the relationship between operating pressure and flux as the volume recovery and feed analyte concentration increased. The results were used to:
The first pilot scale test run was performed at a constant pressure of 3200 kPa. The flux was allowed to decline from an initial flux rate of 45 LMH as the osmotic pressure increased and conductivity, temperature and pH measurements were taken at increments of 10% recovery up to a maximum recovery level. The maximum recovery level was characterized by a significant rate of flux decline.
i. Determine flux behavior over a range of operating pressures for a given solution matrix and assist in determining which membrane type is most suitable for a given application. ii.
Feed and permeate samples were collected at various recovery points and analyzed for Ca, Mg and S by ICP-OES. The concentrated feed sample produced at the maximum recovery level of 80% during the first pilot run was then used for comparisons of neutralization characteristics with untreated sea water by titration of the free alkalinity of a generic Bayer residue water (refer Figure 3). Analytical results from this first pilot run were also used as the input to modeling of the neutralization using Basecon Technology (refer Table 4).
At pilot scale, provide quantitative data that facilitates an engineering design process.
During R-testing, the retentate is recycled to the feed tank such that the concentration of rejected species increases, causing the osmotic pressure to increase toward a maximum value. A schematic diagram showing the layout of the equipment for both the laboratory and pilot scale tests is shown in Figure 1.
A repeat of the previous R-test was performed at a reduced starting pressure of between 1000 and 2400 kPa and then gradually increased to achieve a constant flux rate of 35 LMH. As recovery increased beyond 60%, samples were collected at recovery increments of 5%. P-tests were conducted concurrently during the test at a range of recovery levels to better determine the ideal recovery recommended for a full scale process.
Laboratory scale tests were conducted using 21.5 cm2 flat sheet membranes that represented three commercially available spiral wound NF membranes. The capacity of each membrane to concentrate Ca and Mg into the retentate at suitable NF operating pressures (<4000 kPa) and flux rates (>20 liters per square meter per hour - LMH) were compared and the best performing membrane wasfinallytested in the pilot scale program.
82
magnesium and calcium hydroxides and reactions with carbonate ions also start to occur.
Results and Discussions Table 1 shows that, of the three membranes P-tested in the laboratory, the Dow DK membrane demonstrated superior magnesium rejection performance at relatively low operating pressures and it was selected for use in the pilot scale tests. The TS80 membrane achieved similar magnesium concentrations, however the flux decline was more significant and the applied pressures required were much higher. The TS50 achieved the lowest magnesium rejections. Applied
Recovery
Mg Rejection
Initial Flux
Final Flux
[kPa]
[LMH]
[LMH]
[%]
TS80
2820
35
9.5
60
98
TS50
1700
35.4
22.5
75
92
DK
1425
35.4
18.6
78
97
Type
Pressure
achieved
[%]
(3)
Ca2+ + C032- ->CaC0 3
(4)
Mg2+ + C032" ->MgC0 3
(5)
Once the hydroxide concentration has sufficiently reduced, the hydrotalcite-like compounds precipitate at suitable nucleation sites, typically on the previous precipitates. These compounds are layered double hydroxide compounds with the general formula [M(II)]1.z[M(III)]z(OH)2[A]z.mH20 where M(II) and M(III) are divalent and trivalent metal ions and A is an anion e.g. C0 3 2 \ S0 4 2 \ Cl-, OH\ Typically hydrotalcite, hydrocalumite and paraaluminohydrocalcite can be produced as shown in the reactions below:
Table 1 : Membrane Screening Results Summary I Membrane
Al(OH)4- -> Al(OH)3 + OH-
1
6Mg2+ + 2Al(OH)4- + 80H" + C032" + 4H20 -> Mg6Al2C03(OH)16.4H20 (6)
The sea water neutralization reaction was simulated by titration of 250 ml of a generic Bayer residue water with both untreated sea water, and sea water concentrate (NF retentate). The aim of the titration experiments was to confirm that the increase in magnesium and calcium concentration did not change the reactions from those typically observed during sea water neutralization and the pH was allowed to equilibrate after each addition. Samples were also collected at the start, at the perceived inflection point (-11.5 pH for sea water and -10.3 pH for NF concentrate) and at the end of the titration (~pH 9.4) and details of analytical data are shown in Table 2.
2Ca2+ + Al(OH)4' + 30H" + 3H20 -* Ca2Al(OH)7.3H20 (7) Ca2+ + 2Al(OH)4- + 2C032- + 3H20 -* CaAl2(C03)2(OH)4.3H20 + 40H (8) Determination of the exact stoichiometry is impossible due to the number and complexity of the reactions taking place.
Table 2: Chemical Analysis of Titration Solutions Bicarbonate Carbonate Hydroxide
A|k°..
a
.
[mg Na2C03/l]
Aluminium
Calcium
Magnesium
[mgAI/l]
[mg Ca/I]
[mg Mg/I]
3100
0.5
<0.1
0.34
410
1200
Bayer Residue Water
<1
Sea Water
.
Residue & Sea Water at Inflection
<1
5300
160
5460
620
4
8.4
2800
480
<1
3280
2.4
110
220
0.55
760
6600
6700
4500
<1
11200
40
12
18
9000
720
<1
9720
0.71
240
1900
Residue & Sea Water at end Sea Water NF Concentrate Residue & Concentrate at Inflection Residue & Concentrate at end
8200
46000
54200
.
|
|
Figure 3: Mass Mg Added vs pH for Concentrated Sea Water in Residue Water Inspection of the neutralization curves (Figure 3) shows that there are two areas where a small inflection is observed: around pH 11.5 and pH 10.3. Thefirstinflection point corresponds to the pKsp of magnesium hydroxide of 11.3. This is expected because magnesium is present at three times the concentration of calcium in sea water and in sea water NF concentrate. The second inflection corresponds to the carbonate/bicarbonate equilibrium as the hydrotalcite-type reactions (equations 6, 7 and 8) remove carbonate from solution. Analysis showed that while some hydroxide and aluminate was still present at the upper inflection point they had all been consumed by the lower inflection point. The alkalinity remaining at the lower inflection and the end point was due to the presence of carbonate and bicarbonate ions. These observations were similar, regardless whether sea water or sea water NF concentrate was used for neutralization.
The titration curves (Figure 3) show similar symmetry with two inflection points when the added reagents are normalized to magnesium addition, suggesting that similar reactions are taking place. The causticity in a Bayer residue water can exist as both free hydroxide (OH) and as the aluminate ion (Al(OH)4). Further alkalinity is provided by the carbonate ion (C032~). When this Bayer residue water is reacted with sea water, or sea water concentrate, there are a multitude of reactions that can take place. The initial reactions precipitate magnesium and calcium hydroxides (brucite and portlandite) thereby lowering the pH. 3Mg2+ + 60H- -» Mg3(OH)6 2+
Ca + 20H" -* Ca(OH)2
(1) (2)
Two points of difference between the sea water and sea water NF concentrate neutralization curves are evident: i) at the end point
As the hydroxide concentration reduces, the aluminate ion precipitates, releasing hydroxide ions which then form more
83
where a lower pH was reached using sea water NF concentrate (9.0 cf 9.4 pH), and ii) that the endpoint was achieved much sooner. This is explained by the higher concentration of reactants in the sea water NF concentrate pushing the reaction equilibrium further to the right (equations 1 to 5), thus removing more hydroxide, carbonate and aluminate ions from solution thereby reducing the pH. There is also less dilution if the sea water NF concentrate is applied, which reduces the total amount of unreacted calcium and magnesium in solution and hence improves the magnesium and calcium reaction efficiency.
The suitability of using sea water NF concentrate with Virotec's Basecon Technology was also considered. Untreated sea water and NF concentrate were compared as the feedstocks for neutralization of three different alumina refinery red mud residues in different parts of the world. The sources were: i)
a refinery in Australia, processing bauxite from North Queensland using the Bayer Process; ii) a refinery in North America, processing bauxite from Jamaica using the Bayer Process; and iii) a refinery in China, processing diasporic bauxite from China using the Lime Sinter Process.
Although a faster reaction with concentrate was observed, in good agreement with Virotec's experience, this was not investigated further. Alumina refinery residues have different properties and reaction kinetics need to be specifically evaluated with on-site pilot tests which will determine actual reaction rates and equipment sizing.
The chemical composition of neutralizing streams used in the kinetic model was based on the world average sea water, and the chemical composition of NF concentrate at 70% recovery as analyzed in the first pilot trial. This was the point at which scaling became apparent and therefore the maximum condition for operation. Details of the chemical composition of the two neutralization reagents are shown in Table 3.
The results for the first pilot scale test are shown in Figure 4. In addition to observing the deportment of magnesium in this test, the behavior of calcium was also considered of high importance as the propensity of calcium sulfate to form a scale on the surface of the membrane will lead to rapid flux decline. A highly concentrated layer of calcium and sulfate ions at the membrane surface can form as a function of increasing bulk concentration of each ion in the feed stream and as a function of flux. The formation of this layer is called concentration polarization and was identified in the tests by a sharp decline in flux for a given pressure (Figure 4). The determination of the conditions at which this significant flux reduction occurs is of considerable importance in determining the achievable recovery level.
Table 3: Chemical Analysis of Reagents used for Basecon Technology Computer Model of Sea Water Neutralization Kinetics 1 Chemical Analysis 2+
Sea Water
NF Concentrate |
[mg/1]
412
940
Mg2+ [mg/1
1290
4000
SO/' [mg/1]
910
9400
K+ [mg/1]
400
645
Na+ [mg/1]
10400
13000
Cf [mg/1]
19400
27100
pH
8.3
8.0
Ca
|
1
1
The Basecon Technology neutralization kinetics for each of these three refinery residues was modeled using Virotec's computer model. Table 4 presents the pre- and post-neutralization properties and data of the three alumina refinery residues investigated. Table 4: Basecon Technology Kinetic Model Outputs for Three Different Refinery Residues 1 |
Figure 4: Pilot Test 1 - Analytical Trends
Refinery
North America
Australia
China
Bauxite Source
Jamaica
Nth Qld (Aus)
Central China
1
Process
Bayer
Bayer
Lime Sinter
|
1 1
Properties :
Figure 4 also shows the concentration of each analyte in the feed at each recovery level and the flux at each recovery level. The rejection of chemical species was very high and resembled those achieved in the screening tests recorded in Table 1. At 80% recovery the magnesium rejection was 95% and the sulfur rejection was up to 100% (the limit of detection for these analyses was 100 mg/L). The calcium concentration at 80% recovery was less than that at 70% recovery. This implied that some calcium precipitation occurred at about the 70% recovery level. It can also be seen that the decrease in calcium concentration coincides with a significant increase in the rate of decline in flux and this supports the hypothesis that concentration polarization of calcium sulfate was occurring. As a result recovery levels were reducing because of the resultant rapid rate of flux decline at the applied pressure of 3200 kPa.
1
pH of residue before neutralization
12
12.4
pH of residue after neutralization
9.0-9.5
9.0-9.5
11.6 9.0-9.5
Total alkalinity of residue before neutralization [mg Na 2 C03/L]
19,000
11,370
44,000
<500
<500
<500
5:1
3:1
10:1
2: 1
0.9:1
3: 1
2.5
3.3
3.3
[Total alkalinity of residue after neutralization [mg Na2C03/L] 1 Treatment ratio of Seawater to refinery residue [kL/kL] 1 Treatment ratio of NF Cone, to refinery residue [kL/kL] |
Ratio of Sea Water to NF Concentrate
|
The model supported pilot testing results showing the end point is reached sooner and the volume of neutralizing reagents can be reduced approximately three fold and hence there is significant potential to reduce pumping, piping and size of major equipment if an NF concentrate is applied.
84
The impact of concentration polarization can potentially be reduced by lowering the operating pressure and therefore flux rate, leading to greater recovery levels. This hypothesis was assessed in the second pilot scale test.
reduced to 7.8% at 80% recovery. This is another indication of calcium precipitation on the membrane surface and in the feed. At 85% recovery it was observed that precipitation was occurring in the feed vessel as well.
The applied pressure was varied as required to maintain the flux at a lower constant target flux rate of 35 LMH. The test incorporated a number of P-tests at various recovery points. The operational data and the species transmissions are summarized in Figure 5, and the relationship between flux, pressure and recovery is shown in Figure 6. By comparing with Figure 5 it can be seen that controlling flux and minimizing concentration polarization improves permeate recovery, and therefore calcium and magnesium concentration. It is apparent in Figure 6 that the pressure / flux curve at 75% recovery is almost identical to those at lower recoveries, and the onset of scaling has been delayed by reducing flux.
For these reasons 75% recovery was chosen as the practical concentration limit for a sea water NF plant. Pilot Test No 2 (P-Tests) : Flux vs Applied Presure at Different Recovery Rates
Pilot Test No 2 : Concentrations of Ca, Mg & S in Retenate vs Recovery 1600
2000
2400
Applied Pressure [kPa]
Figure 6: The Relationship between Pressure and Flux at a Range of Recoveries Anti-scalents are commercially available to reduce the rate of nucleation and these products achieve an induction period during which scaling can be avoided. However, they do not indefinitely prevent the precipitation of minerals such as calcium sulfate under supersaturation conditions and the application of anti-scalents was therefore not considered in this investigation. Another advantage of sea water NF is the fact that the NF permeate is largely free of the scaling species and it is therefore an ideal feed to a sea water RO (SWRO) plant producing potable water.
Figure 5: Pilot Test 2 Analytical Trends Predicting the point at which calcium sulfate nucleation occurs is very important in managing a sea water NF process. Precipitation of calcium sulfate is suppressed by the total ionic strength of sea water (in excess of 45 g/L TDS). Residence time in a membrane element is also short (<0.2 sec). As a result, the conditions under which precipitation occurs cannot be readily determined from standard thermodynamic data. The point at which calcium sulfate precipitation occurs on the membrane surface was therefore determined empirically based on the observation of flux versus pressure relationships.
Simulations were performed using Dow's Reverse Osmosis System Analysis (ROSA) simulation software to compare untreated sea water feed with the NF permeate feed. The results are summarized below in Table 5.
Table 5: Comparison of SWRO and Sea Water NF and RO Sea Water RO
Figure 6 shows a linear relationship between flux and pressure at low flux rates. A breakpoint is observed at a flux rate of about 45 LMH, the critical flux. This is the point where the membrane was operated in pilot test 1, explaining why the recovery results were poor. Poor membrane performance in pilot test 2 also occurred around 75% recovery, evidenced by the sideways shift in the 80% and 85% flux / pressure curves as shown in Figure 6
Feed Row Basis NF Permeate Recovery RO Plant Feed RO Permeate Recovery
For each recovery level up to 75% the relationship between applied pressure and flux was very similar and the applied pressure to achieve these recoveries was approximately 2200 kPa. Increasing the pressure beyond 2200 kPa caused a significant rate of flux decline suggesting that resistance at the membrane surface increased significantly with the increase in applied pressure causing a concentration polarization effect. The maximum achievable recovery, while maintaining a flux of 35 LMH, appears to be 75%. Beyond this point, the pressures required to achieve an equivalent flux increased significantly. The transmission of calcium at 75% recovery was 8.3%, however this
Potable Water Production
Sea Water NF &| RO
1000 m3/hr
1000m3/hr
|
N/A
75%
|
1000 m3/hr
750 m3/hr
|
48%
64%
1
3
3
|
480 m /hr
480 m /hr
The precipitation of calcium sulphate is performance limiting in both traditional SWRO and multi-effect distillation (MED) plants. When using sea water NF permeate as a feed to the RO unit, the limiting factor is instead osmotic pressure and the results of this comparison indicate that permeate recovery will increase from 48% to 64%. Hence, a significant reduction in the RO unit capacity can be achieved for equivalent potable water production,
85
reducing capital and operating costs and significantly reducing waste retentate discharge from the process.
and chloride concentrations are maintained close to the original concentrations of sea water.
Meeting fresh and potable water requirements will continue to be a global challenge for alumina refineries. For those refineries in proximity to the sea and with environmental approval to access sea water, the opportunity to combine sea water NF and RO for potable water and neutralization of refinery red mud solid residue using NF concentrate (retentate) offers great benefits.
This NF concentrate improves the reactivity and kinetics of sea water neutralization of alumina refinery residue and the required volumes of NF concentrate are about three times lower than for neutralization with untreated sea water. The size of neutralization equipment like pumps, piping, neutralization tanks, settlers and décantation and drying ponds are significantly reduced, leading to reduced capital costs and footprint for new operations or enabling expansion of capacity for existing alumina refineries using sea water neutralization.
The capital and operating costs for a sea water NF & RO plant will be higher than for an SWRO plant due to the extra membrane installations and pumping costs respectively. However the extra costs for the sea water NF plant need to be compared to the significant cost reductions from the smaller neutralization equipment required, where the volumes of NF concentrate added are about three times lower than for traditional sea water neutralization (refer Table 4). Pumps, piping, neutralization tanks and settler size are significantly reduced for an equivalent amount of neutralized red mud and these changes can be seen in the process schematic below in Figure 7
The low salinity of neutralized residue improves options for reuse in agriculture and soil beneficiation The nanofiltration permeate is also a valuable and improved feed to sea water reverse osmosis (SWRO) and multi-effect distillation (MED) plants. With most of the scaling species removed by the NF membrane, RO process efficiency is much improved, leading to a significant reduction in the RO unit capacity for equivalent potable water production, reducing capital and operating costs and significantly reducing waste retentate discharge from the process. Alumina refinery residues have different properties and reaction kinetics need to be specifically evaluated with on-site pilot tests which will determine actual reaction rates and equipment sizing. References
► Refinery
Le
Eend:
[
Refiner^— '
I Scope Additions I Scope Reductions
A. M. Hassan et al, "Conversion and Operation of the Commercial Umm Lujj SWRO Plant from a Single SWRO Desalination Process to the New Dual NFSWRO Desalination Process", IDA Conference, Manama Bahrain, March 2002.
[2]
A. M. Hassan et al, "A Demonstration Plant Based on the New NF-SWRO Process", Desalination 131, 2000, 157-171.
[3]
L. Fergusson, "The conversion and sustainable use of alumina refinery residues: Global solution examples", Light Metals, TMS, Orlando, 2007, 105-111.
[4]
G. Graham, and R. Fawkes, "Red mud disposal management at QAL". International Bauxite Tailings Workshop, Perth, Australia, 1992, 188-195.
[5]
D. McConchie, P. Saenger, and R. Fawkes, "An environmental assessment of the use of sea water to neutralise bauxite refinery wastes". The Minerals, Metals & Materials Society, 2nd International Symposium on Extraction and Processing for the Treatment and Minimisation of Wastes, Scottsdale, Arizona, October 1996,407-416.
[6]
C. Madden, "Mud, glorious mud?", Science Network Western Australia, 6 October 2006.
[7]
L. Fergusson, "Virotec: A ten-year story of success in environmental remediation". Prana World Publishing, ISBN 978-0-646-53324-7, 2010, 182.
[8]
D. McConchie, M. Clark, and F. Davies-McConchie, "New strategies for the management of bauxite refinery
■ | No Scope Change
Figure 7: Schematic Showing Sea Water NF and RO Process Impact on Bayer Residue Treatment The cost comparisons are dependent on the particular situation but order of magnitude costing estimates have shown that using a sea water NF concentrate to neutralize alumina refinery residue has significant economic potential in situations where: •
[1] Residue Storage Area
An existing refinery using sea water neutralization is looking to expand without spending capital on neutralization equipment and especially if there is limited land area available for décantation and drying ponds.
•
The refinery is not adjacent to the sea but still wants to neutralize its residue with a sea water based reagent.
•
The refinery wants to reduce salinity of neutralized residue to improve options for re-use, and
•
There is a local sea water desalination plant (e.g. SWRO, MED), where the combined benefits of nanofiltration to reduce scaling of desalination equipment and provide a sea water NF concentrate for refinery residue neutralization can be achieved Conclusions
Nanofiltration of sea water has been shown to be a viable method to significantly concentrate calcium and magnesium while sodium
86
resides (red mud)", 6th International Alumina Quality Workshop, Brisbane, Australia, September, 2003. [9]
D. McConchie, F. Davies-McConchie, M. Clark, and L. Fergusson, "The use of industrial mineral additives to enhance the performance of Bauxsol-based environmental remediation reagents", Australian Industrial Minerals Conference, Brisbane, 21 March, 2003.
[10]
D. McConchie, M. Clark, F. Davies-McConchie, L. Fergusson, T. Prowse, and S. Pope, "The use of industrial mineral additives to enhance the performance of Bauxsol-based environmental remediation reagents", Australian Institute of Geoscientists Bulletin, No. 38, 2003, 27.
[11]
M. W. Clark, J. Berry, and D. McConchie, "The longterm stability of a metal-laden Bauxsol reagent under different geochemical conditions", Geochemistry: Exploration, Environment, Analysis, 9, 2009, 101-112.
[12]
C. Hanahan, D. McConchie, J. Pohl, H. Creelhman, M. Clark, and C. Stocksiek, "Chemistry of sea water neutralisation of bauxite refinery residues (red mud)", Environmental Engineering Science, 21(2), 2004, 125138.
87
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
CAUSTIC AND ALUMINA RECOVERY FROM BAYER RESIDUES Gu Songqing Science & Technology Division of Chalco, No.62 Xizhimenbei Street, Beijing China Keywords: Bayer residue, Sintering process, Floatation-Bayer process, Lime-Bayer process, Hydro-process, DSP, Hydrogamet Abstract
Even though China is rich in bauxite resources, only about 20 % of bauxite reserves are of sufficient quality for Bayer processing as shown in Table 1, while bauxites of A/S 4-6 make up about 60% of total reserves.
The Bayer process is not suitable for alumina production from high silica bauxites due to the high caustic and alumina loss into the Bayer residue. In this study, caustic and alumina recovery from the Bayer residue by various processes, such as sintering, lime treatment and hydro-processing etc, is analyzed and compared theoretically. The most important target to treat Bayer residue for the recovery of these products is to find an efficient Desilication Product (DSP) containing less alumina and caustic soda by some suitable and efficient processes.
Most of Chinese bauxite is diasporic with both high alumina & silica content and has very complex silicate mineral composition. However the bauxite reserves per capita are only 10 % of the world average. Furthermore the bauxite mines are usually small in size, and there are lots of underground bauxite reserves in China, which brings about bauxite mining difficulties and higher mining cost. See Table 2.
Introduction
Based on the situation mentioned above, the Chinese alumina industry has to face serious raw materials issues and is forced to put great effort into R&D of new efficient processes to deal with lower grade bauxites.
The biggest challenge for the Chinese alumina industry is the bauxite supply shortage and the reduction of the grade of the available resource, including the ratio of alumina to silica (A/S) and the alumina content in these bauxites. The bauxite A/S supplied to some Chinese refineries has continuously been reducing from 12 to less than 6 in the past 8 years and even to 5.3 last year, which brings a major negative impact on alumina production cost. The bauxite grade change trend is shown in Fig. 1.
Too low a grade of bauxite means higher silica and reduced alumina content, which greatly impacts the cycling efficiency and productivity. The DSP produced in the Bayer process will carry more caustic and alumina to the residues increasing red mud disposal. Consequently, caustic consumption and alumina loss will increase and the operational cost grows quickly. Comparison of various processes to treat low grade bauxites There are many studies looking for new, high efficient processes for the low grade bauxite. The invention of the Rotation - Bayer process in the last decade, tries to pretreat low grade bauxite for silica minerals removal before introduction to the Bayer process. It was a success and applied in some Chinese refineries. But the alumina recovery is too low due to two stage losses in both the flotation and Bayer processes, and the organic substances entering the Bayer process might lead to some behavior changes of desilication and precipitation.
2001 2002 2003 2004 2005 2006 2007 2008 2009
Years Fig. 1 Bauxite grade reduction for some Chinese refineries Table 1 The Grade Distribution of Chinese Bauxites low grade
middle grade
A/S
3-4
4-6
6-7
7-9
9-10
>10 |
%
~8
-49
-11
-15
-12
-5 |
Bauxite grade
high grade
The Lime-Bayer process was developed to reduce caustic consumption by adding more lime to substitute some caustic from the residues. The Lime-Bayer process is simple and easy to apply.
Table 2 The Characteristics of Chinese Bauxite Reserves
Bauxite
1 World China
Reserve
Reserve Base
Reserve/ Population
A/S
23 B t
35 B t
4t
>10
0.54 B t
0.72 B t
0.41
4-7
* B t—billion tons
89
Mine Size
Bauxite Types
Large >0.1Bt Small <0.03Bt
Gibbsite Boehmite Diaspore
Silica Minerals Types Simple Kaolinite, Quartz Complex Kaolinite, Wite etc.
1 1 |
But too low an alumina recovery and high lime additions makes the production cost increase and the amount of residue increase, as well due to its lower desilication efficiency.
DSP analysis for the existing processes to treat low grade bauxite
The "Bayer-Sintering in Series" process is being tested in some Chinese refineries. The core concept for this process is to recover caustic and alumina from Bayer residue by sintering to form calcium silicates as the DSP. It seems that only a small part of the energy can be recovered and the energy consumption is still very high in the "Bayer-Sintering in Series" process compared with the original "Bayer-Sintering Combined" process.
To improve the desilication efficiency in caustic alumina production, the most important factor is to get suitable DSP with low A/S, N/S and C/S, that is, the less A1203 and Na20 content in the residues and the less lime addition, the more efficient for the process.
Fig 2 and Table 3 show the major DSP in the residues for existing processes.
This is the goal of looking for a targeted DSP. The DSP from the sintering process contains no AI2O3 and Na20 so that it is high efficient for silica removal and is energy intensive as well. The DSPfromthe Bayer process has the biggest N/S so it could not be used for high silica bauxites. The DSP from lime-Bayer process contains much A1203 and CaO, but no Na20 at all, so the process can only be used in the refineries with access to cheap lime and bauxite.
The high pressure hydro-process invented by former Russia scientists is a complete hydrometallurgical process without any pyro-process, and offers high alumina recovery and less caustic loss. The major concept is to recover caustic and alumina from Bayer residue by hydro-processing to form calcium sodium silicates as DSP. But this process is too complex and the evaporation is too large for production of a high concentration liquor and sodium aluminate precipitation, and again this process features high energy consumption.
Key technical concepts to develop high efficient alumina production
According to the process review above there is no highly efficient and energy saving process at present to deal with low grade bauxites. A comprehensive summary of the advantages and disadvantages of all the processes and theoretical investigation of the process reactions and products will provide the opportunity to develop new and efficient processes. Caustic Processes
For more economical recovery of caustic and alumina from Bayer residues, it is essential to find an efficient DSP and to apply energy saving processes. The alumina production efficiency can be improved by higher desilication efficiency.
Desilication Elements Ca and Na
DSP 2CaO· SiOv xH 2 0
Sintering
Na 2 0 A1203- 2Si02- 2H 2 0
Bayer Lime-Bayer
3CaOAl203Si02-4H20
High Pressure Hydro
Others
Fig 2 DSP analysis for the residue from different processes Table 3 The chemical compositions of the various DSP |
Processes
DSP in residues
A/S
N/S
c/s
|
Sintering
2CaOSi0 2 xH 2 0
0
0
1.87
|
Bayer
N a 2 0 AI2O3 -2Si02xH 2 0
0.85
0.52
0
|
Lime-Bayer
3CaO · AI2O3· nSiP 2 ·χΗ2Ο(η:0.2-1)
8-1.7
0
14-3
|
A/S—ratio of A1203 to Si0 2 content; N/S—ratio of Na20 to Si0 2 content; C/S-- ratio of CaO to Si0 2 content
90
The only way to greatly reduce energy consumption in the sintering process is by recovering as much energy from sintering as possible, and to enhance the process efficiency so that the total cost of sintering process can be competitive with the Bayer process.
The new highly efficient process to be found must be producing the DSP with relatively lower ratios of A/S, N/S and C/S so as to greatly reduce consumptions of bauxite, caustic soda and lime addition. Another criteria for any suitable process is that it should be as, energy efficient as possible, by adopting hydrometallurgical processing instead of pyrometallurgical processing to save energy.
Conclusions (1) The bauxite A/S for some Chinese refineries has continuously been reducing, which greatly impacts on the technical and economical performance and production cost. (2) The advantages and disadvantages of various existing processes to treat the low grade bauxite are analyzed according to their desilication efficiency and energy consumption. (3) A theoretical study has been carried out by comparison of the chemical compositions of the DSP from the different existing processes. (4) The key technical concepts to develop new and efficient processes for low grade bauxite are summarized. The most important is to produce a DSP with low A/S, N/S and C/S in a hydro-process. (5) It is possible to develop a process to form hydrogarnet structured compounds, which may indeed be the most efficient DSP.
Calcium compounds are much cheaper than other possible elements to replace Na20. But the multi elements compounds containing CaO are very complex and form at quite variety of reaction conditions. So the key solution to get the most efficient DSP will be based on the tests results at various conditions to find suitable compounds. There are numerous compounds produced from A1203, Si02, Fe203 and Ti02 (existing in bauxite) and Na20, CaO and MgO (added into the processes) etc, which can be found from X-ray diffraction tables and the phase diagrams. The most possible DSP in the processes to treat Bayer residues can be seen in Table 4. In this Table, the desilication efficiency increases from top to bottom. Table 4 Possible DSP in Bayer residue treatment
|
DSP in residues
A/S
N/S
c/s
Na 2 02Ca02SiO r xH 2 0 CaO - 2A1203 · 2Si0 2 H 2 0 3CaO · A1203· 2Si02 ·2Η20 3CaO Fe203· 2Si02 ·2Η20 3CaO · 2Si02 · mH20 CaOSiQ2xH20
0 1.7 0.85 0 0 0
0.52
0.94 1
o
0.47 1.4 1.4 | 1.4 0.94
0 0 0 0
| References [1] Gu Songqing, Wu Lichun, Liu Fengqin et al., Progresses of Nonferrous Metals in China (in Chinese), (Changsha, China, Central South University Publishing House. 2007) [2] Bi Shiwen, Yu Haiyan, Alumina Production Processes (in Chinese), (Beijing China, Chemical Industry Press, 2006) [3] Gu Songqing." Chinese bauxite and its influences on alumina production in China". Light Metals 2008, 79-83 [4] Claudia Brunori et al., "Reuse of a treated red mud bauxite waste", Journal of Hazardous Materials, B117, (2005), 55-63 [5] Ma Shuhua, et al., "Recovery of soda and alumina from red mud", Multipurpose Utilization of Mineral Resources (in Chinese), No.l (2008), 27-30
It is found from Table 4 that the DSP from CaO and Si0 2 is the most efficient because there are no A1203 and Na20 losses. Also, the higher the Si0 2 coefficient in the DSP, the higher the desilication efficiency. Laboratory tests show that the DSP from both CaO and Si0 2 usually form at very high temperatures, or in the pyrometallurgical processes. Hydrogarnet is a series of compounds containing CaO, Si0 2 and the oxides of other elements, such as A1203, MgO, Fe203 and Ti0 2 etc. Usually CaO in the hydrogarnet like structured DSP can be replaced by Na20, MgO and K20, while A1203 in the DSP can be replaced by Fe203 and Ti02. The minerals containing Fe203 and Ti0 2 are commonly found in bauxites and will take part in the reactions during the residue treatment so the study on the replacement of Fe 2 0 3 and Ti0 2 to A1203 should be carefully monitored to reduce A1203 content in the final hydrogarnet like structured products. It seems that the hydrometallurgical process is the first choice for getting an efficient DSP since the pyrometallurgical processes such as the sintering process will consume too much energy without much better waste heat recovery.
91
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
INVESTIGATION OF ALUMINA DISCHARGE INTO THE RED MUD POND AT NALCO'S ALUMINA REFINERY, DAMANJODI, ORISSA, INDIA B. K. Mohapatra1, B.K.Mishra2 and C.R.Mishra3 institute of Minerals and Materials Technology (CSIR), Bhubaneswar-751013, Orissa, India institute of Minerals & Materials Technology (CSIR), Bhubaneswar-751013, Orissa, India 3 National Aluminium Company Ltd., Bhubaneswar-751013, Orissa, India Keywords: Bauxite, Alumina Discharge, Red Mud electron microscopy and electron probe micro-analysis with a view to finding out the reason for alumina discharge into the red mud pond during the alumina refinery process.
Abstract Around 14% of alumina gets discharged into the red mud pond from NALCO's alumina refinery. The minerals hosting alumina in red mud have been investigated using Optical microscope, Scanning Electron Microscope (SEM), Electron Probe Micro Analyzer (EPMA) and XRD. The minerals in red mud sample, as identified through integrated techniques, are: gibbsite, boehmite, goethite, silimanite, muscovite, garnet and kaolinite besides minor hematite, sodalite and rutile. The average particle size of red mud is ~8μιη except isolated grains of gibbsite, goethite and garnet which are greater than 20μπι in size. The gibbsite invariably contains up to 60 mole % of boehmite. The nodular goethite contains over 30-mol % of ΆΓ in its lattice and termed as alumo-goethite. Around 34% of the A1203 is present in muscovite/silimanite/garnet. Studies reveal that most of the minerals in red mud have either alumina in their lattices or are undigested aluminium silicates that do not get dissolved during alumina refining, thus a considerable amount of alumina gets discharged through these phases into the red mud.
Materials and Methods For this study, representative red mud sludge was collected from the red mud pond of NALCO alumina refinery. After washing and oven drying, the sample was sprinkled on double sided adhesive conducting tape, placed over the sample holder and studied under SEM (JSM 35CF) equipped with WDS & EDS systems. Compositional x-ray analysis was undertaken for selected minerals. Chemical compositions of various phases were determined by an ARL-SEMQ-II electron microprobe in the Geochemisches Institute, Universität Gottingen, Germany, which is equipped with six spectrometer and four different analyzing crystals. The operating conditions were 15 kV accelerating voltage and 15 nA sample current. ZAF - corrected representative mineral analyses are presented in Table -1.
Introduction Morphology and Mineralogy
Around 2 tons of caustic insoluble waste residue known as 'Red Mud' is generated from NALCO aluminium refinery at Damanjodi, Koraput district, Orissa for every ton of alumina produced. Aluminium metal is extracted from aluminium oxide phases that constitute only 38 to 60% of the original bauxite ore. The rest is made up of Fe203, Si02, Ti02 and many other oxide phases. After dissolution of the bauxite in caustic soda, these impurities remain in suspended form which is separated out after being washed and then pumped in a slurry to the nearby pond called the Red Mud Pond. India accounts for about 2 Million tons of red mud per year of which is more than half the quantity is generated at the NALCO refinery. This industrial waste material poses tremendous environmental and disposal problems. Reduction in the quantity of red mud is possible only through their utilization in one form or another. However, the inherent complexity of red mud poses problem in its bulk utilization. Prior to going for bulk utilization of red mud rejects from the NALCO refinery, it needs in-depth characterization. Rao et al [1] have reported the characteristics of undigested sand rejects from the NALCO refinery. Only limited attempts have been made on the characterization of red mud sludge from this refinery [2-5].
In the red mud sample, about 35% by weight of solids contain less than 5-micron particles and the average size is found to be around 8 micron. However, a few independent grains > 20μπι size (about 10%) are also present. Due to the very fine size, some of the particles appear coagulated. Broad mineral species present in the red mud sample were identified through XRD (Philips). As can be seen from XRD pattern illustrated in Fig.l, the red mud sample contains gibbsite, boehmite and sodalite as alumina bearing; goethite and hematite as iron bearing and rutile as titanium bearing phases. Goethite, hematite, rutile, kaolinite and mica were recognized under optical microscope. Some of these alumina bearing minerals were subsequently analyzed under electron micro-probe. Gibbsite and goethite are observed to be dominating phases in the red mud sample. Gibbsites occur in pseudo hexagonal or prismatic habit. Goethite is either flaky or nodular; the later showing bunches of rotund units. Presence of fine grains of rutile, ilmenite, and zircon are recorded in subordinate amount. In addition, flakes of kaolinite, mica, needles of sillimanite (as litho relics) and sub-rounded grains of garnet are also noticed.
Fine particles of red mud can be characterized only through selected instrumental techniques. The present paper describes the characteristics of NALCO red mud through scanning
93
H 6
H
Table 2. Electron Probe Microanalysis results of Kaolinite and Goethite 2 4 1 Compounds, % 3 43.834 45.859 0.795 0.70 Si0 2
Gb
Cû
GD
Os"
J V
fel v** ^χ,ν
A rv Γ
A1203 FeO MgO MnO Ti02 CaO
UL τ τ τ
ΓΤΓΐτιτιτΓη^ττ ΐΓτΡτ ι ττ ΐΊ y
CuK α
<- 2Θ
Fig.l: XRD pattern of Red-mud. [ H: Hematite, Go: Goethite, S: Sodalite, Gb: Gibbsite, R: Rutile, B: Boehmite]
κ2ο
Na 2 0 Cr 2 0 3 BaO NiO CaO
Mineral Chemistry A few selected mineral grains viz. gibbsite, goethite and kaolinite, muscovite (precursor minerals not completely replaced by alumina) from red mud were exposed under electron probe microanalyser (EPMA) to know the extent of alumina in them (Table 1-3). Most of the idiomorphic crystals appearing as gibbsite under scanning electron microscope are found to have undergone dehydroxylation containing up to 60 mole % of boehmite in solid solution (Table 1).
H20Calc
SI Totali Al Fé
Mg Mn Co Total 2
The nodular type of goethite contains alumina in its lattice (Fig. 2). Such alumina-rich goethite is termed as alumo-goethite by Jonas and Solymar [6] Since goethite grains in the NALCO red mud are found to contain 30 mol % of alumina, it may be termed as alumo-goethite. 30 to 39% of A1203 are recorded in litho relict minerals like kaolinite and muscovite flakes (Table 2 & 3). Compositional maps of kaolinite, silimanite and garnet grains in red mud (Figs. 3-5) indicate that these minerals are not converted to gibbsite.
Na K Ca Ti | Total 3
A1 2 0 3 FeO Si0 2 MgO MnO CaO
κ2ο Na 2 0 Ti0 2
AlOOH Al(OH)3 FeOOH Other Oxides
2
1
3
4
78.30 0.26 0.04 0.00 0.00 0.49 0.04 0.00 0.00
71.82 0.00 0.00 0.00 0.00 0.22 0.00 0.00 0.00
69.94 0.00
65.60 0.00
0.36 0.07 0.00 0.02 0.00 0.00 0.00
0.10 0.00 0.00 0.00 0.00 0.04 0.01
69.44 29.63 0.32 0.93
49.90 49.87 0.00 0.23
35.65 73.69 0.00 0.66
2.35 97.31 0.00 0.34
-
—
38.009 0.049 0.058 0.018
0.054 0.025 0.203 0.076 0.000 0.009 0.236
0.134 0.033 0.064 0.127 0.017 0.000 0.211
--
~
--
--
Based on 08 cation 3.95 4.05 3.95 4.05
0.957 85.012
---
0.013
0.015 0.974
0.30 0.62
--
21.44 12.20 Based on 1 cation 0.01 0.01
0.02 4.02
0.01 3.93
0.03
--
0.01 0.01
-
---
0.03
0.02
~
--
0.01
~
6.12 0.42
~ ----
3.90
--
~
---
3.98 0.01 0.01
--
18.48 52.63
-
-~
0.01 0.06
~
Note: Column 1 & 2: Kaolinite; 3: Goethite; 4: Alumo-goethite
Table 1. Electron Probe Microanalysis Results of Selected Alumina-rich Grains from Red Mud Compound, %
39.228 0.145 0.076 0.000
1
Reasons for Discharge of Alumina Phases into Red Mud After the soluble aluminium containing oxides and hydroxides are recovered in the Bayer process, other metal oxides present in the bauxite are disposed as a red mud material. However, discharge of alumina rich phases into the red mud pond, resulting loss in alumina recovery, is of great concern. Discharge of alumina to an appreciable extent (Av. A1203: 14%) into the red mud pond may be attributed to gibbsite containing boehmite in solid solution, alumo-goethite having 30 mol% of alumina in its lattice, needles of silimanite, specks of mica and garnet (litho-relict minerals) and minute kaolinite platelets recorded in the red mud. All such mineral phases do not get dissolved during the industrial treatment of bauxite with caustic soda and ultimately get into the red mud pond. However, an in-depth study is necessary to improve the efficiency of digestion of all these phases during the Bayer process, so as to recover the alumina value from at least some of them. Volkov et al [7] reported that by increasing the amount of lime at 240° C, it is possible to completely break down aluminogoethite.
Note: Column 1: Boehmite dominating; 2: Mixed boehmitegibbsite; 3: Mixed gibbsite-boehmite; 4: Gibbsite dominating
94
Fig.4. Compositional Map of Si & Al in Silimanite Grain Present in Red Mud. Presence of Silica and Alumina indicates its poor leaching during Bauxitisation Process
Fig. 2. Compositional X-Ray Map of Fe & Al in Alumogoethite found in Red Mud. Image Map of Iron, Alumina confirms the Presence of significant Alumina in the Goethite Lattice
Table 3. Electron Probe Micro-analysis Results of Muscovite Si0 2 A1203 FeO MgO MnO CaO
κ2ο Na 2 0 Ti0 2 BaO
Si Al Totall Al Fe Mg Mn Ti Cr Total 2 Na K Total 3
Fig. 3. Compositional X-Ray Map of K, Al & Si in a Litho Relict Observed in Red Mud. The Silica Rich Phase Still Occupies the Core Region and hence escaped leaching
95
02 03 01 44.529 46.526 46.566 34.074 34.566 33.120 1.651 1.725 1.711 0.482 0.498 0.455 0.000 0.102 0.045 0.012 0.000 0.015 10.734 11.697 10.270 0.441 0.400 0.455 0.214 0.350 0.355 0.000 0.00 0.00 Based on 14 cations 6.22 1.78 8.00 3.60 0.19 0.10
-
0.02
-
3.91 0.10 1.99 2.09
6.28 1.72 8.00 3.78 0.19 0.09 0.01 0.03
6.20 1.80 8.00 3.63 0.20 0.10 0.01 0.04
4.10 0.12 1.78 1.90
3.98 0.12 1.90 2.02
~
~
04 44.407 30.493 1.177 0.375 0.00 0.00 10.304 0.426 0.248 0.076 6.47 1.53 8.00 3.72 0.14 0.08
»
0.03
-
3.97 0.12 1.91 2.03
The mineral chemistry of some of the constituent phases in red mud reveals the presence of lattice bound alumina which escapes the digestion stage in the Bayer's process, resulting in loss of alumina during the refining process at NALCO.
References 1.
M.B.S. Rao, B. K. Mohapatra, B. Das and A. K. Paul, "Characteristics of sand rejects- A case study from an Alumina Refinery Plant," Vistas in Geological Research, UU Spi. Publication in Geology, 2 (1997), 164-172.
2.
. S. Thakur and S. N. Das, "Red Mud Analysis and Utilization," Pubi. & Inform. Directorate & Wiley Eastern Ltd., New Delhi, (1994), 291p.
3.
J. Pradhan, S. N. Das, J. Das, S.B. Rao and R. S. Thakur, "characterization of Indian red mud and recovery of their metal values," Light Metals: Proceedings of Sessions, TMS Annual Meeting (Warrendale, Pennsylvania), Minerals, Metals & Materials Soc, USA., (1996), 87-92.
4.
M. G. Sujana, R. S. Thakur, B. C. Acharya, S. N. Das and S. B. Rao, "Effect of calcination and physico-chemical properties," Light Metals: Proceedings of Sessions, TMS Annual Meeting (Warrendale, Pennsylvania), Minerals, Metals & Materials Soc (TMS), USA, (1986), 93-98.
5.
B.K.Mohapatra, M.B.S.Rao, R.Bhima Rao and A.K.Paul. Characteristics of red mud generated at NALCO refinery, Damanjodi, India. Light Metal.(Warrensdale, Pa), (2000) 161-165
6.
K. Jonas and K. Solymar, "Preparation, X-ray, derivitographic and infrared study of aluminium substituted goethites," Ada Chim. Acad. Sci. Hung., Tomus. v.66 (4), (1970), 383-383.
7.
V. V. Volkov, M.I. Soboi, N. I. Eremin, "Lime addition effect upon Alumina extraction from Alumino-goethite," kv. V.u.z. Tsvetn. Metal, 5 (1984), 22-24.
Fig. 5. Compositional Map of Si & Al in Relict Garnet Present in the in Red Mud Conclusions The red mud sludge from the NALCO alumina refinery at Damanjodi, Orissa, has been characterized using SEM-WDS and EPMA techniques. These are very fine grained muds, the average size being 8 micron. Different alumina containing phases such as alumo-goethite, sillimanite, kaolinite, mica and garnet were identified from their micro-morphology and in-situ chemical analysis, which contribute to the loss of alumina into the red mud. The presence of dehydroxylated grains of gibbsite probably indicates the higher temperature (than that for pure gibbsite) required for their dissolution.
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
PRODUCTION OF ORDINARY PORTLAND CEMENT (OPC) FROM NALCO RED MUD Chitta Ranjan Mishra1, Devendra Yadav2, P.S.Sharma2& M.M.Alli2 former Deputy General Manager (R&D), National Aluminium Company Limited (NALCO), NALCO Bhawan, P-1, Nayapalli, Bhubaneswar-751013, Orissa, India Scientist, National Council for Cement & Building Materials (NCB), Ballavgarh, 34-KM Stone, Delhi-Mathura Road, PIN-121004, Haryana, India Keywords: Red Mud, Limestone, Shale, Coal Ash, Clinker, Gypsum, OPC Abstract A process for production of Ordinary Portland Cement (OPC) from NALCO Red Mud has been successfully developed from a raw mix containing limestone, red mud, shale and fine coal. The raw materials are ground to the required fineness and then blended to prepare the raw mix. The raw mix is fed in to a kiln and fired to a temperature of 1400-1450 degree centigrade to obtain clinker. Clinker was cooled and gypsum was added in to it to obtain OPC. 3-3.5% of NALCO red mud was used for production of OPC.OPC prepared from this clinker conformed to the requirements of three Indian Standard Specifications for 33, 43 & 53 grade of OPC. The process is efficient, cost-economic and effectively addresses the environmental problems associated with the waste red mud generated during the refining of bauxite for alumina manufacture.
Sample Collection & Sampling Five shift samples of red mud, bauxite and fly ash were collected. Red mud samples were designated as RMD-1 to RMD-5. Red Mud: Five shift samples of red mud weighing about 25kgs each were dried in electric oven at temperature of 105±5°C. The dried lumps in the samples were crushed. The samples were reduced to ~3kg by coning and quartering procedure and ground to pass through 100 mesh sieve. The laboratory samples were prepared by drawing ~200gms from each shift samples by coning and quartering procedure and were subjected to chemical and mineralogical characterization. The results of chemical and mineralogical characteristics of these laboratory samples are discussed in the following section.
Introduction National Aluminium Company Limited (NALCO), a Navaratna Company and a Govt. of India Enterprise under Ministry of Mines has set up Asia's largest integrated alumina and aluminium complex in the state of Orissa, utilizing state of the art technology from Aluminium Pechiney, France and is producing 4,8,00,000 tpa of bauxite, 1,5,75,000 tpa of alumina, 3, 45,000 tpa of aluminium metal and 960MW of captive power. During the refining of bauxite for production of alumina employing the Bayer process at NALCO's alumina refinery, at Damanjodi in the district of Koraput, Orissa, on a dry basis ,about 2 million tones /annum of red mud is generated as a by-product. The red mud contains more than 55% water and is disposed off in the form of slurry in to the nearby red mud pond. The accumulated quantity of red mud in the ponds is estimated to be around 18.5 million tones. The water overflowing the pond after the mud is settled, is recycled to the alumina plant for partial recovery of soda. Therefore, the soda content in red mud is minimized. It was thought prudent to explore the possibilities for utilization of this uncausticised red mud dried in air/sunlight for production of value added items. Development of Ordinary Portland Cement (OPC) was one of the most positive efforts undertaken in this direction [1]. NALCO in collaboration with National Council for Cement & Building Materials (NCB), Ballavgarh, and Haryana, India has developed this process.
Limestone and Other Raw Materials: Limestone and other raw materials viz. shale, coal and gypsum samples were ground to pass through 100 mesh sieve. The laboratory samples of these raw materials were prepared by drawing ~200gms from each of the ground by coning and quartering procedure. The results of chemical and mineralogical characterization of these samples are discussed in the following section. Characterization and Evaluation of Samples Red Mud Samples: (a) Quality and quality variation: Five ground samples of red mud (RMD-1 to RMD-5) were subjected to chemical analysis for their major constituents' viz. LOI, Si02, Fe203, AI2O3 and alkalies for assessing their quality and the variation therein. The results are presented in Table-2 indicated the variation in the composition of red mud shift samples with respect to major oxides to be low. In view of this, all the five shift samples of red mud were mixed together to make a composite sample. The composite sample was designated as RMD-C. Further investigations were carried out on the composite sample of red mud, RMD-C only. Table 2. Quality and Quality Variation in Shift Samples of Red Mud.
Table 1. Chemical Composition of NALCO Red Mud. Constituents % By Weight 16.13 A1203 53.92 Fe 2 0 3 Ti0 2 04.82 06.29 Si0 2 Na 2 0 03.45 CaO 01.88 P205 00.11 00.11 v2o5 LOI 11.78
97
SI No
Constituent Determined
1 2 3 4 5 6
LOI Si0 2 FeA A1203 Na20
(%)
κ2ο
Shift I RMD1 11.32 6.32 50.10 20.51 4.07 0.07
Shift II RMD2 11.86 6.35 49.70 20.70 4.11 0.08
Shift III RMD3 11.03 6.48 50.15 21.07 3.40 0.06
Shift IV RMD4 11.13 6.07 50.52 20.14 3.50 0.06
Shift V RMD5 11.64 5.90 49.57 20.55 3.78 0.06
mineral and muscovite, yavapaiite and anorthite as minor minerals.
Red Mud Composite Sample (RMD-C):
Gypsum:
(a) Chemical Characterization: The composite red mud sample, RMD-C was subjected to chemical analysis and the results are presented in Table-3.
(a) Chemical Characterization: The gypsum sample was subjected to chemical analysis and the results are presented in Table-4. The sample is considered suitable as a set retarder in the manufacture of cement. (b) Mineralogical Characterization: X-ray diffraction of the gypsum sample indicated the presence of gypsum as a major mineral and α-quartz and albite as minor minerals.
Table 3. Chemical Analysis of Composite Red Mud Sample (RMD-C) SI No Constituents Results Determined (%) 1 11.45 LOI 2 6.32 Si0 2 49.62 3 Fe 2 0 3 4 20.74 A1203 5 CaO 2.45 6 0.15 MgO 7 0.08 so3 8 Na 2 0 3.66 9 0.07 κ2ο 10 0.33 p2o5 11 4.76 Ti0 2 12 Mn 2 0 3 0.11
Coal: (a) Chemical Characterization: The results of proximate analysis of the coal sample are presented in Table-5. The ash prepared from the above coal was subjected to chemical analysis and the results are presented in Table-4 Table 5. Proximate Analysis of Coal Sample. Characteristics Sl.No. Results 35.34 1 Ash Content (%) 2 Moisture Content (%) 1.82 Volatile Matter (%) 30.04 3 4 Fixed Carbon (%) 32.80 5 Calorific Value(kcal/kg) 5140
(b) Mineralogical Characterization: X-ray diffraction analysis of the composite red mud sample, RMD-C indicated the presence of gibbsite and hematite as major minerals and goethite as minor mineral.
The above results of chemico-minerological characteristics of red mud, limestone and other raw materials indicate that their matrix is compatible with the cement matrix and can be used in the manufacture of cement.
Other Raw Materials: Limestone: (a) Chemical Characterization: The limestone sample was subjected to chemical analysis and the results are presented in Table-4. The sample can be classified as cement grade limestone.
Sl.No. 1 2 3 4 5 6 7 8 9 10 11
Table 4.Chemical Analysis of Raw Materials Constituents Limestone Shale Gypsum Determined
(%)
LOI Si0 2 Fe 2 0 3 A1203 CaO MgO
so3 Na20
κ2ο Ti0 2 cr
37.25 10.78 0.81 1.95 44.90 3.06 0.00 0.13 0.73
-
0.008
2.72 85.93 1.14 7.22 0.58 0.21 0.13 0.18 1.76 0.27 0.01
17.10CW 11.44(+IR) 0.26 NIL 28.85 0.66 39.90 0.24 0.04
-
Technical Suitability of NALCO Red Mud for the Manufacture of Cement Chemical and mineralogical analysis of the NALCO red mud sample indicated that its composition is quite compatible with the cement matrix. Therefore, the NALCO red mud sample was considered prima-facie suitable for its gainful utilization in the manufacture of Ordinary Portland Cement (OPC)clinker in combination of other raw materials viz. limestone, shale and fine coal [2 -5]. The process of making OPC is discussed below.
Coal Ash 0.87 58.64 6.25 25.85 3.19 0.67 1.72 0.08 0.83 1.56 0.017
Raw Mix Design & Its Optimization: The Approach: While designing the raw mixes, attempts were made to maximize the use of red mud, RMD-C as raw mix component in the raw mix design for the manufacture of OPC. Keeping this in mind a large number of different raw mixes were designed to achieve good quality clinker. The proportions of raw material in the raw mixes designed (RM-1 to RM-5) are shown in Table-6. The design parameters optimized in the designed raw mixes are shown in Table-7-11.
(b) Mineralogical Characterization: X-ray diffraction analysis of the limestone sample indicated the presence of calcite as a major mineral and α-quartz, dolomite and muscovite as minor minerals.
Raw Mix No. RM-1 RM-2 RM-3 RM-4 RM-5
Shale: (a) Chemical Characterization: The shale sample was subjected to chemical analysis and the results are presented in Table-4. (b) Mineralogical Characterization: X-ray diffraction analysis of the shale sample indicated the presence of α-quartz as major
98
Table-6. Various Raw Mix Designs Prepared Red Mud Shale Limestone Coal Ash Absorption(%) (%) (%) (%) 3.00 94.70 2.30 4.75 2.14 3.50 94.36 4.75 94.02 4.00 1.98 4.75 4.30 93.81 1.89 4.75 93.34 5.00 1.66 4.75
Raw Materials Proportion
Table-7. Design Parameters of Raw Mix RM-1 Coal Ash 1 Red Mud Limestone Shale Absorption 4.75 3.00 94.70 2.30
(%)
Composition
(%)
Raw Mix
S 35.68
Clinker
MgO: 2.89 Na 2 0: 0.27 K 2 0: 0.49
4.39
5.37
62.69
MgO: 4.30 Na 2 O:0.40 K 2 0: 0.76
SM
AM
1.08
2.17
1.08
Liquid Content(%)
0.93 C3S
2.12
1.22 C3A
C4AF
55.94
17.04
6.80
13.35
2.63
42.61
MgO: 2.91 Na 2 O:0.24 K 2 O:0.73
Raw Mix
21.11
3.68
5.13
63.25
MgO: 4.34 Na 2 O:0.35 K 2 0: 1.13
Clinker
Liquid Content(%)
Modulii Values RawMix
LSF
Clinker 1 Potential Phase Compo1 sition(%)
2.52
1.15
Clinker 1 Potential Phase Compo1 sition(%)
0.93
2.40 C2S
1.40 C3A
C4AF
57.32
17.32
7.37
11.19
c3s
(%)
29.25
35.53
12.09
20.65
PU
£5
Pu
12.23
2.53
2.72
42.47
MgO: 2.90 Na 2 O:0.25 K 2 O:0.73
Raw Mix
20.08
4.03
5.25
62.97
MgO: 4.32 Na 2 0: 0.38 K 2 0: 0.76
Clinker
Liquid Content(%)
Modulii Values RawMix
LSF
Clinker Potential Phase Composition^)
S
9
κ2ο
Modulii Values RawMix
LSF
SM
AM
1.08
2.43
1.08
Clinker Potential Phase Composition^)
0.93 C3S
2.25
1.30 C3A
C4AF
56.64
17.17
7.08
12.27
c2s
c2s
(%)
Composition
MgO Na20
d <
d
d
d
κ2ο
31.17
Table-10. Design Parameters of Raw Mix RM-4 Raw Red Mud Limestone Shale Coal Ash Absorption Materials 4.75 4.30 93.81 1.89 Proportion
Table-8. Design Parameters of Raw Mix RM-2 Red Mud Limestone Shale Coal Ash Absorption 3.50 2.14 94.36 4.75
(%)
Clinker
42.32
2.28
κ2ο
1.08
35.61
2.80
12.37
(S
AM
Raw Mix
2.77
PU
SM
(%)
9
S
d
MgO Na20
d <
9
LSF
Composition
(%)
Composition
MgO Na20
1 Modulii Values 1 RawMix
Raw Materials Proportion
Table-9. Design Parameters of Raw Mix RM-3 Limestone Shale Coal Ash Red Mud Absorption 94.02 1.98 4.75 4.00
d <
O
d
Raw Materials Proportion
(%)
30.03
99
en
MgO Na20 K20
q
9u
2.92
2.85
42.24
MgO: 2.88 Na 2 O:0.28 K2O:0.72
4.62
5.44
62.51
MgO: 4.29 Na 2 O:0.42 K 2 0: 0.75
SM
AM
1.08
2.11
0.98
Liquid Content(%)
0.93 C3S
2.04 C2S
1.18 C3A
C4AF
55.44
17.03
6.63
14.00
S
O
35.49
12.01
20.52
PU
31.86
Raw Materials Proportion
Red Mud
Limestone
Shale
5.00
93.34
1.66
(%)
Composition
(%)
Raw Mix
up to 5% in the raw mixes designed, the raw mix RM-2 was selected to avoid process problems attributed to the presence of a high proportion of fluxing agents (A1203 and Fe203) in such raw mixes. From Table-7 to 11, it is found that on increasing the red mud content from 3.0% to 5.0%, the liquid content was found to be increased continuously. The presence of increased liquid content known to cause process implications in the cement manufacture and especially beyond 30% liquid content is detrimental to the life of refractory lining of the kiln. Generally, the liquid content has been maintained in the range of 27 to 30% in cement manufacture. 3.5% has been selected as the optimal red mud content keeping in mind the maximum utilization of red mud in cement manufacture. Its liquid content is 30.03% and is on border line but is still in an acceptable range. On going from red mud content of 3.5 to 5%, the liquid content increases from 30.03% to 33.45% and the resultant clinkers were found to be sticky to the refractory linings and will reduce the life of the refractory linings and have many process implications. Also, the presence of high liquid content results in the formation of boulders instead of normal sized clinkers which further makes problems in the process. The role of fluxing agents like Fe203, A1203 etc is very critical in the formation of clinker. The general range for A1203 content is 4 to 7% and for Fe203 is 2 to 4% in the clinker. The details of the optimized raw mix RM-2 is given in Table-8. The raw mix RM-2 is capable of yielding good quality clinker at 1450 °C with a retention time of 20 minutes.
Table-11. Design Parameters of Raw Mix RM-5
d
S
cJ5
35.39
Clinker
d Pu
d <
9
Coal Ash Absoφtion 4.75 MgO Na20 K20
11.80
3.26
2.97
Aim
MgO: 2.87 Na 2 O:0.31 K2O:0.48
20.19
5.10
5.61
62.13
MgO: 4.26 Na 2 O:0.46 K 2 0 : 0.75 Liquid Content(%)
Modulii Values RawMix
LSF
SM
AM
1.09
1.90
0.91
Clinker 1 Potential Phase Compo| sition(%)
0.93 C3S
1.89 C2S
1.10 C3A
C4AF
54.53
16.78
6.24
15.51
33.45
Table 12.Burnability Studies of Cement Raw Mixes with Red Mud as a Component (Retention Time 20 Minutes)
Raw Mix Preparation: Raw mixes, RM-1 to RM-5 were prepared by taking weighed quantities of raw materials as per the designs, blending them thoroughly and grinding the mixes to fineness of 10% residue on 90μ(170 mesh) sieve. Nodules of about 1 cm in diameter were prepared by mixing about 12% water and were dried in an electric oven at 105±5° C for about 2 hrs before subjecting them to burnability studies. Burnabilities Studies: Burnability studies were carried out on all the raw mixes. The dry nodules were introduced in to a laboratory furnace at ambient temperature, which was gradually raised to 1450°C. The raw mixes were fired at 1300, 1350, 1400 and 1450°C with a retention time of 20 minutes. The clinkers, CL-1 to CL-5, prepared from the raw mixes, RM-1 to RM-5 respectively were room cooled and their free lime content determined. The results of free lime determination are presented in Table-12 which indicate that all the raw mixes have good burning characteristics and are capable of yielding quality clinkers even at 1450°C with a retention time of 20 minutes. The free lime content in all the clinker samples was found to be < 0.50% at 1450°C.
SI. No. 1
Raw Mix RM-1
2
RM-2
3
RM-3
4
RM-4
5
RM-5
Temperature(°C) 1300 1350 1400 1450 1300 1350 1400 1450 1300 1350 1400 1450 1300 1350 1400 1450 1300 1350 1400 1450
Free Lime % 1.66 0.82 0.32 0.06 1.73 0.88 0.34 0.06 2.44 1.57 0.55 ! 0.13 2.35 1.18 0.44 0.09 1.73 0.61 0.31 0.15
Preparation and Evaluation of Bulk Clinker
Optimization of Raw Mix Design:
10 kg sample of optimized raw mix RM-2, selected for detailed investigations was prepared by taking weighed quantities of raw materials viz. red mud, limestone, shale and coal ash as per its raw mix design, blended them in a ball mill and ground to a fineness of 10% residue on 90μ(170 mesh) sieve. The nodules were prepared in a pan nodulizer and dried in an electric oven at 105±5 °C for 2 hrs before introducing in an electric furnace at ambient temperature and eventually firing it at 1450 °C for 20 minutes. The resultant clinker CL-2 from raw mix RM-2 was
While designing the raw mixes, every effort was made to keep the level of utilization of red mud to the maximum extent possible. All the raw mixes prepared viz. RM-1 to RM-5 yielded good quality clinkers when test fired during burnability studies at 1450 °C with retention time of 20 minutes. But, keeping in view the maximum utilization of red mud, design parameters and process implications there from, raw mix RM-2 was taken as the optimized raw mix design. Although the red mud could be utilized
100
13KWH/T. As this value of Bond Index lies on the middle of the range, the clinker appears to be hard enough in nature.
studied for chemical, mineralogical and grinding characteristics and the OPC prepared there from was evaluated for performance as per relevant Indian Standard Specifications. Three batches of such bulk clinkers, lOkgs each, were prepared by adopting the same procedure to evaluate the reproducibility of the results.
Reproducibility of the Clinker In order to ascertain the reproducibility of the product i.e. clinker, two more batches of the clinker were prepared in the laboratory from the optimized raw mix, keeping the process parameters same as maintained during preparation of bulk clinker sample, CL-2.
Chemical Analysis: The chemical analysis of the bulk clinker, CL-2, carried out as per Indian Standard Specification, IS: 4032-1985 and is presented in Table-13. These results indicated that the quality of the clinker was good and capable of yielding good quality cement.
Preparation and Evaluation of Two Additional Batches of the Clinker: Two additional batches of the clinkers CL-2-A and CL-2-B were prepared by taking 10 kgs of raw mix RM-2 adopting the same procedure as described under Preparation and Evaluation of Bulk Clinker. The resultant clinkers designated as CL-2-A and CL-2-B were studies for chemical and mineralogical characteristics. The results are discussed below.
Table 13. Chemical Analysis of Bulk Clinker
SI. No. 1 2 3 4 5 6 7 8 9 10 11 12
Results Obtained
Constituents Determined LOI Si0 2 Fe 2 0 3 A1203 CaO MgO
(%)
0.35 20.60 4.02 5.30 63.27 4.22 0.27 0.38 0.78 0.003 0.38 0.25
so3 Na20 κ2ο cr
Insoluble Residue Free Lime
Chemical Analysis: The chemical analysis of the two additional batches of clinkers, CL-2-A and CL-2-B, were carried out as per Indian Standard Specification. IS: 4032-1985 and the results are presented in Table-15. Results indicated that the chemical composition of both the clinkers, CL-2-A and CL-2-B, are comparable to the composition of thefirstbatch of the clinker, CL-2. Table 15. Compositional Variation in Three Batches of Clinkers
SI. No.
Mineralogical Analysis: The clinker sample CL-2 from raw mix RM-2 was evaluated for its mineralogical composition by optical microscopy and the results along with the granulometry of the clinker phases are presented in Table-14. Clinker phases are moderately developed and are homogenously distributed. Majority of alite grains are subhedral in shape. Few pseudo hexagonal alite grains are also developed in the clinker. Transformation of belite into alite has also been observed. Some belite clusters were also presented in the sample. Few crystals of free lime were also present. It indicated that the quality of the bulk clinker sample so prepared was good. Table 14. Mineral Phase Analysis of Bulk Clinker Sample by Optical Microscopy SI. No.
Clinker Sample
Phases
Quantity
1
CL-2
C3S C2S C3Al C 4 AFj CaOf
48 33 18
(%)
1
1 2 3 4 5 6 7 8 9 10 11 12
Granulometry (μπι) Min. Avg. Max. Size 12 3 18 3 17 10
-
-
-
1
9
4
Constituents Determined LOI Si0 2 Fe 2 0 3 A1203 CaO MgO
so3 Na20 K20 cr
Insoluble Residue Free Lime
Results Obtained (%) CL-2-A CL-2 CL-2-B 0.35 0.45 0.40 20.60 20.48 20.43 4.02 4.12 4.18 5.30 5.29 5.26 63.24 63.27 63.35 4.22 4.23 4.26 0.34 0.27 0.31 0.37 0.38 0.31 0.78 0.69 0.77 0.004 0.003 0.006 0.34 0.38 0.36 0.25
0.22
0.24
Mineralogical Analysis: (a) Optical Microscopic Analysis: Samples from the two additional batches of clinkers, CL-2-A and CL-2-B, prepared in the laboratory ascertaining the reproducibility were evaluated for their mineralogical composition by optical microscopy. The results along with the granulometry of clinker phases are presented in Table-16. The alite grains were subhedral in shape. Transformation of belite into alite was observed in both the samples. Most of the belite grains were sub-rounded in shape with corroded margins. Few crystals of free lime were also observed in these two samples. The results have revealed that the granulometry and mineralogical characteristics of the two clinkers of additional batches, CL-2-A and CL-2-B, are comparable to the first batch the clinker i.e. CL-2.
The X-ray diffractogram of the bulk clinker sample CL-2 indicated the presence of homogenously distributed and well developed phases. Grindabilitv Studies: In order to determine the energy required for grinding clinker, laboratory grindability test was conducted using standard bond ball mill. The Bond Index for clinker, CL-2, has been found to be 10.9KWH/T which lies in between the range of Bond Index 9-
101
Table 16. Variation in Minerai Phase Analysis of Three Batches ___^ of Clinker Samples by Optical Microscopy
SI. No.
Clinker Sample
Phases
Quantity
1
CL-2
C3S C2S C3Al C4AFJ CaOf C3S C2S
48 33 18
2
3
CL-2-A
CL-2-B
Granulometry^m) Min. Avg. Max. Size 12 18 3 17 10 3
-
-
-
C3A!
1 45 34 17
1 3 4
9 19 17
4 11 11
C3A!
1 46 34 18
1 3 4
9 18 18
4 11 12
1
1
9
4
C4AFJ CaOf C3S C2S C4AEJ CaOf
Table 17.Performance of Ordinary Portland Cement (OPC)
SI. No.
-
-
-
-
1 2 3
-
4
-
Property
Fineness(M2/kg) Setting time (Min.) Initial Final Compressive Strength(N/mm2) 3 Days 7 Days 28 Days Soundness Le-chatelier(mm) Autoclave (%)
Results 305
Requirement of IS: 12269-1987 (53 Grade OPC) Not less than 225
128 254
Not less than 30 Not more than 600
29.0 42.5 60.8
Not less than 27 Not less than 37 Not less than 53
1 0.073
Not more than 10 Not more than 0.8
The above results indicated that good quality clinker can be prepared by using 3.50% red mud, 94.36% lime stone and 2.14% shale. The Ordinary Portland Cement (OPC) prepared from this clinker conformed to all the requirements of three Indian Standard Specifications for 33, 43 and 53 grade of OPC. In addition, OPC made from 3.0% of NALCO red mud, 94.70% of lime stone and 2.30% of shale also confirmed to the above specifications. Complete performance evaluation of different OPC samples prepared from 3.0 and 3.5% of NALCO red mud was conducted.
(b)X- Ray Diffraction Analysis: Samples from the two additional batches of clinkers CL-2-A and CL-2-B, were also subjected to X- Ray diffraction analysis. It indicated the presence of homogenously distributed and well developed clinker phases in these two samples, CL-2-A and CL-2-B, similar to the first batch of the sample, CL-2. The chemico-minerological evaluation of the results of the two additional batches of clinker samples, CL-2-A and CL-2-B, indicated their close resemblance to the clinker, CL-2, prepared in the first batch. Results have clearly established the reproducibility of the product i.e. clinker made from NALCO red mud.
Conclusions Chemical and mineralogical analysis of NALCO red mud indicated that its composition is quite compatible with the cement matrix. The NALCO red mud therefore was considered primafacie suitable as a raw mix component in the manufacture of OPC. While designing the raw mixes, every effort was made to keep the level of utilization of NALCO red mud to the maximum possible extent. Upto 5% red mud could be gainfully utilized in the raw mixes. However, to overcome process problems attributed to the presence of high proportion of fluxing agents (A1203 and Fe203) in other raw mixes, the raw mix RM-2 with 3.5% red mud was considered most suitable for the preparation of bulk clinker, CL-2. Physical performance of the OPC made from the bulk clinker, CL-2, utilizing 3.5% of NALCO red mud conformed to all the requirements of all the three Indian Standard Specifications for 33, 43 and 53 grades of OPC viz. IS.269-1989, IS:8112-1989 and IS: 12269-1987 respectively. Reproducibility of the clinker made from NALCO red mud is also established.
Preparation and Evaluation of Ordinary Portland Cement (OPC) Ordinary Portland Cement ,OPC-2 was prepared by grinding the bulk clinker CL-2 so prepared with 5.0% gypsum to a fineness of ~300m2/kg and tested for setting time, compressive strength, Lechatelier and autoclave expansion tests as per IS :4031-1988 and the results are presented in Table-17. Setting Time: The initial and final setting times of Ordinary Portland Cement prepared in laboratory from bulk clinker were determined as per IS: 4031-1988 and the results are presented in Table-17. The results indicated that the cement sample OPC-2 conformed to the requirements of the standard.
References 1. H.Z. Xu, Technological and Economic Feasibility Study on Producing Building Materials with Red Mud, Gold 17(1996), pp. 17-21(in Chinese). 2. Maneesh Singh, S. N. Upadhyaya and P. M. Prakash, Preparation of Special Cements from Red Mud, Waste Management, Vol. 16, Issue 8, (1996), pp.665-670. 3. M Singh, et al., Preparation of Iron Rich Cements using Red Mud, Cement Concrete Research, 27(7)(1997), pp.1037-1046. 4. P.E.Tsakiridis, S.Agatzini-Leonardou, P.Oustadakis, Red Mud Addition in the Raw Meal for the Production of Portland Cement Clinker, Journal of Hazardous Materials, B 116(2004), pp.103-110. 5. C.Brunori, C.Cremisini, P.Massanisso, V.Pinto, L.Torricelli, Resuse of a Treated Red Mud Bauxite Waste, Studies on Environmental Compatibility, Journal of Hazardous Materials, 117(l),(2005),pp.55-63.
Compressive Strength: The compressive strength of OPC-2 prepared above was determined as per IS: 4031-1988 and the results are presented in Table-17. The results indicated that the OPC-2 sample conformed to all requirements of all the three Indian Standard Specifications for 33, 43 and 53 grades of OPC viz. IS:269-1989, IS:8112-1989 and IS: 12269-1987. Soundness: Autoclave and Le - chatelier expansion tests on OPC-2 prepared above were carried out as per the procedures laid down in IS: 4031-1988. The results indicated high volume stability of the cement sample and conformed to all the requirements of all the three Indian Standard Specifications for 33, 43 and 53 grades of Ordinary Portland Cements viz. IS:269-1989, IS:8112-1989 and IS: 12269-1987 respectively.
102
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
RECOVERY OF METAL VALUES FROM RED MUD P.K.N.Raghavan1, N K. Kshatnya1 and Katarzyna Wawrynink2 Îharat Aluminium Company Limited,Korba (CG) India 2 Warsaw University of Technology,Poland Keywords: red mud, roasting, alumina, iron, titania values. However, there is a very strong argument against this in the countries with large availability of Iron ores, Titanium ores and Aluminum ores. However, the time is not very far off when depletion of mineral resources will make it viable to convert these wastes into useful products. The growing awareness of environmental pollution and concern for protecting ecological balance has created alternative ways of its utilization converting the waste into wealth. The successful use of red mud as a raw material source for metal production depends on the recovery of each metal value present in the particular red mud.
Abstract In processing bauxite for production of alumina by the conventional Bayer process, huge quantities of so called red mud are generated which are disposed of as a waste product. The red mud is normally contains caustic soda and therefore poses pollution hazardous and requires storage in specially made large size ponds. This waste material at present does not find any use, hence, in addition to the pollution hazard, considerable expenditure and wastage of land is involved in disposal of this waste material. Therefore, in view of the increasing problem of waste disposal, it is desirable to recover the oxides of iron, titanium and aluminum from red mud.
Besides the undissloved alumina, red mud contains the alkali insoluble oxides like Fe 2 0 3 and Ti02.The alkali aluminosilicate is formed due to the reaction between combined silica of bauxite with caustic soda. Various formulae have been proposed to represent this desilication compound. Bayer gave the formula for this desilication product as Al203.Na20.3Si02.9H2O.Though the composition of these silicates has not been definitely ascertained, it is certain that it depends upon the content of the reactive silica and digestion conditions adopted.
The process applies to the recovery of alumina, titania and ferric oxide from red mud which comprises the following steps (i) Roasting of red mud in presence of alkali (ii) Extracting the salts produced in step (i), (iii) Precipitation of hydroxide salt and (iv) Converting the hydroxide salt into oxides. Balco's red mud contains 45-47% Fe 2 0 3 , 1820% Ti0 2 and 16-25% of undigested alumina .Many methods have been tried by others to recover the metal values and it is proposed in the laboratory on bench scale. The present papers provide an alternative and more environmentally acceptable process for the recovery of iron, titanium and aluminum from Balco's red mud.
It has been indicated earlier that the alumina is present as an alkali aluminosilicate with minor amounts of undissloved (diasporic) aluminium oxyhydroxide. Experimental The possible method for treating red mud for the recovery of metal values comprises of the following steps: (i) Roasting the red mud with soda ash and lime (ii) Extracting the fused mass (iii) Digestion and precipitation of hydroxide salt (iv) Conversion into appropriate oxide The red mud sample of Bharat Aluminum Company Limited, Korba (India) was in the form of cakes having pale reddish color. This was easily friable and could be crushed with ease. The major constituents of red mud are analyzed by XRF (PW-2440 Philips, Netherland). The sample analyzed as follows as shown in Table 1 :
Introduction To produce every ton of the aluminum metal 5 to 5.5 tonnes of bauxite ore has to be treated by the well known Bayer process. On digestion with caustic soda, bauxite leaves behind about 50% of its own weight of an insoluble residue known as Red Mud. Disposal of such huge quantities of this solid waste by their impoundment in mud lakes causes problems of increasing land cost, storage and pollution. The application of any red mud depends on its Physico-chemical properties. The red mud is highly complex material containing largely the following six oxides namely ferric oxide, Aluminum oxide, Titanium dioxides, Silica, Sodium Oxide and Calcium Oxide in varying quantities besides minor amount of traces of other elements as oxides.
Tablel: The Major Constituents of Red Mud Weight % Constituents 10.48 LOI 5.45 Si0 2 37.33 Fe 2 0 3 Ti0 2 18.88 18.38 A1203 Na 2 0 5.89 CaO 2.71
Furthermore, it is alkaline, fine sized and poor in settling. Dumping of red mud is certainly an environmentally unfriendly activity. However the solid waste, red mud is a store house of wealth because it contains the metal values of Iron, Aluminum and Titanium. Perhaps the best way would be to convert this waste into the corresponding metal
103
Results
XRD analysis was carried out using PANalytical X'Pert Cubix Pro Series diffractometer equipped with a copper target tube, X'celerator detector and operated at 40kV and 30mA.The dried sample were scanned within 2Θ range from 10-70°. Diffraction data were analyzed using PANalytical X'Pert High Score Plus Version 2.1.The main crystalline phase of red mud were Hematite (Fe 2 0 3 ) »Calcite (CaC0 3 ), Goethite (FeOOH), Gibbsite (Al(OH)3), Boehmite (AlOOH), Cancrinite (3NaAlSi04.NaOH), Quartz(Si02), Anatse (Ti0 2 ). The X-ray diffractogram of red mud is shown in Fig.l.
In the roasting experiments burnt lime was used to immoblize the Si0 2 of the red mud. A burnt lime containing 85% of total CaO was used. The effect of the different variables on the recovery of A1203 was studied.
RED MUD SODIUM CARBONATE
Roasting (1100 Of
The method adopted for the recovery of A1203, Ti0 2 and Fe 2 0 3 from red mud is presented in Fig.2. The process for the recovery of metal values from red mud is briefly outlined below: Crushed red mud was thoroughly mixed with (i) burnt lime and soda ash in the different weight ratios and then roasted at the desired temperature, ranging from 900-1100°C.The roasted mass was crushed and extracted with hot water and filtered. The Filtrate, sodium aluminate can be (ii) accommodated by recycling into the aluminate solution in the Bayer process. (iii) The residue is light and can be treated for the recovery of Iron and Titanium. The residue is digested with concentrated sulfuric acid. (iv) The digested slurry was dissolved in water and filtered. The residue containing high silica was separated. Titanium hydroxide was precipitated from the (v) filtrate by hydrolysis. The Titanium hydroxide was filtered and calcined to get pure Ti0 2 . The Filtrate was evaporated to obtain FeS0 4 . (vi)
SODIUM ALUMINATE Digest with H2S04, Fijte$_
Hydrolysis
FERROUS SULPHATE
Fig. 2. Schematic diagram for recovery of metal values from Red Mud.
Effect of Soda ash on Alumina Recovery: The red mud mixture containing burnt lime and soda ash was roasted at 1100°C for 3 hours and subsequently leached. Lime was added to the mixture so that the mass contained CaO:Si0 2 in the molar ratio of 1.5:1.The Na 2 C0 3 weight ratio was varied from 0.3 to 0.7 by the addition of varying quantities of soda ash and the alumina recoveries are indicated in Table 2 and Fig. 3.
Red Mud-14 21 08 10
It is observed that on addition of increased quantities of soda ash, the sinter becomes hard and has a tendency to absorb moisture thereby making dry grinding difficult and wet grinding has to be employed. I,
*VW**»
mm ÌW w
%M
i p
Table 2. Effect of varying sodium carbonate ratio % A1203 Sinter Weight Ratio Extracted Characteristics ( Mud:Na2C03:CaO) 64.98 1:0.3:0.1 Easily extract 69.09 1:0.4:0.1 Easily extract 76.17 1:0.5:0.1 Slightly hard 76.32 1:0.6:0.1 Slightly hard 1:0.7:0.1 76.47 Slightly hard
■ ■ ■ i ■
Position [*2Theta]
Fig.l. XRD Pattern of Balco's Red Mud
104
1 1
Fig.3.Bfect of Na2C03 ratio on alumina
Fig.4. Bfect of CaO ratio on alumia recovery
recovery
Io 80
« 80 O|70
2
1 75 -
<5 60 X
M
o
S 50
M
0.3
0.4
0.5
0.6
0.7
< 70 * 0.1
0.8
Weight ratio of Na 2 C0 3
0.12
0.14
0.16
0.18
Weight ratio of CaO
From Table 2 and fig.3 it is observed that the alumina recovery increases from 64.98% to 76.17% as the Na 2 C0 3 weight ratio increases from 0.3 to 0.5 and no appreciable increase is observed thereafter. This may be attributed to the side reactions taking place during the roasting operation itself, such as the formation of sodium ferrites and sodium titanate. The presence of Fe 2 0 3 & Ti0 2 affects the alumina recoveries. It was observed that use of Na 2 C0 3 and red mud in the weight ratio of 1:0.5, 76.17% A1203 could be recovered and further roasting tests were carried out using this quantity and the effect of other variables.
Table 4. Effect of roasting temperature %Mole ratio of 1 Temperature %A1203 Extracted Si0 2 :Al 2 0 3 dissolved in of roasting,°C solution 900 1000 1100 1150
71.12 74.34 76.30 78.09
1.01 1.09 1.11 1.23
It is observed in figure 6 that on increasing the roasting temperature from 1 hour to 3 hour, the A1203 recovery is increased from 70.25% to 75.95%. Further roasting for 4 hours and 5 hours decreases the A1203 recovery to 74.14%. A retention time of 2 to 3 hours is sufficient at this roasting temperature.
Effect of CaO on alumina recovery: Varying quantities of burnt lime were used for roasting red mud with sodium carbonate in the range of 1:0.5 and the alumina recoveries are indicated in Table 3 and Fig.4 were studied. It is observed from the Table 3 and Fig.4 that as the CaO content is increased, the alumina extraction is decreased and this can be attributed to the formation of a complex compound of lime and alumina.
Discussion When the red mud is roasted with sodium carbonate and lime, the alumina is converted into a sodium aluminate while silica is converted to an insoluble dicalcium silicate as follows: Na 2 C0 3 + A1203 = 2 NaA102 +C0 2 Si0 2 + 2CaC0 3 = (CaO) 2Si0 2 + 2C0 2
Effect of temperature of roasting on recovery of alumina: In order to determine the effect of roasting temperature mixtures containing red mud, sodium carbonate and burnt lime were heated in the weight ratio of 1:0.5:0.1 for 3 hours at temperature ranging from 900°C to 1150°C.
The soda besides reacting with A1203 reacts with Fe 2 0 3 .Ti0 2 .Si0 2 as follows: Na 2 C0 3 + Fe 2 0 3 = Na 2 O.Fe 2 0 3 + C0 2 Na 2 C0 3 + Ti0 2 = Na 2 O.Ti0 2 +C0 2 Na 2 C0 3 + Si0 2 = Na 2 O.Si0 2 + C0 2
Table 3. Effect of varying lime quantity on alumina recovery %A1203 Weight Ratio % Mole Mud:Na2C03:CaO extracted ratio of Si0 2 :Al 2 0 3 (inclusive of lime in mud) dissolved in solution 76.03 1:0.5:0.10 1.08 75.12 1:0.5:0.12 0.97 1:0.5:0.14 74.05 0.89 0.83 1:0.5:0.16 72.66
During leaching, the sodium ferrite and part of the sodium titanate decompose to Fe 2 0 3 and Ti0 2 . Na 2 O.Fe 2 0 3 + H 2 0 = 2NaOH + Fe 2 0 3 Na 2 O.Ti0 2 +H 2 0 = 2NaOH+Ti02 The sodium silicate dissolves in water and thus enters the solution.
105
Acknowledgement
Fig.6.Bfect of time of roasting on alumina recovery * * 80 Ί e« g Z 70 <
-t—
-f—
*
^^
♦
-♦
3
4
5
The authors acknowledge the constant encouragement from our Chief Operating Officer,Mr.Bibhu Prasad Mishra and Chief Executive Officer and Whole Time Director,Mr. Gujan Gupta during the progress of this work.The authors would like to thank the Management of Vedanta Resources for allowing us to publish this paper.
X*50 2
References
6
Time,hours
1. J. D. Edwards, F.C. Frary, The Aluminium Industry, Vol.1, Mcgraw Hill Book Co, p. 127. 2. Edwards, Trans, A.I.M.E., Vol.182, 1949. 3. F.R. Archibald, Trans.A.I.M.E. Vol.159, 1944. 4. R.S. Thakur and S.N. Das, Red Mud Analysis and utilization, Wiley Eastern, Limited, New Delhi, 1994.
The residue resulting from the lime soda roasting process is very light, fine and can be treated for the recovery of Iron and titanium values. The following observations were found on roasting of red mud with Na 2 C0 3 and lime in the weight ratio of 1:0.5:0.1 atll00°Cfor3hours: (i) The alumina extraction efficiency was 76%. (ii) Recovery of Iron as Fe 2 0 3 was 74.81 % (iii) Recovery of Titania as Ti0 2 was 72.13% Conclusion It was observed that 76% of the A1203 could be extracted by roasting the red mud with Na 2 C0 3 and CaO (weight ratio 1:0.5:0.1) at 1100°C for 3 hours. About 0.8-1.1% mole ratio of Si0 2 dissolved simultaneously in solution. The aluminate liquor obtained is unsuitable for precipitation of desired grade of aluminium hydroxide. It will thus be necessary to adopt a separate desilication step. It is suggested that if this silica containing solution from the lime soda process is added, the resulting pregnant solution will contain low Si0 2 concentrations suitable for subsequent precipitation of aluminium hydroxide. The proposed method provides more environmentally acceptable process for the extraction of Aluminium oxidejron Sulphate and Titanium dioxide from red mud.
106
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
RED MUD FLOCCULANTS USED IN THE BAYER PROCESS F. Ballentine, M. E. Lewellyn, S. A. Moffatt Cytec Industries Inc.; 1937 West Main Street; Stamford, CT 06904-0060 USA Keywords: Bayer process, flocculants, flocculation, red mud characteristics of each type of flocculant can ultimately be attributed to the molecular composition and also to the molecular structure. This paper examines chemistry of each type of flocculant, and correlates these differences to the flocculation performance.
Abstract Flocculation and separation of red mud is an integral part of the Bayer process. Over the latter half of the 20th century, flocculant technology dramatically changed from natural starches to use of "rationally designed" polymers. Many of these advancements were due to the introduction of liquid or emulsion based flocculants which enabled elaborate post-reaction chemistry to be done on the polymer backbone. This paper presents a historical overview of milestones of flocculant technology used in the Bayer process up to present day. Discussion of flocculants is based on inventions in the published literature that have gained widespread use throughout the industry and will included the benefits/advantages of different flocculant technology for settling red mud.
Background It is worth noting that following his discoveries, Dr. Bayer received contracts to build alumina refineries in France, Germany, Italy, the United Kingdom, and the United States [1]. Thus, it can be conjectured that the global alumina industry likely originated from the plants that Dr. Bayer helped design and build. In all likelihood the initial design of these plants did not utilize flocculants in the separation of red mud, and mud was removed by direct filtration of digested slurry. Undoubtedly, efforts at process improvement commenced in parallel among the different geographical regions where Bayer helped build refineries. However, for the purpose of this paper and the role of red mud flocculants, discussion will be limited to developments within the North American and Jamaican refineries.
Introduction The Bayer process for the production of alumina has changed little since it was discovered and developed by Dr. Karl Joseph Bayer. The initial discovery (1888) that aluminum hydroxide could be precipitated by seeding a supersaturated sodium aluminate solution, and the subsequent discovery (1892) that aluminum hydroxides and oxides in bauxite could be selectively dissolved in sodium hydroxide under pressure to produce a sodium aluminate solution remain the fundamental steps in the production of alumina today [1]. Obviously, there have been tremendous advances in terms of design, engineering, and control of the process since then, but the basic operations are still based on Dr. Bayer's discoveries.
In the early 20th century, there are numerous references to the use of starch as a flocculant [2-4]. The coal industry was among the first to utilize starch, in the 1930s, as a flocculant for the removal of fine coal particles from washery water [2]. Around this time, 1939, starch was also reportedly used as a flocculant to aid in the filtration of red mud [3]. It is further reported that in 1940, due to the use of lower grade bauxites in the United States, the first cone bottom settlers were installed in the process. This led to the development of counter current décantation (CCD) circuits comprised of primary settlers and washers. At this point it seems that the alumina industry was by far the largest consumer of starch based flocculants [4].
One of the most important advances in the Bayer process has been the introduction of red mud flocculants. Interestingly, it is reported that Dr. Bayer further described the advantages of his process in terms of the purity of the alumina produced, relative to precipitation by C0 2 , and the crystalline form which made it easy to wash and filter [1], However, he also cited the main difficulty in the process was the separation and removal of the red mud; fine, insoluble, impurities that needed to be removed from the sodium aluminate solution. This is where flocculants have played an indispensable role in the process. Considering the history of flocculant use for alumina production by the Bayer process there are arguably three distinct milestones: starch, synthetic polymers, and modified synthetic polymers. Each of these flocculants have unique characteristics which is why the industry gradually moved from starch to synthetic polymers, and then to modified synthetic polymers. The benefits and performance
A significant advance in flocculant technology occurred in the mid-1950s with the introduction of synthetic flocculants for the mining industry [5,6]. The technical literature is rich with information on the development and use of synthetic flocculants of this era as several large chemical manufactures, in addition to governmental and academic researchers, were active in this area. These flocculants, based on acrylamide and/or acrylate offered several advantages over starch. Among these were: significantly lower dosages, higher settling rates, better filtration, improved overflow clarity, and higher underflow density. Within the North American region of the alumina industry, these synthetic flocculants were first used continuously on an industrial scale in Jamaica in the early 1970s. Synthetic
107
flocculants either replaced starch or were used with significantly reduced dosages of starch.
CH 2 OH
CH 2 OH
0-K°»
Over the years significant advancements have been made to synthetic flocculants. Many of these have been described in technical literature and include product form (solution, gel, dry powder, and oil based emulsion), molecular weight, variation in anionicity, structure, and composition. However, the most significant advancement has been the development of hydroxamated polyacrylamide (HXPAM) polymers. Development of these polymers as a flocculant was first announced in 1988 [7], and over the last 20 years they have become the primary flocculant used in red mud settlers around the world.
H
OH
H
,/L< OH
H
OH
Figure 1. Structure of amylose in starch based flocculants
With above as a background, the following sections of this paper explain why the industry moved from starch to polyacrylate/polyacrylamide and then onto hydroxamated polyacrylamides. The argument presented is based on the compositional and structural differences among these polymers and the benefits they offer.
Figure 2. Structure of amylopectin found in starch.
The composition and structure of amylose and amylopectin provide insight into how starch interacts with red mud particles to form floes. The hydroxyl groups extending from the cyclic rings are likely to adsorb on the red mud particle surface. However, because of the sterically hindered structure and the limited charge density, the starch is likely to be poorly adsorbed on the surface. More importantly, the effective molecular weight (< 1 million for amylose) is relatively low for a flocculant, and therefore there is minimal extension of the molecule from the mud particle surface to allow contact with other particles and form floes. Floe or aggregate formation is believed to occur via a a physisorption mechanism, which includes electrostatic, hydrogen and hydrophobic bonding. Although hydrogen bonding can form strong attachment of the starch to the mud particle surface, limited attachment points and molecular extension from the mud surface likely explain why such high dosages of starch are needed to effectively flocculate red mud. Dosages on the order of approximately 16 kg of starch per ton of mud solids have been reported for muds produced from some bauxites [9].
Starch Starch is a set of natural polymeric materials primarily composed of amylose and amylopectin, and belongs to the general class of polysaccharides. It has been commonly derived from sources such as flour, corn, and potatoes. In terms of composition, amylopectin composes the outer shell of the starch grain while amylose is contained within the grain. Before starch can be used as a flocculant, a solution of it must first be made. This is accomplished through a process described as gelatinization whereby the starch grain is heated in water. During this process, the outer shell of the grain swells or expands allowing the amylose to be released. Once in solution, the amylose has flocculating properties. Amylose has an appreciable solubility in water, more so than amylopectin. Literature on the use of starch states that the addition of caustic to the water used to create starch solutions enhanced the flocculation performance. It was conjectured that the caustic was responsible for hydrolyzing phosphoric esters, present in the amylopectin, into amylose and phosphoric acid [2]. Thus, the concentration of amylose in solution increased resulting in enhanced flocculation performance.
Furthermore, this would also seem to explain the significant contribution made by starch towards liquor organics concentration. Starch is very soluble in Bayer process liquor, and because of the relatively weak adsorption mechanisms; a significant portion of the dosage stays in the liquor and reports to the settler overflow. In one study, it was reported that approximately 40-50% of the total organics were due to the use of starch as a flocculant [10].
Structures of both amylose and amylopectin are shown in Figures 1 and 2. Structurally, both of these molecules are composed of joined glucose units. Amylose is described as more linear in its linkage configuration, a more flexible molecule, and to have molecular weights below 1 million daltons [8]. Amylopectin is a more branched molecule, and can reach higher molecular weights on the order of 10-100 million daltons [8].
Despite the drawbacks listed above, starch is still used in a few refineries today. One positive attribute it does have is that due to its structure and size it generally does not blind red mud filter presses that are found immediately downstream of the primary red mud settlers. Furthermore, because of the high amount of residual starch in the Bayer liquor, unflocculated mud that may escape from a settler
108
feedwell can be flocculated in the more quiescent region of the settler before reporting to the overflow. However, despite these features, and a significant economical cost advantage, starch has essentially been replaced by synthetic polymers as the primary flocculant. Synthetic flocculants Synthetic polymers used as flocculants in the Bayer process are derived from acrylamide and/or acrylate monomers. Other noteworthy flocculants have been invented over the years; however the most widely used by the alumina industry have been the aforementioned. The basic structure for polyacrylamide and polyacrylate are shown in Figures 3 and 4, respectively. Furthermore, Figure 5 shows the structure of a copolymer synthesized by copolymerizing monomers of both acrylamide and acrylate. While the characteristics of these polymers have been well described in technical literature [11-15], a common feature to the structures is the flexible linear "backbone" with amide and/or carboxyl groups extending off the backbone. The linear structure enhances the aqueous solubility of the polymer, relative to highly branched polymers. This conformation is also optimal for adsorption of the polymer onto mud particles, thereby improving the flocculation performance of the polymer.
(—CH2—CH—)n C=0 H2 Figure 3. Basic monomer unit of a polyacrylamide.
-^CH2-ÇH^
Oe N a®
Figure 4. Basic monomer unit of polyacrylate. As shown, the sodium exists as a counter ion on the negatively charged carboxyl group. -4-CH 2 —CH4x
fCH2-CH^
C=0
C=0
NH 2
O^Na®
Conventional primary flocculant, x = 5-0%, y = 95-100% Conventional washer flocculant, x = 50-5%, y = 50-95% Figure 5. Copolymer of acrylamide and acrylate. Respective ratios illustrate composition differences used for primary settlers and red mud washer flocculants.
As opposed to natural polymers, such as starch, there is a great deal of flexibility in manufacturing synthetic polymers. It is possible to customize the molecular weight, structure, and the amount and type of charge for a particular application. In terms of flocculation, it is quite possible to create linear polymers with molecular weights on the order of 20 million, much higher than starch or naturally occurring polymers. On the other hand, if a filter aid or dispersant is needed it is possible to reduce the molecular weight to the order of 100,000 or < 100,000, respectively. In terms of charge it is well known that primary red mud settlers require flocculants with a high degree of anionicity, such as found in polyacrylate polymers. However in middle to lower stage red mud washers, moderate (90-70%) to lower (70-50%) charged copolymers provide superior performance. Additionally, synthetic flocculants are made in a variety of product forms. Introduced initially as aqueous solutions, gel logs, or dry powders; these were the only forms until the mid to late 1970s. At this time, the invention of water-in-oil emulsion polymerizations enabled much higher molecular weight polymers to be produced in a much easier-to-use form. While dry powders are still used and offer certain benefits, the vast majority of flocculants used in the alumina industry today are inverse emulsions. While the above information provides the compositional and structural characteristics of synthetic flocculants it does not tell the full story about why they offered such superior performance relative to starch. This is attributable to the high molecular weight and high charge density which affect how, and how much, polymer is adsorbed on mud particle surfaces. In primary red mud settlers polyacrylate is by far the most effective flocculant of the polymers discussed so far. Polyacrylate contains 100% carboxyl groups extending from the backbone of the polymer, and therefore has the highest charge density possible. The repulsive forces between adjacent carboxyl groups help extend, or open up, the conformation of the polymer in solution. This serves to increase the probability that part of the polymer will adsorb to the surface of a mud particle when a collision occurs. Adsorption of polyacrylate and copolymers of acrylamide/acrylate to the mud surface is widely believed to be through a multi-valent ion bridging mechanism whereby multi-valent ions act as a bridge between the carboxyl group on the polymer and the mud particle surface [16]. This type of mechanism would fall under a broad category of chemisorption mechanism. It is conjectured that only part of the molecule is attached to the mud particle surface and that a significant portion of the molecule extends out from the surface. The unattached portion can exist as "tails" or "loops" and serves to attach to other mud particles. Once this happens, a floe or aggregate is formed. The higher molecular weight and linear structure of synthetic flocculants is a distinct advantage in this regard. Larger floes can be formed that have faster settling rates, and provide better mud compaction. Perhaps the most prominent attribute of synthetic flocculants, relative to starch, is the higher settling rates achieved with significantly lower dosage. Settling rates
with synthetic flocculants were reported to be an order of magnitude higher than those achievable with starch; easily exceeding 10 m/hr. Furthermore, effective dosages were reportedly an order of magnitude less with the synthetic flocculants. The higher settling rates produced other benefits, such as increased settler throughput, and increased underflow density. Higher throughput was often an artifact of attaining higher settling rates in large settlers designed with starch as the flocculant. Increased underflow density led to improved soda and alumina recovery in the washer circuit. Interestingly, although synthetic flocculants produced higher settling rates, the floes were more fragile than those made with starch [9, 17]. Floes formed with synthetic flocculants were easily sheared through the underflow pumps so that additional flocculant was required in subsequent washing stages. However, the fragility was also perceived as a positive in that the floes were more compressible and easier to dewater. Much of the "robustness" of starch formed floes was due to small (an artifact of the low molecular weight of starch) but strong floes, and perhaps residual starch in the entrained liquor, which acted to reflocculate the mud after the underflow pump. The benefits of synthetic flocculants listed above far outweighed deficiencies in supernatant clarity, and filtration. While clarity was "good" with polyacrylate or various copolymers, starch was often used in conjunction with these flocculants to achieve better clarity [18]. A general performance characteristic with these flocculants was that as dosage was increased, and consequently settling rate, overflow clarity degraded [19]. This represented an operational issue in that if more flocculant was required to control mud level or increase underflow density, overflow clarity could be compromised. The solution was to supplement the synthetic flocculant dose with some starch to control clarity. One definite drawback to synthetic flocculants that was not an issue with starch was the risk of blinding red mud pressure filters. The extremely high molecular weight of the synthetic polymers would easily blind filters if nonadsorbed, or residual, flocculant escaped the feedwell and reported to the settler overflow. A common indication of this was an immediate decrease in filter throughput, a shiny or glossy appearance of the filter cake, and a "sticky" consistency to the mud on the individual leafs of the press. Again, this difference between starch and synthetic flocculants can be attributed to not only molecular weight, but also structural differences in the molecule. Obviously, the alumina industry has embraced the use of synthetic flocculants in the separation of red mud. This was despite a large "learning curve" associated with determining makeup and application requirements. Something normally anticipated with the introduction of new technology. As reported, the time invested in developing and converting one (of two) Jamaican refineries to synthetic flocculants easily incurred enough savings in raw materials in one year to finance conversion of both plants to synthetic flocculants [9].
Hydroxamated polyacrylamides The latest advancement in synthetic flocculants, that gained widespread use in the alumina industry, occurred in the late 1980s with the introduction of hydroxamated polyacrylamides (HXPAMs) [20]. These polymers, as shown below in Figure 6, are an example of a terpolymer in that there are three different functional groups on the polymer backbone: amide, carboxyl, and hydroxamate. This type of polymer is only available in liquid (i.e. inverse emulsion) form and was made possible by the development of emulsion polymerization technology. It is essentially made by carrying out a post-reaction on polyacrylamide to incorporate the hydroxamate and acrylate groups on the backbone. A post reaction, meaning a separate chemical reaction after the main polymerization reaction to create the backbone, can only be carried out in any practical measure in a liquid or emulsion form. Dry powders do not lend themselves easily to additional chemical reactions, and aqueous solutions are too dilute.
-eC^-CIftr-fCH^H)^— (CH2-ÇH)—
ç=o
NH2
ç=o
0"Na
+
ç=o _
NHO Na+
Figure 6. Hydroxamated polyacrylamide flocculant which contains three separate functional groups: amide, carboxyl, and hydroxamate. These polymers have unique performance characteristics in that they have a positive dosage response in terms of settling rate and clarity. This means that as the dosage of an HXPAM flocculant is increased settling will increase, and clarity will also improve. This is a departure from the performance of traditional synthetic flocculants in that an increase in dosage would produce an increase in settling rate, but at the compromise of clarity. Laboratory data illustrating the different response characteristics is shown in Figure 7. In terms of settler operations, this is of tremendous value. When situations arise where higher settling rates, or higher dosage is needed the operator can make the necessary adjustment without being concerned about degrading the quality of the overflow of the settler. Development of these types of polymers also represent the current era of flocculant research and development. The incorporation of other functional groups into traditional polymer backbones is intentional, by design, to give the polymer improved flocculation capabilities [16]. This could be improved dosage response, or the ability to flocculate selective mineralogy, such as DSP, in red mud [21]. Since the first industrial application [22] of HXPAMs in the early 1990s, these flocculants have slowly become the main flocculant used in primary red mud settlers. Initially, several refineries used these flocculants not only on the red mud settlers, but also on the washer circuit. The reason for this was the exceptional performance, and positive control
formation in the settler, thereby extending the life of the settler [25].
that these flocculants offered relative to conventional polyacrylate and copolymers. Since then, use of HXPAMs in washer circuits has been scaled back significantly, mainly due to economics, and the fact that copolymers deliver better mud compaction. However, a number of plants use HXPAMs on the upper stages (first or second) because of superior clarity (when the overflow goes to precipitation) and reduced auto-precipitation. Occasionally, the use of HXPAMs is extended down the washer train to deal with difficult settling mud.
A
• : t ^
-
B
C
D
E
F
G
H
I
J
K
tlTT
L
M
N
I
&ENETTS Ac I ligner Prwiuctica 8: Improved F$krit»n C: latfrov«i Clarity D: Staxh Elimînaâcn E: Handling Diöicu t Bauxite F: Higher Uff Density G: Lesa Thkkener Scaie H: Una Reduction I: Reduce Na Scttk-s J: Bern* Mud l£vé Control K: Incseased A/C L: Lowei Floe. Costs M: Improved Wasbg- Opciarion Η'. fiter By-hss
Figure 8. Summary of benefits provided by hydroxamated polyacrylamide flocculants on primary red mud settlers. ♦ Hydroxamated polyacrylamide
Conclusions
50
100
150
This paper has summarized the main types of flocculants used by the alumina industry, since the process was first discovered by Bayer. Flocculant technology was shown to have progressed from the introduction of natural polymers, such as starch, to the use of synthetic polymers based on acrylamide and/or acrylate. This lead to modified synthetic polymers such as hydroxamated polyacrylamide, which are used throughout the industry on primary red mud settlers.
200
Floccu Ian t dosage - grams of f loccu lant / ton of mud
Figure 7. Supernatant clarity from lab tests showing the different responses achieved with hydroxamated polyacrylamide and traditional polyacrylate flocculants.
So what does the future likely hold for red mud flocculants? It is unlikely that there will be a change from synthetic flocculants based on acrylamide and/or acrylate in the near future. These types of polymers have clearly demonstrated the ability to provide suitable performance in the separation of red mud for approximately 40 years. Furthermore, it is difficult to compete with the economics of the raw materials used to produce these flocculants, and therefore they should continue to be the primary flocculants in the industry for the foreseeable future.
The unique interaction of HXPAMs with red mud particles is due to the hydroxamate functional group. This group has a very strong affinity for iron bearing minerals and tends to adsorb through a chemisorption (chelation) mechanism. Studies using atomic force microscopy (AFM) have confirmed that surface adsorption of hydroxamated polymers is stronger relative to polyacrylate [23]. Furthermore, the presence of this functional group provides faster adsorption rates for these polymers relative to polyacrylates or starch. This is particularly true in certain liquors with a high organic load. Additionally, it is what also has enabled HXPAMs to flocculate a wide range of red muds, even those considered to be difficult to settle.
Research efforts on flocculants will likely focus on "specialty" or "rationally designed" polymers that have unique capabilities to (selectively) flocculate certain solid phase impurities, such as DSP for example. New flocculants may be utilized as a supplement to the primary acrylamide/acrylate flocculants outlined in this paper, and be used on an occasional basis. Efforts to develop this type of flocculant technology would seem to fit well with the direction that the industry is moving; processing bauxite reserves with deteriorating quality. The ability to separate and remove mud produced from these bauxites, and maintain industry specifications for alumina quality, may rely heavily on the use of processing aids, such as new flocculants.
As described earlier in this paper, the chief benefit of starch was the ability to settle red mud and provide a significant reduction in mud load to the filters. Synthetic flocculants brought significant increases in settling rate, settler throughput, flocculant dosage, and underflow density. Performance improvements associated with the use of HXPAMs have also been reported [24]. These are summarized in Figure 8 and include: settling of difficult mud, increased settler throughput/production, improved filtration, improved overflow clarity, and elimination of starch. Many of these improved characteristics are due to the stronger, faster polymer adsorption on mud particle surfaces. This leads to better scavenging of fine mud particles, and less residual polymer in the liquor that can blind filters. Lower suspended solids and a more distinct mud interface reduces seed sites for gibbsite scale
Ill
References 1. 2.
3. 4. 5. 6.
7.
8. 9. 10.
11.
12.
13.
14.
15.
16.
Fathi Habashi, "A Hundred Years of the Bayer Process for Alumina Production", Light Metals (1988), pp. 3-11. George R. Gardner and Kenneth B. Ray, "Flocculation and Clarification of Slimes with Organic Flocculants", AIME Transactions, (1*939), pp. 146-168. L. Keith Hudson, "Evolution of Bayer Process Practice in the United States", Light Metals (1988), pp. 31-36. Donald A. Dahlstrom, "Liquid-Solid Separation Challenges in the Fast Lane", Challenges in Mineral Processing, (1989) SME, pp. 467-475. Merrill F. McCart and Robert S. Olson, "Polyacrylamides for the Mining Industry", Mining Engineering, January (1959), pp. 61-65. William F. Linke and Robert B. Booth, "Physical Chemical Aspects of Flocculation by Polymers", AIME Transactions, Vol. 217, March 12, (1959), pp. 364-371. D.P. Spitzer, A.S. Rothenberg, H.I. Heitner, and M.E. Lewellyn, "New polymers for the Bayer process", Alumina Quality Workshop, Gladstone, QLD, September (1988), pp. 221-230. Marguerite Rinaudo, "Polysaccharides", KirkOthmer EncycL Chem. TechnoL, 5th Ed., J. Wiley & Sons, New York (2006), 20, pp. 549-586. J.L. Chandler, "Advances in the use of synthetic flocculants", Light Metals, (1976), Vol. 2, pp. 163-171. M.J. Pearse and Z. Sartowski, "Application of special chemicals (Flocculants and Dewatering Aids) for red mud separation and hydrate filtration", Bauxite - Chapter 38, (1984), pp. 788810. L. J. Connelly, D. O. Owen, and P. F. Richardson, "Synthetic flocculant technology in the Bayer process", Light Metals, Vol. 2, (1986), pp. 61-68. M. J. Pearse, "Synthetic flocculants in the mineral processing industry- types available, their uses and advantages", Light Metals, (1984), pp. 101-106. S. Weir and G. M. Moody, "Trends in the development of flocculants as aids to solid/liquid separation", Journal of The Filtration Society, Vol. 1(4), (2001), pp. 11-12. M. J. Pearse, S. Weir, S. J. Adkins, and G. M. Moody, "Advances in mineral flocculation", Minerals Engineering, Vol. 14, No. 11, (2001), pp. 1505-1511. Richard E. Ellwanger, "Use and application of synthetic organic flocculants in the mining industry", AIME Transactions, Vol. 270, (1980), pp. 1812-1815. H. T. Chen, S. A. Ravishankar, and R. S. Farinato, "Rational polymer design for solidliquid separations in mineral processing applications", International Journal of Mineral Processing, Vol. 72, (2003), pp. 75-86.
17. Ronald J. Allain, "Water soluble polymers - the future?", Reagents for a better metallurgy, Chapter 26, SME, (1994), pp. 263-267. 18. T. K. Hunter, G. M. Moody, S. E. Sankey, and C. A. Tran, "Advances with chemical additives for the alumina industry", Light Metals, (1991), pp. 159-165. 19. David O. Owen, Lawrence J. Connelly, and Peter A. Dimas, "Evaluation of downstream effects of specialty chemicals in the bayer process", Light Metals, (1991), pp. 173-176. 20. D. P. Spitzer, A.S. Rothenberg, H.I. Heitner, M.E. Lewellyn, L.H. Laviolette, T. Foster, and P.V. Avotins, "Development of new Bayer process flocculants", Light Metals, (1991), pp. 167-171. 21. M. Davis, Q. Dai, H.-L. Chen, and M. Taylor, "New polymers for improved flocculation of high DSP-containing muds", Light Metals, (2010), pp. 57-61. 22. Maritza Faneitte, Alfredo Galarraga, and Terry Foster, "Utilization of new polymer in Interalumina", Light Metals, (1994), pp. 129-131. K. E. Bremmell and P. J. Scales, "Adhesive forces between adsorbed anionic polyelectrolyte layers in high ionic strength solutions", Colloids and Surfaces A: Physicochemical and Engineering Aspects, Vol. 247, (2004), pp. 19-25. Roderick G. Ryles and Peter V. Avotins, "Superflue HX, a new technology for the alumina industry", 4th International Alumina Quality Workshop, Darwin, Northern Territory, Australia, June (1996), pp. 205-215. Peter V. Avotins, Larry L. Laviolette, Emiliano F. Repetto, and Araujo M. Eli, "The effects of flocculants on thickener scaling", 5th International Alumina Quality Workshop, Bunbury, WA, March (1999), pp. 448-455. © 2010 Cytec Industries Inc. All rights reserved.
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
REDUCTIVE SMELTING OF GREEK BAUXITE RESIDUES FOR IRON PRODUCTION A. Xenidis1, C. Zografidis2,1. Kotsis1, D. Boufounos3 national Technical University of Athens, School of Mining and Metallurgical Engineering, Greece 2 LARCO General Mining & Metallurgical Company; Larymna Lokridos; Fthiotida, GR350 12, Greece 3 Aluminium of Greece SA; Agios Nikolaos, Viotia, GR 32003, Greece Keywords: bauxite, residue, smelting, reduction, iron have been reported [4, 5, 6]. The most common reducing agent utilized is coke while the iron recovery finally obtained is always over 90%.
Abstract The reductive smelting of Greek bauxite residues was investigated for the production of an iron product that meets some crucial industrial requirements as a blast furnace feed. Fine-grained Greek bauxite residue - either as is or in the form of pellets - and solid fuel reducing agents -lignite and coke-, were used as raw materials. The effect of parameters such as the smelting temperature, the amount of the reducing agent in the mixture, the retention time and the addition of fluxes on the quality of the metallic product, as well as the basicity and the desulfurization capability of the slag, was investigated. The results obtained regarding the chemical properties of the metallic product were very promising, providing input for further research on the optimization of the proposed pyrometallurgical method for the production of an attractive ferrous raw material for the iron ore industry.
Experimental The sample of bauxite residue utilized in the present study was provided by the Greek alumina refinery owned by Aluminum of Greece S.A. It is a fine grained material, given that a percentage of 40% by weight is less than 2 urn. Lignite and coke were used as reducing agents. Chemical analysis of the raw materials is given in Table I. The total iron (Fetot) content in the bauxite residue is about 33.5%, which is indicative of the fact that it constitutes an important potential iron source for the iron ore industry. Its main mineral constituents, as determined by mineralogical analysis conducted by X-Ray diffraction, were hematite, gibbsite, diaspore, calcite and hydroxysodalite. Table I. Chemical Analysis of Greek Bauxite Residue and Reducing Agents Lignite Coke Bauxite Component residue
Introduction The application of various processing methods for the effective recovery of major metallic constituents of bauxite residues, has been a matter of scientific interest by numerous researchers [1]. The current work constitutes a part of an integrated research effort which aims to investigate alternative ways of utilizing bauxite residue as a raw material in the iron ore industry, by applying pyrometallurgical treatment methods. Within this framework, roasting reduction, magnetic separation for iron enrichment and pelletization of the Greek bauxite residue have already been investigated [2, 3]. The present study is focused on the reductive smelting of the same raw material for iron production.
Fe 2 0 3 Na20 Si0 2 CaO MgO A1203 K20 Ti0 2 C0 2 S0 2 L.O.I.
The crucial technical requirements that determine the usability of an iron-bearing raw material for the iron ore industry -either as a feed for a rotary kiln or a blast furnace-, are the following: iron content, "gangue constituents" (Si02, CaC03, MgO, A1203), "deleterious impurities" (S, P, Na 2 0, K20 and Ti02), physical properties (porosity, durability, ore structure), reducibility and its amenability to concentration. Given that about 1.61 tonnes of iron ore with an iron content of 60% are required for the production of each tonne of cast iron, and also considering the fact that iron ore prices have almost quadrupled since 2004, it is easy to understand how crucial it is for the industry to make use of an iron bearing raw material with a lower iron content (20-30% Fe) which can substitute part of the iron ore feed. The iron content of bauxite residue, mainly in the form of iron (III) oxide, ranges from 30 to 50%, while at the same time the typical content of gangue material -such as CaO and A1203 - is such that this does not seriously affect its melting point. Thus, much attention has been focussed towards iron recovery in these residues.
C fix
Ash Volatile Matter
(%)
48 3.26 6.96 14.84 0.24 15.85 0.07 7.06 2.24 0.78 11.37
(%)
(%)
4.44 2.83 53.81 4.91 3.39 19.11 0.69 0.79 3.78
29.09 2.67
30.1 20.1 49.75
81.7 5.0 13.3
0.5 2.9
Moreover, a single experiment was conducted regarding smelting reduction of bauxite residue in form of pellets with lignite (bauxite residue/lignite: 1/3 by weight and bentonite: 0.6% by weight as a binding agent). The melting temperature of the bauxite residue was initially determined with a LECO AF 600 apparatus, by employing the Seger cones technique, based on the determination of four characteristic temperature values: the start of softening, end of softening, start of melting and end of melting points. The melting point of the bauxite residue was determined to be 1400°C, as a mean of five values. The determined value is in agreement with the theoretically calculated value from the ternary phase diagram of FeO - CaO - A1203 (the Fe203 content of the bauxite residue is stoichiometrically converted to FeO content).
Reduction smelting constitutes the most significant pyrometallurgical process for the production of cast iron. It is mainly conducted in blast furnaces or electrical furnaces. Many efforts in the smelting reduction of bauxite residue material
The smelting reduction experiments were conducted in two types of laboratory furnaces:
113
The sample fed in both furnaces was 70-80 g. It is noted that two types of furnaces were used for the conduction of smelting reduction of experiments, due to the fact that each one has several advantages and disadvantages. It is noted that the determination of temperature was more precise in the experiments conducted in the Tamman furnace, however, the duration of each experiment was more than 4 hours in Tamman and only half an hour in the induction smelting furnace. Nevertheless, due to the intense eddying of the material resulting from the Foucault current, there was loss of fine material and therefore, it was difficult to conduct mass balance calculations.
i) Electric resistance heating furnace Tamman (Figure 1) (60 KVA, Tmax: 2500°C): its main part is a cylindrical heating tube made of carbon material, surrounded by a thermal insulating material, so that the inside of the heating tube acts as a heat treatment chamber, in which the graphite sample holder is placed. The temperature was recorded by a Raytek optical sensor pyrometer (laser). ii) Induction smelting furnace (Figure 2): This consists of an induction heating system, a water cooled induction coil, a supporting body having a suitable circumference for supporting a graphite crucible. The temperature during the experimental procedure was measured by the same Raytek optical sensor pyrometer.
1. Crucible 2. Heating tube 3. Graphite protecting tube 4. Silimanite mantle 5. Amianthus 6. Water connection 7. Cooling water connecting tube 8. Grains offillermaterial 9. Graphite grains 10. Graphite grains 11. Cooling water 12. Steel mantle
Figure 1. Electric resistance heating furnace Tamman
Figure 2. Heating chamber of the induction smelting furnace
In Figure 4 the effect of the type of the reducing agent and reaction time on carbon content in the pig iron is given. It is seen that increase of the retention time of the bauxite residue melt results in a significant decrease of the carbon content. Moreover, after 15 minutes of smelting reduction, the carbon content of the metallic product fluctuates between 4.6-5.1%, while after 45 minutes the respective values are 2.04-3.01, in compliance with the requirements of commercial cast iron. The use of a solid fuel like coke, which is much more reactive at higher temperatures (>1000°C) favors the reduction, contrary to lignite, which is much more reactive and evolves its thermal energy at lower temperatures.
Results and Discussion The effect of the following parameters on the final result of smelting reduction was investigated: temperature, retention time of the smelted material, bauxite residue/lignite ratio and the addition of fluxes. The determination of the bauxite residue/lignite ratio was based on the stoichiometrically required quantity of carbon required for the complete reduction of Fe203 to metallic iron. Moreover, the addition of fluxes (CaO and Si0 2 reagents), was based on the theoretical determination of the desirable lower melting point, according to the ternary phase diagram of Al203-CaO-Si02, taking into consideration that the aforementioned are the three basic constituents of the bauxite residue slag. Thus, the addition of three different ratios of CaO and Si0 2 was investigated per 100 g of the mixture bauxite residue/lignite (4/1 by weight): i)Si02-CaO: 20-0 g, ii)Si02-CaO: 35-5 g, iii)Si02-CaO: 62-38 g.
1800 1750
|;..;,;; ..:,;;..:Ί
1700
u 1650 e g, 1600 oo
The experimental conditions of reduction smelting as well as the chemical analysis of the metallic products, are presented in Table Π. The most important conclusion deduced is that regardless of the conditions employed, a metallic product with iron and carbon content ranging from 91.3-95.8% and 2.4-5.4 % respectively, was produced. The quality of iron produced can be characterized as satisfactory, since according to a typical analysis of cast iron, the iron content is approximately 94% and carbon content fluctuates between 3.5-4.5%. In Figure 3, the effect of the ratio of bauxite residue/lignite on mixture melting point is presented. It is noted that increase of the aforementioned ratio results in a considerable increase of the melting temperature.
I 1550
I
Ì::-:::'::: ;·■>;:; ■>::;■:■)
1500 1450 1400
4:1 3:1 1.5:1 Bauxite residue/lignite ratio (wt/wt)
Figure 3. Effect of the ratio bauxite residue/lignite on the melting point
114
Table IL Experimental Conditions of Reduction Smelting Experiments and Chemical Analysis of the Metallic Product (%) Content in the Metal | Fluxes BR/ Retention Smelting Raw Reducing Reducing Time Rest CaO Si0 2 s/n Material Agent Agent Fe S C (min) Admixtures (g) (g) Ratio Pellets B.R. * 1 B.R. 2 B.R. 3 B.R. 4 B.R. 5 B.R. 6 B.R. 7 B.R. 8 B.R. 9 B.R. 10 B.R. 11 * BR: Bauxite Residue
0
Lignite Lignite Coke Lignite Lignite Lignite Lignite Coke Coke Lignite Lignite
4/1 4/1 11/1 4/1 4/1 4/1 4/1 11/1 11/1 3/1 1.5/1
35
5
20 62
38
15 15 15 15 15 30 45 30 45 15 15
94.8 91.3 94.2 91.7 91.9 94.5 95.3 95.0 95.8 95.6 95.8
0.1 0.7 0.2 0.4 0.4 0.3 0.2 0.1 0.5 0.3 0.4
3.4 5.4 4.6 5.3 5.2 3.5 3.0 3.2 2.4 2.7 2.8
1.7 2.6 1.0 2.6 2.5 1.7 1.5 1.6 1.2 1.3 1.0
1450°C. On the contrary, the melting temperature value measured the same way of the mixture of bauxite residueAignite (4/1 by weight) was 1560°C. The aforementioned values are in agreement with the theoretically calculated melting temperature values from the ternary phase diagrams in Figures 6 and 7. 4.5
60 40 Time (min) Figure 4. Effect of the reducing agent and retention time on carbon content of the metallic product (reducing agent in 50% excess of the stoichiometrically required quantity) 20
1
2
4
5
6 7 8 9 10 Smelting S/N Figure 5. Basicity index of the slag samples produced Moreover, the slag produced by the addition of 35 g Si0 2 and 5 g CaO (Point B) per 100 g of mixture (bauxite residue/lignite), is in the eutectic region of the slag produced by the blast furnace operation.
The addition of fluxes results in the alteration of the basicity index of the slag. The basicity index is determined within the framework of the current study as the ratio (bases/acids). The oxides FeO, MgO and CaO are considered as bases while Si0 2 is considered as acid. A1203 is considered as an amphoteric oxide, which means that an A1203 (%) content higher than 18% is considered as basic, and lower than 15% is considered as acidic. On the contrary, for A1203 (%) content 15-18, it does not participate in the slag basicity calculations.
The smelting reduction of pellets of Greek bauxite residue with lignite (diameter of pellet: -9.5+6.3 mm, bauxite residue /lignite: 4/1 by weight) resulted in the production of a metallic product of good quality (94.8% iron and 3.1% carbon), something which is ascribed to the better contact of the bauxite residue with the solid reducing agent, favoring in such a way, the iron oxide reduction. This constitutes an essential conclusion taking into consideration that pelletization is a very common agglomeration process used in the iron ore industry.
More precisely, the addition of fluxes results in the alteration of the basicity index of the slag from 4.5 to 0.6, at an addition rate of 35 g Si0 2 and 5 g CaO per 100 g of bauxite residue/lignite mixture (Figure 5). Due to the reduction of slag basicity, the (%) sulfur content in the metallic product increases to approximately 0.7%. It is noted that the sulfur content of a typical cast iron should not exceed 0.04% and this constitutes a crucial parameter of its quality.
A metallographical section of grains of the metallic product as well as the slag samples from the smelting reduction of bauxite residue with lignite (4/1 by weight) with the addition of 35 g Si0 2 and 5 g CaO per 100 g of the mixture, was studied by SEM/EDS (JEOL® JSM-6380LV). Back-scattered electron image of the aforementioned samples, are presented in Figures 8 and 9.
The addition of fluxes in the current work can be characterized as successful, since the melting temperature value measured by the optical pyrometer during the three experiments conducted with the addition of CaO and Si0 2 , fluctuated between 1400-
115
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**l / «£»*>
*g$#*
Ce02Ä%5, WLXfr
CaOt
Jfe
Figure 6. Melting temperature value of the slag sample produced by the mixture bauxite residue /lignite (4/1 by weight)
~ * W
«55» 7 <»4· W»« JCeO-Al^O, / «CoOTÄtp,
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CaO mfa
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Figure 7: Melting temperature values of the slag samples (A,B,Q produced by the addition of: A) Si02-CaO: 20-0 g, B) Si02-CaO: 35-5 g, C) Si02-CaO: 62-38 g per 100 g of the mixture bauxite residue /lignite (4/1 by weight)..
116
Figure 8 shows a typical feature, where the metal matrix and the lamellar structure of iron carbide flakes can be seen. In Figure 9 a typical matrix of an iron ore slag is seen, where the black area corresponds to the slag oxides as determined by the EDS analysis. We have also focused on metal grain 'trapped' in the slag (white area), which is probably ascribed to the intense eddying of the molten bath due to the Foucault current, resulting in non satisfactory phase segregation (metallic product - slag).
reduction smelting with solid reducing agents resulted in the production of a metallic product with an iron content fluctuating between 91-95.8%, which can be characterized as satisfactory considering that a typical cast iron has an approximate analysis of 91-96 % in iron. The addition of solid reducing agents results in a significant increase of the melting point of the bauxite residue, from 1400°C up to 1740°C for a great excess of the reducing agent, due to the formation of the refractory iron carbides. The addition of CaO and Si0 2 reagents as fluxes favored the decrease of the melting point, in agreement with the respective ternary phase diagrams. The type of the solid reducing agent used, proved to be a critical parameter affecting the quality of the metallic product, in terms of the desired sulfur content, the basicity of the slag and the kinetics of the reductive procedure. The results render the future investigation for optimization of the proposed processing method very promising. Apart from optimization of the iron recovery, intensification of the future research is necessary concerning the extraction of impurities, such as S, P and Ti from the metallic product. Acknowledgements The authors would like to acknowledge the financial support of the General Secretariat for Research and Technology of Greece, Programme PAVET 2005, No. 05PAB102. References
Figure 8. Back-scattered electron image of the metallic product from the smelting reduction of Greek bauxite residue with lignite.
1. R. K. Paramguru, P. C. Rath, and V. N. Misra, "Trends in Red Mud Utilization - A Review" Mineral Processing and Extractive Metallurgy, 26 (2005), 1-29. 2. A. Xenidis et al., "Reductive Roasting and Magnetic Separation of Greek Bauxite Residue for its Utilization in Iron Ore Industry" (Paper presented at the Light Metals as held at the 138th TMS Annual Meeting, San Francisco, California, 2009), 63-67. 3. A. Xenidis et al., "Pelletization and Reductive Smelting of Greek Bauxite Residues for Iron Production" (Paper presented at the 3rd International Conference AMIREG: Assessing the Footprint of Resource Utilization and Hazardous Waste Management, Athens, September 2009), 433-439. 4. L. Visnyovszky, "Complex utilization of Hungarian bauxites", Femip. Kut. Intez. Kozlem, 9 (1971). 5. N. I. Eremin, "Complex processing of bauxites" (Paper presented at the Proceedings of the 2nd International Symposium ICSOBA Bauxites, Alumina, Aluminium, Leningrad, 1969).
Figure 9. Back-scattered electron image of the slag phase from the smelting reduction of Greek bauxite residue with lignite. Conclusions Smelting reduction of Greek bauxite residue revealed that it can be used successfully for direct cast iron production. The
6. A. I. Zazubin et al. "Complex processing of red mud", Tr. Inst. Met. Obogashch., Akad. Nauk. Kaz. SSR, (1967), 25.
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
Light Metals 2011 ALUMINA and BAUXITE
Precipitation, Calcination and Properties SESSION CHAIR
Hans-Werner Schimdt Outotec GmbH Oberursel, Germany
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
EFFECT OF TECHNOLOGICAL PARAMETERS ON PSD OF ALUMINUM TRI-HYDROXIDE FROM SEED PRECIPTATION IN SEEDED SODIUM ALUMINATE SOLUTIONS Yusheng Wu, Mingchun Li Yanping Qu School of Materials Science and Engineering, Shenyang University of Technology, Liaoning 110178, P. R. China Keywords: Bayer process, Sodium aluminate, Particle size distribution, Aluminum tri-hydroxide extremely slow growth rate. Agglomeration consumes smaller particles and creates larger particles, representing the additive volume of the smaller particles consumed. Nucleation is the formation of new, minute particles which replenish the particle count of the system. Particles are continuously lost through agglomeration or product removal. Thus for efficient operation, nucleation must provide enough particles to maintain the total particle count, and to provide surface area for continuing growth. Particle size distribution of products mainly depends on all of the above four precipitation steps.
Abstract Periodic attenuation of particles, which interferes seriously with normal alumina production, is a characteristic of the Bayer process. In order to construct a mathematical model of total number of particles and particle size distribution (PSD) which is helpful to predict and control the PSD of aluminum tri-hydroxide for an alumina refinery, the PSD of aluminum tri-hydroxide from the seeded precipitation process with different technological parameters, has been investigated under industrial conditions modeled in the laboratory. The results show a move of PSD to smaller particle sizes with decreasing initial precipitation temperature and increasing molar ratio of Na20/Al203. The volume fraction of particles below 44um first increased and then decreased with increasing initial concentration of Na20. The PSD has the opposite variation tendency with different initial seed amounts, compared with the initial concentration of Na20.
Most alumina plants nowadays have already developed, or are in the process of developing, their own precipitation model for the simulation of the aluminium hydroxide precipitation mechanisms. In order to construct a mathematical model of total number of particles and particle size distribution (PSD) which is helpful to predict and control the PSD of aluminum tri-hydroxide for alumina refinery, the PSD of aluminum tri-hydroxide from a seeded precipitation process under different technological parameters has been investigated under industrial conditions, modeled in the laboratory.
Introduction The most common industrial process for alumina production from bauxite is the Bayer process. Seed precipitation is the key stage, significantly affecting the yield and specification of product [1-3]. Lots of work has been done on the precipitation process for enhancing precipitation ratio in the sodium aluminate liquor, and improving product specifications [5-8]. However, the periodic attenuation of grain size, that mass percent of particle size above 44um and attenuating periodic change from 50% to 90% and 3 to 5 months respectively, inherently exists in the Bayer process in China. During attenuation, the functioning of filters deteriorates and product quality drops. Meanwhile, the electrical consumption and the dust density in the exit of electrofilter increase [9-12]. Although some work has been done to solve the problem, such as using sandy alumina production technology invented by France and investigating the particle size distribution (PSD) of production alumina in some branches of China Aluminum Co., Ltd, the explosive attenuation of alumina particles can't be avoided.
Experimental The main apparatus for seed precipitation study included a bladepaddle mixer tank and a Model LB-801 super constant temperature bath. The supersaturated sodium aluminate solutions, with industrial concentration and initial gibbsite seed for all experiments, were provided by the Shandong branch of China Aluminum Co., Ltd. Precipitation time is 40 h. After adding seed, the precipitation commences by stirring at 260 rpm and the temperature decreased uniformly over the desired ranges. Solution samples (precipitate suspension) were analysed by acidbase titration and complexometric titration methods. The precipitated aluminum tri-hydroxide samples obtained were washed with hot deionised water and dried at 60° C for 24h before observations with Scanning Electron Microscopy (S-3400N). The PSD and average size were measured by laser diffraction (Malvern Mastersizer 2000).
In general, the precipitation of aluminium hydroxide from seeded sodium aluminate solutions (SAS) is comprised of crystal growth, nucleation, agglomeration, and attrition [13-15]. By the mechanism of crystal growth, dissolved alumina hydrate transfers from the aqueous phase to the solid phase, depositing on the surface of existing particles. Thus the growth rate determines the overall liquor/solid mass balance. Growth also impacts the particle population balance because particles will increase in diameter. The crystal growth rate in gibbsite precipitation has been shown to depend on three factors: the supersaturation of sodium aluminate in solution, the temperature, and the solid surface available for reaction. Agglomeration is essential in gibbsite precipitation because it is the only economically feasible method of producing particles with the required size, given the
Results and discussion Effect of technological parameters on particle size distribution In our study, the question was to understand how all the experimental parameters usually considered in the Bayer process influenced PSD. In order to focus our attention on the effect of other parameters, only one parameter was varied in the experiments. The PSD of the products with different parameters are displayed in Fig. 3. It is found that all particle size distribution curves have a similar shape, but show different trends.
121
Figure 3(a) shows that PSD moves to smaller particle sizes with decreasing initial precipitation temperature. Fig. 3(b) shows that 100 0s-
a
PSD moves to coarser particle sizes with decreasing initial molar ratio Na20/Al203. The results we obtained are in agreement with 100
80 60 40
o > 20
0 50 100 Particle size, μπι
50 100 Particle size, μΐΏ
150
50
100
150
50 100 Particle size, μηι
150
Particle size, μπι
150
Fig. 1. Variety of Particle size with (a) temperature; (b) molecule ratio of Na20/Al203; (c) concentration of Na 2 0; (d) fine seed content. three fine seed concentration loading tests have the same degree of supersaturation. That is, have the same driving force for seeded precipitation. Therefore, the secondary nucleation takes place with the fine seed content at 10%, and the agglomeration degree decreases with the fine seed content at 30%.
of those of Yamada [15]. Yamada indicates that the agglomeration of Al(OH)3 particles, for similar supersaturations, is more liable to take place at higher temperatures. He supposes that, at high temperature, the dispersion of the growth species across the particle surface, is faster, and provides a higher probability for the successful cementation of the particles. At the same temperatures, higher initial supersaturations also favour agglomeration, due to the higher concentration of adsorbed growth species. We have worked with fine seeds which are more liable to participate in the agglomeration process. At the same concentration of Na20, as the supersaturation of sodium aluminate solution increases, this favours agglomeration as the initial molar ratio of Na20/Al203 decreases.
Table 1 shows the effects of technological parameters on crystal growth. It is found that when the initial concentration of Na20 was varied from 140g/L to 160g/L. The volume percentage of product particles below 44μπι is increased by 6.05% and then decreases by 3.88%. The surface area is increased by 0.015 m2/g and then decreased by 0.009 m2/g. However, the average volume size is decreased by 3.77μπι and then increased by 1.828μηι. Compared with the initial concentration of Na20, the volume percentage of particles below 44μιη, surface area, and the volume average size have opposite variation tendencies with different initial fine seed content.
Fig. 3(a) shows that the PSD moves first in the direction of coarser and then to finer particle sizes with increasing initial concentration of Na20. Compared with the initial concentration of Na20, the PSD has the opposite tendency with different initial fine seed content. In a seeded precipitation process, the fine particles need more aluminium tri-hydroxide deposition from SAS to transform into coarse particles. However, the seed surface area must respond to the decreasing degree of supersaturation. If not, this is harmful to the agglomeration process. In the tests, the supersaturation of SAS favours agglomeration with Na20 at 140g/L and that can't match agglomeration with Na20 at 160g/L. However, with further increase in concentration, the precipitation solution has a higher viscosity, which increases the chance of collision between fine particles, thereby increasing the agglomeration efficiency and reducing fine particle levels. The sodium aluminate liquors in
The volume percentage of particles below 44um and the surface area, decreased with increasing initial temperature. This varied from 32.17% to 12.41% and 0.148 m2/g to 0.107 m2/g respectively. That is, higher precipitation temperatures favour particle size coarsening. For example, the average volume increased by 11.51 lpm with initial precipitation temperature increase from 60°C to 75°C. The product characteristics therefore show the opposite variation tendency with different o^ compared with initial precipitation temperature.
122
140g/L
12.03
78.171
surface area /m2/g 0.104
150 g/L
18.08
75.401
0.119
160 g/L
14.20
77.229
0.110
1.35
13.65
78.047
0.111
1.45
15.72
76.223
0.115
Technological parameter Nk
Table 1. Characteristics of product volume fraction volume average of<45um/% size /urn
ak
1.55
18.27
75.740
0.120
60°C
32.17
66.890
0.148
65°C
20.92
74.453
0.120
70°C
15.54
77.016
0.110
75°C
12.41
78.401
0.107
10%
23.24
64.683
0.111
20%
18.38
74.611
0.115
30%
29.65
62.130
0.120
Fine
Note: Nk - concentration of Na20; ak - molecule ratio of Na20/Al203; T- temperature; Fine - fine seed content. 8.1. Seyssiecq, S. Veesler, D. Mangin, J. P. Klein, and R. Boistelle, "Modeling gibbsite agglomeration in a constant supersaturation crystallizer," Chemical Engineering Science, 55(2000) , 5565-5578. 9. Y. S. WU, D. Zhang, M. C. Li, S. W. Bi, and Y. H. YANG, "Periodical attenuation of Al(OH)3 particles from seed precipitation in seeded sodium aluminate solution," Transactions of Nonferrous Metals Society of China, 20(2010), 528-532 10. C. B. Zhang, and P. Zhao, "Studied on polarizing particle size of aluminate hydroxide in Bayer process," Light Metals, (1999), 17-19. (In Chinese) 11. G. Zhang, and J. N. Yang, "Studied on particle size fluctuating of aluminate hydroxide in Bayer process," (2002), 9-12. (In Chinese) 12. B. Garner, B. Cristol, and A. Soirat, "Precipitation particle size control," Light Metals, (1999), 71-76. 13. B.K. Satapathy, and T. Padhi, "Determination of grain-size distribution of sandy alumina using electron sensing zone method," Light Metals, (1990), 185-191. 14. T. K. Hunter, G. M. Moody, and S. E. Sankey, "Advances with chemical additives for the alumina industry," Light Metals, ( 1991 ), 159-162. 15. K. Yamada, "Nucleation and agglomeration during crystallization of aluminium trihydroxide in sodium aluminate solution," Light Metals, 32(1980), 720-726.
Conclusions A systematic study has been made on the PSD of aluminum trihydroxide from a seeded precipitation process with different parameters. The PSD moved to smaller particle sizes with decreasing initial precipitation temperature, and increasing molar ratio of Na20/Al203 . The volume fraction of particles below 44um first increased, and then decreased with increasing initial concentration of Na20. The PSD has the opposite tendency with different initial seed amount compared with the initial concentration of Na20. Acknowledgments The authors are grateful to appreciate the financial support of National Natural Science Foundation of China (NO. 50804031) and the educational department of Liaoning Province of China (NO.L2010395). References 1. R. Chester, F. Jones, M. Loan, A. Oliveira, and W. R. Richmond, "The dissolution behavior of titanium oxide phases in synthetic Bayer liquors at 90 °C," Hydrometallurgy, 96(20093), 215-222. 2. H. Watling, J. Loh, and H. Gatter, "Gibbsite crystallization inhibition 1. Effects of sodium gluconate on nucleation, agglomeration and growth," Hydrometallurgy, 55(2000), 275-288. 3. F. Farhadi, and M. B. Babaheidary, "Mechanism and estimation of Al(OH)3 crystal growth," Journal of Crystal Growth, 234(2002), 721-730. 4. N. Brown, "A quantitative study of new crystal formation in seeded caustic aluminate solution," Journal of Crystal Growth, 29(1975), 309-315. 5. Z. Wang, S. W. Bi, Y. H. Yang, and Z. F. Yuan, "Evolution of particle size and strength of hydragillite from carbonization in seeded sodium aluminate liquors," Journal of Crystal Growth, 274(2005), 218-225. 6. A. M. Paulaime, I. Seyssiecq, and S. Veesler, "The influence of organic additives on the crystallization and agglomeration of gibbsite," Powder Technology, 130(2003), 345-351 7. H. LI, J. Addai-Mensah, J. C Thomas, and A. R. Gerson, "The crystallization mechanism of Al(OH)3 from sodium aluminate solutions," Journal of Crystal Growth, 279(2005), 508-520.
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Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
Methods to Reduce Operating Costs in Circulating Fluidized Bed Calcination Cornells Klett, Bernd Reeb, Michael Missalla, Hans-Werner Schmidt Outotec GmbH, Ludwig-Erhard-Str. 21, 61440 Oberursel, Germany Keywords: Circulating Fluidized Bed, Calcination, Energy Efficiency, Product Quality, Particle Breakage, Dry Hydrate the free moisture of the hydrate, with typical values of 6 to 8% moisture, is evaporated and the hydrate is dried.
Abstract The calcination of gibbsite or hydrate to alumina is one of the most energy consuming process steps in every alumina refinery based on the Bayer process. Approximately 30% of the thermal energy input is used for the calcination process step.
The dried hydrate is then separated from the gas in an electrostatic precipitator and fed to the second preheating stage, where it is mixed with the 950°C hot off-gas of the furnace. Here a significant step of the reaction is already taking place with approx. two thirds of the hydrate water released. The solids are separated again from the gas and transported to the CFB furnace.
Over the years new technologies such as Circulating Fluidized Bed (CFB) Calciners by Outotec (formerly Lurgi) have reduced the energy consumption for the calcination step significantly (see also recent Energy Efficiency Award, given in 2010 by the German Energy Agency). The CFB technology was introduced as early as 1961. Up till then rotary kilns were the standard technology for the calcination of alumina. This innovation reduced the consumption of fuel by up to 30%. Since then, CFB calciners have been constantly improved, and methods have been developed to reduce fuel consumption even further.
Finally under the high solids density and homogeneous mixing and temperature field in the CFB furnace, the precalcined material is calcined to the specified alumina quality. The good mixing and high solids density, with a sufficient residence time of several minutes, allows very homogeneous calcination of the material and the optimum product quality, with very low alpha alumina and no residual gibbsite, as well as low particle breakage.
Not all, but some, of the methods to reduce fuel consumption have resulted in increasing process complexity and operator and maintenance demand. However, a number of measures have also been introduced to mitigate the negative effects of the increased process complexity and to even improve on operability and maintainability. In this paper the different options and methods introduced for reduction of fuel consumption, but also to increase of operability and maintainability are compared and evaluated with regard to their effects on installation and operating costs.
Cooüng Sfege U
Cooling Wafer
Introduction Calcination of hydrate to alumina is the last step in the production of smelter grade alumina from bauxite in an alumina refinery. The energy consumption for the production of alumina varies between 6.5 and 18 GJ/t alumina for the whole refinery. Depending on the technology, approx. 30% of this is consumed in the calcination step.
Additional Atr Secondary Air Primary Air Cooling Wafer
Figure 1: CFB Calciner Flowsheet of recently installed calcination plants
Until stationary calciners, such as flash and CFB calciners were introduced, rotary kilns were used. Rotary kilns have a thermal energy consumption of around 4.5 to 5.5 GJ/t alumina based on lower calorific heating value. With the introduction of stationary calciners and the first CFB calciner by Outotec (at that time known as Lurgi) in 1961, the consumption dropped instantly to approx. 3.3 GJ/t. Thus operating costs dropped significantly, which gave a huge incentive to invest in this new technology.
The final product is discharged from the CFB calcining stage and then cooled down in two direct gas solids contact cooling stages and one indirect cooling stage, which is the fluid bed cooler. In the first two cooling stages the hot alumina is mixed with air and the air heated up as the alumina is consequently cooled down. In each stage the alumina is separated from the heated air by means of a gas cyclone and transported further to the next stage, whereas the air is flowing in direction of the CFB furnace to serve as heated combustion air.
Since then the technology has matured and improved with current energy consumptions of 2.79 GJ/t for the state of the art flowsheet as shown infigure1.
At last in the fluid bed cooler all residual heat, which is not easily recovered in the calcining process, is extracted from the alumina to enable discharge with safe alumina temperatures for handling. The fluid bed cooler consist of several sequential compartments, where primary fluidizing air is preheated for the nozzle grate of the CFB furnace and finally cooling water lowers the alumina temperature to safe handling levels.
The CFB calcination process utilizes maximum heat recovery by having air and furnace off-gas flowing in a countercurrent scheme to the hydrate feed or alumina product, respectively. In the current flowsheet the hydrate is preheated in two stages with the off-gas of the CFB furnace. In the first preheating stage
125
thus defining the general temperature level in the cooling stages and consequently has an influence as described on the performance of the cooling cyclones.
Key Consideration for Fuel Efficient Calciner Design Essential for the success of the CFB calcination flowsheet for optimized performance is not only the design of the Fluid Beds (CFB and Fluid Bed Cooler [1]), but also the cyclones in the system. In the presented calcination process heat recovery is done by counter current solid gas heat transfer. The means of separating solids and gas are the cyclones. Any amount of solids in the cyclones off-gas is counteracting the counter current heat recovery scheme. Further, the fines in the off-gas of one cyclone will be fed to the cyclone in the upstream recovery stage and subsequently reduce the separation efficiency as well.
The process calculation results show that in fact with a reduced furnace temperature, the fuel consumption of the calciner is less than with a higher furnace temperature. This can then only be compensated by more cooling stages. Not surprisingly reduced cyclone separation efficiencies also lead to increased fuel consumption. Measures for Further Reduced Fuel Consumption
The cyclone design is not only critical for best separation performance and thus for the process efficiency, but also for the amount of fines in the system. The fines are both fed to the process with the hydrate feed, and also generated in the process by particle breakage. A major part of the particle breakage is generated in the cyclones. This needs to be taken into account in the process and specifically cyclone design [2]. Cyclone design, e.g. geometry, velocity profile, etc. needs to be carefully optimized to have max. separation efficiency not only for one cyclone, but also for the overall process as breakage and fines influences all of the cyclones at the same time. This phenomenon is not only known for alumina calcination [3]
In the above sections the basic flowsheet of Outotec's CFB calcination technology is described and assessed. As the technology is now well proven and lot's of research and development has been undertaken in recent years, there are several developments and options available in this flowsheet, which further improve energy consumption, availability and operability. Which option is best suitable depends on client's requirements, fuel prices and project structure such as investment budget and predicted plant life time and life cycle. In the following sections, the different options available for reduced fuel consumption and improved plant performance are described and the impact on investment and operating costs are discussed. The discussions further distinguish between options for greenfield plants or revamps and upgrades of existing units.
■ 100 m90
Cyclone Separation Efficinecyin%
Di
Based on the above basic flowsheet several further options and variations of the flowsheet have been developed to reduce fuel consumption and hence reduce operating costs. Naturally these options will have an impact on the investment costs as well. There are several options available to reduce mei consumption of the calcination process. The available measures are targeting the optimization of the heat recovery and energy utilization within the calcination process inside the battery limits of the calciner. The two main options possible are discussed in this publication.
2 Cooling Cyclones 2 Cooling Cyclones 3CoolingCyclones 3CoolingCyclones with 1100°C with 950°C Furnace with 1100eC with 950°C Furnace Furnace Temperture Temperture Furnace Temperture Temperture
The available options are: • Hydrate bypass, where preheated material is bypassed around the CFB furnace and calcined with heat from CFB discharge alumina • Pre-drying of hydrate feed with waste heat from fluid bed cooler
Figure 2: Fuel Consumption depending on Cyclone Efficiency and Calcination Temperature Besides velocity and solids concentration in the cyclone inlet, the operating temperature also has a strong influence on the performance [4]. With increase of gas temperature, the gas viscosity increases and the drag on the particles inside the cyclones increases. The higher wall friction also causes reduced tangential velocities for the centrifugal field. As a result the separation efficiency decreases for a cyclone with constant inlet velocity. In order to compensate the inlet velocity needs to be increased, which will then increase breakage and generation of fines. As described above this again will have an impact on the overall process performance.
In figure 3 the first option the hydrate bypass is shown. This option has been developed in the 1990s by Outotec [6] and [7]. In the bypass flow sheet a part of the dried hydrate from the first preheating stage is taken and directed to the discharge of the CFB furnace. The bypassed solids are mixed with the hot alumina from the CFB discharge and allowed to react in a so-called mixing pot, which will provide sufficient residence time for the reaction. With this method, heat from the CFB alumina discharge is utilized for the calcination reaction of the bypassed hydrate, rather than for heating up of combustion air. Hence the heat is utilized more efficiently and secondly the amount of hydrate or alumina respectively, which needs to be heated to the reaction temperature
In figure 2 the influence of cyclone performance on the fuel consumption of a stationary calciner is shown. For comparison two calciners one with 2 and one with 3 cooling cyclones and two different calcination temperatures are assessed. CFB calciners are typically built with two cooling cyclones and operated at 950°C in the furnace. Stationary calciners are often equipped with 3 cooling cyclones and operated at 1100°C [5]. The furnace temperature is
126
P< «ïïmïng
ïvige !
Secondary Air Primary Air CBnaWst*r ■«·■*■
Figure 3: Flowsheet with the Hydrate Bypass implemented. of the CFB of more than 950°C is reduced. This reduces the required amount of fuel significantly. The Bypass was developed in the 1990s and has already been installed in more than 6 plants worldwide since then.
Figure 4: CircoCal flowsheet with bypass, hydrate dryer and 3rc preheating stage As a result, less heat is utilized in the preheating stage 1fromthe calciner off-gas. Consequently the off-gas temperature would rise. To utilize this heat, a third preheating stage can be installed, which then recovers the heat. Finally this means that less heat is rejected with the cooling water to atmosphere and there is also reduced heat loss with the calciner off-gas.
With the hydrate bypass in operation, the temperature profile in the cooling stages changes, as the solids enter the first cooling stage with a lower temperature than before. Furthermore, due to the reduced fuel consumption, less air is required and also less combustion off-gases are produced. This allows the design of the preheating stages, CFB stage and cooling stages to be smaller, with less construction material for the same nameplate capacity.
The flowsheet for this process is shown in figure 4 and forms the CircoCal process by Outotec. Advantages are that all extra equipment does not need to be refractory lined. Also there is only one more cyclone installed on the preheating side and no further cooling cyclone is needed to increase heat recovery. Cyclones are one of the major sources for particle breakage [2] and the most sensitive is the alumina on the product side (cf. section above). Therefore the impact of product quality is minimal for this approach, compared to the classical approach to expand the counter current heat recovery scheme.
To further reduce the fuel consumption of the calcination process, a closer analysis of the process is required. A thermodynamic heat and mass balance analysis of the calcination flowsheet shows that CFB with its fixed and homogenous temperature profile is a "pinch point" for the heat flow and exchange within the process. This means that with the existing counter current heat exchange and recovery scheme it is very difficult to transport heat from the back end of the process through the CFB stage into the front end of the process.
Alternatively the dried hydrate can also be transported to a dry hydrate storage silo instead of feeding to the calciner if dry hydrate is desired as an extra product from the refinery. However this option is not investigated further in this publication.
One way could be to install more preheating and more cooling stages to increase heat recovery at both ends. However this would mean more investment costs for very big and refractory lined vessels.
3.00
Another possibility is to collect the heat in one end and transport it around the "pinch point" to the other end to utilize it there. A method for this approach has been developed by Outotec in the past [8].
.2 O 2.95 i £> 2.90
Heat can be recovered from the fluid bed cooler by means of a heat carrier, such as water under pressure or thermo oil. The heat carrier is then pumped to the front end of the calciner. There a part of the moist hydrate feed is taken and fed into a fluid bed hydrate dryer (for details see the section below). In the hydrate dryer the heat from the heat carrier is used to dry the hydrate before it is fed to the preheating stage I. Then the heat carrier is pumped back to the fluid bed cooler. Hence the waste heat from the cooler is used to evaporate the free surface moisture of the hydrate.
Base Case
Hydrate Bypass
CircoCal™
Figure 6: Specific energy consumption of different flowsheet variations
127
Base Case
Hydrate Bypass
CircoCal™
Figure 7: Impact on investment costs Figure 8: Schematic of fluid bed hydrate dryer design
Evaluation of Measures for Reduced Fuel Consumptions
costs / benefit ratio can possibly be further improved if carbon projects are being included.
In the previous section, different options are presented to further reduce the fuel consumption of CFB calciners. The base calcination flowsheet has a specific energy consumption of-2.79 GJ/t with a feed hydrate moisture of around 7%. The results of the fuel consumption comparison are shown infigure6.
Drying Hydrate in a Fluidized Bed Key equipment of the CircoCal flowsheet is the hydrate dryer. There are a lot of refineries, which also produce dried hydrate. However in most cases direct heat transfer with hot off-gas from combustion is used to dry the hydrate. Here the waste heat from the fluid bed cooler is used. As the heat carrier, water or oil is used in a closed circuit (cf. figure 4). To transfer the heat to the wet hydrate from the hydrate filtration, indirect heat transfer via heat transfer surface is required.
In the comparison the hydrate bypass reduces the specific energy consumption by - 3 % and the CircoCal flowsheet has the lowest specific fuel consumption of all the options discussed. The reduction compared to the base case is about 6%. For all options, the reduction in fuel consumption can be seen as direct reductions in operating costs. There will be an increase in labor and maintenance costs for the extra equipment in the plant. However these can be seen as marginal. None of the options include maintenance intensive equipment such as extra rotating equipment or huge refractory lined vessels. The extra fluidizing air requirements can be covered with the existing blower concepts and Outotec's technology for solids mass flow control does not require any moving parts inside the solids streams.
Outotec has developed for this a hydrate dryer based on a bubbling fluidized bed similar to the fluid bed cooler. In the fluid bed dryer, heat transfer bundles are submerged in the fluidized solids. A schematic of the fluid bed dryer is shown in figure 8. In this fluid bed, large transfer areas can be installed in very small vessel volumes. For optimum bundle design the same design criteria are applied as for the fluid bed cooler with same success as reported in 2007 [1]. To ensure easy maintenance the bundles can be removed from the side and cleaned.
Investment costs of the plant will be impacted as presented in figure 7. It shows that an improvement in energy efficiency does not necessarily mean also an increase in investment costs.
The implementation philosophy of the dryer into the overall calcinerflowsheetis shown infigure9.
Figure 7 shows the changes of plant weights normalized to the base case. Again here the boundary condition is that all plants have the same nameplate capacity. The changes in weight here are considering vessel plate work, refractory and steel structure. Instrumentation, piping and other equipment such as blowers, screws, etc. do not change significantly.
Hydrate from Filtration
Preheating Stage I Offgas to ESP
I
Dryer Off gas to ESP
Due to the changed temperature profile in the hydrate bypass option the vessels can be reduced significantly in size as described above. This then leads to reduced overall plant weight and does in fact reduce investment cost for the plant when nameplate capacity is maintained asfigure7 demonstrates.
A
Heat Carrier from Fluid Bed Cooler
The reduced fuel consumption of the CircoCal flow sheet is offset by a potential increase of investment cost of about 15% if based on weight. This is the highest increase of investment of all options. On the other hand the savings in fuel can give a return of investment in 2 years based on an oil price of 80US$/barrel. The
Fluidizing Air for Dryer
Fluidizing Air for Dryer -►
Hot Offgas from Preheating Stage II
Figure 9: Schematic of hydrate dryer implementation in preheating stage I
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The base case includes 5 main control loops, which are: • Hydrate feed • Furnace temperature • Furnace inventory • Fluid bed cooler level • Emergency water sprays
The wet hydrate is fed into the calciner feed bin. The feed bin has two separate bottom discharges, which are connected to screw feeders. One of the screw feeders is feeding wet hydrate directly into the venturi of the preheating stage I. The second screw feeder is feeding into the fluid bed hydrate dryer. The dryer screw feeder is controlled by a temperature measurement inside the dryer fluid bed and ensures that wet hydrate is only fed into the dryer, when the temperature is above 100°C. This guarantees that the fluid bed is always dry and easily fluidizable. Secondly, the speed of the screw feeder is controlled to maintain stable dryer temperature and ensures that all the heat available from the heat carrier is used.
All these loops are tied together in a governing overall control scheme with smart feed forward and troubleshooting mechanisms. In the recent years the improvements of process control and automation have led to very easy operation and fast ramp up performance [10]. As result there is little difference anymore between the different calcination technologies available and the limitations for fast process changes, ramp ups and shut downs are only in the speed in which process equilibrium can be achieved. Figure 10 demonstrates the comparability in ramp up performance for different calcination technologies.
Furthermore the inventory of the fluid bed dryer needs to be kept stable to keep the bundles submerged during normal operation. This is achieved by means of a lifting seal pot in the same way as it is used to maintain the level of solids in the fluid bed cooler. The solids level in the fluid bed dryer is detected by means of a differential pressure measurement across the fluid bed height. The differential pressure is then controlling the fluidizing air to the fluid bed dryer seal pot, which discharges the dry hydrate directly into the venturi of the preheating stage I.
In 2008 the influences on availability and consequences from unplanned outages of CFB calciners have been discussed [11]. Although not the major reason for unplanned outages, the consequences of refractory failures are very time consuming and costly. With the latest installations of CFB calciners the impact of plant trips and subsequent cooling and reheating cycles on the refractory could be explored. The stable operation, as reported in [2] and [12], proved to have very positive impacts on refractory life, thus not only decreasing operating costs by reduction of fuel consumption, but also reducing maintenance costs and loss of production by increasing reliability and plant integrity.
Finally in the preheating stage I both hydrate streams (wet and dry) are mixed with the hot off-gas of the preheating stage II. With the mixture of wet and dry hydrate, the utilization of the available heat in the heat carrier and thus in the fluid bed cooler is maximized. The vapor from the hydrate dryer is then added to the preheating stage I off-gas and leaves the calciner together with the combustion off-gas through the ESP and stack. Alternatively the vapor can also used to recover water if required. The off-gas from the hydrate dryer will have very high amounts of water with small amounts of air.
The above mechanisms and approaches will also be applied to the options described above, to further reduce fuel consumption on CFB calcination. For the bypass option only one more control loop is required which is the mixing pot temperature. For CircoCal the following control loops are needed: • Hydrate feed • Furnace temperature • Furnace inventory • Fluid bed cooler level • Emergency water sprays • Mixing pot temperature • Hydrate Dryer temperature • Hydrate Dryer level
Measures for Improved Operating Stability and Availability Besides the energy optimized flow sheet and plant layout, stable operation of the process is critical for the achievement of low operating costs. It is not only important for the optimized performance with minimum consumption of utilities, but also for maximum lifetime of plant equipment and low maintenance requirements. For example the typical it experience has been that stable process operation and low frequencies of unplanned plant downtime is one key for long refractory life.
With the base case, a thermal efficiency of -88% is achieved. In order to get closer to the theoretical fuel consumption of 2.45 GJ/t more initiatives need to be taken. Therefore the complexity of the plant will likely also increase. For the operator, the complexity is only visible in the amount of information he has to process and the attention he has to pay to each individual unit.
A further advantage for the refractory lifetime of the CFB system is the fact that it operates at significantly lower calciner reactor temperatures, compared to other stationary calciners while achieving similar product quality [9]. In the recent past, Outotec hasfrequentlyreported on the progress in plant operation and stability. With new technologies available, and with the experience of more than 60 plants built, the controls and automation has been developed to an extent that constant operation is achieved under nearly all conditions. This leads to a minimized number of unplanned plant downtimes as the available control philosophies now prevent plant drop outs by automated reaction of the system. Further the controls now assist in automated start-ups, change of plant loads and shut downs [9].
In this case it can be argued that the complexity grows significantly from the base case from the Bypass options to the CircoCal process, when one looks at the number of control loops to supervise. However the extra control loops are be integrated in the overall control philosophy to achieve same temperature stability as in the base case. The relationships between temperature changes and solids mass flows to different areas in the plant are well known, which makes the controls easily manageable.
129
consumption, but also on product quality such as fines in the final product. E.g. precipitation and calcination need to be looked at together in order to optimize investment and operating costs to achieve the required product specification. Acknowledgement Outotec herewith thanks their customers Alunorte S.A. and AOS Stade GmbH for their support and cooperation. The support from Alunorte led to the recently given energy efficiency award 2010 from the German Energy Agency.
—CFB Calciner 1 —CFB Calciner 2 - -GS Calciner
References
time, min Figure 10: Ramp Up of stationary calciners. Data for GS Calciner performance adapted from [5]
[I] M.Missalla, C.Klett, R.Bligh, "Design Developments for Fast Ramp-up Easy Operation of New Large Calciners" TMS Light Metals, (2007)
Therefore it can be suggested that the increase of complexity lies only in the increased demand for instrumentation and control maintenance, which can be handled via the general refinery maintenance scheme and relieves the operator.
[2] C. Klett, M. Missalla, R. Bligh, "Improvement of Product Quality in Circulating Fluidized Bed Calcination", TMS Light Metals, (2010)
Retrofits to existing CFB Calciners
[3] J. Reppenhagen, A. Schetzschen, J. Werther, "Find the optimum cyclone size with respect to fines in pneumatic conveying systems", Powder Technology 112, (2000)
The measures presented for reduced operating costs and increased availability, can also be applied to existing units. The improved control schemes are implemented easily with only limited costs involved, depending on the existing instrumentation infrastructure.
[4] M. Missalla, "Calculation Method for Highly Loaded Cyclones", Ad. libri Hamburg, Dissertation Technische Universität Braunschweig (2009)
Implementing a hydrate bypass or turning an existing unit into CircoCal operation is also feasible, but will involve installation of extra hardware. However this would not only reduce specific energy consumption and improve operability, etc. it also leads to an increase of plant production capacity. This is in contrast to the cases assessed above, where the plant capacity is deliberately kept constant. In a retrofit situation, the blower capacities and existing vessel and cyclone geometries are kept unchanged. This means that with changed temperature profiles, the throughput can be increased to maintain the velocity profile in the plant before and after the modifications. This also means that particle breakage can be kept constant or even improved depending on the existing unit.
[5] S. Wind, B.E. Raahauge, "Energy Efficiency in Gas Suspension Calciners (GSC)", TMS Light Metals, (2009) [6] H.-W. Schmidt, W. Stockhausen, A. N. Silberberg, "Alumina calcination with the advanced circulating fluid bed technology", TMS Light Metals (1996) [7] D.J. Brodie, H.W. Schmidt, "Custom Design Fluid Bed Calciner for Nabalco Pty Ltd." Proceeding of 5th International Alumina Quality Workshop, (1999) [8] H.-W. Schmidt, W. Stockhausen, "Latest developments in circulating fluid bed calcination based on operating experience of large calciners", Proceeding of 6th International Alumina Quality Workshop, (2002)
Conclusions In the above sections different options have been presented as to how to make CFB calciners more fuel efficient, how to reduce the operating costs and the impact on investment and process complexity and control.
[9] A. Saatci, H.W. Schmidt, W. Stockhausen, M. Ströder, P. Sturm, „Attrition Behaviour of Laboratory calcined Alumina from various Hydrates and ist influence on SG Alumina Quality and Calcination Design", TMS Light Metals, pp 81-86 (2004)
With these different options, thermal efficiencies of up to 92% can be achieved and fuel consumptions as low as 2.65 GJ/t. The risks involved for particle breakage, product quality and process complexity is minimal and can easily be mitigated.
[10] M. Missalla, J. Jarzembowski, R. Bligh, H.-W. Schmidt, "Increased availability and optimization of calciner performance due to automation", TMS Light Metals , (2009)
The options presented are not only of interest for new installations, but also for older ones and even viable for units of the very first generation of CFBs.
[II] P. Hiltunen, R. Bligh, C. Klett, M. Missalla, H.-W. Schmidt, "How to achieve high availability with large calciners and avoid unforeseen downtime", TMS Light Metals, (2008)
However for some options, the integration of the calciner in the overall refinery energy balance can give extra benefits. This needs to be the focus of future investigation on how to reduce overall operating costs in the production of smelter grade alumina from bauxite. These investigations shall not only focus on fuel
[12] M. Missalla, H.-W. Schmidt, J. R. A. Filhio, R. Wischnewski, "Significant Improvement of Energy Efficiency at Alunorte's Calcination Facility", TMS Light Metals, (2011) to be published
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Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
PRESSURE CALCINATION REVISITED F. S. Williams, CMISCORP, Longmont, CO 80503 and C. Misra, AluminaTech, Pittsburgh, PA 15239 Bayer Process, Calcination, Pressure Alcoa Pressure Calcination Process
Abstract tth
The pressure calcination process developed at Alcoa Laboratories addresses both the questions of energy use and fines generation. It enables a substantial recovery of calcination energy together with the production of a stronger alumina.
Twenty-five years ago at the TMS 100 anniversary meeting of the Hall-Heroult Process, two Alcoa scientists, S. W. Sucech and the co-author of this paper, C. Misra presented a paper on an improvement to the alumina calcination process - pressure calcination.
In this process, alumina hydrate is calcined in two stages. In the first stage alumina hydrate is heated indirectly to 500°C in a decomposer vessel under a steam pressure of about 8 arm. About 85% of the combined water is released and is recovered as process steam. There is also a partial transformation of gibbsite to boehmite during decomposition involving alteration of the crystal morphology. The product from the decomposer has about 5% LOI and is calcined by direct heating to 750-850°C to obtain smelting grade alumina. The alumina product has high attrition resistance.
The improved process offered an opportunity for a net energy reduction in fluid flash calciners of 1.6 GJ/ton alumina (and the subsequent green house gas reductions). This could be retrofitted into existing fluid flash calciners and produce an alumina meeting smelting requirements with the added advantage of high attrition resistance and thus low dust generation. The question is: Why hasn't this improvement been incorporated into today's alumina plants?
Heat Effects in Hydrate Decomposition
Technical data in the previous paper will be reviewed and updated in the present paper.
The process of thermal decomposition of alumina hydrate can be studied from X-ray diffraction, DTA and TGA data [27]. The TGA plot for a typical American Bayer hydrate is shown in Figure 1 for atmospheric pressure calcination.
Introduction Energy requirements for alumina calcination have decreased from 4.5 GJ/t A1203 to less than 3.0 GJ/t A1203 following the replacement of rotary kilns by stationary calciners. Since the commercial introduction fluid flash calciners in the 1960s, there have been numerous papers on the development and improvement in design as well as increases in installed throughput capacities of this class of calcination equipment [3, 5, 6, 7, 8, 9, 12, 15, 18, 19, 20, 28, 30]. One major drawback of the new calciners was that the type of alumina trihydrate particles being produced in alumina refineries using the new calcination technology tended to break down and create a significant amount of fine dust during calcination and that dust, in addition to being an environmental problem, led to a range of feeding problems for alumina smelters using the flash calcined aluminas [1].
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This problem of dust creation has been mitigated over the years by the alumina refineries changing precipitation practices to create "sandy" "mosaic" crystal habit alumina trihydrate particles [22, 23, 24, 26, 31], reduction in internal velocities and sharpness of bends in calciner design [8, 28], and methods of measuring or characterizing the potential attrition resistance of alumina trihydrate particles [2, 4, 10,13, 14, 16,17, 21,22, 23, 28].
Figure 1, TGA Analysis of Hydrate As can be seen, more than 85% of the water is released below about 520°C. Some boehmite is formed due to hydrothermal conversion within the crystals. This is shown by the exothermicity at 270°C and endothermicity at 520°C in the DTA plot (Figure 2).
Today, the majority of the calcined alumina produced has been calcined in some form of fluid flash calciners as opposed to rotary kilns. The problem of alumina fines, while reduced, remains a continuing problem for handling and smelter cell feeding systems.
Heat requirements have been estimated from DSC measurements andare: For water removed between 200-350°C = 3660 J/g H 2 0 For water removed between 350-1000°C = 1686 J/g H 2 0 Overall = 1128 J/g hydrate
In 1986, an Alcoa paper at the TMS meeting introduced the concept of pressure calcination [25]. The present paper will reintroduce the data presented in that paper.
Similar tests carried out under eight atmospheres (120 psig) steam pressure show, as expected, larger amounts of boehmite formation and, unexpectedly, an 11% decrease in heat requirements to reach thefinalLOI value of about 1%.
131
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Figure 2. DTA Analysis of Hydrate Characteristics of Pressure Decomposer Product Bench scale decomposition experiments were carried out in a modified autoclave system with provision to inject dry alumina hydrate. In a typical run the autoclave was brought to pressure and temperature by starting with a small amount of water. Hydrate was then charged from the injection vessel by pressurized gas. A venting arrangement maintained constant steam pressure in the autoclave during the test. Range of conditions studied were: Pressure: 4-30 atmospheres Temperature: 250-650°C Residence Time: 30-120 minutes At the end of the test the autoclave was vented fully and cooled. The product was removed and examined for properties.
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Figure 4. Surface Area of Pressure calcined Aluminas 36
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Figures 3, 4, and 5 show boehmite content, surface area, and LOI values of decomposition products for various test conditions. Table 1 shows particle size analyses (by sieving) of the feed hydrate and product after decomposition.
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The decomposed products were then subjected to the second stage atmospheric calcination in an electric furnace for one hour at the desired temperature. The products were again analyzed for LOI, surface area, attrition resistance and crystalline phase. Morphological features were examined by SEM and also by sectioning the alumina particles.
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132
Table I. Sieve Analysis( ' Comparing Hydrate Particle Size distribution to Pressure Calcined Alumina(2) Particle Size Distribution Sieve Size +100 +150 +170 +200 +270 +325
Hydrate 1 Hydrate Alumina 9 32 44 61 85 93
Hydrate 2 Hydrate Alumina
9 35 50 67 89 96
17 54 76 89 98 99
18 59 77 90 99 100
Hydrate 3 Hydrate Alumina 5 31 53 74 95 98
6 34 59 80 98 100
Notes: (1) Values represent cumulative weight percent retained on indicated sieve size. (2) Average of all aluminas pressure calcined at 8 atmosphere 1st stage and 2nd stage at 1 atmosphere and 850°C. Results of these tests, which have a direct bearing on the development of a practical pressure calcination process, are summarized as follows: a. More than 50% transformation of gibbsite to boehmite occurs at 120 psig pressure and 500°C for a residence time of 60 minutes. A part of the boehmite subsequently converts to Y-A1203. b. This transformation affects the development of surface area. There is no rapid rise in surface area; the surface area increases to about 70 m2/g under the above conditions. This surface area value remains practically unchanged after second stage calcination at 750-850°C. c. The above decomposer conditions result in a product having a LOI of about 5%. A final LOI of <1% (corresponding to smelting grade alumina) can be attained by carrying out the second stage calcination at the relatively low temperature of 750-850°C. The final product consists largely of γ- A1203 with the complete absence of a - A1203. d. The formation of boehmite has considerable impact on the internal structure of the alumina and consequently on the attrition behavior. Figures 6 and 7 show SEM pictures of internal structures of atmospheric and pressure calcined (after second stage) aluminas. The absence of cracks and fractures in the pressure calcined material is remarkable.
Figure 7. Section of Pressure Calcined Alumina Attrition test results confirm the above observations. First, there is little particle breakdown during the pressure decomposition stage. Results actually show a slight coarsening effect. Second, the product from the second stage has very high attrition resistance as measured by the usual modified Forsythe-Hertwig test. Some results are given in Table Π. Table II. Sieve Analysis(1) and Attrition Indices, of Bench Pressure Calcined Aluminas(2) Umos. 8 atmospherei$ Hydrate 1 Atmos Sieve 300°C 400°C 500°C Mesh 12 17 19 17 17 =100 48 56 60 55 54 +150 64 74 74 78 76 +170 78 87 91 90 89 +200 90 98 96 99 99 +270 99 93 98 100 100 +325 Notes: (1) Numbers represent cumulative weight percent retained on indicated sieve size. (2) Second Stage Calcination at 850°C for 1 hour. (3) Atmospheric calcination performed in laboratory flash calciner for 30 minutes at 850°C.
Figure 6. Section of Atmospherically Calcined Alumina
Results also suggest that the residence time of the hydrate in the decomposer, and hence the extent of transformation to boehmite,
133
strongly influences later attrition behavior. Higher attrition resistance corresponded with longer residence time. Pilot Scale Continuous Decomposer
In general, pilot plant findings fully confirmed bench scale results with respect to product properties. Some results are displayed in Table III. Heat transfer results showed the gas side coefficient to be limiting with overall heat transfer coefficients in the range of 8-10BTU/ft2/hr/°F.
A pilot continuous decomposer was designed and operated to test some concepts of indirect heating of the hydrate and compare product quality results from batch operated bench tests with continuous decomposition tests. The aim was to obtain information useful for the final design and construction of a fullscale pressure calcination system. This must include geometry to incorporate a large amount of surface area for indirect heat transfer from hot combustion gases, practical feed and discharge schemes for the decomposer, and recovery of clean steam.
Table III. Sieve Analysis(1) and Attrition Indices of Pilot Pressure Calcined Aluminas Sieve Testi Test 2 Test 3 Mesh Feed Prod. Feed Prod. Feed Prod. +100 3 3 2 3 8 9 +150 27 28 22 26 36 39 +170 42 44 38 42 50 52 +200 73 75 71 74 73 76 +270 94 96 94 95 93 97 +325 97 98 97 98 96 99 A. I. 3 2 4 3 3 4 Notes: ( 1 ) Numbers represent cumulativeweight percent retained on indicated sieve size
Several practical decomposer designs were considered. Observations in an externally heated- transparent quartz tube showed that the decomposition of the hydrate bed is selffluidizing due to release of steam. This fluidizing effect can be utilized to improve heat transfer to the hydrate bed. On this basis, the most economical design concept for the decomposer resembles a vertical shell and tube heat exchanger with a selffluidized bed of hydrate flowing down the tubes and hot combustion gases flowing counter currently upwards in the shell.
Modeling of the fluidizing behavior inside the tube showed that the bed behaves as a bubbling fluidized bed in the top 80% of the tube and as a packed bed in the bottom section.
A picture of the pilot pressure decomposer is shown in Figure 8. The design is based around a single tube identical to what would be used in the plant calciner. It resembles a double pipe heat exchanger, with solid hydrate flowing down the tube and hot combustion gases flowing upwards in the annulus. Continuous feed to and discharge from the system were through pressurized lock-hopper arrangements. System steam pressure was maintained through a backpressure regulator. A gas-fired combustion system supplied hot combustion gases. A computerized data acquisition system connected to sensors provided flow, temperature and pressure data and computed mass and energy balances, heat transfer coefficients and pressure drop measurements along the column.
Heat Balance for Pressure Calcination Process A simplified heat and mass balance for the pressure calcination process is displayed in Figure 9.
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Figure 9. Pressure Calcination Process Flowsheet Calculations show that 1020 lbs of steam at 120 psig and 400°F are produced when calcining 1 t of alumina (0.5% LOI). The gas burning rate is 115 lbs, which works out to be 2.8 GJ/t A1203. Subtracting the heat available in recovered steam (1.25 GJ) the
Figure 8. Pilot Pressure Calciner
134
effective fuel consumption for the process is only 1.55 GJ/t A1 2 0 3 . Design Of Industrial Pressure Calciner
3. D. J. Brodie and H. W. Schmidt, Custom designed Fluid Bed Calciner for Nabateo ΡΊΎ LTD, Presented at 5 th International Alumina Quality Workshop, Bunbury, Aus., 1999.
At the time of the original paper the preliminary design of a 76 t/hr AI2O3 capacity pressure calciner was worked out based on bench and pilot plant experience. The design allows retrofitting of an existing Alcoa fluid-flash calcination system to operate in the pressure calcination mode. The design utilized a pressurized lockhopper system for feeding dry hydrate to the decomposer vessel. Further development work and implementation of this technology, however, was stopped at this point.
4. P. Clerin and V. Laurent, Alumina Particle Breakdown in Attrition Test, TMS Light Metals, 2001, 41-47.
Discussion
7. E. Guhl, and R. Arpe, Nearly 30 Years of Experience with Lurgi Calciners and Influence Concerning Particle Breakage, TMS Light Metals, 2002, 141-144
5. J. Fenger et al., Experience with 3x4500 TPD Gas Suspension Calciners (GSC) for Alumina, TMS Light Metals 2005. 6. W. M. Fish, Alumina Calcination in the Fluid-Flash Calciner, TMS Light Metals, 1974, vol. 3, 673-682.
An even more energy efficient unit, not discussed in the previous paper, would be to build a stand-alone pressure calciner. Wet filter cake would be fed into a pressurized dryer/decomposer unit where the alumina would exit at 850°C. A tubular design with self-fluidized hydrate/alumina on the inside of the tubes and POC on the outside, similar to the retrofit unit discussed above, is visualized. The hot alumina product could than go to a two-stage atmospheric indirect cooling section using air and water to cool the alumina to handling temperature and preheat air for fuel combustion. In this case all of the steam latent heat from hydrate moisture and water of hydration would be recovered as usable steam. Dust collector flow volume and particle load sizing would be significantly less than in current flash calciners.
8. Vladimir Hartmann, et al., Upgrade of Existing Circulating Fluidized Bed Calciners at CVG Bauxilum without Compromising Product Quality, TMS Light Metals 2006, 125-130. 9. J. B. Henin, Flash Calcination by Fives-Cail Babcock, TMS Light Metals, 1984, 1669-1696. 10. H.P. Hseih, Measurement of Flowability and Dustiness of Alumina, TMS Light Metals 1987, 139-149. 11. Cornells Klett, et al., Improvement of Product Quality in Circulating Fluidized Bed Calcination, TMS Light Metals 2010, 33-38.
Conclusions By recovering the water released during alumina calcination as process steam, the Alcoa pressure calcination system decreases effective heat requirement for alumina calcination to less than 1.6 GJ/tAl 2 0 3 .
12. E. W. Lussky, Experience With Operation of the Alcoa Fluid Flash Calciner, TMS Light Metals, 1980, 69-80. 13. Valerie Martinent-Catalot et al., A New Method for Smelting Grade Alumina (SGA) Characterization, TMS Light Metals 2004, 87-92.
The large gibbsite - boehmite transformation occurring during pressure decomposition results in a strong attrition resistant product and enables the production of < 1 % LOI smelting grade alumina at the low calcining temperature of 750-850°C. The surface area of the product is around 70 m2/g suitable for effective F recovery.
14. James Metson et al., Evolution of Microstructure and Properties of SGA with Calcination of Bayer Gibbsite, TMS Light Metals, 2006, 89-93. 15. Michael Missalla, et al., Increased Availability and Optimization of Calciner Performance Due to Automation, TMS Light Metals, 2009, 241-245.
The process was studied in bench and continuous pilot units and an initial design for a 76 t/hr alumina pressure calciner was developed but not implemented.
16. Dag Olsen, Alumina Dustiness Related to Physical Quality Parameters - User Experience and R&D in Hydro Aluminum, Presented at 5 th International Alumina Quality Workshop, Bunbury, Aus, 1999.
With the energy saving, ability to retrofit into existing fluid flash calciners and the ability to produce a better SGA for smelters, it is very surprising that this technology has languished without being implemented! References
17. S. Perra, Measurement of Sandy Alumina Dustiness, TMS Light Metals 1984, 269-286.
1. Evan W. Andrews and Barry J. Welch, Alumina Quality Requirements for Improved Cell Feeding, Presented at 5th International Alumina Quality Workshop, Bunbury, Aus., 1999.
18. Andre' Pinoncely, and Karim Tsouria, FCB Flash Calciner Technology: Ten Year Performances, TMS Light Metals, 1995, 113-120.
2. T. R. Barton et al., Flash Calcination of Alumina: An NMR Perspective, TMS Light Metals 1995, 71-74.
19. B. E. Raahauqe et al, Energy Saving Production of Alumina With Gas Suspension Calciner, TMS Light Metals, 1982.
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20. L. Reh and H. W. Schmidt, Application of Circulating Fluid Bed Calciners in Large Size Alumina Plants, TMS Light Metals, 1973, vol. 2, 519-532. 21. G. I. D. Roach, and J. B. Cornell, Dust and Dustiness Testing of Smelter Grade Alumina, Presented at 2 nd International Alumina Quality Workshop, 1990. 22. G. I. D. Roach et al., Charge Contrast Imaging -Applications to the Bayer Process, Presented at 5 th International Alumina Quality Workshop, Bunbury, Aus., 1999. 23. Alpaydin Saatci et al., Attrition Behaviour of Laboratory Calcined Alumina from Various Hydrates and its Influence of SG Alumina Quality and Calcination Design, TMS Light Metals 2004, 81-86 24. J. V. Sang, Factors Affecting the Attrition Strength of Alumina Products, TMS Light Metals, 1987, 121-127. 25. S. W. Sucech and C. Misra, Alcoa Pressure Calcination Process for Alumina, TMS Light Metals, 1986, 119-124. 26. O. Tsamper, Improvements by the New Alusuisse Process for Producing Coarse Aluminium Hydrate in the Bayer Process, TMS Light Metals, 1981,103-115. 27. K. Wefers, and C. Misra, Oxides and Hydroxides of Aluminium, Alcoa Tech. Paper No. 19, 1987, Alcoa Technical Center. 28. Suzanne Wind and Benny E. Raahauge, Energy Efficiency in Gas Suspension Calciners (GSC), TMS Light Metals, 2009, 235240. 29. Suzanne Wind et al., Development of Particle breakdown and Alumina Strength During Calcination, TMS Light Metals, 2010, 17-24. 30. K. Yamada et al., Development of a Fluid Calciner With Suspension Preheaters, TMS Light Metals, 1983, 159-172. 31. J. D. Zwicker, The Generation of Fines Due to Heating of Aluminium Trihydrate, TMS Light Metals, 1985, 373-395.
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
DYNAMIC SIMULATION OF GAS SUSPENSION CALCINER (GSC) FOR ALUMINA Susanne Wind1, Mengzhe Wu2, Torsten Vagn Jensen2 and Benny E. Raahauge1 F.L.Smidth A/S Minerals, automation Denmark Vigerslev Allé 77, DK-2500, Denmark Keywords: Dynamic Simulation, Operator Training, Alumina Calcination This is especially so, if the Bayer circuit can process more bauxite, liquor and red mud as illustrated in Table 1 below, assuming:
Abstract Training of plant operational personnel is becoming more important today than ever, to sustain high availability and productivity of high capacity equipment. The Gas Suspension Calcination process for production of Smelter Grade Alumina is very easy to operate and control regardless of calcining capacity. But increasing calcining capacity of single GSC units exceeding 4500 tpd of SGA makes it increasingly costly to lose operating time. The process dynamics of GSC units are very fast with some true response times in fraction of seconds. To train GSC operators, FLSmidth has developed a dynamic Calciner Simulator which serves the primary purpose of training both new operators, as well as maintaining the skills of experienced ones. The dynamic Calciner Simulator has been developed from supply of more than 75 Pyro process simulators by FLSmidth to the global cement industry. The first Calciner Simulator for Alumina will be commissioned in Australia later this year.
1) 2) 3)
SGA production of 1,125,000 tonne per year; SGA Sales Value: 275 US$/ton; SGA production cost: 230 US$/ton; Calcination Capacity 3,500
Capacity
ton/day S&%
3,087
3,190
3,259
3,430
3,602
1 Availability - o r [Operating Factor
55%
2,993
3,092
3,159
3,325
3,491
|
9B%
2,930
3,027
3,092
3,255
3,418
1
2,867
2,962
3,026
3,185
3,344
1
|
|profrt/{Uïs)
(236)
4,394
7,481
15,199
22,916
(4,4891
0
2,993
10,474
17,955
|
1 Operating Factor
93%
(7,324)
(2,930)
0
7,324
14,648
1
91% US$/yr
(10,159)
(5,859)
(2,993)
4,174
11,340
]
(84,507)
1,571323
2,676,043
5.436,593
347
(1.556.474)
0
1,037,649
3,631,773
6,225,896 |
33$
{2,486,047}
(594,419)
0
2,486,047
4,972,094 J
332
(3374,229] (1,946,067)
(993359)
1,386,311
3,766,581 |
1 Availability in 1 Days per Year
If the Operating Factor decreases from 95% to 93% (at 93% Capacity Utilization) due to lack of operator training, the loss of just 8 days production, will incur a loss of US$ 994,419 to the refinery per year.
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8,197,143|
Table 1: Gross Profit and Loss Opportunities.
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JE
US$/d 3 y |
95%
I Profit /(Loss)
As demonstrated over many years the Gas Suspension Calcination (GSC) process for production of Smelter Grade Alumina is very easy to operate and control regardless of calcining capacity.
1
1 Availability - or
1
Introduction
Capacity Utilisation (CU)
1 Nominal Design
jt ·
Or, if due to sufficient operator training, the Operating Factor can be increased from 93% to 95% and the Capacity Utilization can be increased to 100% based on improved market conditions as well, the gross profit will increase with US$ 3,631,733 per year.
1 - *JÎ
The above examples illustrates that to sustain high availability and productivity of high capacity equipment, training of plant operational personnel is more important today than ever before.
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Figure 1: 3 x 4500 tpd GSC Units at QAL, Australia
Furthermore, the training needs to be done without exposing the refinery to undue risk which can be done by using a Dynamic Process Model Simulator.
However, when the GSC operating factor exceeds 98% as is the case in some GSC units, very little opportunity is left for the operator to gain actual hands on shut-down and start-up experience.
In summary improved operator training can provide the following benefits:
With calcining capacity of single GSC units now exceeding more than 4500 tpd of SGA a day, it also becomes more and more costly to lose operating time owing to unnecessary calciner down time.
> > > >
137
Reduced and improved handling of incidents such as plugging; Improved Safety of operation with less stops; Maximize Capacity Utilization to meet demand; Maximize the Operating Factor at all times;
> >
develop a model that reliably predicts the quality of the calcined alumina in response to the operating conditions imposed. The FLS calcination model is built on the experience and know how gathered since the first GSC was put into operation in 1986, and incorporates modeling of the degree of calcination as well.
Improved quality by more stable operation; Minimize down time and cost;
Needles to say, in addition to operator training - proper plant and equipment maintenance is also required in order to maximize return on the assets installed.
The cooling section (COI, C02, C03 and C04) is used for recovery of heat from the hot calcined alumina. This is efficiently done by using four cyclones. The counter-current flow obtained with four (4) cyclone stages of co-current flow in series is providing a high thermal efficiency with respect to cooling the alumina and simultaneously pre-heating the combustion air. Again modelling heat transfer as well as solids separation becomes paramount.
To train GSC operators, FLSmidth has developed a dynamic Calciner Simulator which serves the primary purpose of training both new operators, as well as maintaining the skills of experienced operators. The dynamic Calciner Simulator described in this paper has been developed from operational experience, theoretical principles and the supply of more than 75 dynamic calciner process simulators by FLSmidth to the global cement industry.
The last section of the GSC system comprises a fluid bed cooler, which will reduce the alumina temperature from approx. 180 °C to a temperature low enough for the alumina to be transported with a belt conveyor. However, this process will not have any significant influence on the overall heat consumption of the GSC system.
CEMulator Unit Sale Volume
DUST RECOVERY
HEAT RECOVERY
Figure 2. CEMulator unit sales volume: 1986 to 2009.
jj~
As seen in Figure 2 the sales of such simulators to the cement industry have increased over the past decade. FLSmidth did experience a sales volume reduction in 2008 and 2009 due to the global economic conditions, but sales are expected to be at the 2007 level in the near future.
*""** — i
fis »MM
»*V Figure 3. FLS Gas Suspension Calciner (GSC) flow sheet.
The first dynamic Calciner Simulator to be described in the remainder of this paper, will be delivered to Australia for a new 3500 tpd GSC unit before it is started-up and commissioned later this year.
The Gas Suspension Calciner is straightforward to control with the following control loops: > > > >
Gas Suspension Calciner Process Flow Sheet Many processes are involved in the calcination of alumina hydrate. When the moist hydrate enter the GSC system it encounters hot gases from combustion of fuel and the calcination reactions. The purpose of the venturi and the first two cyclone stages is to utilize this heat for drying, pre-heating and precalculation of the hydrate. The foremost processes are heat- and mass-transfer and solids separation.
>
Production Rate controlled by Hydrate Feed Rate; Alumina Quality controlled by Calcination Temperature; Calcination Temperature controlled by Fuel Flow Rate, Alumina Discharge Temperature controlled by Cooling Water Flow Rate to Fluid-bed Cooler; Excess Oxygen controlled by Exhaust Gas ID-fan;
The control parameters for the closed-loop process simulator can be altered as the parameters in the plant DCS. Tuning these loops will enhance the closed-loop dynamic response of the process even further in the direction of truthful plant dynamic behavior.
The hydrate is partly calcined in the riser duct to the second preheater cyclone, but will continue to react in the next section, which is the calciner furnace and holding vessel. The fuel is injected into the furnace and combusted to achieve the necessary furnace temperature. The hot combustion gases entrain the partially calcined alumina into the Holding vessel, where sufficient retention time is provided to reach the final degree of calcination and alumina quality (LOI, SSA and alpha phase). Again heat transfer is important, but in combination with a calcination model and a combustion model that can handle both natural gas and heavy fuel oil. The central point is however to
The control strategy is to keep the Hydrate Feed Rate and thus Production as stable as possible and then modulate the Fuel Flow to maintain the Calciner Furnace temperature. This offers the advantage of a rapid control response as the manipulated variable (Fuel Flow) and the measured control variable (Furnace temperature) is situated in the same vessel. Dynamic Process Modeling of the GSC
138
reaches its steady state velocity (= gas velocity less its terminal velocity) in less than 1 second.
The aim of the simulator was to create a dynamic model that also corresponded with our steady state model when solving the timedependence mass, energy and momentum equations. Building the model thus allowed us to simulate starting and stopping the equipment, but also to introduce changes in any variable to model different operation scenarios.
Finally the operator observed process dynamics are further slowed down by the sampling frequency of the DCS. While the dynamic response of thermocouples is measured in minutes, the alumina quality like SSA is measured in hours, owing to the turnaround time of alumina samples through the laboratory.
In general constructing a model can be approached in different ways. It can be either: • • •
Fully theoretical, derived from first principle and theoretical properties Semi- empirical, using first principles and derived model parameters by comparison with plant data Fully empirical, derived by correlating input and output data.
Still other parts of the plant never reaches a true steady-state such as the outer part of the refractory lining being exposed to the ambient temperature cycle covering 24 hours per day and seasonal cycles over the entire year. In all, the FLS model is a mathematical analogy of the process generated by applying the principle of conservation of mass, energy and momentum. It can be used in real time simulations, thus improving the predictability and flexibility of plant operation.
The FLS model is semi empirical, solving the dynamic balance equations to reach steady state, but utilizes the considerable FLS know-how about process and operation of the GSC to obtain the right model response in comparison with observed behavior.
Model equations - An Example The mass and heat balance models are derived from first principle:
Some of the process dynamics of Gas Suspension Calciners are very fast with some true response times in fraction of seconds, such as accelerating the particles discharged from a cyclone into the down-stream riser duct as described below. Once particles are dispersed into the flowing gas stream in the Cyclone Riser Duct, each particle in the gas suspension is accelerated by the flowing gas until each particle reaches its steady state velocity which depends on the gas velocity and the particles terminal velocity. The time needed for each particle to reach their steady state velocity can be estimated by solving the equation of motion for each single sized particle. The vertical upwards flowing particle velocity relative to the gas velocity, at time t>0, assuming that the Stokes drag law applies, is:
Accumulation = Input - Output + Generation
In other cases a force balance is used as in the example of the pressure drop over a cyclone stage including its up-stream riser duct (heat exchange & separation stage): 0 =
Pressure drop due to acceleration Δρ,
Particle Velocity
v /// /
/
__
'
/
/
100 um/
/
/
(FgM + Fdust + Fs)-v2
{FgM +
FdustyVl
(4)
Pressure drop in the riser pipe bend
^~^~
/""
/
(3)
Pout " 4 P
drop components associated in the separation vessel:
The calculated particle velocity (please see list of symbols below) is shown in Figure 4.
/
Pin'
The pressure drop Δ/7 is determined by summing five pressure
Uo = [dp2 (pp - pg) g /18μ] (1 - exp[-18t/dp2pp]) + Uio exp[-18t/dp2 pp] (1)
45 ym
(2)
/ /ί
Δρ φ6 =0.5/^(1+%±£-)Λ 2 ν 2 2
/
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//
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Pressure drop due to rise pipe area change Apm = 0 . 5 ( 1 + ^ ^ - ) / ^ (v32 -vl )
/
(6)
Pressure drop due to friction
χ έ £ ^
Apcyc=0.5kcycpg^
Time, sec
(7)
Pressure drop due to gravity
Figure 4. Particle Acceleration in a Vertical Riser Duct.
(8)
From the above figure it can be seen that a 45 micron particles reaches its steady state velocity (= gas velocity less its terminal velocity) in less than 0.1 second, while a 150 micron particle
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The total pressure drop will thus be expressed as the summation of the contributors:
4 -
Δ
0 -
Ρ = 4 P , « + Δ / V + APrPa + APcyc + 4 P , ™
2
<9>
-2 -
Below the dynamic response to a step change in the ID-fan speed can be seen on the outlet suction pressure from the dedusting filter and cyclone POI respectively. 95 94 93 92 91 90 89 88 87 86 85
• ID-Fan »POI press. » BH press.
1n
-7.5 -8.5
0
100
200
300
400 500 Time (s)
-10 -10.5 r
i
600
700
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However, matching the stationary steady-state response shown above is a necessary criteria, but insufficient to qualify the dynamic model as applicable for training operators!
-9.5
—i
1 1 1 1 1
Figure 7. Pressure Profile at 50% Turn Down ratio with 0,2% deviation relative to absolute pressure of 101,325 kPa.
-8 -9
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1 I
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-6.5 -7
^_
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-6
T
«Simulation «SteadyState
T
Needless to say, if a dynamic simulator model does not respond dynamically in a way that the operator can recognize from the real plant dynamic response observed from the DCS, the dynamic model will surely be abandoned and not used.
-11
800
Figure 5. Dynamic response in relative suction pressure to change in ID fan speed.
This raises the very important issue of proper model validation. We are of the opinion that the response of the dynamic process model must be tested and verified by our experienced GSC commissioning engineers in order to be qualified as a valid training tool - and thus applicable in the Training Simulator.
Implementation of Dynamic Model Results and Validation The overall pressure profile model prediction at steady-state is compared below with the resulting pressure profile calculated by the independent steady-state design model, which in turn has been verified against practical GSC plant operation:
Training Simulator System Structure To cover the needs for conducting simulation based training the FLSmidth training simulator consists of 4 program modules. The modules that are illustrated in Figure 8 covers: • • • •
Figure 6. Pressure Profile at Nominal Production with 0% deviation relative to absolute pressure of 101,325 kPa. As can be observed from Figure 6, a reasonable accurate convergence is obtained from the model when applied at nominal production of the GSC. But to be of use to the operator, the model must also provide a reasonably true response over the entire GSC capacity range as demonstrated below in Figure 7 at 50% Turn Down ratio.
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Dynamic model and solver, simulating the dynamic process behavior; Soft PLC, simulating control loop actions and equipment unit/sequence interlocking behavior; Simulator, control of the training communication between instructor and trainee, etc. DCS (Distributed Control System), enablesriskfree trainee operator interaction;
Simulator Dynamic solver Process simulation
CERTIFICATE
Session control
Session evaluation
ACTION REPORT
ACTION REPORT
DCS Operator Access
Unit simulation
Soft PLC
ACTION REPORT
Automated tutorials
Data transfer
Figure 8. Simulator program structure split in program modules (dashed lined boxes can be standard FLSmidth setup or Honeywell).
Figure 9. Example of action reports and session plots representing the offline session evaluation.
In the Dynamic solver module the process dynamics are simulated. This includes the heat, mass and momentum balances and will represent the fast as well as the slow dynamic behavior in the GSC system based on mass, component and energy holdup. At FLSmidth we use the gPROMS, solver which is a state of the art dynamic solver delivered by PSE based in UK.
Next in the Simulator module is the automated Tutorials. Tutorials are prerecorded sessions including session control events like the action described in online session control. These can be used to uniform the training as events will have exactly the same timing for all Trainees, which results in that evaluation of two Trainees can be compared. The messages included in Tutorials are used to support the training and can therefore be used to ensure that all Trainees learn the same procedures covering startup, hot standby and shutdown.
The Soft PLC module is where unit simulation is hosted. As in a true plant control system setup the PLC holds the process and plant information defining the operating state of the process. For this purpose PLC programs are uploaded in the Soft PLC. These programs are modified to simulate the physical equipment responses such as a starting motor. Besides the FLSmidth standard Rockwell Soft PLC setup a solution supporting Honeywell Soft PLC s are also supplied.
Finally the Simulator module includes the data transfer between the other modules. The main data transfer is synchronizing the Soft-PLC and Dynamic solver so that all process data is updated in the PLC as well as all motor states and set-points in the dynamic solver. This module also ensures that the Soft PLC and Dynamic solver are aligned when restoring a saved session.
In the Simulator module all the programs used to set up the simulation is hosted.
In the Distributed Control System (DCS) objects are presented to the Trainee just as in a real DCS. Here object trending, alarm handling, set-point changes and the remaining features covered by the DCS will be available.
The first area covered by this module is the Session control, where the Instructor can interact with the simulation session allowing control of the following: • • • • • •
Start/Stop/Pause session Simulation speed - accelerate process responses if needed Disturbances - offset on demand Input parameters - external inputs/definitions On line evaluation Text messages
In the FLSmidth standard solutions the in-house developed DCS named ECS (Expert Control and Supervision) is used, but the simulation setup also supports a solution using Honeywell Experion DCS as operator interface. Common for all setups are the need for interaction from the Trainee and Instructor through their different stations. Figure 10 and 11 illustrates the accessibility needed for these two stations.
Secondly the Simulator module includes online and offline Session evaluation. In the online Session evaluation the Instructor can monitor the current performance of the Trainee throughout the session. In offline evaluation the completed sessions are presented as reports representing evaluation objective trends and action reports. These reports can be used to evaluate the performance of the Trainee and dig into what learning areas that needs focus of the instructor.
141
In addition to the increased use of alternative fuels and focus on emissions, the current situation also raises demand for the operators to get the right skills by simulator training. Subsequently, this training results in more continuous and optimal plant operation with economical benefits to follow.
Trainee Station
In the cement industry we have successfully developed an array of standard simulation solutions, which can be purchased as off the shelf applications for fast delivery. The trend is that these are sufficient to cover the needs for training in most organizations, while specific plants tend to need minor adjustments for the process to look more like the specific plant. Back in the pioneers days in the 1980's the simulator was sold to specific plants. This trend has changed today where we still deliver half of our simulators to specific plants, but an increased interest from global companies is to purchase simulators for corporate training centers.
Figure 10. The program module access from the Trainee station. From the trainee station the Trainee can make some of the basic session controls such as start/stop/pause and send messages. To ensure focused training the Trainee must be able to monitor the evaluation throughout the session. As the natural environment for an operator is in the DCS the Trainee's main access to the simulation will go through this interface to do control, monitoring and error tracking.
Apart from the CEMulator units sold externally there are 70 internal licenses spread across FLSmidth where the use varies from training to assistance in development of high level control standard solutions. Conclusion
Instructor Station
Increase in calcining capacity of single GSC units, sometimes exceeding 4500 tpd of SGA, makes it costly for the refinery to lose operating time. When at the same time the GSC operating factor exceeds 98%, as experienced in some GSC units, very little opportunity is left for the operator to gain actual hands on shutdown and start-up experience. The Dynamic Calciner Simulator for SGA production developed by FLSmidth, will provide the Alumina refinery with the option of a risk free training of its new and experienced operators. In the cement industry the trend is that sufficient and documented training of operators is becoming a growing demand from local authorities as more and more focus is placed on sustainable production and ensuring environmentally constrained emissions.
Figure 11. The program module access from the Instructor station.
Consequently FLSmidth has experienced a surge in supply of training CEMulator units to the global Cement Industry, a demand that may come from the global Alumina Industry in the not so distant future.
The instructor station will act as a base for the Instructor. Here the full palette of session control must be available so that the Instructor in addition to the Trainee control can introduce disturbances, change input parameters and evaluation objectives on line. During training preparation the Instructor can use the tutorial editor to tailor make training sessions to the upcoming training needs. While undertaking the training the Instructor will also need to access the Operator DCS to fine tune the tutorials. The DCS can also be used during the session to monitor the current Trainees sessions during training.
List of Symbols Symbol Fs ^g.in Fdust
υι υ2 υ3
FLS Experience with Training Simulator In the cement industry the trend is that sufficient and documented training of operators is becoming a growing demand from local authorities. A recent example is the requirement for documented training in cement industry in USA. The reason for these demands is not only for the purpose of plant and staff safety, but also a requirement for implementation of a sustainable production.
Ai Août Krpb Kcyc Pg,2 Pg,3
142
Definition Solid Mass flowrate Gas Mass flowrate Solid entrained in gas flow Gas velocity at inlet of riser pipe Gas velocity at outlet of riser pipe Gas velocity in cyclone inlet area of riser pipe outlet area of cyclone Friction constant Friction constant Gas density at outlet of riser pipe Gas density in cyclone
Unit kg/s kg/s kg/s m/s m/s m/s m2 m2
-
kg/m kg/m
M Δρ Pin Pout
H g
Δρ 8 1 1 8 Aprpa Aprpb Apcyc Apgrav
u0 dp
PP Pg
t μ
ui0
Solid hold up Pressure drop Inlet pressure Outlet pressure Height of cyclone Gravity Pressure drop of acceleration Pressure drop of changed area Pressure drop in the riser pipe bend Pressure drop in the cyclone Pressure drop through gravity lift Rei. particle velocity at t>0 Particle diameter Particle density Gas density Time Gas viscosity Rei. particle velocity at t=0
kg kPa kPa kPa m kg/m2 kPa kPa kPa kPa kPa m/s mm kg/m3 kg/m3 s kg/s m m/s
References 1.
2. 3. 4. 5. 6.
B.E. Raahauge,et al.: "Application of Gas Suspension Calciner in Relation to Bayer Hydrate Properties", The Australasian Institute of Mining and Metallurgy, 1981 Annual Conference, Sydney NSW, 1981. B.E. Raahauge,et al.: "Energy Saving Production of Alumina with Gas Suspension Calciner", 111th AIME Annual Meeting, Dallas,US,1982. T.A.Venugopalan,"Experience with Gas Suspension Calciner for Alumina", Proceedeings 1st International Alumina Quality Workshop, pp 53-66, (1988). S.Wind and B.E.Raahauge: "Energy Efficiency in Gas Suspension Calciners (GSC)",TMS Light Metals, pp 235- 40, (2009). J.Ilkjaer, L.Bastue and B.E.Raahauge:"Alumina Calcination with the Multi Purpose Calcine", TMS Light Metals, pp 107111,(1997). R.K. Jonas:"Design and Training Improvements through the use of Dynamic Simulation", TMS Light Metals, pp 63-67, (2004).
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Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
Physical Simulation and Numerical Simulation of Mixing Performance in the Seed Precipitation Tank with a Improved Intermig Impeller Zhang Ting-an1, Liuyan1, Wang shuchan1, Zhao hongliang1, Zhang chao2, Zhao qiuyue1, Lv guozhi1, Dou zhihe1 (1 School of Materials and Metallurgy of Northeastern University, Key Laboratory of Ecological Utilization of Multi-metal Intergrown Ores of Education Ministry, Shenyang, 110004; 2 Shenyang Aluminum & Magnesium Engineering & Research Institute, Shenyang, 110001)) Keywords: Seed precipitation tank; Mechanical stirring; Numerical simulation; Solid-liquid phase flow Abstract A PC6D fiber optic reflection probe was used to measure local solid concentrations distributions in the seed precipitation tank with a new-style Improved Intermig impeller, and the commercial software FLUENT 12.0 on parallel computing graphic workstation was used to simulate numerically its flow fields. The physical simulation results show that at high rotation and viscosity, high-speed particles are more conducive to uniform distribution. When //=3.50cp, «=172rpm, the multiple relationships of the average value of the experiment and theory is 1.011 times. The numerical simulation results show that the too long or too short impeller off-bottom clearance height(C) is not conducive to the suspension of Al(OH)3 particles. Enlarging the blade diameter is good for suspending of Al(OH)3 particles. The blade diameter has a big influence on the stirring power, and the longer blade diameter needs more power consumption, but the C has little effect on the stirring power, which can be ignored. The physical simulation results compare well with numerical simulation.
could reduce the sediment significantly, was devised by the Shenyang Aluminum & Magnesium Engineering & Research Institute. Power consumption also declined from 200kW to 75kW. The difference between the traditional Intermig impeller and the improved impeller has been reported in the literature141.
At first, we want to find the best impeller off-bottom clearance height and impeller diameter just in single impeller experiments. And in next step, we will study the multiple impellers' performance based on the results of single impeller experiments.
Cold water experiments The profile of local solid concentrations on the water-glass bead particles in liquid-solid system, were measured by using a PC6D fiber optic reflection probe and analyzed under different conditions. The PC6D particle concentration measuring instrument is manufactured by the Institute of Process Engineering, Chinese Academy of Sciences, the measuring range is 0~800g/L. The instrument uses an optical fiber as the measuring probe. The reflected light of materials at the fiber end comes back to the photo detector in the instrument through the same fiber and is converted into a voltage signal proportional to the concentration of materials. The average concentration of solid particles is obtained from the voltage signal.
In order to fully understand the impact of the structure of the blade on the solid suspension characteristics and power consumption, the seed precipitation tank with Intermig impeller was studied in our laboratory delegated by the Shenyang Aluminum & Magnesium Engineering & Research Institute. The study includes mixing uniformity, suspension of Al(OH)3 particles and power consumption, etc. By this research we intended to improve the internal structure of the seed precipitation tank, solve the problem of sedimentation, and improve the yield and quality of alumina. This has important theoretical and practical significance for the alumina industry. Research Method
Introduction Technological revolutions have recently hit the alumina industry. As one of the key steps in the Bayer process for the production of alumina, the seed precipitation step has a great influence on alumina product output and quality, and also has an indirect effect on other processes111. So it is necessary to optimize and improve this step. Currently, the mechanically agitated precipitation tank has become the major equipment of seed precipitation step, because it has many advantages which include low power consumption, better mixing effect and less sediment; avoiding "Short-circuiting" of slurry; high reliability, etc[2]. But the application of mechanically agitated precipitation tanks was late in China, and we have less information on the flow field in the tank.
The cold water experiment is based on the similarity principle. The ratio between water model and industrial seed tank is 1:33, and the experimental equipment was made with plexiglass 043.5cm x 52cm. Liquid height was 35cm and material system was glass beads-water (with sugar added to change the viscosity). Table 1 shows the comparisons of physical parameters. Glass beads were used instead of Aluminum hydroxide and an amount of 3800g added in the cold water experiment. The syrup viscosity in the cold water experiment was consistent with the viscosity of the NaA102 solution in the seed precipitation process.
By comparing the power consumption of a traditional Intermig impeller and a pitched blade impeller in industrial tests, the former is much lower than the latter, but using the traditional Intermig impeller produced a large amount sediment133. So a new-style impeller, - an improved Intermig impeller - which
145
n=172rpm, it was 1.340 times; when r|=3.50cp% n=172rpm, it was 1.011 times. The results showed that the Experimental concentration was closer to theoretical concentration under this condition, which was more evenly mixed. Turbulence model: Realizable k-ε; Multiphase model: Eulerian; Interphase drag: Gidaspow; Rotate model: Multi-Reference Frame; Numerical solution: SIMPLE algorithm for pressure velocity coupling; Discretized with the second-order upwind; All Residual converges to 10"3
Table I Comparisons of physical parameters Physical Parameters
Density /kg/m3
Viscosity /cp
Sodium Aluminate Solution
1330
3.5
1000 1149
1.52 3.50
2430
Granularity ΙΟΟμπι
Running Water Syrup Aluminum Hydroxide Glass Beads
2380
Numerical simulation Some research has been done on solid - liquid phases numerical simulation15"71, that may verify the feasibility of study on seed precipitation tank by CFD. (1) Grid Generation In order to reduce computing time and improve the convergence and stability of solution process, we should reduce the grid number and improve the mesh quality. So we took the following methods: A: Ignored the thickness of baffle. B: Changed grid from a tetrahedron into a polyhedron. Figure 1 shows the grid model of the seed precipitation tank. The grid number was about ten million.
Fig.l Grid model of the seed precipitation tank
Results and analysis of numerical simulation
(2) Simulation Strategies From the literature about simulation of solid-liquid phase flow[8"11], we chose the simulation strategies are as follows:
The numerical simulation research is for the seed precipitation process, the liquid-solid system of NaA102 solution and Al(OH)3 particles (for detail about material parameters see attached Table 1). At first, we validated the simulation strategy based on cold water experiments and then we found the best conditions by simulating the seed precipitation process.
Results and analysis of cold water experiment Experimental measurement points were set to 7.5 % 9.5 * 11.5% 13.5% 15.5x 17.5 s 19.5> 21.5cm in radial distance (the center of slot as the 0 point that was indicated as 1); 2x 6^ \0S 14% 18% 22% 26> 30cm in axial distance (the bottom of slot as the origin that was indicated as d).
Simulation validation based on cold water experiments In this section, the liquid-solid system was composed of water and glass beads. By comparing the results of water model experiment and numerical simulation on the solid particles concentration field, we got the following results:
Figure 2 showed the concentrations along the radial and axial distance. A uniform distribution of particles in the tank was assumed to be the theoretical concentration. When r|=1.52cp% n=96rpm, the multiple relationships of the average value of the experiment and theory was 0.962 times; when r|=1.52cp>
90 r
10 12 14 16 18 Radial distance(cm)
10 15 20 25 Axial distance(cm)
22 (a)f/=1.52cp, w=96rpm
146
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O
8
10 12 14 16 18 20 22 Radial distance(cm)
10 15 20 25 Axial distance(cm)
(b)»pl.52cp, n=172rpm 90
90
80 l·
80 70l· ^60 3 50 40 30 8
20
10 12 14 16 18 20 22 Radial distance(cm)
10 15 20 25 Axial distance(cm)
30
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the same trend on particle concentration gradient, so the simulation strategy and results were credible. Keeping other conditions the same, we increased the viscosity of the solution, and then we got the following results: 210
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consistent with the experiment results. The effect of distance between the impeller and the bottom(Ci on particle suspension and power consumption
148
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Figure 7 shows the distribution of particle volume fraction with different C. (y= 0 section). The results showed that too long or too short a value of C may lead to aggregation of Al(OH)3 *»·** S1sr**t
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The results showed that C had little effect on the particle suspension at low stirring speed, while a large C may make the particle distribution more uniform in the whole tank at high stirring speeds.
0.132 h
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The power consumption with different C at 150rpm can be seen in Figure 9. With the increase of C, the stirring power at first increased, and then began to stabilize, but the C had little effect on the stirring power. In this section the best value of C was 111mm.
M C
0.126 h
on 0.123h 40 80 120 Distance between impeller and bottom(nim)
The effect of impeller diameter (D) on particle suspension and power consumption Figure 10 shows the distribution of particle volume fraction with different D/T (T was the tank diameter, y= 0 section), and Table 2 showed the power consumption under the same conditions.
160
Fig.9 Power consumption with different C by numerical simulation, 150rpm
149
(a)D/T=0.5 (b)D/T=0.6 (c)D/T=0.7 Fig. 10 Concentration field with different D/T by numerical simulation, 150rpm 2. Zhang Chunyang, Fu Yan. Alumina Plant Design, (Beijing: Metallurgical Industry Press, 2008), 115-141. 3. Wu Jianqiang, Dong Baocai, "Evaluation of Operating Results of Large-scale Precipitation Tanks with Mechanical Agitators", Linght Metals, 2000(03). 4. Wang Rui, "Numerical Simulation of Multi-Phase Flow in the Alumina Seed Precipitator" PhD Thesis of Northeastern University, Shenyang, China, 2009:29-30. 5. G Montante, "Experiments and CFD Predictions of Solid Particle Distribution in a Vessel Agitated with Four Pitched Blade Turbines", Chemical Engineering Research and Design, 2000,79(8):1006-1010. 6. R. J. Weetman, "Automated Sliding Mesh CFD Computations for Fluidfoil impellers", 9th Euro.Conf.On Mixing, 1997:195-202. 7. I. Naude, C. Xuereb, J. Bertrand, "Direct Prediction of the Flows Induced By a Propeller in an Agitated Vessel Using an Unstructured Mesh", Can Chem. Eng., 1998,76:631-640. 8. Montante Q "Experiments and CFD Predictions of Solid Particle Distribution in a Vessel Agitated with Four Pitched Blade Turbines", Chemical Engineering Research and Design, 2000,79(8):1006-1010. 9. Weetman R J, "Automated Sliding Mesh CFD Computations for Fluidfoil Impellers", 9th Euro. Conf. on Mixing, 1997:195-202. 10. Naude I, Xuereb C, Bertrand J, "Direct Prediction of the Flows Induced By a Propeller in an Agitated Vessel Using an Unstructured Mesh", Can Chem. Eng., 1998,76:631-640. 11. Bakker A, Fasano J B, Myers K J, "Effects of Flow Pattern on the Solids Distribution in a Stirred Tank"(8th European Conference on Mixing, Cambridge, U.K), IChemE Symposium, 1994,136(329):l-8.
Table II Power consumption with different DAT by simulation(unit: W) C=71mm
D/T=0.5
D/T=0.6
D/T=0.7
40rpm
0.0472
0.127
0.258
150rpm
2.025
6.060
13.690
Figure 10 and Table II show that enlarging D may promote the suspension of Al(OH)3 particles, but the power consumption increased substantially. So we should choose moderate D to get a better efficiency in particle suspension, with low power consumption. In this section the best value of D/T was 0.6. Conclusion (1) In terms of overall distribution, the particle concentration with radial distance increases, while the axial distance reduces. The high rotation and viscosity, high-speed particles are more conducive to uniform distribution. When r|=3.50cp> n=172rpm, the multiple relationships of the average value of experiment and theory is 1.011 times. It shows that the experimental concentration is closer to theoretical concentration under this condition, indicating the particles are more evenly mixed. (2) By comparing simulation and cold water experimental results, the feasibility of the CFD method is verified to simulate the flow and solid-liquid mixing characteristics. (3) At a constant stirring speed, there is an optimum value of D and C which can get a better effect of particle suspension with low power consumption. In this study, the best values of C and D/T are 111mm (C/T=0.55) and 0.6 respectively. Acknowledgement The authors acknowledge the support of the National Natural Science Foundation of China (NO.50974035, 51074047). References 1. Yang Zhongyu. Alumina Production Technology, (Beijing: Metallurgical Industry Press, 1993.10),133.
150
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
Two Perspectives on the Evolution and Future of Alumina 1 Linus M. Perander, Barnes B. Metson, 2Cornelis Klett ^ight Metals Research Centre, The University of Auckland, New Zealand 2 Outotec GmbH, Frankfurt, Germany
Alumina, Bayer process, Hall Heroult process, Emissions, Energy, Calcination continuous way to feed the alumina resulted in the development of so called Break and Feed systems. The first implementation of this way of feeding was the Bar Breaking method. Using this method, a breaker beam is lowered to break the layer of crust (which covers the cell and anodes) and then a specified amount of alumina (determined by volume) is dumped into the cell. Along with the alumina a large amount of broken crust and cover material also enters the cell. Dissolution is slow and large amounts of undissolved alumina/crust are deposited on the surface of the cathode and gradually dissolved until the next feed cycle. The bar breaking method has given way to Point Feeding. Up to five of these point feeders (shot size 0.5 - 3 kg of alumina determined by volume per feeder) operate every few minutes, changing the operation from a semi-batch type to a more continuous nature [3].
Abstract Over the 125 year history there have been a number of stepchanges in the Hall-Heroult process, despite a remarkable adherence to the original concepts of the invertors. In addition to the steady increment in scale, most noteworthy perhaps have been the introduction and impending disappearance of Soderberg technology, the introduction of magnetically compensated cell design, changes in dynamics of alumina feeding and the introduction of dry-scrubbers for HF control and fluoride recovery. The Bayer process has also seen some significant advances, driven by the demands of energy and environmental imperatives and the steadily narrowing window of product specifications, driven in turn by refinements in the Hall-Heroult process. Demands for coarser particle size distribution, higher specific surface areas and lower attrition index have been accompanied by changes in precipitation strategy and conversion to more energy efficient stationary calcination processes.
As a result, particularly of these feeder changes, the demands on alumina quality have also changed. The so called Floury Alumina (with fine particle sizes, low specific surface areas and high alpha alumina contents) has poor flow and dissolution characteristics and has been displaced by Sandy Alumina (with larger particle sizes, high surface areas and low alpha alumina content). The impact of alumina properties on transport, feeding and dissolution characteristics, the ability to form a stable crust and anode cover as well as the adsorption of HF in the dry-scrubbers, have been of particular interest. Of note is the continued evolution of these properties (for example the inexorable demand for higher surface areas), at times with a limited understanding of the net process impact of such changes.
The properties of a "typical" metallurgical alumina have thus changed. Indeed the term "alumina" is now more indicative of stoichiometry than it is of structure, and even in this, it is less than precise. In this paper we discuss how new scientific tools and insights are changing the way we define (and perhaps should specify) this material. The History of Alumina Calcination The early history and development of the alumina and aluminium processes and industries are summarised elsewhere [1, 2] Fundamentally these processes remain the same, although a number of technological breakthroughs and step-changes have occurred. The past 50-years in particular have fundamentally changed what a smelter demands in terms of properties and performance of the primary raw material, the alumina fed to the reduction cell. This paper addresses the impact of the calcination step in the Bayer process on the development of alumina properties and our understanding of how these properties impact on the smelter operations.
Stationary calciners, as an alternative to the Rotary Kilns were introduced in the 1950s to 60s. The gains, both in increased production volumes, reduced maintenance and in energy savings, resulted in considerable research efforts into these technologies, and as a result variations of thefluidisedbed technology emerged. Today these are known as: Circulating Fluidised Bed (CFB), Gas Suspension (GS) and Fluid Flash calciners. For the main calciner technologies typical energy consumption, residence time, production capacity, alpha alumina content, calcination temperature and free heat transfer surface are presented in table 1. With typical production capacities between 2500 to 4000 tons per day for modern Gas Suspension or Circulating Fluidised Bed calciners, significant effort goes into minimising the downtime of these installations which has also led to other process and control improvements. Most noteworthy, however, is the reduction in energy consumption compared even to best practice Rotary Kilns.
The use of dry scrubbers in the aluminium smelters started in the 1960s as a response to therisingemissions concerns, and potential fluoride losses associated with this process. The dry scrubbers make use of the high specific surface area and the reactivity of partially calcined alumina (LOI < 1 wt-%) to capture the volatile fluorides emitted from the electrolyte in the reduction cell. Around the same time as the dry scrubbers were introduced, another significant change also occurred. Traditionally, the primary raw material, alumina, was fed to the electrolysis cell periodically (often manually) and in relatively large doses. Environmental requirements to enclose the cells as much as possible and the clear process control benefits of a more
The energy reductions can mainly be attributed to more rapid heat transfer into particles, heat recovery in the cooling stages and from the waste gas heat, resulting in common features such as the pre-heating stages and cooling stages with direct gas solid heat transfer as well as indirect heat recovery in the fluid bed coolers for both Gas Suspension and Circulating Fluidised Bed calciners. The energy consumption should however be compared to the
151
theoretical energy for gibbsite calcination which lies around 1.98 to 2.40 MJ kg"1 for dry and moist gibbsite, respectively [4], indicating that there is still margin for improvements. With the calcination stage amounting to approximately a third of the energy consumed in the Bayer process the potential energy savings becomes significant.
stationary calciners will inevitably span the diagram of reaction pathways and challenge the value of any type of quantitative phase analysis. Nevertheless such an analysis is one of the few comparative tools we have in explaining differences in alumina properties and performance. <€| 100 — — i
Table 1. Typical energy consumptions, residence times, production capacities, alpha alumina contents, calcination temperatures and free heat transfer surface for different calciner technologies [4-8].
Specific energy (MJ / kg product) Material residence time in the hot zone in the furnace (s) Typical calcination temperature (°C) Free heat transfer surface (m 2 /g product) Typical alpha alumina range (wt-%) Typical production capacity (tons per day)
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Circulating Fluidised Bed
Rotary Kiln
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1000
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Figure 1. Thermal decomposition pathways of gibbsite, adapted from Wefers et al. [2]. High intensity, high resolution synchrotron (and neutron) diffraction data has been shown to be extremely useful in terms of monitoring gibbsite crystal growth mechanisms from Bayer liquors in-situ [15], for identifying minor crystallographic impurities [16] as well as for examining the transition aluminas and their transformation reactions [17-20]. However, for deducing information on the structures of the poorly crystalline transition alumina phases, even these techniques face some limitations [21], particularly in complex mixed phase environments such as the MGAs.
Differences in the alpha alumina contents for the different technologies are directly tied to the residence times and calcination temperatures, whereas the higher free heat transfer surface for the modern calciners is related to the high gas velocities and large surface areas. It has been proposed that the heating rate during the calcination may influence the reaction pathways [9, 10], and it is therefore hardly surprising that material from rotary kilns often display different properties [11, 12]. Surprisingly limited effort has however gone into understanding how these differences arise and more specifically what impacts differences in the precursor material have on the calcination reactions and resulting product properties. The speed of the modern calciners, together with the variation of particle size and morphology, is expected to produce a structurally more disordered and heterogeneous material than those produced in rotary kilns. The disorder and co-existence of phases, often within a single particle, poses significant challenges for characterisation and understanding behaviour of these materials. It also challenges our conventional definition of what we call alumina.
Due to the short residence times in modern calciners, deviations from the average transition alumina structures are observed. This is caused by local disorder, which results in significant peak broadening which, combined with the overlapping peaks for the transition aluminas, complicates structural analysis using traditional diffraction methods. Despite the improvements in fitting diffraction data on MGA, made possible through the works by Ashida et al. [22] and Paglia et al. [19], conventional diffraction techniques are limited by the need for crystallographic long range order. Indeed from the analysis of a typical fluid bed MGA, Ashida proposes that what we call A1203 is better described as H2Al10O16 [22]. This is consistent with previous views of the structure of gamma alumina [2] and provides a clear insight into the origin of HF generation in the reduction cell as examined by Patterson et al. [23]. The hydrogen (as residual -OH groups) is a consequence of the increasing undercalcination of alumina, necessary to meet the demand for increased surface area. Thus we gain HF adsorption capacity at the scrubber, at the direct expense of an increased HF burden circulating with the cell gas.
The Science of Alumina Calcination Several excellent reviews on transition alumina phase changes during the calcination of gibbsite and boehmite exist; particularly noteworthy are those of Levin et al. [13], and the overview of Wefers et al. [2]. The reactions have been found to be influenced by several parameters (such as temperature, heating rate, residence time, particle size and morphology, crystallinity, impurities and atmospheric conditions) some of which are directly related to Bayer operations [2, 9, 10, 12, 14]. Figure 1 summarizes the possible reaction pathways for gibbsite dehydroxylation. It should be pointed out that, apart from the initial decomposition to chi alumina, rho alumina or boehmite, the reactions proceed slowly; even if the energetic barriers are overcome prolonged heating is required to reach equilibrium for any of the meta-stable transition alumina forms. Thus the particles formed in modern
The best practice in phase analysis is exemplified in the Rietveld refinement in figure 2 for which the diffraction data was obtained at a synchrotron source. Apart from the obvious overlap of several peaks in the diffraction pattern, the broad and diffuse peaks from the transition aluminas, the presence of additional X-ray amorphous components and the incomplete structural models for the transition aluminas, results in the discrepancies between the fitted (black) and observed (red) spectra. In the distinction
152
between the transition alumina phases (or forms) it is generally agreed that theta alumina is more ordered than delta alumina which again is more ordered than the gamma alumina phase [2426], however, exactly how the transformation into the more ordered forms proceeds is still debated [2, 13, 27]. The formation of gamma - gamma' - delta and theta alumina can also be seen as waypoints on the gradual transformation into a fully ordered state (represented by the thermodynamically stable alpha alumina).
5
13
21
29
37
45
53
61
65
77
85
Figure 2. Synchrotron X-ray powder diffraction patter of a CFB calcined MGA sample and resulting Rietveld refinement results. Note that the shape of the background is a result of the amorphous nature of the quartz capillary used for mounting the sample. The move to stationary calciners has apparently swapped the historical problem of the broad distribution of residence times, and thus phase composition for individual particles, for the problem of distributions of phases within single particles due to the heat transfer constraints in these technologies. As reported elsewhere, Environmental SEM can be used to observe alpha alumina directly in cross sectioned alumina grains through the Charge Contrast phenomenon [28]. The different dielectric properties of the structurally more ordered alpha alumina, compared to the transition alumina forms, result in a contrast difference in the ESEM. This has allowed for new insights into the alumina phase distribution and phase transformation mechanisms within single particles. The results indicate that alpha alumina formation is closely tied to growth morphology, and seems to follow the same pattern as the growth rings revealed in gibbsite cross sections (figure 3). This suggests that local structure and impurities are important in the nucleation and transformation reactions and ultimately the formation and location of alpha alumina within particles. The phase inhomogeneities between and even within particles exemplifies a wider challenge in understanding alumina behavior. The properties reported on a alumina specification sheet represent at best a heavily averaged view, where the outlier populations may be of more significance in, for example influencing alumina dissolution and flowability, than the mean value specified. The dominance of over and under calcined material in the fine particle size fraction is a good example of this [29].
Figure 3. Top left: Alpha alumina observed around the edge of a cross sectioned GS calcined alumina particle. Top right: Alpha alumina formation in a soak calcined alumina grain following the gibbsite growth ring pattern. Bottom left: gibbsite growth rings as revealed by Environmental SEM and corresponding Na distribution (bottom right), obtained using a ToF-SIMS instrument. The same consideration applies to specific surface area, LOI, and the rarely reported, but important, pore size distribution. As discussed more extensively elsewhere [30], the low order transition aluminas (gamma, rho, chi alumina) contain more residual hydroxyls (confirmed by LOI measurements) which inevitably results in more HF being generated upon dissolution. At the same time, the fine pore size in these under-calcined components might be expected to influence their ability to capture HF in the dry-scrubber. It seems that the narrow pores restrict access to internal porosity and readily become blocked (when HF reacts to form oxy-fluorides). This further restricts access to internal sites, thus reducing the capacity and rate of HF absorption. An example of a distinctly bi-modal pore-size distribution in a SGA sample is provided in figure 4. and corresponding HF generation and emissions data are presented and discussed in elsewhere [30].
I 0,0052 °~ 0.004-
0.10 0.08
importance and interest [34]. Another critical performance criterion for the alumina is rapid dissolution in the molten cryolite based electrolyte. To achieve this, sufficient dissolution power (or superheat) is needed, but also the method, amount and frequency of the alumina additions and the quality (dissolvability) of the alumina are of importance [35-38]. The operational stability and feed strategy relies on a consistent alumina quality. However, variations between, and even within, alumina shipments is often a reality. Such variations are frequently not reflected in the specifications of the alumina, making it difficult to anticipate and make process adjustments to accommodate the raw materials variations. Typical outcomes are process fluctuations, sludge formation or other feed related instabilities and on occasion, emissions problems.
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The alumina specification sheet currently represents a contractual quality index in terms of a number of (often historically) defined key properties. It is frequently less helpful as a predictive tool in terms of assisting process optimization and informing the smelter in terms of how the alumina is expected to behave.
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This is then exacerbated by an incomplete understanding at a smelter level of how alumina properties impact on smelter performance. Good examples of this are: the impact of microstructure (particularly specific surface area and pore size distribution) or phase distribution between particle size classes on HF generation, dissolution and dry scrubbing.
The Future of Alumina Calcination With the increasing number of green- and brownfield alumina refineries it is no surprise that research today focuses primarily on production gains and energy savings through process and control improvements. Typical energy consumption is around 11.6 GJ per metric ton of A1203. However, the broad range of energy consumption across the industry [31], indicates that there is still room for significant energy savings in a number of operations and process areas. As the higher grade bauxite reserves are being depleted the industry increasingly turns to more energy intense and difficult to process diasporic and boehmitic ores.
These properties are primarily influenced by precipitation and calcination strategies but these relationships are complex. Most critical is the "averaging" impact of numbers reported in the specifications sheet where the outlier populations may be the dominant contributor to process fluctuations and instabilities in the smelter. References 1.
In the past decades alumina production volumes have increased dramatically, with calciner operating capacities up to 4000 metric tonnes per day. As a result of the push towards faster and more energy efficient, but also more abrasive, stationary calcination processes, attrition continues to receive a lot of attention [32, 33]. Many options and improvements have been developed to the calcination technologies to reduce breakage and improve product quality. Not all, but some of these improvements come with the price of higher operating costs. Thus research programmes are increasingly focusing on a better understanding of the role of the precipitation step and resulting particle morphology and gibbsite strength, for the alumina attrition behaviour during calcination and product handling. There is also a wider consensus emerging that new definitions for how attrition is defined, and indeed measured, are needed, as again the specification sheet is not always informative as to smelter experience. The next stages of this discussion should include the processes downstream of calcination in alumina handling, and also at the smelter to get a more holistic view. This informs the optimum compromise between costs, product quality and stable operation.
2. 3. 4. 5. 6.
7.
In the smelter surprisingly few studies systematically examine the impact of alumina quality (or properties) on operations, although the impact of alumina on HF emissions is an area of increasing
8.
154
Edwards, J.D., F.C. Frary, and Z. Jeffries, The Aluminum Industry - Aluminum and Its Production. Chemical Engineering Series. 1930. Wefers, K. and C. Misra, eds. Oxides and Hydroxides of Aluminum. Alcoa Technical Paper No. 19. 1987, Aluminum Company of America: Pittsburgh, PA. Andrews, E.W., A Controllable Continuous Mass-feed System for Aluminium Smelters. PhD Thesis, The University of Auckland, New Zealand, 1998. Hudson, L.K., ed. Alumina Production. Alcoa Research Laboratories. 1982, Aluminum Company of America: Pittsburgh, PA. Jenkins, B. and C. Bertrand, Improvements in the Design and Operation of Alumina Flash Calciners. IFRF Combustion Journal, 2001 (November): p. 1-18. Hiltunen, P., R. Bligh, C. Klett, M. Missalla, and H.-W. Schmidt, How to achieve high availability with large calciners and avoid unforeseen downtime. Light Metals (Warrendale, PA, United States), 2008: p. 63-68. Raahauge, B.E., Advances in gas suspension calcination technology. Aluminium (Isernhagen, Germany), 2007. 83(1/2): p. 40-42. Mclntosh, P., R. Greenhalgh, and P. Mills, Advanced control techniques for alumina calcination rotary kilns.
9.
10.
11.
12.
13.
14. 15.
16.
17.
18.
19.
20.
21.
22.
Light Metals (Warrendale, PA, United States), 1987: p. 59-63. Ingram-Jones, V.J., R.C.T. Slade, T.W. Davies, J.C. Southern, and S. Salvador, Dehydroxylation sequences of gibbsite and boehmite: study of differences between soak and flash calcination and of particle-size effects. Journal of Materials Chemistry, 1996. 6(1): p. 73-9. Whittington, B. and D. Ilievski, Determination of the gibbsite dehydration reaction pathway at conditions relevant to Bayer refineries. Chemical Engineering Journal (Amsterdam, Netherlands), 2004. 98(1-2): p. 89-97. Metson, J., T. Groutso, M. Hyland, and S. Powell, Evolution of microstructure and properties ofSGA with calcination of Bayer gibbsite. Light Metals (Warrendale, PA, United States), 2006: p. 89-93. Yamada, K., T. Harato, S. Hamano, and K. Horinouchi, Dehydration products of gibbsite by rotary kiln and fluid calciner. Light Metals (Warrendale, PA, United States), 1984: p. 157-71. Levin, I. and D. Brandon, Metastable alumina polymorphs: crystal structures and transition sequences. Journal of the American Ceramic Society, 1998. 81(8): p. 1995-2012. Bennett, I. and R. Stevens, Calcination and phase changes in alumina. British Ceramic Transactions, 1998. 97(3): p. 117-125. Loh, J.S.C., A.M. Fogg, H.R. Watling, G.M. Parkinson, and D. O'Hare, A kinetic investigation of gibbsite precipitation using in situ time resolved energy dispersive x-ray diffraction. Physical Chemistry Chemical Physics, 2000. 2(16): p. 3597-3604. Latella, B.A. and B.H. O'Connor, Detection of minor crystalline phases in alumina ceramics using synchrotron radiation diffraction. Journal of the American Ceramic Society, 1997. 80(11): p. 2941-2944. O'Connor, B., D. Li, B.K. Gan, B. Latella, and J. Carter, Time-resolved studies of alumina ceramics processing with neutron and synchrotron radiation data. Advances in X-Ray Analysis, 1999. 41: p. 659-667. Neissendorfer, F., U. Steinike, B.P. Tolochko, and M.A. Sheromov, On the decomposition of hydrargillite investigated by synchrotron x-ray diffraction. Nuclear Instruments & Methods in Physics Research, Section A: Accelerators, Spectrometers, Detectors, and Associated Equipment, 1987. A261(l-2): p. 219-20. Paglia, G., CE. Buckley, A.L. Rohl, R.D. Hart, K. Winter, A.J. Studer, B.A. Hunter, and J.V. Hanna, Boehmite Derived Gamma-Alumina System. I. Structural Evolution with Temperature, with the Identification and Structural Determination of a New Transition Phase, Gamma'-Alumina. Chemistry of Materials, 2004.16(2): p. 220-236. Zhou, R.S. and R.L. Snyder, Structures and transformation mechanisms of the eta, gamma and theta transition aluminas. Acta Crystallographica, Section B: Structural Science, 1991. B47(5): p. 617-30. Billinge, S.J.L. and I. Levin, The Problem with Determining Atomic Structure at the Nanoscale. Science (Washington, DC, United States), 2007. 316(5824): p. 561-565. Ashida, T., J.B. Metson, and M.M. Hyland, New approaches to phase analysis of smelter grade
23.
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aluminas. Light Metals (Warrendale, PA, United States), 2004: p. 93-96. Patterson, E.C., Hydrogen Fluoride Emissions From Aluminium Electrolysis Cells. PhD Thesis, The University of Auckland, New Zealand, 2002. Jayaram, V. and C.G. Levi, The structure of deltaalumina evolved from the melt and the gamma -> delta transformation. Acta Metallurgica, 1989. 37(2): p. 56978. Wilson, S.J., The dehydration of boehmite, γ-ΑΙΟΟΗ, to γ-alumina. Journal of Solid State Chemistry, 1979. 30(2): p. 247-55. Wilson, S.J. and J.D.C. McConnell, A kinetic study of the system boehmite/alumina (γ-ΑΙΟΟΗ/Α1203). Journal of Solid State Chemistry, 1980. 34(3): p. 315-22. Wolverton, C. and K.C. Hass, Phase stability and structure of spinel-based transition aluminas. Physical Review B: Condensed Matter and Materials Physics, 2001. 63(2): p. 024102/1-024102/16. Perander, L., C. Klett, H. Wijayaratne, M. Hyland, M. Stroeder, and J. Metson. Impact of Calciner Technologies on Smelter Grade Alumina Microstructure and Properties, in Proceedings of the 8th International Alumina Quality Workshop. 2008. Darwin, Australia. Perander, L.M., Z.D. Zujovic, T.F. Kemp, M.E. Smith, and J.B. Metson, The Nature and Impacts of Fines in SGA. Journal of Metals, 2009. 61(11): p. 33-39. Perander, L.M., M.A. Stam, M.M. Hyland, and J.B. Metson, Towards Redefining the Alumina Specifications Sheet - The Case of HF Emissions. Light Metals (Warrendale, PA, United States), 2011. Henrickson, L., The need for energy efficiency in bayer refining. Light Metals (Warrendale, PA, United States), 2010: p. 173-178. Klett, C , M. Missalla, and R. Bligh, Improvement of product quality in Circulating Fluidized Bed calcination. Light Metals (Warrendale, PA, United States), 2010: p. 33-38. Wind, S., C. Jensen-Holm, and B.E. Raahauge, Development of particle breakdown and alumina strength during calcination. Light Metals (Warrendale, PA, United States), 2010: p. 17-24. Iffert, M., M. Kuenkel, M. Skyllas-Kazacos, and B. Welch, Reduction of HF emissions from the trimet aluminum smelter (optimizing dry scrubber operations and its impact on process operations). Light Metals (Warrendale, PA, United States), 2006: p. 195-201. Dando, N., X. Wang, J. Sorensen, and W. Xu, Impact of thermal pretreatment on alumina dissolution rate and HF evolution. Light Metals (Warrendale, PA, United States), 2010: p. 541-546. Liu, X., S.F. George, and V.A. Wills, Visualization of alumina dissolution in cryolitic melts. Light Metals (Warrendale, PA, United States), 1994: p. 359-64. 0stb0, P., N, Evolution of Alpha Phase Alumina in Agglomerates upon Addition to Cryolitic Melts. PhD Thesis, Norwegian University of Science and Technology, Trondheim, Norway, 2002. Welch, B.J. and G.I. Kuschel, Crust and alumina powder dissolution in aluminum smelting electrolytes. Journal of Metals, 2007. 59(5): p. 50-54.
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
Significant Improvement of Energy Efficiency at Alunorte's Calcination Facility Michael Missalla1, Hans-Werner Schmidt1, Joaquim Ribeiro Alves Filhio2, Reiner Wischnewski2 outotec GmbH, Ludwig-Erhard-Str. 21, 61440 Oberursel, Germany 2 Alunorte-Alumina do Norte do Brasil S.A., Rodovia PA 481, km 12, Distrito de Murucupi, CEP 68 447-000 Barcarena, PA Brazil Keywords: Circulating Fluidized Bed, Calcination, Energy Efficiency, Product Quality, Particle Breakage Abstract The Alunorte refinery produces 6.3 million t/a of alumina with seven Circulating Fluid Bed (CFB) Calciners. The calcination facility needs about 3 GJ per ton of alumina of the total energy of 8 GJ per ton for the refinery. Until the introduction of CFB Calciners by Outotec (formerly Lurgi) in 1960, rotary kilns were the standard technology for the calcination of alumina. Since then, stationary calciners such as CFB s are the preferred technology for new installations due to their superior energy efficiency and uniform product quality.
To achieve significant lower energy consumption than rotary kilns, CFB Calciners utilize much more energetic means of gassolids contact and gas solid separators (such as cyclones) with significantly higher velocities and hence strong mechanical stress on the particles. With rising energy prices it became also more favorable to increase the number of heat exchanger stages. The number of heat exchanger stages it is very important to have an efficient particle mixing into the gas stream [9]. Further a high separation efficiency of the cyclones in the heat exchanger stages is necessary to achieve low energy consumption.
The CFB calcination system has implemented several preheating and cooling stages with cyclones for the separation of gas and solids. The inefficiency of these cyclones is leading to a reduction in the heat recovery and increased energy consumption.
The cyclone inlet velocity can be increased to achieve improved separation efficiency. However because of attrition, this leads to higher generation of fines in the process and has the result that the hydrate produced by the alumina refinery must be stronger and contain less fines in order to meet the final product quality specification. While this has certainly been achieved in many refineries, nevertheless, there are additional constraints placed on refinery performance indicators (e.g. yield) [11].
The paper presents a procedure to improve the efficiency of these cyclones and the solid dust load significantly by using a new method of simulation technology. With the optimized process considerably lower energy consumption figures are achieved. The results of these improvements regarding fuel and electrical energy consumption, as well as return on investment are presented. Alunorte has installed the improvements already at two of their seven Calciners and has received the energy efficiency award 2010fromthe German Energy Agency.
Since the introduction of CFB Calciners, a major part of Outotec's calcination research has been dedicated to reducing fines generation without compromising on energy consumption or other aspects of product quality (e.g. attrition index)[2]. While Outotec has been steadily improving its technology to reduce fines generation for many years, the latest installations represent a significant step improvement, with the result that the particle breakage in CFB Calciners has been reduced to only marginally higher than in rotary kilns and much lower than in earlier installations. Also the energy consumption has been reduced significantly and the latest alumina calciner installed in Alunorte has received the energy efficiency recognition prize at the industrial technology exhibition the Hannover Messe 2010 in Germany.
Introduction Product quality is always a major focus for every industrial operation, with only production safety being on the same level or even more important. This applies to the production of alumina calcined from Aluminium Hydroxide with the same intent. The calcination step is the last step in the chain of process steps to produce alumina from bauxite. Until the 1960s the calcination reaction was always performed in rotary kilns. In 1961 Outotec (then known as Lurgi) invented the circulating fluidized bed technology, and applied it to the calcination of alumina in CFB Calciners. With this new technology, energy consumption for alumina calcination was immediately cut by approx. 30% and fluidized bed has been the technology of choice ever since.
The following main improvements have led to the energy reduction in the calcination process: • •
The new technology had quite some impact on product quality. While criteria such as LOI, BET, pick-up of iron and silica kept constant, the production of alpha alumina was significantly decreased while fines generation was increased. Although reduction of alpha alumina is basically considered to be an improvement in quality, increased fines is not. The mass fraction of particles smaller than 45μπι is a common measure of fines content and should be as low as possible but certainly within the limits specified by the smelter customers.
•
Improved Cyclone design to reduced the recirculation of fines and dust in the calcination system Improved Cyclone design to reduce the generation of fines Optimization of process stability for stable operating conditions Improvements at Alunorte
Alunorte S.A. was founded in 1978. Production started in 1995 and since then its alumina production capacity increased continuously [14]. Currently Alunorte S.A. is the world largest alumina refinery with a total production capacity of 6.3 million tones of alumina per year.
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hot gas. In the cyclone, which is arranged downstream of the venturi preheater, the flow of gas and solids are separated. From the cyclone the pre-calcined alumina is delivered into the fluid bed furnace through the material feeding line. The waste gas leaving the cyclone of the second preheating stage is conveyed to the first preheating stage where it assumes the above-mentioned function. The final calcination of the preheated and partly dehydrated hydrate takes place in the fluid bed furnace as shown in figure 3.
Figure 1: Alunorte S.A. 6.3 million tons per year alumina refinery Figure 1 shows the alumina refinery of Alunorte S.A. located in Barcarena, Brazil. With its latest expansion Alunorte S.A. added two alumina calcination plants from Outotec each with a name plate capacity of 3300 tons per day (tpd).
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Figure 3: CFB Calciner Flowsheet of recently installed calcination plants at Alunorte S.A. The required heat for calcination is generated by direct combustion of fuel oil in the fluid bed. Part of the air required for combustion is introduced through the nozzle grate as primary air, and the remaining air is added above the grate as secondary air. Due to the intensive mixing and heat exchange that takes place in the fluidised bed, the furnace temperature adjusts itself as a mixed temperature between the combustion temperature and the solids temperature, and is kept steady at a pre-set level. In the lower furnace zone, between the grate and the secondary air inlet, a fluid bed of high solids concentration is adjusted. It favors the combustion of the fuel and increases the mean retention time in the calcining furnace. In the upper furnace zone the internal recirculation of the solids causes a continuous reduction of solids concentration until a relatively low concentration is reached. With this solids concentration, the hot gases enter the recycling cyclone where they are separated from the solids. The hot alumina, which is separated in the recycling cyclone, passes through the seal pot and re-enters the fluid bed furnace [12],[13].
Figure 2: View of the seven CFB calciners at Alunorte S.A. In figure 2 the seven alumina calcination plants of Alunorte S.A can be seen. The wet aluminium trihydrate is conveyed by belt conveyors to the hydrate feed bin of the calcination plant. From the hydrate feed bin, the hydrate is discharged by a screw feeder, which delivers the material into the venturi preheater of the first preheating stage. There, the solids are mixed with the waste gas, which leaves the cyclone of the second preheating stage. The heat contained in the waste gas evaporates the entire surface moisture of the hydrate.
The recirculation of the solids leads to uniform and practically identical product and gas temperatures in the entire calcination stage, which consists of the fluid bed furnace, recycling cyclone and seal pot. The alumina, which is discharged from the calcining stage is cooled in two direct cooling stages each of which consists of a liftduct and a secondary air cyclone. The third cooling stage is designed as a fluid bed cooler and mainly relies on indirect heat transfer. The alumina coming from the calcining stage is first mixed with pre-heated air in the liftduct and conveyed into the secondary air cyclone 1. At the same time, the air is heated from the temperature of the second cooling stage to the temperature of the first cooling stage. This air is routed from the outlet of the secondary air cyclone 1 to the CFB system as secondary air and
Then the preheated hydrate entrained with the waste gas is conveyed into the two-stage electrostatic precipitator (ESP). The solids precipitated in the ESP are conveyed to the venturi preheater of the second preheating stage. The conveying air is dedusted in the airlift cyclone and delivered via ducts to the secondary air cyclone and used finally as combustion air. The hydrate entering the venturi preheater of the second preheating stage is mixed with the hot waste gas leaving the recycling cyclone and is dehydrated by the heat contained in the
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To calculate the grade efficiency curve for a cyclone, a model was used which is shown in figure 4. The cyclone is divided into different separate regions which are:
used for combustion. The same applies for the second direct cooling stage. In order to achieve low specific heat consumption during the calcining process, there is further heat recovery in the air-cooled part of the fluid bed cooler. Primary air is pre-heated by indirect heat exchange in a counter-current flow to the alumina. Secondary air is also preheated by direct contact with the alumina. Further heat recovery occurs in the water-cooled section of the fluid bed cooler as secondary air is preheated in the air-cooled section. Any part of the heat content of the alumina, which cannot be utilized for preheating of air or liquid media is removed by means of cooling water.
o o o o o
For the above mentioned regions, differential particle mass balances are solved to calculate the grade efficiency curve for the cyclone. The model considers the different separation efficiencies for the main flow V and the secondary flow Vsek over the top and the wall of the vortex finder. In the model the re-entrainment of already separated particles at the solid outlet is also considered at the re-entrainment region 3.
With the process described above the specific energy consumption at Alunorte has been reduced to 2790 kJ/kg Alumina.
Australian Average
3,900
Previously lowest recorded value m World
I
3,000
1,000
1,500
2,000
2,500
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3,500
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kJ/kg Alumina
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Cyclones are firstly named in 1886 in a patent of Knickerbocker Company, USA [5]. Since then the design and calculation of cyclones have been improved significantly. Basically cyclones are used to separate particles from a fluid or gas by centrifugal force. Therefore the dust loaded gas stream is fed into the cyclone where a rotary motion is induced. The solid particles, due to their higher density, are separated to the wall of the cyclone and flow downwards to the solid exit, while the gas leaves the cyclone via the vortex finder. The flow field in a cyclone is highly complex and can be differentiated in:
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w
m
In Alunorte S.A. the outstanding performance with respect to the specific energy consumption was reached by improved design of the cyclones which are used for the separation of the solids from the offgas and the combustion air as described above. Furthermore, the process stability was improved significantly which also contributed to the reduction in specific energy consumption.
•
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if I
Figure 3: Specific energy consumption according to CSIRO National Framework for Energy Efficiency, 2007 including Alunorte S.A. calciner.
• •
Inlet region e Separation region 1 and 11 Secondary Flow regions at the top d and at the vortex finder tr Separation region 2 and 4 Re-entrainment region 3
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In CFB calciners as well as in flash calciners the typical solid loading in the preheating and cooling cyclones inlets is high. Due to the very high solids concentration, it dominates the flow field in the cyclone. It has been found that with higher solid loading the tangential velocity decreases significantly in a cyclone [8].
strand formation. Figure 6 shows the effect of the formation of strands in the inlet of a cyclone. It can be seen that increasing the load in the inlet of the cyclone enhances the separation due to the formation of strands. loo i 0.90
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Figure 7: Calculated and measured particles in the offgas of a cyclone at the Alunorte calciner Figure 5: Relative tangential velocity of a loaded to an unloaded cyclone as function of the solids loading μ* in the inlet stream
Figure 7 shows the measured particle size distribution in the inlet as well as in the offgas of a cyclone in the Alunorte calcination plant. It can be seen that the measured particle size distribution fits quite well with the calculated curve from the model, thus allowing the optimization of the cyclones.
Figure 5 shows that the tangential velocity of a cyclone with high solid loads in the inlet drops significantly. A cyclone with a solid load of μ*=1 in the inlet has just 60 % of the tangential velocity of a cyclone with no solid load (μ*=0) in the inlet. Thus it becomes very difficult to separate fine particles since the tangential velocity is reduced substantially. Muschelknautz [10] showed that the separation in high loaded cyclones is mainly driven by the
Figure 6: Formation of Strands in a cyclone at different inlet loads: a) μ*=0,01 b) μ*=0,1, c) μ*=1, d) μ*=10 [10]
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The process flowsheet shown in figure 3 has four cyclones of which one is in the preheating stage for the hydrate, one is in the CFB and two are in the preheating stages for the combustion air. The integration of the ESP in the overall CFB process ensures that all dust is collected and returned to the process. In other stationary calciner systems the dust returns to the fluid bed cooler [9]. The disadvantage of the latter method is that it cannot be ensured that all the hydrate is completely calcined.
Operating Stability In order to achieve the best results for the separation in a cyclone the operating stability of the process becomes more and more important. Over the last few years, [3] [4] the stable operation of CFB calciners has been in focus and great improvements have been made up to the point of complete automation of the plants. Figure 8 shows the ability to maintain calcination temperature in the furnace under all circumstances. The modern automation approaches enable the operators to change the load of the plant in a flexible manner and also to keep the furnace temperature within arangeof+/- 10tol5°C.
According to the grade efficiency curve of a cyclone, certain fractions of a specific particle size will not be separated. Assuming two cyclones in series, as in the flowsheet in figure 3 for the secondary air cyclones, the fine particles not separated by the second secondary air cyclone serve as additional solid feed for the first secondary air cyclone. So the first secondary air cyclone is fed with the fresh solid feed from the CFB and with the fine particles from the second secondary air cyclone. Thus the efficiency of the first is reduced.
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Furnace Temperature
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Applying this logic further, then due to the fines in the offgas, the other upstream cyclones are reduced in their separation efficiency. In table 1 a comparison between CFB systems with two and three cyclone cooling stages with different total cyclone separation efficiencies ηιοί is made. A total separation efficiency of r|tot =100 % means that no particles are leaving the cyclone with the offgas while ritot = 90 % means that 10 % of the particles been fed to the cyclones are leaving it with the off gas.
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Table 1 : Specific energy consumption normalized to the system 1 with the highest energy consumption Number of Furnace Temp. °C Normalized nto %t Cooling Energy Stages Consumption 2 950 100 953% 90 96.1% 3 950 92.4% 100 93.4% 90 80 94.8%
The temperature stability in the CFB calciner is very important for the separation efficiency of heat exchanger cyclones in the flowsheet, since the separation efficiency is negatively affected by increasing temperature. Also an increase in particle breakage due to higher furnace temperatures can be observed [2]
* 45 S 40
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The results in table 1 show that cyclone efficiency is of utmost importance to reduce recirculation in the process and thus the energy consumption. Comparing a CFB system with two cooling stages and a ηιοί =100 % with a CFB system of three cooling stages with not optimal total cyclone separation efficiency of r|tot =80 % the result for the specific energy consumption is almost the same.
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200
· 400
600
800
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With the model as described in figure 4 the optimum vortex finder length can be found to increase the separation of fine particles. With increasing height H of the cyclone its separation efficiency is improved. Finally also increased recirculation of fines in a calcination process leads an increased amount of alpha alumina in the system. Due to its stickiness, this type of alumina can lead to processfluctuationswhich finally affect the separation efficiency of a cyclone negatively.
Figure 9: Friability of alumina calcined at different calcining temperature. From Figure 9, between 800 and 1000°C the particle breakage changes approx. 1.2% for every 10°C increase. With a value of 40% in the Saatci Test [2] at 1000°C this means a ratio of 3% on the actual breakage. For a plant with a normal particle breakage of 3.5% this would correspond to a variation of 0.1% for each temperature variation of 10°C. To control the particle breakage means to control the stability of furnace temperature.
The energy efficiency of a calcination process does not only depend on the number of cooling stages but also on its total separation efficiency. An investment comparison of the different system can only be done when the weights of the cyclones are known.
However the dependency could also be more than proportional, which would mean a disproportionate increase of breakage with increasing temperatures. In this case the fluctuations of calcination temperature around the set-point would be a
161
demonstrates Outotec's capabilities to design optimum calcination technology for energy efficiency and product quality.
disadvantage to the overall breakage, as every increment of temperature would cause more damage than a decrease could compensate. The effect of increased particle breakage in the calciner and thus the generation of more fines is an increased specific energy consumption due to the negative effect on cyclones performance.
References [I] J. D. Zwicker, "The Generation of Fines due to Heating of Alumina Trihydrate", TMS Light Metals (1985)
Future Developments
[2] S. Saatci, H.W.Schmidt, W. Stockhausen, M.Ströder, P.Sturm, " Attrition behaviour of Laboratory calcined Alumina from various Hydrates and its influence on SG Alumina Quality and Calcination Design", TMS Light Metals pp 81-86 (2004)
Beside designs for low fuel consumption [6], [7] the existing plants can be further improved by adding a Hydrate Bypass as shown infigure9.
[3] P. Hiltunen, R. Bligh, C. Klett, M. Missalla, H.-W. Schmidt, "How to achieve high availability with large calciners and avoid unforeseen downtime", TMS Light Metals pp 63 - 68, (2008) [4] M. Missalla, J. Jarzembowski, R. Bligh, H.-W. Schmidt, "Increased availability and optimization of calciner performance due to automation", TMS Light Metals , (2009) [5] Knickerbocker Company, " Staubsammler", in DRP Nr. 39219 (1886)
Primary Air
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[6] C. Klett, M. Missalla, R. Bligh, M. Stroeder, P. Hiltunen, M. Graham, "Outokumpu Technology's State-of-theArt CFB Calciners - A Review", Proceedings to EMC (2007)
»
Cooling Wartw - # »
1
Figure 10: CFB Calciner Flowsheet with Hydrate Bypass
[7] C. Klett, M. Missalla, R. Bligh, "Improvement of Product Quality in Circulating Fluidized Bed Calcination", TMS Light Metals pp 33 - 38, (2010)
Figure 10 shows a CFB Calciner Flowsheet as it is installed at Alunorte S.A. but with the addition of a Hydrate Bypass. The upgrade with a hydrate bypass will further decrease the specific energy consumption of the already for its lowest consumption awarded plant. With the hydrate bypass pre-dried hydrate from before preheating stage 2 will be bypassed around the CFB to a mixing pot. In the mixing pot the hydrate is calcined with by the hot alumina leaving the CFB. Beside using the energy from the CFB the efficiency of the cooling cyclones is improved since they are operating with decreased temperature.
[8] M. Missalla, "Calculation Method for Highly Loaded Cyclones", Ad.libri Hamburg, Dissertation Technische Universität Braunschweig (2009) [9] S. Wind, B.E. Raahauge, "Energy Efficiency in Gas Suspension Calciners (GSC)", TMS Light Metals pp 235 - 240, (2009) [10] U. Muschelknautz, „Hart bleiben Verschleißschutztechnik in schüttguttechnischen Anlagen" Chemie Technik Nr. 10 (2005), pp 14-18
However, the by-pass is a major part of the energy saving methodology and can operate to up to 15% of the total dry hydrate flow. As a result, a few effects happen simultaneously. Firstly, only the remaining part of the hydrate flow will actually pass through the CFB and undergoes breakage there. Secondly the recombined flows of alumina originating from the furnace material and from the by-pass material have now been calcined at different temperatures. According to the temperature effect from above, the by-pass material is significantly less brittle and will lead to a reduction of breakage in the cooling stages. This effect is even large enough to compensate for any temperature increase in the furnace to achieve the target LOI. The specific energy consumption is improved by approx 3-5%
[II] E.Guhl, R.Arpe, „Nearly 30 years of experience with Lurgi Calciners and influence concerning particle breakage", TMS light Metals (2002) [ 12] L.Reh," New and Efficient High Temperature Process with Circulating Fluid Bed Reactors", Chem. Eng.Technol. 18 (1995) page 75-89 [13] A. Squires, "Origins of the Fast Fluid Bed, Advances in Chemical Engineering", Vol. 20, Fast Fluidisation (Ed. Kwauk, M.), Academic Press (1994), pp 4-35
Conclusions
[14] Daryush Albuquerque Khoshneviss, Luiz Gustavo Correa, Joaquim Ribeiro Alves Filho, Hans Marius Berntsen, Ricardo Rodrigues de Carvalho, "Alunorte Expansion 3 - The New Lines Added to Reach 6.3 Million Tons per Year", TMS Light Metals 2011.
With the recent research activities and developments Outotec has made a significant effort to understand cyclone design, process stability and particle breakage in CFB Calciners. The recent success with the latest CFB Calciners Outotec has received the energy efficiency award 2010fromthe German Energy Agency. It
162
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
ATTRITION OF ALUMINA IN SMELTER HANDLING AND SCRUBBING SYSTEMS Stephen J. Lindsay Alcoa, Inc.; Primary Metals; 300 N. Hall Rd. MS S-29, Alcoa, Tennessee, 37701-2516, USA Keywords: Alumina, SGA, Attrition, Attrition Index, Gas Treatment Center, GTC Abstract
Alumina Attrition Mapping
Customers place various levels of importance upon the Attrition Index of Smelter Grade Alumina. The concerns are generally associated with the content of fines in fluorinated alumina as it arrives to the reduction cell. In this paper the author discusses factors of importance related to the design and operation of alumina handling systems from the refinery to the reduction cells. Examples are given in which the actual attrition of alumina particles has been minimized. Techniques are shared on how to separate the contribution of fine particles of bath evolved by the pots from the attrition of alumina itself.
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Introduction Many producers of smelting grade alumina, SGA, include data on % Attrition Index, or %A.L, in each certificate of analysis. This is one metric for assessing alumina particle "toughness". Although suppliers maintain strict standards to determine values of attrition index, the methods used to determine %A.I. do vary. At this time there is no single industry standard for measurement of % Attrition Index.
Figure 1 - Generic example of Attrition Mapping A sampling regimen was prescribed for the participating smelters. The goal was to be able to gather composite samples that would be as representative as possible of alumina at various locations along its flow path toward the pots. Some of these were straightforward sampling plans that focused on a few key sampling locations. Other sampling plans were more detailed. Complexity beyond the basic sampling points was left to each participating location.
There have been contributions to the literature that help to illustrate the topic of attrition, or particle toughness. Particular significance has been given to particle velocity at the time of impact against an obstacle such as a steel plate [1]. But, what does the %A.I. mean to a smelting customer of SGA? Can it be used with the particle size distribution, PSD, of an alumina source to predict the PSD of secondary alumina at the reduction cell? Is it of greater significance for some screen mesh fractions than others? By definition %A.I. is a measure of change in the %+325 mesh. But, is it meaningful for the %-20 micron fraction, or superfines?
Analysis of the gathered samples was made in cooperation with the Technology Delivery Group of Alcoa World Alumina. This included the determinations of particle size distributions using a Coulter Particle Size Analyzer model LS100Q laser diffraction instrument. Interpretation of superfines content was based upon electroformed sieve calibration samples that are periodically restandardized. Measurements of flowability were also made by use of a standard Alcoa flow funnel with a 6 mm orifice.
The best answer was probably inscribed on the portico of the Temple of Apollo at Delphi, "Know thyself. The particle size distribution at reduction cells appears to have more to do with the design and operation of alumina handling systems and dry scrubbers than with the source of the alumina and its particular morphology or particle toughness. When a smelting customer understands their system quite clearly, then some detailed meaning may be able to be ascribed to the %A.I. of an individual alumina source.
This analytical capability allowed the study to go beyond other studies in the literature that have relied upon dry sieving analytical methods. This enabled many of the conclusions that follow on the nature of pot fume and of the amount of attrition that occurs in various types of alumina handling and processing equipment. Results were summarized according to typical particle size fractions. These were mapped to visually illustrate changes in the percentages of each fraction along the alumina flow path. Particular attention was given to superfine fractions. This approach was used to enable conclusions about alumina attrition. It was understood from the outset that increases in superfine fractions might be confounded by thefineparticles of alumina and bath that are found in reduction cell exhaust.
Experimental Design In mid-2008 more than a dozen Alcoa smelters participated in an "Attrition Mapping" exercise. The exercise is generally described by Figure 1. The activity was designed to allow its participants to understand how much alumina particles break down as they pass through their alumina handling and dry scrubbing systems.
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No effort was made to determine the chemical composition or structure of each sample or fraction. This additional dimension of study is left for future consideration. It was not within the boundaries of the test design for this Attrition Mapping exercise.
Sampling at or after point feeders in pots was not advised due to the impacts of dust collection and gas flow patterns in cells. This may have introduced significant bias in superfine fractions that could not be easily controlled by the experimental design.
The design of the study was to create composite samples from various places along the alumina flow path. The minimum requirement to participate was to sample:
Composite samples were then placed into plastic containers with screw down lids. They were clearly marked prior to shipment for analysis.
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Particle size determinations were all made in the same laboratory, using the same equipment, people, and analytical methodology.
Fresh alumina as received Fresh alumina prior to the inlet of the dry scrubber Secondary alumina after the discharge point of the dry scrubber silo Secondary alumina at, or in close proximity to, the reduction cell
Results Study #1 - The first handling system to be analyzed included a nearly ideal configuration for minimal attrition. Handling systems included; air gravity conveyors, silos, one air lift for fresh alumina and another for secondary alumina, a fluid bed type gas treatment center and air gravity conveyors to pot bins.
As indicated, additional sampling locations were added to this baseline if so desired by participating locations. Sampling
The %A.I. of the alumina source was between 15% and 17% during the period of study. This moderately high %A.I. was determined by the Alcoa method for attrition. See equation 1.
Obtaining representative samples of alumina in an industrial setting can be challenging. Fine and superfine fractions can easily be biased via improper sampling technique. As this study was conducted over an array of smelting locations care was taken to prescribe detailed sampling methods. However, direct oversight of sampling at each location was not conducted by any one person or group.
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Sampling was conducted by taking approximate 100 gm samples over multiple days to build composite samples of approximately 1 kg from a minimum of 10 grab samples. When possible these samples were to be taken directly from a conveyor belt or by using a simple grain thief device. Refer to Figure 2.
Results indicated no significant degradation of particle size diameters at the 10th, 50th or 90th percentiles of the PSD. These are otherwise referred to as dlO, d50 and d90. See Figure 3.
Figure 3 - Attrition mapping results of study location #1 Most surprisingly there was essentially no change in dlO at -40 microns diameter along the flow path to the pots. Calculation of the actual %A.I. in the field yielded a result of only 0.7%. This result was roughly 1/20* of the % A.I. reported for fresh alumina.
Figure 2 - Example of a grain thief sampling device It was advised to avoid taking samples from conveying equipment that was under negative pressure while using any open sampling devices such as scoops or cups. In these circumstances a sampling thief was to be used with a fabric seal around the tube at the sampling access point. An alternative approach was to briefly shut down the system to take a sample. This required shutting off dust collection simultaneously with conveying equipment.
Even though the handling systems and fume controls were known not to be harsh on alumina particles, such low results were not expected. There was no apparent attrition of alumina particles along the flow path through the smelter. However, these results were not in concert with the general observation of reduced
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In Study #1 48%+/-2% of the sub-2(ty material was sub-ΙΟμ in diameter and 30%+/-3% was s u b ^ in diameter. These proportions were consistent over the entire flow path from fresh alumina to the pots. Note that 72% of the sub-2(^ material found in alumina at the pot was originally present in the fresh alumina.
flowability according to Alcoa flow funnel test results for Study #1. See these results in Figure 4. Study #i; Alcoa Flow Funnel Time
Further, the data indicates that 28% of the sub-2(^ material at the pot has a similar PSD to the sub-2(^ material found in the fresh alumina. Note that alumina "at the pot" may be a bit of a misnomer. There were no attempts made to actually capture and analyze alumina that was being delivered to a feed hole in the pot crust. The reason is that some dust generated during feeding is lost to the pot exhaust and is not delivered into the pot per se.
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Study #2 - Attrition mapping Study #2 also had a nearly ideal configuration for minimal attrition. Handling systems with this study included; air gravity conveyors, silos, one air lift for fresh alumina and another for secondary alumina, a fluid bed type gas treatment center and air gravity conveyance to pot bins.
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As with Study #1 the dlO and d50 particle diameters remained essentially the same. However, there was a more pronounced difference at d90. Refer to Figure 7 to see this difference.
Figure 5 - SuperfinefractionsfromStudy #1 A number of interesting observations are associated with this data. There is an apparent increase in superfines content prior to the gas treatment center, or GTC. This may be attributed to factors such as: variation in sampling, segregation in the fresh alumina silo, or perhaps minor attrition in the airlift. Note that these flow path studies did not follow specific lots of alumina along the flow path. Samples were gathered at various points in the flow path all but simultaneously. The patterns of %-20μ, %-10μ and %-5μ in secondary alumina match each other quite well as is shown in Figure 5. This infers that the increases have been driven by the same mechanism. The most likely source of these particles is the fine particulate evolved with pot fume. This is illustrated in a typical PSD for fume particulate that is shown in Figure 6 [2].
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Study #3 - Attrition mapping Study #3 was also performed in a modern pre-bake smelter with various types of dense phase conveying systems in its alumina flow path. This location also uses injection type dry scrubbers. Delivery to the pots is via modern transport systems with low particle velocity.
As with Study #1 the distribution of superfmes in secondary, or reacted alumina, were similar to the PSD in the literature for fume particulate in the range of 2 to 50 microns. Refer to Figure 8 for the cumulative distributions of superfmes in fresh and secondary alumina.
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Calculation of actual %A.I. in thefieldwas 0.6% for Study #2. This is approximately l/25th of the % A.I. reported infreshSGA.
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These first two studies establish a baseline for what might be expected with a nearly ideal alumina handling configuration. "Ideal" implies a system with little pneumatic transport to move particles onward at any substantial velocity. See Figure 9 for the approximation of the PSD for fume particles in Study #2.
Note that the attrition index for this SGA is the lowest of the three studies presented thus far. However, actual attrition was more pronounced across the entire PSD. The primary difference at this location is the use of various dense phase transport systems over both long and short distances. Note the significant decrease in average particle diameter prior to fresh alumina arrival to the GTC. A short path dense transport system is used in conjunction with an airlift to move alumina from unloading into the main silo. This is followed by a dense phase transport system over a moderate distance to the fresh alumina silo of GTC#1. This dense phase system includes multiple 90 degree turns. Sub-4^ was increased by -6% in these systems.
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in superfine fractions may be attributed primarily to fine bath and alumina particles carried by the pot exhaust gas stream.
Passing through the GTCs there was only a minor increase in sub45μ, -2%, or 1.1% A.I. across the GTC. This was followed by a second dense phase system for fresh alumina from the fresh silo of GTC#1 to the fresh silo of GTC#2. This caused to another increase in 8υ^45μ of -4.5%.
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Unlike Figure 8, note that there is no convergence of the fresh and secondary PSDs above 45μ particle diameters in Figure 11. This implies that attrition of alumina particles is combined with fume particulate in the βυο-45μ fraction. Sorting out the relative contributions of alumina particles broken by attrition and that of pot fume can require chemical analysis under such circumstances. However, it is possible to draw upon the results of Studies # 1 & #2 to estimate thefractionthat originated with pot fume assigning the balance to the attrition of particles.
The shape of the curve for the injection type GTC in Figure 12 is similar to that of Study #3, but with lower overall percentages of superfines in alumina. The implication is that the energy imparted as alumina makes multiple passes through the process gas stream of the dry scrubber fractures particles. This is true even with a low value of 7.4% A.I. for SGA. The increase in the %-45 micron across the injection type GTC was 2.7% more than with the fluid bed type scrubber. While this amount of attrition is not abnormally high it does illustrate that various types of dry scrubbers in various operating conditions can also have a range of impact upon particle attrition.
For study #3 %A.I. in the field was 7.9% after GTC#1 and 12.8% after GTC#2. These are much closer to the value reported by the refinery for fresh alumina, 10.5%, than with the first two studies. The significance of this is that alumina handling system design is important to the actual amount of attrition observed in the field.
Fluid Bed type - Study #1 ~0%A.I. across the GTC Fluid Bed type - Study #2 ~0%A.I. across the GTC Injection type - Study #3 1.1%A.I. across the GTC Fluid Bed type - Study #4 0.6%A.I. across the GTC Injection type - Study #4 3.5%.A.I across the GTC
Study #4 - In all, more than a dozen attrition mapping studies were performed on both pre-baked and S0derberg smelters. A portion of Study #4 is included here to illustrate the impacts of various types of dry scrubbing systems. The location of Study #4 primarily utilizes conveyor belts, airlifts and air gravity conveyors to transport alumina. But, this location has two types of gas treatment centers. One is a fluid bed type and the other is a modern injection type dry scrubber.
Discussion Note that all studies that have been presented here are the product of multiple grab samples taken according to strict procedures and made into composites. In all cases composites were replicated at each sampling site between two and four times. Each composite was then analyzed using the same equipment in the same labs for determinations of particle size distribution and flowability.
The %A.I. of the alumina in this study, a fourth source, was 7.4% during the period of study as determined by the Alcoa method. The %A.I. in the field, across the dry scrubbers only, was 0.6% for the fluid bed type vs. 3.5% for the injection type. This result is consistent with similar comparisons that have been made at Alcoa locations that use two types of dry scrubbers.
There is margin for error in any industrial scale study. When dealing with fine granular material the risk of introduction of bias is high. This may come either from segregation issues in the handling system or with inadequate sampling technique. Precautions have been made in the presentation of this data to validate it against the findings of other sampling points along each alumina flow path and against other locations that utilize similar equipment or handling system configurations.
Refer to Figure 12. Note that increases in superfines in secondary alumina from the fluid bed scrubber closely mimic fresh vs. secondary alumina from Study #2. Studies #1, #2 and #4 all utilized identical fluid bed dry scrubbers. One again, the PSD converges with that of fresh alumina near particle size diameters of 5 and 50 microns. This appears to confirm that such increases
Conclusions
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2. Grjotheim and Welch "Aluminium smelter technology, 2nd edition, 1988, p. 202
Unfortunately, the output of this study cannot draw a specific conclusion about what may happen when the %A.I. of an SGA changes. Changes in the attrition index with any one source of alumina and changes in sources of alumina passed through alumina handling systems were not included in the scope of this study. At this point it is still uncertain that changes in the magnitude of %A.I. for an alumina source may be used to accurately predict changes in the PSD of alumina prior to the reduction cell.
Acknowledgements The author would like to acknowledge the following people for their insights and contributions: - Jack Sorenson, Warrick Operations (retired) for great dedication and perseverance to the study of alumina attrition and flowability. These studies could not have been completed without his invaluable assistance. - Travis Baroni, Alcoa World Alumina Technology Delivery Group for particle size analysis and support for this study. - Participants in Attrition Mapping studies from more than one dozen Alcoa locations worldwide. Their efforts have provided the raw materials to understand and confirm the results of this study. - The support of Alcoa Primary Metals and Alcoa World Alumina to share these findings. - Merino, Dr. Margarita R. (Ph.D. - Florida State University) - for her encouragement, dedication and support.
The major factor that impacts actual attrition of particles in the field is the configuration of an alumina handling system. It has been observed that Actual %A.I. at a smelter may be roughly equal to the %A.I. reported by a refinery. It may also be much lower with Actual %A.I. being only 1/25* of the value reported by the refinery. Of particular concern are pneumatic transport devices that may rapidly move air to move alumina. In situations such as those described in Study #3 there were high rates of attrition at dlO, d50 and d90. In such cases it appears that %A.I. has similar general meaning for a wide range of particle diameters, not only at 45 microns. Differences in particle attrition were also observed between fluid bed and modern injection type dry scrubbers. This was not unexpected, but some systems did demonstrate higher rates of attrition than others, even when the %A.I. of the fresh alumina was relatively low. Some of these differences may be ascribed to system operating conditions and some may be ascribed to GTC design factors. The most common system presented here, fluid bed GTCs with low impact handling systems, did not produce significant amounts of particle attrition. These type systems were observed across a wide range of attrition indexes, from 7.4% to 17%, as measured by the Alcoa method for %A.I. The comparison of fresh and secondary alumina from Study #2, shown in Figure 8, is typical for low impact handling system configurations. The differences between these two distributions produce a sub-distribution shown in Figure 9 that is quite similar to the literature reference provided for fume particulate that is shown in Figure 6. With low impact systems it appears that the primary contributors to changes in PSD are fine bath and alumina particles carried by the pot exhaust gas. Observations indicate that fume particulate has its greatest impact in the sub-40 micron range. Thus, standard methods for measurement of %A.I. will not be easily extrapolated into superfine particle diameters. With the wide range of alumina handling and dry scrubbing systems that exist in our industry it is most important to "Know thyself before taking differences in the %A.I. for various sources of SGA into serious consideration. References 1. Audet, D., Clegg, R.E., "Development of a New Attrition Index Using Single Impact", Proceedings of the 8th International Alumina Quality Workshop, September, 2008, pp. 117-120
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Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
Light Metals 2011 ALUMINA and BAUXITE
Energy and Environment SESSION CHAIR
Benny Raahauge F. L. Smidth Copenhagen, Denmark
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
PERSPECTIVE ON BAYER PROCESS ENERGY D.J. Donaldson P.E, ChE 1416 Village Center Drive Medford, Oregon 97504 Keywords: Bayer Process, Energy Pintle
Abstract The three most important cost items in the production of alumina are bauxite, caustic soda, and energy. Alumina energy cost will likely rise more than other costs because of energy price increases and energy related environmental issues. As energy cost increases, refineries will need to adjust the process to maintain or improve their position on the world cost curve.
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INTRODUCTION Energy is one of the three largest costs in the production of alumina from bauxite. In the future, energy could be two or three times today's cost (relative to other costs) because of availability and environmental issues. The better the understanding of the factors that make up the usage of energy in the alumina refinery, the better prepared process engineers and designers will be to evaluate refinery design changes in event a significant change in energy supply price or availability.
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This paper on energy in the Bayer Process considers only the interrelated areas of Powerhouse (Boilers), Digestion, and Evaporation. Excluded are Calcination and peripheral energy use areas such as Seed Wash, Liquor Purification, Causticization, etc. As shown in Figure 1, the source of energy to the Bayer Process is the Powerhouse. Energy is distributed essentially as steam to Digestion and Evaporation and electrical power to the refinery. The amount of energy required is determined by four interrelated factors:
Digester temperatures range from about 140 °C for trihydrate alumina (THA) bauxite up to about 280 °C for monohydrate alumina (MHA) bauxite. Because of the flash cooling and feed slurry heat recovery system used in Bayer Process refineries, the Digester temperature itself does not have a significant affect on Boiler Steam usage as long as there are no other streams entering or leaving (e.g., Blow Off). That is, theoretically, it will take nearly the same amount of Boiler Steam for a high temperature digest as for a low temperature digest. This is illustrated in Fig. 2.
1. Type of Bauxite. Trihydrate bauxite (low temperature digest or monohydrate bauxite (high temperature digest).
In practice there are other heat demands that do require additional Boiler Steam to Digestion:
2. Productivity. The amount of alumina produced for each unit volume of liquor circulating (see Fig. 1)
Heat of Reaction (HR). This is the endothermic heat of caustic attacking bauxite alumina hydrate to put the hydrate in solution. The HR is higher for THA than for MHA.
3. Evaporation. The facility that removes the net amount of water added to the process.
Sensible Heat Loss (SHL). This loss will be higher for MHA digestion because there will be over three times as many flash tanks (FT) and heat exchangers (HX) of which 2/3 will be at significantly higher temperatures than for THA Digestion.
4. Refinery Facilities. The physical design of the refinery with respect to heat recovery facilities.
Blow Off Vapor. By design, Blow Off vapor is kept small because it is a direct loss of steam (heat) to the atmosphere.
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Following is an example of the distribution of Boiler Steam in a flash/heater system. The example does not cover the whole range of refinery designs but it does indicate the relative importance of the factors listed above. % Distribution FT/HX (See Fig. 2) Heat of Reaction Sensible Heat Loss Blow Off Washer Overflow
evaporation (Digestion plus Evaporation) in a THA digest refinery would require a large Evaporator facility at about a 3.6 economy and much more total Boiler Steam than the comparable MHA refinery.
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The thermal energy required for Digestion and Evaporation in a MHA Digestion refinery will be less than for a comparable THA digestion refinery because of the higher average efficiency (economy) for total evaporation.
MHA digest 275°C -110°C = 165 °C evaporation THA digest 140°C -110°C = 30° C evaporation
When energy required for refinery electric power is included, the THA Digestion refinery will have a significantly lower electric power requirement. MHA Digestion will require more electric power in order to raise the bauxite slurry to digest pressures as high as 80 atm. (plus the frictional pressure drop in the large number of heat exchangers) compared to THA Digestion pump pressure of less than 10 atm. This will raise the MHA refinery electric power 10% -15% higher than the THA refinery.
For the 275 °C MHA Digestion, the evaporation economy (Fig. 2) is near 7 tons water/ton of steam. To achieve the same total
However, that is not the most important difference in providing energy for electric power. With a 60 atm Boiler in the THA
THA and MHA Digestion widely differ in the amount of evaporation that occurs. For instance, compare the 275°C digest (MHA) to a 140°C digest (THA) with the same product dissolved solids concentration where the temperature difference is indicative of the amount of evaporation:
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process, electric power can be generated by first passing Boiler Steam through an extraction turbo-generator. In a MHA process an extraction turbo-generator can be used only to a limited extent, and most, if not all, electric power will be generated by condensing turbo-generators at nearly three times the energy required by extraction turbo-generators. Another factor is the type of fuel supply. The discussion so far has a Powerhouse design of any fuel (coal, oil, natural gas) generating steam to process and generating electric power by extraction or condensing turbo-generators. In the case of natural gas fuel, electric power can be generated by gas turbo-generators with waste heat boilers for steam supply. Natural gas fuel to the Powerhouse can avoid the use of condensing turbo-generators for electric power and provide power at the same low energy usage of extraction turbo-generators.
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Productivity or Yield is important in design of a refinery. A productivity increase from 75 to 90 gpl alumina is an increase of 20% in production for the same pipes, pumps, and vessels in the refinery. This is very significant in both new and expanding refineries. Higher productivity will give an overall lower Boiler Steam and electric power demand per ton of alumina produced. However, for MHA Digestion, there will be proportionately lower evaporation in Digestion and more mud wash water which will require more evaporation in the Evaporator facility at a lower economy than Digestion. The heat of reaction (HR) per ton of alumina will be the same, but the sensible heat loss in Digestion will be proportionately lower. Overall, perhaps a 10% decrease in Boiler Steam usage per ton alumina for a 20% increase in productivity. The subject requires a more in depth evaluation than can be covered in this paper but it is certainly an important factor in refinery energy usage.
» T ' t and LSH Feed T ° C
superheated (BPE) 10 °C and must be cooled 10 °C before it can condense in the Heat Exchanger. Cooling of superheated vapor to condensing temperature contributes little to heat transfer (less than 2%) compared to the heat transferred by condensation. In this example, the term 'approach temperature' is the difference of the steam condensing temperature above the slurry heater discharge temperature. Shown in the table in Fig. 3 is the sum of the three temperature drops (38 °C) which is also the temperature rise across the LSH. If the temperature of the slurry reporting to the last HX is lower than 182 °C, Boiler Steam will increase. In the example of Fig. 3 there will be more Boiler Steam than required by the heat balance described in Fig. 2. This indicates that if there were no other requirement for heat in the system (such as HR, SHL, etc.), there will be Blow Off vapor. Fig. 4 depicts Fig. 3 graphically and at the bottom shows the effect that HX area and number of stages will have on Boiler Steam to the LSH.
EVAPORATION An Evaporator facility is usually designed by a contractor with boundary limits provided by the alumina process engineer. This typically results in an evaporation economy of around 3.6 tons water/ton Boiler Steam. As liquor caustic concentrations increase for high yields (productivity), achieving a 3.6 economy becomes more difficult because of a corresponding increase in BPE. Increasing the Evaporator LSH temperature to over 200 °C and adding more stages (effects) could increase the economy to over 7 tons water/ton steam, a very significant reduction of Boiler Steam. Erosion/corrosion issues would need to be addressed. Silica scale on heater tubes at the higher temperatures can be controlled by chemical additive. FACILITY DESIGN An important factor that directly determines the minimum Boiler Steam to Digestion is the nature of the Flash Cooling/ Heat Exchange system itself. This is illustrated in Fig. 3 showing an example of temperatures across the first stage following the digester. Digester discharge at 250 °C immediately flashes to 230 °C, the temperature of both liquor and vapor. The vapor is
173
The driving force for heat transfer is the temperature difference across Digestion and across Precipitation. As indicated in Fig. 5, in order to reduce Boiler Steam to Digestion, the delta T across Digestion must be reduced. However, this will not be practical or efficient without correspondingly lowering the delta T across Precipitation This will require more recuperative cooling in Precipitation in place of heat loss by cooling water.
Digester Temperature (250 ^ )
Flash Tanks Temperature BPE
1Û°C
Approach ♦
§βς
Temperature to LSH {212 °C) *** Heot Exchangers Temperature
Design Case
Doubla Area
Double Stages
Double Both
Hash A T , * C BPE, ° C Approach, C C
20 10 8
20 10 2
10 10 8
10 10 2
LSH ΔΤ, ^C % of Bate
38 100
32 84
28 74
22 50
Fig 4, Erteci of Heat Transfer Arsa and Number of Flash Stages
The Bayer Process is one large heat exchanger. This is illustrated in Fig. 5. There is practically no temperature drop in liquor across Evaporation and sensible heat loss across Clarification will be small. The temperature drop across Precipitation is composed of induced cooling (cooling water) to lower the process temperature of liquor to the precipitators, sensible and evaporative heat loss during the long holding time in Precipitation, and induced cooling during precipitation to further lower the temperature.
•i.30 90Δ !^T
T |
SETTLER
Design for MHA bauxite can reach very high temperatures in the Digester (>260°C). The LSH can elevate Boiler pressure above 70 atm. and up to 90 atm depending on Digest temperature. The corresponding temperature of steam will be above the limit to use process condensate for feedwater because of serious corrosion in Boiler tubing. For this reason, Digestion design using Boiler Steam at high pressure (>70 atm), must return LSH steam condensate to the Boilers. For instance, Boiler Steam to LSH at 100 atm and 315°C will be nearly doubled because almost half the steam heat will be returned as condensate to the Boiler. The alternative of routing the condensate to the Digestion Heat Exchangers instead would require using demineralized Boiler feed water at higher cost and, most likely, an increase in Digestion heat exchange area or number of FT stages. Theoretically, the net thermal affect of returning LSH condensate is the same; it's just the issue of physically returning high temperature/pressure condensate to the Boiler. The return of condensate can be avoided by using a different method of heating liquor before the Digester, such as a molten salt HX (2) in place of the LSH. FINAL COMMENTS There should be increased interest in investing in heat recovery facilities in the event of a doubling or tripling of energy cost relative to other alumina costs. The possibilities that immediately arise are: 1.
Use of natural gas to reduce the heat required to generate electric power in MHA refineries. Refineries that do not have the availability of natural gas may consider using LNG.
2.
Increase the number of stages in Digestion Flash Heat Exchange and at the same time increase recuperative cooling in Precipitation to lower the refinery delta T (see Fig. 5).
3.
With high temperature MHA Digestion (>260 °C), it should be feasible to eliminate the Evaporation facility.
4.
Route Washer Overflow directly to Precipitation. It appears wasteful to heat Washer Overflow for Settler operation and then essentially cool the same stream in Precipitation.
5.
Increase the evaporator LSH temperature and the number of stages (effects) to increase the economy in Evaporation.
I
Fig 5. Temperature Difference across Digestion
REFERENCES
and Precipitation
1. D.J.Donaldson, Journal of Metals, Sept. 1981, page 37. 2. Vereint am Werk (VAW house organ) July 1968
174
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
OPTIMIZATION OF HEAT RECOVERY FROM THE PRECIPITATION CIRCUIT Michael Simpson1, Sushant Hial1 and Rashmi Sangram Singh1 Vedanta Aluminium Ltd., Lanjigarh, Kalahandi Dist., Orissa - 766027, India Keywords: Interstage Coolers, Heat Transfer, Steam Savings solid Al(OH)3 in the precipitation reaction as per the following equation :-
Abstract In the Bayer process, temperature profile across the precipitation circuit plays a major role in maximization of the precipitation yield while maintaining product quality. For this reason, plate heat exchangers are used both at the inlet to Precipitation and between precipitation stages at Vedanta Aluminum's Lanjigarh alumina refinery. The cooling medium for the Heat Interchange Department (HID) is spent liquor, while for the Interstage Coolers (ISCs) of Precipitation, both spent liquor and cooling water are used.
Digestion (Heat) Al(OH)3 + NaOH 4 ""► Al (OH)4" + Na+ Precipitation (Cool ) The primary factors which affect the precipitation yield are well documented by authors such as Hond, Hiralal and Rijkeboer [1], They are: > Initial Liquor Alumina(A) and Caustic(C) Concentration > Holding Time > Temperature > Seed Surface Area > Liquor Impurities such as carbonates
A simple model was built using existing heat exchanger performance data, along with heat and mass balances across the Heat Interchange Department (HID) and Precipitation. This was then used to determine feasible modifications for improving heat recovery in the Precipitation circuit. The results obtained have indicated a process steam reduction of 2 % can be achieved with minor modifications.
To maximise precipitation yield, temperature drop is an important parameter that can be controlled within the Precipitation Department. For this reason, interstage coolers have been provided between the precipitation tanks to reduce the temperature of the slurry. Approximately 60% of current precipitation train flow passes through the ISCs, in order to maintain velocity in the correct range through the coolers. The cooling medium for the 1st ISC is spent liquor from hydrate seed thickener overflow, while the cooling medium for other ISCs is cooling water from the alkaline cooling tower. The flow of these cooling media are monitored and controlled to achieve the desired temperature profile across the precipitation circuit.
Introduction In the Bayer process, the temperature profile across the precipitation circuit plays a major role in maximizing the liquor productivity. The supersaturation of the liquor increases as the temperature decreases. Thus the enhancement of precipitation rate is assisted with the help of inter stage coolers across the precipitation unit. The cooling media used in the ISCs are spent liquor and alkaline cooling water. Presently, heat gained by the cooling water is not utilized as the water loses its heat in the alkaline cooling tower. This loss can be prevented by replacing cooling water with spent liquor as an alternative cooling media in ISC - 2. Further heat can be recovered from the precipitation circuit by drawing the inlet slurry to ISC - 1 from further upstream tanks such as the last agglomerator. The recent installation of new ISCs as part of a debottlenecking project will enable the spent liquor temperature to be kept below 55°C consistently. This liquor therefore becomes a good heat sink for hot precipitation slurry, compared to original design when it was at 60°C.
Heat Balance Determination By utilising the well known heat balance equations below [2, 3, 4], and a series of iterations, the final spent liquor temperature can be determined for given pregnant liquor to HID and spent liquor to ISC temperatures Q=UxAxdTLM (1) Q=Mh xCphXiThiTho) (2) Q= Mc xCpcx(Tco.Tci) (3)
Process Description
Where for any heat exchanger: Q=Heat transferred between the streams (kW) U= Effective overall heat transfer coefficient (kW/m2/°C) A= Heat transfer area(m2) dTLM= Log-mean temperature difference (°C) Mh= mass flow rate of hot stream (kg/s) Cph= Specific heat capacity of hot stream (kJ/kg/°C) Τω = Inlet temperature of hot stream (°C) Tho = Outlet temperature of hot stream (°C) Mc = mass flow rate of cold stream (kg/s) Cpc= Specific heat capacity of cold stream (kJ/kg/°C)
The objective of the Precipitation unit is to produce Alumina "Hydrate" (Alumina tri-hydrate or gibbsite) suitable for downstream calcination to produce sandy metallurgical grade alumina. This is achieved by prolonged contact between pregnant liquor and suspended hydrate seed, under conditions conducive to crystal growth and precipitation of alumina from solution. The chemistry of precipitation, which is the reverse of digestion, is quite simple. In digestion, Al(OH)3 dissolves in NaOH to form a complex anion and this complex ion comes out of solution as
175
Tci = Inlet temperature of cold stream (°C) Tco = Outlet temperature of cold stream (°C)
their appropriate heat transfer coefficients (U value). The different scenarios considered are as follows:
Existing Heat Balance
i) ISC - 1 slurry input from last agglomerator Tank:High precipitation temperatures are only required to aid the agglomeration process. Instead of taking slurry input to ISC-l from the second growth stage, we can utilize slurry from the last agglomerator without impacting on the precipitation sizing control. In this case there will be 6.7°C rise in spent liquor temperature across ISC - 1 as shown in Fig 3. The inlet slurry temperature will be 77.9°C and the outlet spent liquor temp will be 61.7°C. The net spent liquor temp after HID will be 83.7°C as shown in Fig 4. A further advantage of this move will be a reduction in the solids concentration through the ISC which is expected to improve its life. The liquor will be of a higher scaling potential however and caustic wash frequency might need to be adjusted. New piping will be required from second precipitator tank (normally last agglomerator) along with third precipitator (alternate last agglomerator) to the ISC-l.
Slurry
Slurry 1054
m3/hr
1054
m3/hr
69.5
DegC
64.2
DegC
ISC-l Spent l i q .
Spent Liq. 1250
m3/hr
59.2
DegC
<— - -
1250
m3/hr
55.0
DegC
Fig 1 : Spent liquor temperature increased by 4.2°C after passing through ISC - 1
Slurry
Slurry 1054
m3/hr
1054
m3/hr
77.9
DegC
69.6
DegC
250 rnB/hrL. 59.2 1250
DegC
I I
irâ/hrl/
i
83.0
SL
I
10Ö0
φ
1250
m3/hr
88.9
DegC
55,2
DegC
1086
iii3/hr
m
DegC
ISC-l
Spent Liq.
1250
Spent Liq.
m3/hr|
t—~~
DegCj
1250
m3/hr
55.0
DegC
HID PGL
PGL 1200
rn3/hr
1Ö1E
DegC
™™
ii
*»
>
>
Fig 3: Spent liquor temperature increased by 6.7°C after passing through ISC - 1
114 nn3/hr 1Ö2.8
DegC
1250
Fig 2: Spent liquor temperature increased to 83°C after passing through HID
83.7
In the existing system, slurry feed for ISC-l is from the fourth tank which is normally the second growth tank. The spent liquor temperature is raised by 4.2°C across the ISC as shown in Fig. 1. As shown in Fig 2., the spent liquor temperature also increases by 23.8°C in passing through the HID. In the existing scenario, there is also a temperature raise of 10 °C in the alkaline cooling water which is used as the cooling medium in ISC - 2.
m
m3/hr
61.7
DegC
1000
φ(
1250
ntf/hr
89.2
DegC
61.7
DegC
φ\
SL
SL 1
HID PGL
PGL
Proposed Modifications with Schematic Diagram
1200
rn3/hr
1086
ί#Γ
102.8
DegC
76.Ö
DegC
'
To recover the heat through spent liquor it is proposed to: (1) Use spent liquor as cooling media in ISC-2 (2) Feed slurry from an upstream precipitation tank having higher temperature Detailed analysis was done across HID and Precipitation by using the historical performance data of the heat exchangers and hence
>
114 πή/hr 102.8
DegC
Fig 4: Spent liquor temperature increased to 83.7°C after passing through HID
176
Fig 7: Spent liquor temperature increased to 83.7 °C after passing through HID ii) Spent Liquor as cooling media in both ISC - 1 & 2:For this scenario, there will be 6.7°C rise in spent liquor temperature across ISC - 1 and 2 as shown in Fig 5&6.The net spent liquor temp after HID will be 83.7°C, shown in Fig 7. Here the spent liquor flow will be in series i.e. the outlet from ISC - 2 will be the inlet to ISC - 1. Spent liquor piping already passes by the ISC-2 to feed ISC-1 and hence this modification will be simple. Analysis of the additional pressure drop from passing through two ISCs shows that the net positive suction head required for the HID pumps will still be met easily. Slurry
iii) Spent Liquor as cooling media in both ISC - 1 & 2 and ISC 1 slurry input from last agglomeratonThis is a combination of the actions from cases(i) and (ii). There will be 9.2°C rise in spent liquor temperature across ISC - 1 and 2 as shown in Fig 8&9.The net spent liquor temp after HID will be 84.8°C , shown in Fig lO.The spent liquor flow will again be in series ( i.e. the outlet from ISC - 2 will be the inlet to ISC - 1) and the ISC - 1 slurry input will be from last agglomerator. Slurry
Slurry
1054
m3/hr
1054
m3/hr
66.4
DegC
62.3
DegC
Slurry
1054
m3/hr
1054
m3/hr
64.9
DegC
613
DegC
y
t
ISC-2 ÌSC-2 Spent Liq.
Spent Liq.
1250
m3/hr
58.3
DegC
<--»-"-
<—. —.
125?
m3/hr
55.0
DegC
Slurry 1054
m3/hr
1054
m3/hr
69.9
DegC
65.6
DegC
<----
1250
m3/hr
5S.3
DegC
Degc| SL k-»1
SL 1000
m3/hrf··-
89.2
DegC]
m3/hr|
64.2
Degcf
m3/hr
51.7
DegC
1018
— 1
I I I SI I
1000
nâ/hff
1250
m3/hf
90.0
DegC
64.2
DegC
m3/hrL/
HID
ρα m3/hr 76.0
Degcl
114
mllk\
DegC
DegC|
PGL
ra3/hrr
m3/hr
64.2 84.8 1250
1250 57.9
c— - » — ■ ·
HID 1200
m3/hr DegC
Spent Liq.
1250
1250
P6L
DegC
1054 71.6
250 m3/hrk
m3/hr|
83.7 JegCJ
m3/hr
55.0
Slurry m3/hr| DegC]
250 m3/hrL 1250
1250
Fig 9: Spent liquor temperature increased by 6.3°C after passing through ISC - 1
Fig 6: Spent liquor temperature increased by 3.4°C after passing through ISC - 1
61.7
<--«-
ISC-1 Spent Liq.
<-
DegC
<- . . .
Spent Liq.
Spent Liq. DegC
5X9
79.3
ISC-1
61.7
m3/hr
Slurry 1054
Slurry
m3/hr
1250
Fig 8: Spent liquor temperature increased by 2.9°C after passing through ISC - 2
Fig 5: Spent liquor temperature increased by 3.3°C after passing through ISC - 2
1250
Spent Liq.
Spent Liq.
102.8
177
PGL
1200
m3/hrl·
1086
m3/hr
102.81
Degc|
m
Degcj
114
m3/hr
102.8
DegC)
Fig 10: Spent liquor temperature increased to 84.8 °C after passing through HID
At a price of $11.1/t steam, the cost savings vary from $147,000-$375,000 per annum and should be adequate to give a good return on investment as only minor modifications are anticipated to be required.
Expected Benefits
1.
Table 1: Cost saving analysis with different modification options Existing
A
2.
m 1
R
f5C-l&2wfià SL series flow Cenditiöß
ÌS€-l*Éh
$t
isc-iyet
SC-l&2«äJi
3.
iSCÖöing
from Last a^Q SL series flow media with ISC
4.
i slurry inlet from last aggio |
12»
1250
83.0
83.7
83.7
84.8
129294
13Ö447
130M5
132225
Heat Gain w,r.t existing, KJ/Sec
1153
1150
2331
Net Mass of Steam Saving, T/hr
152
L52
3.86
0
404
403
1027
0
147467
147126
St Flow, Mi/k
1250
SLTemp.afterHID.DegC HeatAKi/Sec
Safing,$/d3y 1 Saving, $/Annum
1250
374867 1
SAVINGS/ANNUM
m
Existing
Graph 1 : Savings / annum for different scenarios.
Conclusions Up to 3.86 t/hr of steam can be saved by modifying ISC-2 to utilise spent liquor at 55°C as the cooling medium and taking slurry input to ISC-1 from the last agglomerator. Further detailed analysis of the mechanical and operational feasibility along with the cost of the modifications for each scenario will be carried out.
178
References Roelof Den Hond, Iwan Hiralal and Ab Rijkeboer " Alumina yield in the Bayer Process - past, present and prospects" , Light metals 2007 , 3 7 - 4 2 Don W. Green, James O. Maloney, "Perry's Chemical Engineers" Handbook - seventh Edition - Sections 5-12,1114 Nicholas P. Cheremisinoff, "Handbook of Chemical Processing Equipment" , 1 - 6 2 Max S. Peters, Klaus D. Timmerhaus and Ronald E. West, "Plant Design and Economics for Chemical Engineers" Fifth Edition, 643 - 748
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
ALUNORTE GLOBAL ENERGY EFFICIENCY Reiner Wischnewski, Cleto Maues de Azevedo Jr., Emerson L. S. Moraes, Arthur Barros Monteiro Alunorte - Alumina do Norte do Brasil S.A., Rodovia PA 481, km 12. Distrito de Murucupi, CEP: 68.447-000. Barcarena-PA. Brasil Keywords: Bayer process, Energy efficiency, Heat integration, Water balance Abstract
18
Alunorte is the largest alumina refinery in the world with a production capacity of close to 6.3 Mtpy. The plant has a specific energy utilization of 8 GJ per ton of alumina and is one of the most energy-efficient plants in the world. The high energy efficiency is achieved by good process design, the utilization of state-of-the-art technologies, good operation and the processing of high quality bauxite. The technologies which are applied in Alunorte to contribute to the global energy efficiency of the plant are reviewed and heat integration and water balance are discussed.
16
1u
c •2 12 to
>»
2> 8 5 !
3rd quartite
Squartile
/inquartile
6
2
Alumina is the principal raw material for aluminium. In 2009 the world-wide average utilization of electrical power for primary aluminium production was 15,215 kWh per metric ton of aluminium [1]. The energy for alumina production is not included and has to be added. The average energy utilization for alumina production was 3,311 kWh per ton of alumina in 2009 [2]. In theory 1.9 tons of smelter-grade alumina are required for the production of one ton of primary aluminium. Hence, energyefficient alumina production is an important contribution to reduce the energy utilization of primary aluminium production.
50
75
100
% of alumina world-wide production outside China Figure 1. Specific energy utilization for alumina production. are discussed which in sum result in high global energy efficiency of Alunorte. Main elements are bauxite of high quality, the precipitation concept and the resulting high liquor productivity. Alunorte's design with regard to heat integration, water balance and utilities for steam and power generation is discussed as well as technology and equipment applied. Some factors are mentioned which work against an improvement of energy efficiency in the future. Current projects to address these issues are mentioned and some possibilities for a further increase of energy efficiency are presented. Alunorte has the aim to be among the most energyefficient alumina refineries in the world also in the long-term.
Table I. Energy used for 2009 alumina production [2], Africa and South Asia North America South America East Asia and Oceania Europe Weighted Average Total
Ί i /
'«ft
Introduction
spec, energy (MJ/t) 14,768 11,449 9,319 11,252 16,842 11,922 499,355 TJ
Squartile Alunorte (8.0 GJ/t)
alumina produced (t) 3,225,778 2,804,849 12,226,990 16,511,664 7,117,522
Energy Utilization for Alumina Production Alunorte started operation in 1995 with a production capacity of 1.1 Mtpy. Since then the plant was expanded three times to a total production capacity of close to 6.3 Mtpy. In Figure 2 the development of production capacity is shown and compared to the energy utilization of the plant. In 2009, the first year after Expansion 3, the actual production remained slightly behind the nameplate capacity of the plant. The Expansion 3 project, Alunorte's current performance and an outlook about the expected development of Alunorte are presented in [3].
-
41,886,803
Table I shows the 2009 specific energy utilization for alumina production by region. On average the production in South America utilizes less energy than in any other region of the world. Alunorte contributes significantly to this result. The production capacity of Alunorte is close to 6.3 Mtpa after its third expansion in 2008 [3]. Alunorte produces close to 50 % of all alumina in South America. Its energy utilization was 8.0 GJ per ton of alumina in 2009. This is significantly less than the average of 9.3 GJ per ton in South America and 11.9 GJ per ton world wide. In Figure 1 Alunorte's performance is compared to the energy utilization of other alumina refineries. The graph includes alumina plants for the production of smelter-grade alumina outside China. The statistic illustrates that the world-wide benchmark for alumina production is at about 8 GJ/t. There are few plants only with a similar good performance. In the following, the different factors
After the first two years of operation, the plant uses between 7.5 GJ and slightly above 8 GJ per ton of alumina. In 2006 the best result was achieved. In the following years from 2006 to 2009, i.e. since Expansion 2, the specific energy utilization trends upwards. With Expansion 2 some new technologies were installed. Bauxite is received from Paragominas through a pipeline and dewatered at Alunorte in hyperbaric filters. The filters use compressed air and thus compressor power. Two more turbogenerators were installed to increase the electrical power which is co-generated at site. The losses of cogeneration of electr-
179
Bauxite Quality & Liquor Productivity
J2 E
Ì1
n
The bauxite quality, in particular the content of available alumina, reactive silica and organic components, significantly affect the performance of an alumina refinery. Alunorte processes high quality bauxites from two different sources, namely Trombetas and Paragominas. Table II compares the main characteristics of these bauxites.
■ 1111117.57.0 SI
h- co σ> O) O) G) G) σ> σ>
mimi
8 8 8 8 8 8 8 8 8
Table II. Typical bauxite quality.
S
6.5
m
CNCMCNtNCNCNCNCNCN
ipacity Figure 2. Development of Alunorte's alumina production capacity and specific energy utilization. icity decrease Alunorte's global energy efficiency. Losses of cogeneration are included in Alunorte's overall efficiency while electricity from the grid is accounted as is without adjusting for losses during generation. The plant availability reduced somewhat which is associated with some negative impact on the specific energy consumption as Alunorte became more complex.
Table ΙΠ. Spent Liquor Analysis. A1203
NaOH (as Na2C03) Na2C03 causticity organic carbon NaCl Na2S04 Na2C204
total energy utilization
g
\pii
1
electricity m
m
Il
(%)
(g/i) (g/1) (g/i) (g/i)
112 280 5.6 98 4.4 2.20 0.28 <0.1
The concept of bauxite residue treatment and disposal is an important factor contributing to high liquor purity. Bauxite residue is filtered in drum filters and disposed at high solids concentration by dry stacking in the residue disposal area (RDA). Collected water from the RDA is neutralized and clarified in a water treatment station and released as effluent to the nearby river [4]. There is no water returning from the RDA to the process. This
digestion steam
E
(g/i) (g/i) (g/i)
The content of organic material in the bauxite is very low. This results in a very low concentration of impurities in the plant liquor. In Table III a spent liquor analysis of Alunorte is shown. The concentrations of organic carbon, sulfate and oxalate are low and the causticity of the liquor is very high. The plant is operated with a high caustic concentration. A precipitation productivity of up to 89 g/1 in the newest process lines can be achieved [3].
other . /
Figure 3. Breakdown energy used £ ofcogenerated | at 1Alunorte.
i
Paragominas 48.2 % 4.6% <0.05% 14.9 %
Although the composition is very similar certain differences are observed during processing. The high content of available alumina results in a low specific bauxite consumption. This is beneficial in terms of energy utilization, since the amount of the non-gibbsite fraction of bauxite, which does not contribute to the production of hydrate, is small. A low amount of digestion live steam is required only to heat up this material with a positive effect on the overall energy efficiency. A low temperature digestion process is used at Alunorte as both bauxites contain minimal amounts of boehmite. Their reactive silica content, however, is relatively high. The reactive silica reacts under consumption of caustic soda to desilication product (DSP) and increases the amount of bauxite residue. Generally, some heat is lost from the process to the environment due to the bauxite residue having a temperature of about 60 °C upon disposal. This heat loss at Alunorte, however, is moderate due to the relatively low mud factor, e.g. mass of mud per mass of alumina produced, which caused by the high content of available alumina in the bauxite and the bauxite residue beingfiltratedbefore disposal.
In Figure 3 a breakdown of the energy utilization of Alunorte is shown. It is divided into electrical power received from the national grid, energy required for steam and power generation and energy for calcination. The energy for steam and electricity generation and calcination is based on the lower heating value (LHV) of the fuels. Electrical energy received from the national grid is accounted as received and is not adjusted for losses during generation. Steam at two pressure levels can be generated with the installed high pressure or low pressure boilers. Three turbogenerators work between the high and the low pressure level and can cogenerate up to 65 % of the required electrical power of Alunorte. Process steam is supplied to the process from the low pressure header. In digestion about 1.0 t steam per ton of alumina is consumed and about 0.2 t/t in evaporation. 3 GJ per ton alumina is used for calcination. Other consumers include losses due to inefficiencies of boilers and turbogenerators, internal steam consumption in the boilers and miscellaneous steam consumers in the process such as for filter aid preparation.
Ζ ΖΖ;Γ,Ζ~ evaporation steam
Trombetas 49.1 % 4.1% < 0.05 % 11.5%
available A1203 reactive Si0 2 organic carbon moisture
f
electricity T
180
to dilution with steam. The upper limit for the caustic concentration is given in the spent liquor live steam heaters due to scaling and corrosion. At lower caustic concentrations less bauxite can be charged to the liquor and liquor productivity decreases. These negative effects are minimized at Alunorte due to avoiding direct heating by steam injection. Furthermore, the heat exchange area for spent liquor heating is well chosen, so that the energy for regenerative spent liquor heating is high.
approach avoids the contamination of plant liquor with additional impurities and contributes to maintain clean liquor with high liquor productivity. Liquor impurities are purged from the process with the bauxite residue. No caustizisation unit for carbonate removal is installed. Process Design The concepts for heat integration, water balance and generation of steam and electrical power are the main contributors for direct savings of energy.
Indirect heating of the bauxite slurry is another important contributor to low digestion steam consumption. The digestion temperature is a mixing temperature determined by the temperature and mass flows of the bauxite slurry and the final spent liquor temperature. The higher the temperature of the bauxite slurry the lower the final spent liquor temperature after the last live steam heater can be. This reduces the utilization of digestion live steam and increases the global energy efficiency.
Heat Integration Alunorte has a low temperature dual stream digestion unit. A simplified flowsheet is shown in Figure 4. The pregnant liquor stream is expanded and flashed in five stages from the digestion temperature and pressure to atmospheric conditions. The liquor is released from the blow-off tank (BOT) at a temperature close to the liquor boiling point. The spent liquor is heated in countercurrent flow by regenerative heating. Heat is transferred from the pregnant liquor to the spent liquor stream with vapor from the flash tanks which is condensed in four spent liquor regenerative heaters. Vapor from the blow-off tank is released to the environment. Live steam heaters are used to heat the spent liquor to the required temperature before it is fed to the digestors. flash tanks
life steam
life steam condensate
Figure 5. Indirect Bauxite Slurry Heaters (IBSH).
life steam heaters — Wl·*—j
Two stages of Indirect Bauxite Slurry Heaters (IBSH) are installed at Alunorte. The aim is to achieve the highest possible bauxite slurry temperature. The IBSH units are vertical shell-andtube heat exchangers (Figure 5). The bauxite slurry is heated indirectly with live steam heater condensate. Alunorte operates the units without major problems. As part of Alunorte's continuous efforts to increase heat recovery a project has been launched to add a third stage of IBSH and maximize the bauxite slurry temperature before digestion. This will reduce digestion live steam consumption due to a lower required final spent liquor temperature. Furthermore, a positive effect on the silica scaling in the spent liquor heat exchangers is expected. A good maintenance program is in place to keep the heat transfer coefficients of the heat exchangers in the digestion area at a high average level.
bauxite slurry heaters bauxite
Figure 4. Simplifiedflowsheetof digestion and slurry heating. A conceptual problem of dual stream digestion technology is the difference between heat sink (spent liquor) and heat source (pregnant liquor). The mass flow of pregnant liquor is larger than that of spent liquor by the amount of bauxite slurry which is charged to the digestors. It is not possible to condense all vapor which isflashedfromthe pregnant liquor stream in the regenerate heat exchangers and transfer all latent heat of the vapor to the spent liquor. The excess steam leaves the blow-off tank and is a loss of heat to the environment. In plants with large amounts of excess steam some of it can be transferred to other process areas.
The efficiency of the vacuum flash cooling units is another major contributor to the low energy utilization of Alunorte. Pregnant liquor is cooled in counter-current flow against spent liquor in a flash train. The pregnant liquor inlet temperature is close to the boiling point and the outlet temperature shall be as low as possible. Additional plate heat exchangers are used to achieve the final fill temperatures to agglomeration and cementation tanks. A simplified flowsheet of Alunorte's vacuum flash unit and plate heat exchangers is shown in Figure 6.
At Alunorte the difference of heat sink and heat source in the digestion area is small and with it the loss of heat to the environment. This is based on a good design of Alunorte. There is no direct heating of the spent liquor by steam injection and indirect bauxite slurry heaters are installed. Direct heating increases the difference between heat source and heat sink in digestion, an additional amount of water must be evaporated and requires additional energy and the caustic concentration drops due
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slaking, gland water, rain, etc. Alunorte's digestion concept which avoids dilution of liquor and uses indirect heating only is one of the main reasons for a comparably small liquor dilution.
cooling water to aggi.
pregnant liquor
Trombetas bauxite is transported to Alunorte by ship while Paragominas bauxite is pumped through a 244 km pipeline [5]. At Alunorte the Paragominas bauxite is dewatered in hyperbaric filters [6-8]. The residual moisture of the dewatered bauxite is about 15% and slightly higher than that of Trombetas bauxite with about 11.5 %. The moisture content of the dewatered Paragominas bauxite is determined by the performance of the filters and the characteristics of the bauxite, in particular its particle size distribution. Work is underway to optimizing the operation of the filters in cooperation with the beneficiation plant of Mineraçao Bauxita Paragominas (MBP). The potential for improving the water balance in bauxitefiltration,however, is small. The same is valid for water entering with fresh caustic. It is supplied at a concentration 50 wt.-% NaOH. The caustic consumption is mainly determined by the non-controllable loss due to the content of reactive silica in the bauxite and the controllable losses which are affected by the efficiency of mud washing and the residual moisture in the bauxite residue. Controllable losses are already very small with less than 8 kg of caustic per ton of alumina. However, some potential improvements in the bauxite filtration area identified.
to cem, lotest 1
1
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• spenf liquor
condensate Figure 6. Vacuum flash cooling and plate heat exchangers. The better the performance of the vacuum flash cooling units the lower is the pregnant liquor inlet temperature to the precipitation plate heat exchangers. This reduces the requirement of cooling water to the plate heat exchangers and recovers more energy for heating the spent liquor. A high spent liquor temperature reduces the required amount of digestion live steam and improves the performance of the evaporation unit. Alunorte has a spare vacuum flash cooling unit so that cleaning and maintenance can be performed without major performance losses in three of the seven lines. A high average heat transfer coefficient of the heat exchangers of the vacuum flash cooling units can be achieved and process disturbances due to cleaning are small in these lines. As measure for higher average heat transfer coefficients Alunorte optimizes the cleaning cycles of the vacuum flash cooling units and will study the option to install spare vacuum flash cooling units for the other four lines as well. For efficiency improvement their capacity could be increased such that lower pregnant liquor and higher spent liquor temperatures can be reached.
Regenerative condensate is used for washing bauxite residue and hydrate. Both together are the biggest source of plant liquor dilution. The net wash of mud is generally good with 0.8 - 1.2 ton of water per ton of mud. The potential for further improvement is small. The situation for hydrate washing is similar good and the potential for improvements small. Dilution of plant liquor from other sources such as rain, water hosing or gland water is moderate. All tanks are closed and water hosing is done with care. There exist some possibility for further reduction of water input to the liquor but the situation is generally good and room for improvement also small.
Water Balance The water balance directly affects the energy utilization of an alumina refinery. It is unavoidable that water from different sources dilutes the plant liquor. This water is mainly removed by evaporation. With higher rates of liquor dilution the demand for evaporation increases and at the same time energy used in form of live steam.
Small water sinks are product and bauxite residue while the largest part of the water is removed by evaporation in three plant areas - digestion, vacuum flash cooling and evaporation. A good maintenance program is crucial to ensure high average heat transfer coefficients in all heat exchangers of these areas. A high heat transfer coefficient is associated with a high rate of water removal from the liquor and low consumption of live steam in the digestion and evaporation areas. Furthermore, spent liquor return temperature increases with higher performance of the vacuum flash cooling units. The demand for digestion life steam reduces and the efficiency of the evaporation area increases which is associated to a lower requirement for evaporation live steam. Work is underway to further improve the performance of the flash trains in the digestion, evaporation and vacuum flash cooling.
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water entering
Steam and Power Generation
water leaving
The design of the utilities area for steam and power generation has to match the demand of steam and electrical power. Inadequate design can lead to high losses in this area and potential savings in steam or electrical power in the Bayer process would be compromised by additional losses in the utilities area. Alunorte has chosen a design which allows a sufficient flexibility to achieve overall high efficiency in steam and power generation.
Figure 7. Volume of water entering and leaving the plant liquor. In Figure 7 the water balance of Alunorte is shown. Due to a high liquor productivity and no direct heating with steam in digestion dilution of liquor with water per ton of alumina can be kept at around 2.3 m /t. Main sources are the moisture content of the bauxite, water for mud and hydrate washing, a small amount entering with caustic and various sources such as hose water, lime
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at high caustic concentrations and at a high liquor to precipitation (LTP) A/C ratio. Seed filtration is used in five of the seven lines and in two lines cyclones for seed concentration are installed. With the precipitation concepts chosen in Alunorte high liquor productivities are achieved.
electrical power
In addition, it is important that the individual pieces of equipment of the plant have a high efficiency, operational procedures are good, high availability can be achieved and a good maintenance program is in place. High efficiency motors for agitators, blowers, etc. are installed and pumps with variable speed or variable frequency drives. An electrical power utilization less than 0.2 MWh per ton of alumina is achieved.
digestion evaporation +> other condensate make-up water
Steam consumption is significantly affected by the process design as explained above. But also individual units, in particular the evaporation area, have to perform well. Alunorte's evaporation units have an efficiency of about 4.2 ton evaporated steam per ton of live steam. Some potential for improvement was identified and Alunorte works on a project for further performance increase.
Figure 8. Simplified flowsheet of utilities area for steam and power generation. A simplified flowsheet is shown in Figure 8. Alunorte receives some of the electrical energy from the national grid and cogenerates the remaining electrical energy with turbogenerators (TG) which work between the high pressure (HP) and low pressure (LP) steam headers. The fraction of the high pressure steam (87 bar and 485 °C), which is not required for cogeneration, passes through pressure reducing valves (PRV) and is subsequently cooled to saturated conditions by adding water in desuperheaters (DSH). This amount of high pressure steam can be replaced with steam from the low pressure boilers. Thus, the generation of steam and electrical power are not directly coupled but allow some operational flexibility. The operation of the utility area for steam and power generation can be optimized under consideration of overall efficiency and operational cost. The high pressure boilers have an efficiency of 84-86 % while that of the low pressure boilers is 88-90%. The second advantage of decoupling of steam and power generation is that a design could be chosen which avoids the installation of condensing turbines, resulting in an overall higher efficiency of the utility area.
A major operational issue is scaling. It affects the heat transfer coefficient in heat exchangers and has an impact on the overall availability of the plant in dependence of the frequency for descaling and cleaning. Both are associated with lower global energy efficiency. Alunorte works on the one hand to optimize the cleaning and descaling work and on the other hand to reduce scaling rates. Measures to reduce the solids concentration in spent liquor are one example. Another is the modification of digestion side entry flash tanks to bottom entry which was done before Expansion 1. Today there are no side-entry flash tanks installed at Alunorte. Carry-over of caustic reduced and longer cleaning intervals could be achieved. Benefits can also be achieved from chemical agents to reduce scaling. Calcination is a major contributor to the energy used for alumina production. The theoretically achievable limit is 2.4 GJ/t, when the enthalpy of formation for aluminium trihydrate and alumina are compared. This value does not consider energy used for the evaporation of remaining moisture in the filtered trihydrate and other technical limitations, such as a minimum stack temperature of the flue gas. Alunorte uses circulating fluidized bed calciners which use on average 3 GJ/t of energy. Alunorte received the energy efficiency award 2010 for the two newest calciners by the German Energy Agency which operate at 2.79 GJ/t [10].
A high boiler feed water temperature is beneficial for the overall efficiency for steam generation. Some of the boiler feed water is preheated in the calcination area. Potential for further improvement is identified in the rate of live steam condensate from the process. The fraction which does not return is replaced by make-up water. Due to its lower temperature the power plant efficiency decreases with the fraction of make-up water. Alunorte works on a project which will increase the condensate return rate. Some more potential for efficiency improvement exists when higher steam parameters are chosen. 160 bar and 540 °C are typical ranges for circulating fluidized bed boilers [9]. However, this can only be implemented as part of the construction of a new boiler. Furthermore, it needs to be carefully studied if these boilers are suitable as utility boilers of an alumina refinery.
Conclusions Alunorte's is one of the most energy-efficient alumina refineries in the world. The main contributors to this performance are discussed, such as the utilization of high quality bauxites, a good process design, the use of energy-efficient calciners and efficient steam and power generation. The heat integration concept and water balance of the refinery are presented. A good heat integration concept is chosen and Alunorte uses indirect heating only and avoids dilution of plant liquor with steam. Other liquor dilution is generally small. Net wash for example is small. The layout of the utilities area for steam and electrical power generation allows a good operational flexibility and has a generally high efficiency. Alunorte is aware of some factors with a negative impact on energy efficiency. Slowly decreasing bauxite quality, a higher rate of cogeneration of electricity or the ageing of the plant are examples. Alunorte addresses these issues and works
Technology, Operation & Maintenance The potential for direct savings of energy in precipitation (from the first precipitation tank onwards) is limited. Indirectly, however, energy is saved when the productivity of the liquor is maximized. High liquor productivity is important for low specific energy utilization. The higher the liquor productivity the lower is the amount of liquor which has to be heated up in the digestion area and cooled down in the precipitation area per ton of alumina produced. The specific steam consumption in digestion and thus specific consumption of boiler fuels is low. Alunorte is operated
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on a number of projects to implement process improvements, e.g. in bauxite dewatering, bauxite slurry heating, precipitation and bauxite residue filtration. Furthermore, cleaning cycles for descaling of heat exchangers are optimized. Alunorte will continue to identify potential for improvement of energy efficiency and aims to be one of the most energy efficient alumina refineries in the world - also in the long term.
Metals 2011, Lindsay, S., Ed., TMS (The Minerals, Metals & Materials Society).
References 1. International Aluminium Institute (IAI), "Electrical Power used in Primary Aluminium Production", online content, "https://stats. world-aluminium.org/iai/stats_new/formServer.asp?form=7", date of issue Sept 2, 2010. 2. International Aluminium Institute (IAI), "Energy used in Metallurgical Alumina Production", online content, "https://stats. world-aluminium.org/iai/stats_new/formServer.asp?form=8", date of issue Sept 2, 2010. 3. Daryush A. Khoshneviss, Luiz Gustavo Correa, Joaquim Ribeiro Alves Filho, Hans Marius Berntsen and Ricardo Carvalho, "Alunorte Expansion 3 - The new lines added to reach 6.3 million tons per year", In Light Metals 2011, Lindsay, S., Ed., TMS (The Minerals, Metals & Materials Society). 4. Jorge Aldi, "Achieving Excellence in Liquid Effluent Treatment at Alunorte", In Light Metals 2009, Bearne, G., Ed., TMS (The Minerals, Metals & Materials Society). 5. Ramesh Gandhi, Mike Weston, Mani Talavera, Geraldo Pereira Brittes and Eder Barbosa, "Design and Operation of the World's First Long Distance Bauxite Slurry Pipeline", In Light Metals 2008, De Young, D.H., Ed., TMS (The Minerals, Metals & Materials Society), 95-100. 6. Ayana Oliveira, Juarez Dutra and Jorge Aldi, "Alunorte Bauxite Dewatering Station - A Unique Experience", In Light Metals 2008, De Young, D.H., Ed., TMS (The Minerals, Metals & Materials Society), 85-87. 7. R. Bott, T. Langeloh and J. Hahn, "Filtration of Bauxite after Pipeline Transport: Big Challenges - Proper Solutions", In Proceedings of 8th International Alumina Quality Worhhop, Armstrong, L, Ed., AQW Inc.,2008, 319-323. 8. A. Campos, R. Bott, "Determination of a suitable dewatering technology for filtration of bauxite after pipeline transport". In Light Metals 2008, De Young, D.H., Ed., TMS (The Minerals, Metals & Materials Society). 9. Stephen J. Goidich and Ragnar G. Lundqvist, "The Utility CFB Boiler - Present Status, Short and Long Term Future with Supercritical and Ultra-Supercritical Steam Parameters", Proceedings of Circulating Fluidized Bed Technology VII Proceedings of the 7th Int. Conf. on Circulating Fluidized Beds, Grace, J.R., Zhu, J-X., de Lasa, H.I., Eds. Can. Society of Chemical Engineering, Ottawa, May 2002. 10. Michael Missalla, Hans Werner Schmidt, Joaquim Ribeiro Alves Filho and Reiner Wischnewski, "Significant Improvement of Energy Efficiency at Alunorte's Calcination Facility", In Light
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Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
OPPORTUNITIES FOR IMPROVED ENVIRONMENTAL CONTROL IN THE ALUMINA INDUSTRY Richard Mirtina, John Kildea, Everett C. Phillips, Wayne Carlson, Bruce Keiser, and John Meier Nalco Company; 1601 W. Diehl Rd., Naperville, IL 60563-1198, USA Keywords: Bayer process, mercury, emissions, alumina, coal Abstract Alumina production from bauxite offers a unique set of environmental concerns that affect air, water, and solids. Governments and industry have recognized that reductions in plant emissions and environmental impacts are necessary. The alumina industry is not the only industry that has been subjected to, and responded to, such regulatory scrutiny over the past decade. A number of industry sectors are actively developing innovative ways to control a broad range of potential environmental hazards. Several of these technologies may have direct application in alumina refineries. Methods to significantly reduce mercury emissions in both air and water have recently been developed for use in a range of industries. This paper reviews some of these new technologies now in commercial use in non-alumina plants and considers how they may be applicable within the alumina industry. Introduction The emissions to air and water from alumina refineries are coming under increasing scrutiny by both the industry itself and the communities in which they operate. Such environmental concerns are common to a variety of industries1"6 and alumina producers can learn much by looking outside the industry to other industrial processes where emissions problems are being successfully addressed. In recent years, Nalco has developed a number of strategies designed to reduce air and water emissions of a variety of hazardous contaminants. These technologies have been initially targeted at processes unrelated to alumina refining. Nonetheless, the issues that many of these technologies address - emissions of mercury, sulfur oxides (SOx), and nitrogen oxides (NOx) - are also commonly found as issues in the alumina industry.
bauxite, 500-700 ppb levels are common in Jamaican bauxites, and mercury levels as high as 1200 to 2000 ppb have been reported in some bauxites in Suriname.8 Regardless of the initial level, due to the large volumes of bauxite that are processed in the production of alumina, significant quantities of mercury can accumulate at various points in the Bayer process, ultimately resulting in mercury emissions to air and water of up to several tons per year. Additionally, depending on the source of the lime and raw caustic that are also added to the Bayer process, these too can contribute significant inputs of mercury to the system. Fate of Mercury in the Bayer Process Figure 1 outlines the major sources of mercury input and output in a typical Bayer plant. It is interesting to note mat the potential output streams are dominated by gas phase emissions. While clearly there are some differences, emission to air is also the predominant issue in most coal-fired power plants. As a result, the potential for the use or adaptation to Bayer plant operations of the existing control technologies used in coal-fired power plants is apparent. The mercury contained in bauxite exists as a mixture of different inorganic and organic compounds and oxidation states. Regardless of its initial form, virtually all of the mercury is converted to elemental mercury (Hg°) by the highly reducing environment of the digestion phase of the Bayer process.10 A portion of this mercury can then be emitted to the air via digestion vent gases, especially within the first few stages of flash cooling, and some mercury may be collected in the downstream condensate systems, Figure 1.
This paper focuses on the issue of mercury in the alumina industry and then considers a number of recent developments in the area of mercury emissions control and reviews the benefits and downsides of their implementation in the coal-fired power industry. Consideration is also given to how some of these technologies, which are based on both chemical and engineering solutions, might be applied or adapted within the alumina industry to resolve specific problems associated with mercury in the Bayer process.
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The major source of mercury in Bayer process operations is the bauxite ore used. While trace levels of mercury are distributed throughout the geosphere in soils, ores, and mineral deposits at concentrations typically in the range from 20 to 150 ppb,7 mercury levels in bauxite can vary quite drastically between different sources and even between different deposits from a single source. Mercury levels as low as 20-100 ppb have been found in Boke
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Figure 1. Potential Hg Inputs and Outputs in the Bayer Process
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Following digestion, a large portion of the mercury may be carried within the red mud slurry. The extent to which the mercury will partition to the red mud solids or to the liquor will be influenced by many factors such as the digestion conditions used, the liquor chemistry and the extent to which mercury oxidation will occur, and the type and quantity of process additives used. These will ultimately influence the concentration of the mercury present in the pregnant liquor. Mercury losses to red mud residues have been estimated to be as high as 0.5 lb/day or more, and these losses may account for 15-60% of the mercury input to the process.7 Upon precipitation of the alumina trihydrate from the liquor, trace levels of mercury can become incorporated in the trihydrate crystals which will be vaporized and emitted into the air during the downstream calcination process. Similarly, trace levels of mercury may be incorporated into oxalate solids that are either coprecipitated with the trihydrate or precipitated by side-stream oxalate removal. Thus, depending on the oxalate removal operations in a specific refinery, additional mercury can exit the process through residue disposal or potentially via air emissions if the oxalate residues are sent to a downstream furnace for lime recovery. The oxalate cake at one refinery was reported to have as much as 2000 ppb mercury and it contributed to a loss of 15% of the input mercury.7 Completing the cycle, small amounts of the mercury may be stabilized in the liquor, which will then be recirculated to evaporation and then back to digestion where it can be lost via air emissions through the barometric condensers or, as mentioned before, it may become a contaminant in the process condensate streams. Digestion condensates have been reported to have soluble elemental mercury concentrations in the order of 20 μg/L.11 Mercury Control Options in the Bayer Process In the course of processing the very high volumes of vent gases and or condensate that are produced from evaporation and digestion at a typical refinery, elemental mercury often accumulates in the traps on condensers and heat exchangers to the point where it can be drained off and collected. This can yield several hundred grams of metallic mercury (Hg°) per day which can be collected and isolated for proper disposal.
stabilization of mercury in solution decreases as excess sulfide is slowly converted to sulfate in spent liquor and residual mercury species will be reduced back to the elemental form in digestion. Zinc, mercury, and other heavy metals may also be precipitated from digestion condensates or pregnant liquor by treatment with a series of Nalco dithiocarbamate or dithiocarbonate products.13 For example, the additive(s) could be added just prior to mud filtration and the resulting residues may be removed from the process with the filter cake through the red mud disposal operations. This is similar to current practice in power plants for the control of mercury levels in industrial wastewater as outlined below. Furthermore, the use of these additives could substantially reduce or eliminate the need for sulfide addition and the corresponding large quantities of organics (and often liquor poisons) that are introduced to the process as a result of their use. There also exists the potential to treat calciner stack gases with the injection of mercury sorbents in between the point where the alumina is collected and the final particulate control device. In addition, the co-injection of a mercury oxidant product, again such as currently used in other industries, could also be considered as a means to lower the sorbent feed rate. The mercury sorbents would of course become mixed with the alumina fines and this would prohibit the recycling of the fines back into the calcination cycle as is usually done. However, such a strategy could still prove to be the most economically and technically feasible, especially for refineries where the mercury tends to partition to the alumina. Some of these control options may likely require consideration of the overall needs of the Bayer process to enable appropriate implementation. They may also require modification for specific plant conditions. However, the successful implementation of such strategies in other industries indicates the potential for improved control using these techniques. In particular, a number of these strategies have been successfully implemented within coal-fired power plants and the lessons learned from this industry can provide useful insights for alumina producers looking to reduce emissions. Mercury Control in Coal-Fired Power Plants Mercury is naturally found in coal in concentrations ranging from 20 to 1000 ppb and coal-fired power plants account for 30% of the global anthropogenic mercury emissions.14 All forms of mercury present in the coal decompose during combustion into the highly volatile elemental form (Hg°) which can readily evade capture by existing air quality control devices typically found at power plants. However, in contrast to alumina refineries where mercury is emitted almost exclusively in the elemental form, in coal-fired power plants a significant portion of the mercury can subsequently become oxidized to an ionic form (Hg2+) depending on the coal type and process conditions. In this oxidized state mercury is more readily captured by fly ash and sorbents and/or scrubber liquors. As a result, it is advantageous to maximize the conversion of elemental mercury to the oxidized form to enhance capture. However, this is not the only strategy. The extent to which the mercury becomes oxidized and is removed from the flue gas is highly dependent both on the type of coal being fired, and the operational conditions employed at the plant.15 This means that mercury emission rates can be highly plant specific and an intimate understanding of the factors that affect mercury transformations and partitioning is crucial to developing a successful control strategy for any given plant.
In addition, gas volumes from digestion vents tend to be much smaller than those from calciners and oxalate furnaces, such that these can potentially be much more readily treated by standard techniques to remove mercury. For example, the vent gases can be chilled to further condense mercury vapors and then passed through an appropriate adsorbent, such as activated carbon columns to remove the mercury.7"9 Indeed, the removal of elemental mercury from digestion condensates by passage through a column containing mercury sorbents has already been demonstrated on a lab scale.11 With respect to mercury in Bayer process liquors, it has been discovered that the use of sodium hydrogen sulfide to precipitate problematic zinc salts can dramatically affect the partitioning behavior of the mercury in the red mud and liquor.10 If a sufficiently high concentration of sulfide ion is maintained in the liquor, the mercury will remain in solution, presumably in the form of [HgS2]2",12 which will help to prevent it from precipitating with the alumina and the oxalate residues. While this can potentially reduce mercury emissions in calcination or in oxalate removal, the
186
Flue Gas Electrostatic Precipttator or Baghouse Preheater Ajr
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Mercury Removal from Wastewater
Figure 2. Mercury control technologies and their respective points of application for the control of mercury emissions from coal-fired power plants. As a result, like alumina refineries, there is no single, universal solution for controlling mercury emissions from coal-fired power plants. Rather, a wide range of complementary control technologies can be employed to cater to the broad range of process conditions and plant operations in facilities where better control of mercury emissions is required. By understanding the process, where the problems are and what form they take, a suite of various control measures can be selected and employed together, as necessary, to effectively target the needs of each specific site (Figure 2). While the details of the various control strategies can be complex, a brief overview of some of the key technologies that have been developed by Nalco and are now in use within the industry are presented below.
state (Hg2+) rather than the elemental form (Hg°). To facilitate the conversion to the oxidized state, boiler additives can be applied to increase the relative proportion of oxidized mercury in thefluegas. For example, Nalco has developed a solution product that can be applied directly to the coal prior to combustion. It can be added to the coal at the pulverizers, or injected directly into the furnace. In numerous commercial trials, the fraction of mercury in the oxidized state has been increased to as high as 90%.17 By using such mercury oxidants the performance of mercury sorbents, including activated carbons, is significantly enhanced. This can be a major benefit to plant operations since, by using this technology, a reduction in sorbent feed rates (and thereby operating costs) can be achieved. In addition to reducing sorbent costs, this can also result in the preservation of a plant's ash sales since high carbon feed rates will render the fly ash unsuitable for use as a pozzolan in concrete manufacture. By reducing sorbent feed rates or eliminating the need for sorbents entirely, the value and quality of the fly ash sold by the plant to the concrete market can be maintained.
Coal pretreatment One option to limit the emission of mercury and other contaminants from a power plant is to prevent them from entering the process in the first place. While beneficiation of coal to enhance thermal output has been long practiced, the removal of minor elements such as mercury is more problematic. Nalco has developed a technology known as MagMill™ which can remove a large fraction of many of the heavy metals (e.g. Hg, Se, As) present in coal. This is completed in a dry process before the coal is burned. A combination of crushing, belt separation, and high powered magnets removes abrasive minerals like pyrites. By targeting the removal of pyrites, a high degree of mercury removal can be achieved since mercury in coal is known to be associated with pyrite occurrence.16 The result of processing the coal with MagMill is a product that is a cleaner, less abrasive coal with lower sulfur and heavy metal content.
While the fly ash is nominally a "waste product", the value can be quite substantial for those plants that are able to sell their fly ash. For example, a typical 500 MW plant generating 150,000 tons of fly ash annually that can be sold for $20/ton, risks $3 million per year in lost revenue by injecting activated carbon as a sorbent to control mercury emissions. In addition, the plant may incur an additional fly ash disposal cost of at least $1.5 million per year as a result. Mercury sorbents One of the most promising technologies to recently emerge for the control of flue gas mercury emissions is the use of mercury sorbents. Powdered sorbents such as activated carbon or engineered inorganic sorbents can be injected into the ductwork upstream of a plant's particulate control device to capture mercury.19 In this way, mercury is removed from the flue gas and
Mercury oxidation / speciation control Capture of mercury in downstream particulate control devices and/or scrubbers can be enhanced if the mercury is in the oxidized
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becomes commingled with the fly ash which can then be safely disposed of or, depending upon the quality of the ash, used in various applications including concrete, pavement, structural backfill, and brick manufacture. Activated carbon injection (ACI) is by far the most established technology in this area with over 60,000 MW of commercial bookings at over 150 coal-fired plants in North America as of June of 2010.20
mercury capture, however the target of 90% capture was only reached using an ACI rate well above 3 lb/MMacf.
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The application of mercury oxidation technology as described above is especially well suited to those plants that have wet flue gas desulfurization (FGD) units installed for SOx control, where the ionic form of mercury readily dissolves in the aqueous liquor phase. With the bulk of the mercury shifted from the gas phase to the aqueous phase, a coagulant product can be applied to sequester and precipitate the mercury from FGD liquors without any concomitant measurable increase in the mercury content of FGD byproducts. This also has the effect of suppressing the phenomenon of FGD mercury re-emission whereby oxidized mercury in the aqueous phase is reduced back to the elemental form and emitted from the stack.21 Mercury removal from wastewater In order to meet challenging industrial wastewater mercury discharge limitations, companies often turn to precipitation aids. For this application Nalco has developed polymeric chelants with an exceptionally high affinity for mercury. Upon binding mercury from solution, it forms large precipitates thatflocculate,settle, and are readily filtered to consistently attain extremely low mercury levels in the parts-per-trillion range. Case Studies The various technologies outlined above can be applied, either individually or in many cases in concert, to reduce and control mercury emissions. These strategies have been applied in a number of commercial power plant operations with excellent results. Two examples of how these technologies have been applied, each outlining how different control strategies can be used to achieve the desired results, are presented as case studies. Case study #1: Site description: 600 MWe pulverized coal-fired boiler firing subbituminous coal with low NOx burners, overfire air, and cold-side electrostatic precipitator (ESP).
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The overall aim of the operation was to increase mercury capture without impacting plant operations. Therefore, simply increasing the activated carbon injection (ACI) rate was not an option, as this would have compromised the quality of the fly ash produced at the plant, which was being sold for use in concrete. Thus, a 90% mercury capture with an ACI rate of less than 2 lb/MMacf was sought.
In this case, two separate issues needed to be addressed. First, improved capture of the mercury from the gas phase was required. Similar to the previous example, MerControl 7895 was applied to the coal feed to help achieve this. Additionally, the discharge of mercury in the plants wastewater was strictly controlled. While the improved capture from gas could be achieved, further containment and control of the mercury from the aqueous environment was required. For this purpose, a polymer chelant product (e.g., Nalmet® 1689) was used to treat the FGD wastewater.
The "normal" operation involved use of a halogenated carbon added to the outlet of the air pre-heater. Under such an operating mode, using an ACI rate of 2 lb/MMacf, the reduction in mercury was less than 80% (Figure 3). Increasing the ACI rate improved
188
Figure 4 shows the stack emissions of total mercury (in both elemental and oxidized forms) plotted across a range of dose regimes of the MerControl 7895 product. The same data are plotted as a percent of mercury capture. Application of MerControl 7895 at a rate of 265 mg/kg of coal resulted in greater than 85% mercury capture and achieved the target of 0.008 lb Hg/GWh. The net result of this treatment was that the bulk of the mercury reported to the FGD liquor. 100%
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Figure 5 shows the measured total mercury concentration in the clarifier effluent as a function of increasing mercury capture from the air. Despite the increasing load of mercury being captured and delivered to the scrubber liquor, the chelant treatment was able to maintain the effluent concentration well below the allowable target concentration. As shown in Figure 6, the mercury content of the clarifier solids increases with increasing air mercury capture. One can conclude then that as the MerControl 7895 oxidizes the mercury, it shifts from the gas phase to the scrubber liquor. The chelant addition in turn transfers the added mercury load from the liquor to the clarifier solids. With the mercury partitioned to the clarifier solids, it is then easily removed by standard wastewater treatment equipment. The application of this customized, low capital program enabled regulatory compliance for air and water mercury emissions, thereby shrinking the plant's environmental footprint.
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Figure 6. A plot of total mercury concentration of the FGD wastewater clarifier solids as a function of mercury capture from air. As capture of mercury from air increases, the concentration of mercury in the solids increases proportionately due to the application of Nalmet 1689. The use of coal as a power source for Bayer plants is common with coal accounting for approximately 20% of the energy used in the production of alumina.22 In many cases, the Bayer refinery operates their own coal-fired power plant as part of the overall refinery operations. Therefore, as noted above, such power plants can be a significant source of mercury emissions and in such cases, the control technologies described above, will clearly have a direct fit to the industry.
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Atmospheric emissions from alumina refineries are increasingly a problem with mercury being a common issue across the industry. However, similar issues are being faced by a number of other industries and a range of potential solutions are being developed and used for air emission and wastewater control generally, and in particular for mercury capture and control.
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Figure 5. FGD waste water mercury concentration plotted as a function of percent mercury capture from air. As more mercury is captured from the air (by application of MerControl 7895) the
189
9. Alcoa,'Emssions & Waste', (2010). http ://www. alcoa. com/sustainability/en/info_page/operations_e nvemissions.asp
Studies of various Bayer refineries have shown that both the amount of mercury emissions and the points in the process at which they are emitted vary quite drastically from refinery to refinery.7 The complexity of the problem stems from the various sources of bauxite and the differences in operating conditions employed at particular refineries. As is the case with coal-fired power plants, the issue of mercury emissions is so highly plant specific that a detailed knowledge of the various process steps and chemistries as well as the ability to accurately measure mercury in various process streams will be essential to delivering a customized solution for each individual refinery. It is expected that the aforementioned mercury control technologies in use at coal-fired power plants can be adapted to meet the needs of the alumina industry. Whether this will involve the removal of mercury from digestion liquors or the oxidation and sorption of gas phase mercury from digestion vents and calciner stacks, or some combination thereof, will ultimately depend on the particular needs of the refinery in question.
10. A. B. Livingston, 'Estimation of a mass balance for mercury at a Bayer alumina refinery, using thermal decomposition, amalgamation, and atomic absorption as an assessment method, M. S. thesis', (Texas A&M University, 2005). 11. M. Mullett, J. Tardio, S. Bhargava and C. Dobbs, 'Removal of mercury from an alumina refinery aqueous stream', Journal of Hazardous Materials, 144 (2007), 274-82. 12. H. G. Wildman, Process of recovering mercury', US Patent 1,896,876, (1933). 13. J. T. Malito, 'Process for removing heavy metals from a caustic fluid stream, US Patent 6352675', (2002). 14. A. P. Jones, Hoffmann, J. W., Smith, D. N., Feeley, T. J., Murphy, J. T., 'DOE/NETL's Phase II Mercury Control Technology Field Testing Program Updated Economic Analysis of Activated Carbon Injection, Preparedfor the U. S. Department of Energy', (2007).
References 1. D. G. Streets, Q. Zhang and Y. Wu, 'Projections of Global Mercury Emissions in 2050', Environmental Science & Technology, 43 (2009), 2983-8.
15. A. Kolker, C. L. Senior and J. C. Quick, 'Mercury in coal and the impact of coal quality on mercury emissions from combustion systems', Applied Geochemistry, 21 (2006), 182136.
2. A. Cain, 'Mercury Releases from Industrial Ore Processing', US EPA, (2005). http://www.epa.gov^s/reports/stakesdec2005/mercury/Cain2. pdf.
16. B. Toole-O'Neil, S. J. Tewalt, R. B. Finkelman and D. J. Akers, 'Mercury concentration in coal-unraveling the puzzle', Fuel, 78 (1999), 47-54.
3. E. D. Thoma, C. Secrest, E. S. Hall, D. Lee Jones, R. C. Shores, M. Modrak, R. Hashmonay and P. Norwood, 'Measurement of total site mercury emissions from a chlor-alkali plant using ultraviolet differential optical absorption spectroscopy and cell room roof-vent monitoring', Atmospheric Environment, 43 (2009), 753-7.
17. B. A. Keiser, G. Finigan, J. Meier, J. Shah and J. Lu, '#145, Mercury Emissions: Demonstration of Air and Water Quality Management', MEGA Symposium (Baltimore, MD, 2010).
4. J. B. Milford and A. Pienciak, 'After the Clean Air Mercury Rule: Prospects for Reducing Mercury Emissions from CoalFired Power Plants', Environmental Science & Technology, 43 (2009), 2669-73. 5. D. Wolf, 'Understanding the Industrial Boiler MACT Rule', HPAC Engineering, (2010).
18. K. H. Pedersen, A. D. Jensen, M. S. Skjoth-Rasmussen and K. Dam-Johansen, Ά review of the interference of carbon containing fly ash with air entrainment in concrete', Progress in Energy and Combustion Science, 34 (2008), 135-54. 19. H. Yang, Z. Xu, M. Fan, A. E. Bland and R. R. Judkins, 'Adsorbents for capturing mercury in coal-fired boiler flue gas', Journal ofHazardous Materials, 146 (2007), 1-11. 20. Institute of Clean Air Companies (ICAC), 'Updated Commercial Hg Control Technology Bookings (June 2010)'. http://www.icac.com/files/members/Commercial_Hg_Booking s_060410.pdf
6. US EPA, 'EPA Sets First National Limits to Reduce Mercury and Other Toxic Emissions from Cement Plants', (2010). http://yosemite.epa.gov/opa/admpress.nsfe77fdd4f5afd88a385 2576b3005a604ffef62balcb3c8079b8525777a005af9a5!Open Document
21. G. Blythe, J. Currie and D. DeBerry, 'Bench-scale Kinetics Study of Mercury Reactions in FGD Liquors', Final Report — DE-FC26-04NT42314, (2008).
7. C. Dobbs, C. Armanios, L. McGuiness, G. Bauer, P. Ticehurst, J. Lochore, R. Irons, K. Ryan and G. Adamek, 'Mercury emissions in the Bayer process - an overview*, Proceedings of the Seventh International Alumina Quality Workshop (Perth, Western Australia, 2005), 199-204.
22. International Aluminium Institute, 'Form ESO 12 - Energy Used in Metallurgical Alumina Production', (2009). https://stats.worldaluminium.org/iai/stats/formServer.asp?form=8
8. Bauxiet Instituut Suriname, (2006). http://www.bauxietinstituut.com/files/Environment%20%20an d%20Technology%20-%20Suralco%20LLC.pdf
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Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
ALUMINA REFINERY WASTEWATER MANAGEMENT: WHEN ZERO DISCHARGE JUST ISN'T FEASIBLE.... Lucy Martin1 and Steven Howard1 ^echtel Australia, 100 Brookes Street, Brisbane, Queensland 4006, Australia Keywords: "Alumina Refinery, Waste water, Environmental Design Criteria" of the refinery will be strongly influenced by the investment required for materials transportation. Project NPV will in turn be impacted by the operating costs associated with that location. As commodity prices increase and accessible deposits become depleted, the need to develop inland ore bodies, and consequently refineries, is increasing.
Abstract Management and treatment of liquid effluents are determinant considerations in the design of alumina refineries. Rainfall, evaporation rate, proximity to the coast, process design and layout, ore mineralogy, the local environment, and potential impact on contiguous communities are all integral to the development of an appropriate refinery water management strategy. The goal is to achieve zero discharge of liquid effluent to the environment. However this is not always the most feasible solution under the extreme rainfall conditions in tropical and subtropical locations. This paper will explore the following issues for both inland and coastal refineries: • • •
Bauxite is formed through weathering of lateritic rock under high rainfall in tropical and subtropical environments. Figure 1 shows regions currently undergoing lateritic weathering. These regions tend to have well defined wet and dry seasons, which forces refinery designers to address the "feast or famine" issue associated with water availability. Zero discharge of effluent is ideal for sustainability, but is often impractical for water impoundment.
Methods to reduce and control refinery discharges Treatment design criteria Socioeconomic aspects relating to surface water use in settlements adjacent to the refinery
Bauxite also contains organic and inorganic impurities which increase the costs and environmental risks associated with its processing into alumina. This is primarily due to the solubility of the various minerals and organic compounds in the Bayer process liquor, and the tendency of the impurities to accumulate therein. Demonstrated treatment methodologies for the large scale effluent neutralization required for discharge into surface or marine waters primarily involve the use of either seawater alone, or seawater supplemented by spent sulfuric acid, a byproduct of the refining process. When considering refineries close-coupled to inland bauxite deposits, the cost of pumping seawater must be set against the cost of bauxite and alumina transportation.
Introduction With an increased global focus on sustainable development from governments, international financing institutions, mining companies, and communities, new and expanding mining operations can face a higher level of environmental and sociological regulation and scrutiny than in the past. Many legislative frameworks allow existing operations to continue to meet a less stringent standard than would be expected on a greenfield facility. Therefore this paper will pay most attention to those regions expected to become future alumina sources and the associated technical barriers to sustainable and viable projects.
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In summary, this indicates that: 1. Water availability for refinery consumption will be highly variable from season to season 2. Water discharge will be required during both "extreme" or "normal" rainfall events. 3. In inland locations, water effluents may discharge into streams potentially used for drinking and primary industries (farming, animal husbandry, and fishing) 4. Typical refinery waste water treatment techniques, such as seawater neutralization, may not be feasible.
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Refinery Water Balance Water supply and demand An alumina refinery that utilizes the Bayer process will consume 2.0-2.3 tonnes of raw water per tonne of alumina produced. The actual rate will depend upon the bauxite quality, the process design, demand for nonprocess applications (e.g, for potable water), and the extent water is recycled within the facility. About 10 percent of the total intake is accounted for by free moisture in the bauxite feed, and in the 50 percent caustic soda solution, the primary process reagent.
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Figure 1. Lateritic Bauxite Sources (modified from Freyssinet et al [l]) The relationship of a refinery location to its bauxite deposit and transport infrastructure directly affects material handling equipment costs and resultant project NPV. The optimum location
Figure 2 depicts the inputs and outputs of water across the outer system boundary between the refinery, including the residue
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disposal area, and the external environment. This assumes one hundred percent diversion of potential 'run-on' to the site.
various facilities. These measures minimize the probability of releasing highly concentrated, potentially toxic material which can only be discharged under the most extreme circumstances. Environmental control measures for an alumina refinery must also include facilities to handle contaminated runoff, due to rainfall catchment within the facility perimeter. One or more stormwater ponds (SWPs) would also be provided, and must be considerably larger than the process spill ponds. An SWP is intended to impound rainfall runoff collected from nonprocess areas, which will mobilize relatively minor amounts of surface contaminants during the initial onset of rain. After a short period, typically about one hour, the runoff quality is similar to that found in neighbouring areas outside the refinery perimeter. The impounded SWP water is impure, but is suitable for recycling into the process for various duties, reducing the intake of raw water. Finally, the impoundment area constructed for the permanent storage of bauxite residue and other solid wastes, the bauxite residue disposal area (BRDA), also accumulates contaminated water, which may be returned to the process in order to recover the soda content.
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Figure 2 Refinery Water Balance There are only two sources of water that supply the remaining ninety percent of the total demand: raw water from natural surface and subsurface sources, and rainfall. This means that a significant proportion of the water intake is noncontrollable, and under certain circumstances may far exceed the capability of the refinery to control the quantities involved. With the notable exceptions of the People's Republic of China (PRC) and Russia, most refineries have been established in tropical and subtropical regions, in close proximity to the principal lateritic bauxite provinces. These are not only exposed to high seasonal rainfall, but are also at risk of extreme and unpredictable flood events. New capacity planned for Brazil, Guinea, and SE Asia will be faced with the same issues.
Under normal operating conditions, the Bayer plant liquor inventory is controlled within narrow limits by varying the input rates of caustic soda, raw water and in-plant evaporation. Operating procedures should be aimed at keeping the process spill ponds empty, and the water levels in the SWP and BRDA as low as possible, during periods of little or no rain. The onset of heavy and/or sustained rainfall, typical in tropical locations, gives rise to large volumes of site runoff, and potentially run-on if this is not effectively diverted. This can rapidly fill the available impoundments, which have a finite limit despite the risk analysis and major investment involved in their provision. In order to handle situations that exceed the maximum impoundment capacity, the regulatory option often exists to impound only the so-called "first flush" runoff, after which all subsequent catchment would be allowed to bypass the SWPs and be released to the environment. Exceptional rainfall events would defeat even this strategy (assuming that it were permissible) and the refinery would then be forced to release contaminated water to the environment. Such a release may be diluted to mitigate the alkalinity of the effluent, but would still result in contaminating receiving waters above background levels, probably in violation of licence provisions.
Environmental control requirements It has long been recognized that effective environmental management is critical to the viability of any project. Stringent environmental control standards and industry best practice with regard to operations and maintenance must be reflected in the refinery design criteria.
Community demand, including agriculture, for water may become a significant component of total consumption. Any shortage or degradation of the community supply may become a significant issue. These factors need to be considered in the overall conceptual design of the water supply system.
The focus of this paper is the management of liquid effluent, any release, either treated or untreated, of which is potentially harmful to the environment and consequently deleterious to the wellbeing of the community. The Bayer plant itself handles a large volume of process liquor, the bulk of which is an aqueous solution of caustic soda containing dissolved aluminium, silica, and many other organic and inorganic impurities, including trace metals such as Molybdenum and Vanadium, which occur in the bauxite.
The conclusion is that despite careful scenario planning and investment in major infrastructure, sooner or later an extreme event will cause an unacceptable environmental incident. Therefore, the location of the refinery must be such that the receiving waters are able to withstand the impact in a sustainable manner. This explains why most major refineries are located near the ocean, which also facilitates the application of seawater neutralization, the only current demonstrably effective liquid effluent treatment technology.
The refinery is designed to contain the live liquor inventory, with minor spills within operating facilities being promptly returned into the process. Provision is also made to intercept larger, accidental spillages which could result from equipment failure or abnormal events, such as the loss of electrical power. Spill ponds are installed at strategic locations, and have sufficient capacity to handle the contents of one or more of the largest tanks in the
Refinery design principles The unavoidable accumulation of contaminated water due to rainfall, reclamation constraints, and severe limitations on release to the environment have the following consequences for the design of the refinery:
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1.
areas is perceived to improve employee safety by minimizing the hazard of caustic liquor exposure, but this may be achieved by proper safety procedures and training, whilst minimizing water usage.
Refinery environmental policy that mandates compliance with relevant regulatory standards becomes a major driver for selection of the refinery and BRD A locations.
This policy also has a profound impact on refinery design features and the operational procedures that enable the standards to be met.
4.
A policy mandating compliance with liquid effluent standards is consistent with that for particulate emissions or any other form of environmental impact. The difference between implementation for particulates (e.g., the installation of dust collectors) and that for release of impounded water is one of relative complexity for the latter and requires coordination of the total refinery design.
Water consumption may be allocated to two broad categories: noncontrollable, governed by the combined requirements for essential services, and controllable, which allows for some degree of flexibility as to the quantity used. Most of the process requirements fall into the noncontrollable category, for example, boiler feedwater, flocculant, and lime preparation. All are examples of end uses directly related to the refinery production rate and to the demand for raw materials and additives. Domestic water demand is closely related to the number of employees. Effective control depends upon the installation of water saving devices such as tap restrictors, employee education regarding conservation, and the prohibition of usage for nondomestic purposes.
Facility design criteria must address the (water) environmental policy, ranked among safety and NPV considerations, and evaluate this before detailed design can proceed. Some objectives may be compatible, for example minimizing the caustic content of final stage mud washer underflow results in a significant reduction in operating cost. 2.
Departures from conventional refinery and waste impoundment design become necessary, and can add to the mitigation costs of potential effluent management problems.
The final disposition of water usage in this category is important in limiting the potential environmental impact. Applications such as dust mitigation can help to dispose of excess impounded catchment under dry season conditions. Others, such as vehicle washing, may create additional problems caused by runoff turbidity or contamination by hydrocarbons and other chemical agents.
A refinery location that will experience extreme rainfall events should incorporate features to minimize both the volume effects and possible contamination of rainfall catchment. This facilitates the segregation of concentrated process liquor (that cannot be released) from relatively innocuous nonprocess runoff that can be impounded, recycled during dry periods, and released subject to meeting licence conditions.
The supply of water to end users beyond the outer system boundary should be from an independent source of supply. This measure imposes a physical constraint on consumption and avoids uncontrolled influence on the infrastructure provided for the refinery itself.
Waste impoundment design must satisfy all of the following criteria: • Effective, permanent sequestration of red mud waste that will continue to generate alkaline leachate over the longer term • Recovery of supernatant liquor and impoundment drainage back into the refinery process. • Sequestration of solid-phase organic and inorganic waste from process liquor purification facilities • Disposal of other environmentally sensitive wastes (e.g., ash, scale, and waste acid) 3.
Introduction of measures to enforce economic use of controllable process and domestic water can significantly reduce usage.
5.
Several tiers of impoundments must be established to segregate process liquor, contaminated effluent and rainfall catchment.
• Process spill containment A "process spill" is defined as any form of liquor, slurry, or solids released during routine operational tasks within the plot limit of any facility. Spills of this nature are normally of low quantity and are promptly collected and returned into the process. The concentrations of process chemicals are too high to permit release to the environment.
Measures must be adopted to minimize the controllable intake of raw water, particularly during the wet season, thereby reducing the total inventory to be managed.
There is a possibility that large process spills may occur due to equipment failure, such as a pipeline rupture. The refinery design should incorporate process only spill ponds for the purpose of intercepting and returning material that may overtop the limited facility containment capacity. The contents of spill ponds cannot be released, and should be recycled to the refinery via large sump relay tanks (SRTs), which provide additional surge capacity and operational flexibility.
Due to the probable excess of rainfall over evaporation in the regions under discussion, raw water intake must be strictly controlled to minimize the total volume under management. It follows that the consumption of raw water in the Bayer process should be minimized by using as much water as possible that is already within the outer system boundary. Raw water is needed for a number of essential process duties, treated for domestic (potable) water, and for the fire control infrastructure. The number of water entry points into the refinery must be minimised, and all volumes accurately metered. Strong design discipline is necessary to ensure consistent water control policies and practices are enforced. Some measures may conflict with custom and practice employed elsewhere, but capital and operating costs associated with treatment of this water may be excessive. For example, using hoses for housekeeping in process
Process design modelling should incorporate provision for large process spills, to ensure total containment facilities and return rates into the process are adequate, and that defined events of this nature do not compromise the refinery's rated production and other key parameters. •
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Runoff management
"Runoff is defined as rainfall catchment from the refinery plot limit that must be monitored for contaminants and handled accordingly. Runoff falls into two categories: •
•
temperature and turbidity) associated with the effluent itself and the background conditions in the receiving waters. • Final Stage Mud Washer Underflow Slurry The residue slurry contains solid phase compounds that will release highly alkaline leachate over a long time frame due to unavoidable contact with supernatant liquor and rainfall. This tendency cannot be controlled at source, due to the presence of sodium aluminium silicates and calcium compounds formed in the Bayer process. The only control, at the margin, is to minimize the concentration and quantity of liquor disengaged from the residue, and to reduce direct contact of the consolidated mud with liquor or rainwater.
Catchment within process facility plot limits. This may be relatively dilute, but is still far too contaminated for release. It must be returned to the process via the SRT system, and will have negative impacts on the process and energy demand. These effects must also be considered for the refinery mass and energy balance. The impact must be minimized by reducing the process catchment footprint, i.e., by roofing large tanks and routing rainwater catchment to surface drains outside the facility plot limit. Catchment from nonprocess areas. Runoff will mobilize soluble contaminants such as oil, dust, and dirt, the concentrations of which may be appreciable depending upon such factors as paving and drainage design, control of fugitive dust from the refinery, and general housekeeping standards. Under tropical rainfall conditions, the initial (or "first-flush") nonprocess runoff will exhibit short term contaminant loadings that will be sampled and the flow directed to a large Storm Water Pond (SWP).
The major controllable variables for residue disposal are the density of the slurry, and the alkaline concentration of the liquor discharged with the slurry into the BRDA disposal area. Paste thickening is the most effective means of preparing very dense, immobile slurry which will release little or no liquid as it further consolidates under its own weight. The remaining design issue is to minimize the final stage mud washer underflow total soda, usually measured as g/L Na2C03. A typical target is 5 g/L, estimated from the number of washing stages to be installed, the ratio of wash water to mud, and other factors. The target is estimated by the steady state refinery mass balance, and by making certain assumptions related to equipment reliability, washer stage efficiency, etc. Under actual conditions, major departures from steady state operation and other assumptions cause excursions in last washer soda, by as much as a factor of 10, to 50 g/L or more. This implies that the number of installed washing stages and the sparing of pumps should go beyond that suggested by steady-state modelling; otherwise the environmental control objectives will be unachievable.
In practice, the volume of runoff from "average" rainfall intensity can be far greater than the SWP capacity, as costs to construct these facilities to collect the rainfall volumes experienced in tropical and sub-tropical locations can be prohibitive. This can mean that all but a small fraction of nonprocess runoff must be released, and the refinery must be designed with this in mind. The water diverted to the SWP will be reclaimed as "process water", diverted to the BRDA buffer zone, or (as a last resort) treated and released. There is a tendency to work with "average" rainfall data that can be highly misleading, and could result in erroneous assumptions underlying the planning for the capacities of spill ponds, SRTs, the SWP, and the BRDA. A credible environmental control strategy should be based upon scenario planning for abnormal events, such as 1:100 year return rainfall events.
Discharge Criteria Where bauxite mining and alumina refining are emerging, local and national legislation and associated compliance monitoring are often not sufficiently developed to address contaminants of interest associated with these activities. Consequently, drinking water, primary industries and ecosystems may not be adequately protected by the existing environmental legislative framework.
• BRDA design The BRDA is one of only two impoundments from which lowlevel contaminated water may be released (the other being the SWP), provided that the regulatory and/or best environmental control practices can be addressed. It may be possible to optimize the total capital investment for the two impoundments by linking them, such that the BRDA provides the final and somewhat larger capacity for low level effluent.
Many international mining and minerals processing companies have implemented sustainable development policies which require them to not only meet the requirements of in-country legislation, but to also consider the use of the best available technologies and international guidelines. For projects seeking external financing, the International Financing Corporation (IFC) has developed specific Environmental Health and Safety (ES&H) guidelines that are "technical reference documents with general and industryspecific examples of Good International Industry Practice (GIIP)". These guidelines are used by the IFC for project appraisals. IFC EH&S guidelines have been developed for both mining and base metal smelting and refining [2,3]. However these may not include all contaminants of interest relevant to alumina refinery effluent discharge. For example, the base metal and smelting guidelines state that toxicity should be considered on "a case specific basis" and do not provide guidelines for discharge of heavy metals associated with alumina refining.
The design of the BRDA must prevent contamination of impounded runoff outside the deposition areas for red mud and the other solid wastes that have special sequestration criteria. Liquor that is either released from the residue slurry or accumulates from rainfall directly onto the mud deposition area must be recycled to the refinery. A buffer zone between the active solid waste deposition zones and the outer BRDA perimeter should be provided to allow for abnormal rainfall events that may temporarily exceed the active zone's capacity. This liquor, and any other leachate from other waste disposal areas, should not be allowed to mix with impounded low level water.
Assuming that local legislation is not sufficiently defined, guidelines from international organizations and other countries with highly developed legislativeframeworkscan also be used to develop robust design criteria for effluent discharges from the project site.
The remaining area of the BRDA may be used to store and reclaim, or to treat and release, excess water. The final treatment and disposition method will be site-specific, and will depend upon a number of factors (alkalinity, toxic ionic species, BOD,
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An alumina refinery will generate large volumes of liquid and solid phase alkaline effluent from a variety of sources, predominantly waste mud and other process byproducts discharged with the mud, including:
Design criteria for effluent disposal must consider the environment to which the effluent will be discharged - a primary consideration is whether the discharge is to fresh water or the marine environment. Discharge quality requirements to marine water are often less stringent than the discharge to fresh water due to various factors, including higher naturally occurring baseline concentrations of compounds in marine water compared to freshwater, increased dilution of effluent when discharged to open water, and less sensitive receptors from both an ecosystem and downstream user perspective in saline water.
• • • •
Surface water is used for drinking, recreation, and primary industries in many countries. In developing countries it is more likely to be used for these purposes without treatment. Strict and enforced industrial zoning legislation is required where industrial effluent is discharged to control population influx and resultant water usage; however this is often lacking in the regions currently experiencing growth in bauxite mining and associated refining. In these situations it must be considered that human and animal ingestion of surface water could take place at any point outside of the industrial fenceline, therefore drinking water quality standards, as a minimum requirements, must be attained at inland refineries discharging to streams and rivers.
Diluted process liquor entrained with the bauxite residue (red mud) Insolublefractionof the bauxite (iron and titanium oxides) Desilication product (DSP) - hydrated sodium aluminium silicates Calcium compounds, e.g., tri-calcium aluminate, calcium oxalate.
The other effluent streams will be generated by contamination of rainfall runoff from the refinery site, and possible spillage and release of process materials to impoundments. Effluent that is released from the waste mud stream will invariably exhibit pH levels above 12. Both the liquid and the solid phase sources of alkalinity (hydroxide ions) must be reacted with a neutralizing agent. The rate at which the neutralization reaction proceeds will vary greatly depending upon the reactants involved, the pH, concentration, and temperature. Principles of Effluent Treatment The primary objective is to reduce the immediate and longer term environmental risk of the (solid or liquid) waste stream by reducing the alkalinity to the minimum practicable level. However, low alkalinity is not the only consideration. Potentially serious soluble pollutants such as aluminium and other elements (e.g., molybdates, vanadates and arsenates) must also be targeted.
An alternative is to provide a separate secure drinking water supply to the local population and educate them regarding withdrawal of water, which may limit the treatment costs for effluents. If this can be achieved and is accepted by permitting authorities, then it may be possible to discharge at a higher concentration and allow for mixing until drinking water, or other international standards, are achieved at a compliance point downstream of the discharge location. Mixing zone calculations need to take into account the lowest surface water flowrate when calculating dilution, as this can often mean that discharge criteria to achieve compliance at a the edge of a mixing zone need to be more stringent during the dry season. This applies to impurity concentrations, temperature, and turbidity (visual impact)
Options for neutralization include reaction with seawater, dilute sulfuric acid (or a combination of both), and carbon dioxide. A third option, carbonation, is in the early stages of development and little process performance data is available to suggest this as a viable option at this phase of development. Application of either seawater or acid neutralization must recognize that a certain fraction of solid phase alkalinity is released over long periods of time, weeks or months, depending upon the prevailing conditions and the composition of the residue. In practical terms, there is no such thing as complete neutralization, due to the limited treatment time available.
Non-Governmental Organizations (NGOs) are an increasingly prominent force in the enforcement of environmental and sustainable development policies. There is an increasing public demand for independently verified information and action, due to the poor performance of some regulators and companies. Therefore, in countries without sufficient environmental legislation to protect the health of local communities, their livelihoods, and the surrounding flora and fauna, large corporations are often under pressure to implement stringent environmental and sustainable development policies to protect their reputation.
Treatment with sulfuric acid invariably involves the utilization of spent dilute acid used for cleaning refinery heat exchangers. During the cleaning process, the acid dissolves scale deposits which may contain additional pollutants. A corrosion inhibitor will also be employed, the nature of which must be assessed if the reacted acid is to be released into the environment.
In summary, development of project specific effluent discharge criteria to meet sustainable development policies will require a vigorous social, health, and environmental impact assessment of existing and potential future uses, and a "one size fits all" solution cannot be applied. However, generally speaking, discharge quality requirements to marine rather than aquatic environment are less stringent and consequently more achievable.
The chemical and physical properties of the particular bauxite to be processed have a determinant impact upon the process design of the Red Side of the refinery, upon the BRDA design, and upon the technology selected for effluent treatment. Significant effort and cost must be invested in the characterization of the bauxite, and the same attention must be paid to the environmental control requirements.
Effluent Neutralization Technologies
Basis of Design The Refinery Basis of Design must specify the following fundamental criteria: Refinery location. . • Will it be practicable to utilize seawater neutralization? Climatic conditions - rainfall and evaporation.
In this context, the term neutralization refers to the reduction of the level of alkalinity in the stream to be treated. Neutralization processes are required for effluent that is to be permanently sequestered, or is to be released to the environment where there is no other alternative.
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hydrotalcite precipitation will occur at the outfall, creating a visible plume.
•
Will it be necessary to release effluent to the environment, and what quantities are likely? Receiving waters for effluent release. • What are the social, agrarian and environmental factors associated with effluent release? • What effluent discharge standards should apply? • What are the critical parameters for the outfall point? Waste streams to the BRDA • What tonnages of solids and liquid waste are to be discharged? • What will be the composition of the red mud? • What variability in waste loadings and concentrations may be expected? • What is the probable toxicity profile, and does mitigation technology exist?
It should be noted that attempts have been made to augment seawater neutralization with sulfuric acid. This significantly alters the chemistry such that a much lower pH is necessary to remove aluminium from the solution. Removal of oxalate is favoured at high pH values, so that the addition of acid may be counterproductive and costly. Alternative Technologies Other technologies for industrial water treatment, such as membrane treatment and ion exchange, are untested on alumina wastewater chemistry and at the scale discussed in this paper. If these technologies were proven to be effective, the volumes that would require treatment in tropical and sub-tropical locations would result in significant increases in capital and operating costs for these facilities.
Acid Neutralization Residue is mixed with dilute acid which reacts immediately with soluble alkalinity, producing a rapid but temporary drop in pH. This is commonly achieved by adding (waste) acid after the last mud washing stage. Attack by residual acid (if any) on the solid phase alkaline content occurs over a much longer time frame, and may lead to a gradual increase in pH. It is therefore impractical to neutralize the solid phase component prior to residue disposal. Post-neutralization of waters released from the BRDA may be necessary.
Conclusion In tropical and subtropical climates, coastal and inland refineries will discharge effluent under normal and extreme circumstances. To minimize the amount of effluent requiring treatment and discharge, the environmental design features and operational controls for management of raw water intake, and the water balance detailed in this paper should be incorporated into the alumina refinery design.
Acid neutralization produces a dilute sodium sulphate solution, which, if released, may give rise to environmental impact in its own right, such as algal blooms or local concentrations exceeding background levels, or that specified for potable water (<250 mg/L). Careful assessment must be made to establish that sufficient acid will be available for primary/secondary residue or effluent treatment, that probable peak effluent discharge rates can be handled, and potential heavy metal contamination of spent acid is acceptable.
Inland refineries pose additional challenges due to the stringent discharge criteria applied to inland waterways. Discharge criteria can be imposed, such as government legislation or financial institution requirements, or prescribed by company internal sustainable development and environmental policies. In either circumstance, effluent discharge requirements are often defined according to the expected use of the water body by contiguous communities. The most demonstrated treatment method for large scale refineries is seawater neutralization. Transport of seawater to an inland refinery and the reverse for discharge can put a strain on the project capital and operating costs, and may not be feasible.
Seawater Neutralization The important reaction in seawater neutralization is the precipitation of hydroxyl ions by reaction with magnesium (Mg++) ions present in the seawater. Again, the rate of reaction varies greatly - rapid in the liquor phase, much slower with calcium compounds and DSP. The presence of sulfate ions will inhibit the reaction with DSP.
Other water treatment methodologies are not proven on large scale refineries required to meet stringent inland water requirements and may require expensive reagents to treat the effluent. These can increase the risk profile of the project, drive up capital and operating costs and may result in breaching of discharge design criteria, with consequential downstream impacts.
A major advantage of seawater neutralization is the precipitation of aluminium ions during the formation of hydrotalcite, the primary reaction product from the soluble alkalis and magnesium. Hydrotalcite formation has also been shown to remove vanadium, molybdenum and phosphorus provided that the pH is in the range 8.0-10.0.
Acknowledgements Thanks to Bob McCulloch and Bill Imrie for their valuable comments and Rosemary Dotlic, Daniel Hayes & Eric Tlozek from the Bechtel Communications team for their assistance in formatting and editing this paper.
Seawater neutralization of residue allows immediate effluent discharge, provided turbidity criteria can be attained, eliminating the need for separate containment and management of liquor. By providing excess seawater, alkaline runoff generated by the slow dissolution of alkaline compounds can be neutralized and released.
References 1. Ore-forming Processes Related to Lateritic Weathering; P.H Freyssinet, C.R.M Butt, R.C Morris and P. Piantone. Society of Economic Geologists Inc. Economic Geology 100th Anniversary Volume p681-722 2. IFC General Environmental, Health, and Safety Guidelines: Mining, December 10, 2007 3. IFC Environmental, Health, and Safety Guidelines Base Metal Smelting and Refining April 30, 2007
System design must ensure that the seawater supply and discharge capacity always exceeds the magnesium demand of excursions in residue alkalinity, due to process problems. If the pH is allowed to rise, some of the trace metals in the hydrotalcite will revert to the soluble phase and impact on receiving water quality. If excess alkalinity cannot be precipitated within the BRDA, additional
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Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
HIGH PURITY ALUMINA POWDERS EXTRACTED FROM ALUMINUM DROSS BY THE CALCINING—LEACHING PROCESS LIU Qingsheng u , ZHONG Chunming1, FANG Hui1, XUE Jilai 2 2
falculty of Material and Chemical Engineering, Jiangxi University of Science and Technology, Ganzhou 341000 , CHINA; School of Metallurgical and Ecological Engineering, University of Science and Technology Beijing, Beijing 100083, CHINA Keywords: aluminum dross, alumina, extraction, calcining-leaching process order to decompose the mullite phase present in the Al dross. Due to the low cost and stability of sulfuric acid, this acid is used as a lixiviant during the recovery of A1 2 0 3 from sintered mixed dross. The impurity ions in the leaching liquor were removed by means of EDTA complexation, and washing with distilled water. This is a low temperature process that does not consume too much energy, and high purity A1 2 0 3 is obtained.
Abstract A processing technology was developed for alumina extraction from Al dross with the calcining-leaching process. The aluminum dross was mixed with soda and sintered at 1173K to yield soluble aluminates. Subsequently the sintered dross was leached with sulfuric acid to produce a solution containing aluminum . The unwanted metal ions, including Fe 3+ and Na+, were removed by ethylene diamine tetraacetic acid (EDTA) and water washing. Then NH4HCO3, was added controlling the crystallization of NH4A10(OH)HC0 3 , and the drying and calcining process was carried out, resulting in ultra fine A1 2 0 3 powders with high purity. The characteristics of the A1 2 0 3 powders were examined by means of XRD and SEM. The extraction efficiency of A1 2 0 3 can
Experimental The primary raw materials of the experiment were Al dross, soda and H 2 S0 4 . The compositions of the Al dross are listed in table 1. Al 39.77
surpass 98 % by optimization of the calcination and lixiviation processes. Well-dispersed fibriform A1 2 0 3 powders were obtained by calcining at 1000°C and the purity of the ultra fine A1 2 0 3
Table 1 Chemical Composition of the Al dross(wt%' M K Ca Fe Ti Si O 8 Na 1.33 0.56 1.59 4.98 4.57 0.55 0.40 41.59
The recovery of A1 2 0 3 from Al dross is based on application of hydrometallurgical processes such as acid or base leaching, purification, precipitation and calcination. The procedures of the experiment are illustrated in Fig.l. The crystalline phase and microstructure were characterized by X-ray diffraction (XRD) and scanning electron microscope (SEM).
powders was more than 99.36% . Introduction Billions tonnes of Al dross produced during the casting or remelting of aluminum have accumulated over the years. The majority of this dross is disposed off in landfill sites, causing serious pollution of the environment!!]. Finding methods for producing useful materials from the dross is a very important task for society.
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Aluminium dross contains mostly aluminium, silicon and other metal oxides and salts such as KC1 and NaCl. Thus it has the potential to be a rich alumina source, and is a good raw material suitable for production of aluminum and its chemicals. Different types of dross have been employed as raw materials by several authors. Dross includes considerable amounts of aluminum alloys, and it has been utilized as a deoxidsing agent in steel making, as a refractory material, and as a cement material [2-5]. Attempts have been made to extract alumina from the alloy dross by adopting either pyro- or hydrometallurgical methods [6,7]. Amer carried out the dissolution of waste dross in two steps, in order to produce a highly pure aluminum sulfate[8]. However, these processes have found little practical application because of the highly corrosive nature of the concentrated acid and alkali involved, and because these processes also constitute an environmental hazard.
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The objective of this work is to facilitate the recycling of Al dross and produce high purity A1 2 0 3 . In this paper, the calciningleaching method was used to recover A1 2 0 3 from Al dross. In the proposed process, Al dross is mixed with soda prior to sintering in
Α12Ω2. Fig.l Flow chart of extraction of alumina from Al dross
197
b. Sintered mix powders Fig.2 XRD patterns of raw fly ash and calcined ash
Results and discussion
SEM analysis of the sintered raw Al dross and sintered dross also gave similar results, as shown in Fig.3. SEM image of sintered dross exhibited mainly as clumps. They were formed by reactions of sodium hydroxide, aluminum, and silicon oxide in the sintering process. These clumps were irregular in shape. However, Glassy material was found in the SEM image of raw Al dross, as shown in Fig.3(a). This was the evidence of existence of a mullite phase in the raw material.
Sintering the mixture of Al dross and soda The main objective of sintering the mixture of Aluminum dross and soda was to break the crystalline mullite phase (3Al203-2Si02) and the mesh framework of the glass phase, rendering free aluminate for leaching [9]. The following reactions occur during the calcining of fly ash and soda[10]: NqCQ=NqO+Cqî
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Figure 2 shows the XRD patterns of Al dross and the sintered mixture of Al dross and soda. From this figure we can see there are characteristic peaks of mullite in the raw material, while after sintering, these peaks disappeared and many new characteristic peaks were found. So the stable mullite crystalline phase was successfully broken-up by the sintering. NaA102 is the major component of the sintered mix. 2CaOSi02 and Na2OFe203 are the minor components. The soluble aluminum materials are dissolved. The process also produced insoluble residues called red mud.
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The sintering parameters such as sintering temperature and sintering time were considered to find the optimal activation efficiencies (rendering the maximum free aluminates). As shown in Fig.4, 1173K was the optimum temperature. At this temperature, the sintered material had the highest alumina extraction efficiency. Sintering below or above this temperature would produce abnormal sintered products, which led to lower alumina extraction efficiency.
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The purpose of the leaching process is to leach the soluble Al ions (in the form of sodium aluminates). The possible reactions that take place in the leaching of the raw Al dross are as follows :
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The leaching parameters such as the concentration of sulfuric acid, leaching time and leaching temperature were considered to find the optimal extraction efficiency of AI2O3.
80 °C the extraction efficiency is low. Increasing the leaching temperature can accelerate the progress of the leaching reactions and improve the aluminum extraction efficiency.
The results presented in Fig.5 and Fig.6, indicate that the efficiency of aluminum species extraction increases with an increase in both acid concentration and leaching time. Increasing the acid concentration can accelerate the leaching reaction and achieve high extraction efficiency. However, when acid concentration reached a certain value, the extraction efficiency may not improve. The optimum acid concentration was 0.9mol/l at which the alumina extraction efficiency reached the maximum. To some degree, prolonging the leaching time can increase the aluminum extraction efficiency,as prolonging the leaching time will allow the completion of the leaching reactions. Dissolution temperature/K" 100
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Fig.7 Effect of dissolution temperature on alumina extraction efficiency
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The leaching liquor obtained from the previous processes contained Al , Na+, Fe3+, Ca2+, Mg2+, and other impurities. In order to obtain high purity A1203, the impurity ions had to be removed. The selective precipitation of ions by pH value adjustment was found to be unsuccessful in separating Fe3+ and Al3+[11]. We chose ethylene diamine tetraacetic acid (EDTA), a complexing agent, to remove the Fe3+ and other impurity ions. Selectivity arises because the complexing ability of each metal in the leaching liquor with EDTA is different [12,13].
Φ
«
20
H2S04 concentration/(/T7oi/m/)
Fig.5 Effect of H2SO4 concentration on alumina extraction efficiency The dissolution time was defined as the period that started at the time when the sintered mixture was introduced into the acid and ended when the liquid was separated from the red mud. This was the time period that the dissolution solution actually contacted the sintered mixture product. The longer the dissolution process is run, the greater the loss of the alumina extraction efficiency. This was due to the dissolution of Na2OFe203 as well as the side reactions. Since the particle size of the sintered mixture was very small, a short dissolution period actually gave the best result. The exact relationship between the dissolution time and aluminum extraction efficiency is shown in Fig.6.
The pH value of the leaching liquor was adjusted to 3.0 before the EDTA was added, and then the solution was continuously stirred for 30 minutes. The color of the solution turned from dull yellow to bright yellow. NH4HCO3 was subsequently added into the solution to precipitate the aluminum. At the same time the solution was quickly stirred. After filtration, white precipitates and yellow solution were obtained. These results show that the Fe3+ contained in the leaching liquor was effectively removed. The other soluble impurities absorbed in the precipitates could be removed by three cycles of washing. A white gelatinous precipitate was obtained and then dried in a microwave oven for 5 min and the powder calcined at 1000°C for 2 h. Active γ-Α12θ3 powder was obtained. The XRD pattern and SEM images of the A1203 product are shown in Figs.8 and 9. The purity of the A1203 product was as follows: A1203 99.36%, MgO 0.51%, CaO 0.06%, Fe203 0.04%, and Si0 2 0.03%. The result indicated that further investigation of strategies for Mg removal is necessary to achieve higher purity A1203.
Dissolution time//)
Fig.6 Effect of dissolution time on alumina extraction efficiency Temperature was the most important factor affecting alumina dissolution from the sintered mixture. From the results presented in Fig.7, it can be seen that when the temperature is lower than
199
3. H.N. Yoshimura and A.P. Abreu, "Evaluation of aluminum dross waste as raw material for refractories," Ceramics International, 34(3)(2008), 581-591. 100 [■
4. El-Katatny, E.A., Halany, and S.A., Mohamed, "Surface composition, charge and texture of active alumina powders recovered from aluminum dross tailings chemical waste," Powder Technol, 132, 2003 , 137-144. 5. LIU Qingsheng, "Preliminary research on preparation Al-Si-Ti alloy with aluminum dross as electrolysis materials". TMS,EPD congress, 2010, 639-644. 6. B. Dash and B.R. Das, "Acid dissolution of alumina from waste aluminium dross," Hydrometallurgy, 92(2008), 48-53.
10
20
30
40
50
60
70
80
90
7 B. R. Das, et al., "Production of alumina from waste aluminium dross," Minerals Engineering, 20(2007), 252-258.
2Theta/deg
8 Amer, A.M., 2002, "Extracting aluminum from dross tailings," /. Metals, 54(2002), 72-75.
Fig.8 XRD pattern of A1 2 0 3 product
9. MIKITO UEDA and SHIRO TSUKAMOTO, "Recovery of aluminum from oxide particles in aluminum dross using AlF-NaFBaCl2 molten salt," Journal of Applied Electrochemistry, 35(2005), 925-930. 10. XIAO Y. Y AN, "Chemical and electrochemical processing of aluminum dross using molten salts," METALLURGICAL AND MATERIALS TRANSACTIONS B, 348, 2008, 348-363. 11. MATJIE R H, et al., "Extraction of alumina from coal fly ash generated from a selected low rank bituminous South African coal," Min Eng, 18(2005), 299-310. 12. AKIRA O, et al., "Leaching of various metals from coal into aqueous solutions containing an acid or a chelating agent," Fuel Process Technol, 85(2004), 1089-1 102.
Fig.9 SEM images of α-Α12θ3 Conclusions
13. IYER R S and SCOTT J A, " Power station fly ash—a review of value-added utilization outside of the construction industry," Resour Conserv Recycl, 31(2001), 217-288.
(1) Al dross can be used can be used as a bauxite substitute in the soda sintering process for alumina production. Optimum conditions for a desilicated ash sintering were 1173K and 60 min. (2) Optimum conditions for sintered for Al dross dissolution were an H 2 S0 4 concentration of 0.9mol/l at 358Κ for 1.5h. (3) The AI2O3 purity was improved by means of EDTA complexation of other impurity metals and washing with distilled water. Finally, High purity (Al 2 0 3 >99%) active α-Α1203 was obtained. References 1. O. Manfredi and W. Wuth, "Characterizing the physical and chemical properties of aluminum dross," JOM, 49(1997), 48-51. 2. F.A. Lopez et al., "The recovery of alumina from salt slags in aluminium remelting," Canadian Metallurgical Quarterly, 33(1994), 29-33.
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Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
Effect of Calcium/Aluminium ratio on MgO containing calcium aluminate slags 1
Wang Bo1, Sun Hui-lan1, Guo Dong1, Bi Shi-wen2 Hebei University of Science and Technology; 70 Yuhua East Rd; Shijiazhuang, Hebei, 050018, China; 2 Northeastern University; No.l 1, Lane 3, Wenhua Road; Shenyang, Liaoning, 110004, China Keywords: calcium aluminate slag, CaO ratio, MgO, C20A43M3S3, alumina leaching 12CaO»7Al203 and to inhibit the formation of C2oA13M3S3 [1315].
Abstract MgO is the main impurity in calcium aluminate slags. The presence of MgO will change the nature of the alumina bearing phases and decrease the alumina leaching properties of the slag. In order to remove or reduce the negative effect of MgO, the method of changing the C/A (CaO/Al203, molar ratio, excluding CaO in 2CaOSi02) of the calcium aluminate slag was studied and the effect on leaching mechanism was also analyzed. The results showed that the formation of the quaternary compound 20CaO13Al2O3-3MgO-3SiO2 (C2oA13M3S3) is inhibited with increasing C/A in the calcium aluminate slag, and the MgO crystallized in the form of periclase under these conditions. There is an optimal C/A to improve the alumina leachability of the calcium aluminate slag. The increase of C/A could not remove the negative effect of MgO on leachability completely, and the optimal C/A of the slag increases with increasing MgO content.
Previous research demonstrates that the alteration of the CaO content could improve the alumina leaching properties of the slag. So the alteration of C/A (CaO/Al203, molar ratio, excluding the CaO of 2CaOSi02) of the calcium aluminate slag, to remove the negative effect of MgO, was studied and mechanism of this effect are discussed in this paper. Experimental 2.1 Materials CaC03, Na2C03, NaOH, Si0 2 and MgO used in the experimental studies were chemical pure reagents. Al(OH)3 used in the experimental was an industrial pure reagent. The A/S (Al203/Si02, mass ratio) of the calcium aluminate slag was 1.3.
Introduction
2.2 Smelting of calcium aluminate slag
Calcium aluminate slag is obtained from blast furnaces when smelting iron-bearing bauxite. The ideal components of calcium aluminate slags are 12CaO-7Al203 and y-2CaOSi02 [1, 2]. The slag can react with a sodium carbonate solution and its alumina leaching rate can reach 85%. This method realizes the comprehensive utilization of iron and alumina values in the ore [3-5].
Samples with different C/A were smelted in a MoSi2 resistance furnace, and the container was a graphite crucible. Another corundum crucible was used outside of the graphite crucible in order to prevent the break-out of the slag which could disintegrate and destroy the graphite crucible during cooling, thus contaminating the furnace. The smelting temperature was 1500°C and the holding time was one hour. The sample was taken out at 400 °C from the resistance furnace, and its cooling rate was 5°C/min.
The existence of impurity MgO which comes from iron-bearing bauxite and lime, will affect the alumina leaching rate of calcium aluminate slag during the industrial process [6-8].
2.3 Leaching of calcium aluminate slag
Eremin [9] studied the effect of MgO on calcium aluminate slag and pointed out that the compound 6CaO-4Al203MgOSi02 (CwAxMySz) would be formed in the slag when MgO was present. This compound does not react with the sodium carbonate solution. Wang Bo[10, 11] studied the mechanism of MgO effects on such slags systematically. The results showed that MgO will dissolve into the crystal lattice of 12CaO-7Al203 and forms limited C2oA13M3S3 when the MgO content is less than 1.0%. Above this level, a large amount of C2oA13M3S3 was formed which decreased the alumina leaching rate. The results also showed that this quaternary compound could react with sodium carbonate solution but its alumina leaching rate was very low [12].
The sodium aluminate solution obtained from leaching the slag was treating using the carbonization precipitation process, and the circulating mother liquid is then used to leach new calcium aluminate slag. The feasible conditions for alumina digestion are: leaching temperature 75 °C, leaching time 2h, L/S ratio 4.5, caustic alkali concentration 7g/L and sodium carbonate concentration 120 g/L. The leaching experiments are carried out in magnetically stirred constant temperature water bath. After leaching and dry filtration, the filtrate is used to analyze the composition of the solution, and the filter residue is washed and dried for analysis.
Therefore, it is clear MgO will affect the alumina leachability. The presence of MgO could not be avoided because it is the main impurity of the slag, so how to remove or reduce the negative effect of MgO became the focus of this research.
2.4 Analysis The contents of A1203 and Na20 in samples and filtrate were analyzed by a chemical method. Phase components of the calcium aluminate slag were identified by X-ray diffraction (PANalytical PW3040/60). A Malvern laser particle analyzer was used to analyze the particle size of the slag.
There are two methods to remove the negative effect of MgO from the phase components of the calcium aluminate slag. One is to make MgO form a more stable component which contained no, or little A1203. The other method is to stabilize the compound
201
Results and discussion 3»! Effect of C/A on self-disintegrating property of slag The self-disintegration of calcium aluminate slag can reduce the energy consumption during leaching and alumina production, and this is a very important characteristic of calcium aluminate slags. The percent content of granularity which is lower than 74um in samples is defined as the self-disintegrating rate. Slags with different MgO contents were cooled slowly and then well-mixed in order to analyze the granularity of the slag. The particle size results analyzed by Malvern 2000 are shown in Fig. 1. 100 w
98 S
96
-
^^^
■^^'* [■ f^-"-""
^„---'"^
ö0
%
92
a
^<^C—===·
^χ^
J^-"-""
-m- MgO 0% * Mg01% - A - MgO 3% ψ MgO 5%
\-
90 1
1.4
1
1.5
1
1.6
1
1.7
1
1
1
1.8
1.9
2.0
OA
Fig. 1 Granularity of slag with different C/A and MgO content The results of granularity showed that the self-disintegrating rate of calcium aluminate slag with different MgO content and C/A were good, and they were basically higher than 95%. The selfdisintegrating rate was lower than 95% only when C/A was very low. It can be seen that when the C/A of slag was fixed the selfdisintegrating rate decreased with the increase of MgO content, and also when the MgO content was fixed the self-disintegrating rate increased with the increase of C/A. The self-disintegration of slag was caused by the crystal transformation from ß-2CaOSi02 to y-2CaOSi02. With the increase of MgO content, self-disintegration of slag decreased because of the increase of C20A13M3S3 content. A certain amount of Si0 2 was consumed when the quaternary compound C20A13M3S3 was formed. Therefore, the content of 2CaOSi02 decreased in the slag and the self-disintegrating rate of the slag decreased. The content of C2oA13M3S3 decreased with the increase in C/A when the MgO content wasfixed,so the self-disintegrating properties were improved. But the content of Si0 2 in C2oA13M3S3 was low, so the decrease of 2CaOSi02 was also low and had little effect on self-disintegration rate of the slag. Therefore, the selfdisintegration rate of the slag was still sufficient to meet the demand of the alumina leaching process. 3.2 Effect of C/A on alumina leaching property of slag The alumina leaching experiment on the calcium aluminate slags was carried out in order to study the effect of C/A on the alumina leaching properties of MgO containing slags. The leaching conditions were as in section 2.3, and the results are shown in Fig.2.
Fig.2 Leaching rate of containing MgO slag with different C/A Fig.l shows that the alumina leaching rate of slag increased at first and then decreased with increase of C/A. The highest alumina leaching rate of a slag which did not contain MgO appeared when C/A was 1.7. The highest alumina leaching rate of slag whose MgO content was 1% appeared when C/A was 1.8. The higher alumina leaching rate of slag whose MgO content was 3% appeared when C/A was between 1.8 and 2.0, and then the leaching rate decreased smoothly. The highest alumina leaching rate of slag with MgO content of 5% appeared when C/A was 2.0, and a maximum rate did not appear under these conditions. Therefore, the alumina leaching rate of calcium aluminate slag decreased with increasing MgO content when C/A was 1.7. The increase of C/A could improve the leaching property of slag to some extent, and the optimal C/A of the slag also increased with increasing MgO content. In addition, the variation of C/A could not remove the negative effect completely, because the optimal alumina leaching rate decreased with the increase in MgO content. 3.3 Effect mechanism of C/A on MgO containing calcium aluminate slag The increase of CaO can reduce the negative effect of MgO on leachability of calcium aluminate slags, so the mechanism of effect of C/A on MgO containing slags was analyzed by XRD. The existence of the quaternary compound C20A13M3S3 was also investigated. The XRD spectra are shown in Fig.3~Fig.5. Fig.6 and Fig.7 showed the relative intensity of independent characteristic peaks of 12CaO«7Al203 and C2oA13M3S3 as a function of MgO content and C/A.
was 1.8. Crystalline MgO or its mineral forms, was not found in the slag under these conditions. - - 20CaQ13AI 203-3MgO3Si 02 90
C/A=1.8
-
»
·
■-
80 70
^
60
c "03
C/A=2.0
LOLAJI 10
20
30
50
- ■ - C/A=1.8 - · - C/A=2.0
40
40
50 60 70 80 2 Thet a Fig.3 XRD spectra of slag with MgO=l%
30
90
■
20 L_
,
1
1
_l
MgO content /%
Fig.6 Relative intensity of characteristic peaks of 12Ca07Al 2 0 3
2 CaO· Si 0 2 12CaO-7AI 2 0 3 20CaO13AI 203-3MgO-3Si 0 2
14
MgO
C/A=1.8
-·
12 A 10 •>9
1
8
d>
6 4
Fig.4 XRD spectra of slag with MgO=3%
0 - 2CaOSi 0 2
C/A=1.8
30
40
50
60
70
1
..1
_J
Fig.7 Relative intensity of characteristic peaks of Q0A13M3S3 withC/A=1.8
- 20CaO13AI 2 0 3 -3^ÖO3Si 0 2
20
1
MgO content/%
- 12CaO7AI 2 0 3
10
•
80
Fig.6 and Fig.7 show the relative intensity of characteristic peaks of 12Ca07Al 2 0 3 and C2oA13M3S3. Relative intensity is defined as a peak intensity ratio between the strongest independent characteristic peaks of 12Ca07Al 2 0 3 or C20A13M3S3 of slag and the strongest peak of slag. The main phase in the calcium aluminate slag whose MgO content was 3% and C/A was 1.8, were y-2CaO«Si02,12CaO»7Al203, and C2oA13M3S3. The intensity of the characteristic peaks of Q0A13M3S3 is strong showing that a large amount of C2oA13M3S3 was generated. A great deal of 12CaO»7Al203 transformed into C2oA13M3S3 under these conditions. This is why alumina leaching rate decreased. When the C/A of slag increased to 2.0, compound C2oA13M3S3 disappeared and 3CaO»Al203 was not found either. Meanwhile periclase was found in the slag. So the extra-addition of CaO decomposed the quaternary compound C2oA13M3S3 and that was the reason why the alumina leaching rate increased. The reaction between C20A13M3S3 and CaO is shown below:
90
2 Thet a
Fig.5 XRD spectra of slag with MgO=5% The compound C2oA13M3S3 which is difficult to leach was found in the slag whose C/A was 1.8, but it could not be found in the slag whose C/A was 2.0. The intensity of characteristic peaks of 12CaO7Al203 were clearly enhanced when the C/A of the slag increased. Thus some of the C2oA13M3S3 transformed into 12Ca07Al 203 which is easy to leach, with the increase of C/A of the slag. Some 3CaOAl 2 0 3 was also found in slag when the C/A was 2.0, and that was the reason why the alumina leaching rate of slag whose C/A was 2.0 was lower than that of a slag whose C/A
7(20CaO13Al2O3-3MgO-3SiO2)+58CaO-+ 13(12CaO»7Al203)+21(2CaOSi02)+21MgO
203
The analysis of a slag whose C/A MgO content was 5% was similar to that of a slag whose MgO content was 3%. The only difference between them was the intensity of the characteristic peaks of periclase (MgO).
Northeastern University: Natural Science, 29(11)(2008), 1593-1596. [7]
Therefore, the increase of C/A of calcium aluminate slag improved alumina leaching properties and decreased the negative effect of MgO on teachability of the slag, but it could not remove the negative effect of MgO completely. Thus the mechanism of this effect needs to be studied further.
TONG Zhi-fang, BI Shi-wen, YU Hai-yan and WU Yusheng, "Leaching kinetics of non-constant temperature process of calcium aluminate slag under microwave radiation," The Chinese Journal of Nonferrous Metals, 16(2)(2006), 357-362.
[8]
D.J.Connor, Aluminium extraction from non bauxitic materials, (Sydney: Aluminium-Verlag Gmbh, 1988), 230250.
[9]
N.I.Eremin, "Investgations on the complex processing of bauxites," (Bauxite-Alumina-Aluminum, Symposium of ICSOBA . Budapest: Research Institute for Non-Ferrous Metals, 1971), 329-335.
[10]
WANG Bo, YU Hai-yan, MIAO Yu, SUN Hui-lan, BI Shiwen and TU Gan-feng, "Effect of MgO on leaching and self-disintegrating property of calcium aluminate slag," Light Metals, (4)(2008), 11-13.
[11]
WANG Bo, YU Hai-yan, SUN Hui-lan, BI Shi-wen, TU Gan-feng and GENG Hong-juan. "Effect of MgO on crystal and leaching property of 12CaO-7Al203," Minning and Metallurgical Engineering, 28(05)(2008), 68-71.
[12]
WANG Bo, YU Hai-yan, SUN Hui-lan, BI Shi-wen and TU Gan-feng, "Synthesis and AI2O3 leaching property of 20CaO»13Al2O3eMgO«3SiO2," The Chinese Journal of Nonferrous Metals, 19(2)(2009), 378-382.
[ 13]
I.Kapralik and F.Hanic, "Studies of the system CaO-A1203MgO-Si02 in relation to the Quaternary phase Q," Transation of British ceram society, 1980, 128-133.
[14]
JIANG Feng-hua and XU De-long, "Influence of trace compositions on the formation of Q phase in high-alumina cement system," Journal of the Chinese Ceramic Society, 33(10)(2005), 1276-1279.
[15]
MENG Tao, YANG Li-qun and XU Xian-yu, "Studies of CaO-Al203-MgO-Si02 system in relation to the formation and hydration of phase Q," Bulletin of the Chinese Ceramic Society, (3)(1998), 21-34.
Conclusions 1) MgO content and C/A of calcium aluminate slags had little effect on self-disintegration rate of the slag. The selfdisintegrating rate of most samples was above 95%. 2) The increase of C/A could decrease the negative effect of MgO on leachability of the slag, but it could not remove the effect completely when the MgO content was less than 5%. The optimum C/A of the slag increased with the increase of MgO content. 3) The addition of CaO could decompose the quaternary compound C2oA13M3S3, and MgO crystallized in the form of periclase. Acknowledgements The authors greatly acknowledge the financial support of the National Nature Science Foundation of China (Project No: 50674028), and the foundation of Hebei University of Science and Technology (Project No: XL200921). The authors express their profound gratitude to the editors and reviewers of TMS. References [1]
L.K.Barr, "Alumina Production from Andalusite by the Pedersen Process," (Stockholm: Almqvist & wiksell international, 1977), 64-70.
[2]
J.Grzymek, A.Derdacka and Z.Konik, "Method for obtaining aluminum oxide," U.S Patent : 4149898 , 19782-21.
[3]
ZHANG Jing-dong, LI Yin-tai, BI Shi-wen and YANG Yihong, "Research on integrated utilization of high-ferrum bauxite in Guigang Guangxi," Light Metals, (8)(1992), 1618.
[4]
BI Shi-wen, YANG Yi-hong, LI Yin-tai, ZHANG Jingdong and DUAN Zhen-ying, "Study of alumina leaching from calcium aluminate slag," Light Metals, (6)(1992), 1015.
[5]
J.Grzymek, "Complex Production of aluminium oxide and Iron from laterite raw materials applying the calcium aluminates polymorphism," TMS Annual Meeting and Exhibition, Light Metals, 1985, 87-99.
[6]
WANG Bo, YU Hai-yan, SUN Hui-lan and BI Shi-wen, "Effect of material ratio on leaching and self-disintegrating property of calcium aluminate slag," Journal of
204
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
Study on extracting aluminum hydroxide from reduction slag of magnesium smelting by vacuum aluminothermic reduction Wang Yaowu, Feng Naixiang, You Jing, Hu Wenxin, Peng Jianping, Di Yuezhong, Wang Zhihui School of Metallurgy, Northeastern University, Shenyang, Liaoning 110004, P.R.China Key words: magnesium smelting; aluminum hydroxide; reduction slag; Ca02Al203; Abstract Magnesite + dolomite Aluminum powder
The reduction slag from magnesium smelting by vacuum aluminothermic reduction using dolomite and magnesite as raw materials consists mainly of A1203 and CaO. The alumina is more than 67% of the slag and is mainly in the form of CaO · 2A1203. The process of producing aluminum hydroxide from the slag by leaching with a mixture of sodium hydroxide and sodium carbonate was studied. This showed that the leaching rate of alumina is more than 86% when leaching temperature is 95°C, and L/S is 5, and leaching time is 2h. The magnesia-alumina spinel which generated in the reduction process and 3CaOAl203#6H20 which is generated in the leaching process, are the two main causes of loss of alumina. The chemical composition of aluminum hydroxide products obtained after carbonation precipitation can meet the quality standard of the alumina plant.
C02^
] Calcination] I Mixture \* |Briquetting| i A , I Vacuum reduction \—*> Magnesium Reduction slag_ Na2C03 NaOH
T
I Leaching I -
7
Decomposing Al(OH)3 Figure 1 The flowchart of the new magnesium production method
Introduntion A new magnesium production method has recently been introduced, which is based on two steps, the first of which is producing magnesium with vacuum aluminothermic reduction, and the second step is the extraction of aluminum hydroxide from the reduction slag. Figure 1 gives the flowchart of the new magnesium production method. The process of producing magnesium uses dolomite and magnesite as materials and aluminum powder as the reductant. After the reduction process, magnesium and a reduction slag can be obtained. The main phase of the reduction slag is Ca02Al 2 0 3 . The content of A1203 is more than 67% and the content of Si0 2 is below 1%. For a mass ratio of A1203 to Si0 2 is more than 60, the slag is suitable for leaching to produce aluminium hydroxide. This paper describes the study of extracting aluminum hydroxide from the reduction slag of this magnesium smelting process.
Experimental Materials The experimental materials is a reduction slag of magnesium smelting using dolomite and magnesite as materials and aluminum powder as reductant, the addition of materials and reductant follows the reaction equation (1). The magnesium smelting experiment was carried out at 1140°C, and the aluminum power excess is 5%. The chemical composition of the reduction slag is listed in Table 1 and Figure 2 shows the XRD pattern of the reduction slag. CaO+6MgO+4 Al = 6Mg + CaO · 2Al203
Table 1. The chemical composition of reduction slag (%) MgO
CaO
Si0 2
A1203
Al
Fe 2 0 3
6.13
20.86
0.89
67.54
3.50
1.08
205
(1)
£
4000 l··
40
2θ/(°)
Figure 2 The XRD spectra of reduction slag As shown in Table 1 and Figure 2, the main phase of the reduction slag is Ca02Al 2 0 3 , and there are also small amounts of CaOAl203, MgOAl203, MgO and Aluminum powder. The main chemical composition is A1203, CaO, unreacted MgO and Al.
The leaching experiments were carried out in a constant temperature water bath, and used an Erlenmeyer flask which is sealed with a condenser and stirred with a magnetic rotor. Before the experiments began, the leaching solution of mixed NaOH and Na2C03 was heated to a given temperature, then added into the reduction slag. After leaching for a given time, the leaching solution was filtered rapidly and washed three times by hot deionized water, and the alumina concentration of the filtrate was measured by chemical titration, and the alumina leaching rate of reduction slag was calculated.
Experimental principle Alumina in the reduction slag is mainly in the form of Ca02Al 2 0 3 , and it cannot be leached by a sodium hydroxide solution because in the leaching process, the CaO of the reduction slag would react with sodium hydroxide to form Ca(OH)2 which would then react with NaAl(OH)4 to form tricalcium hydroaluminate (3CaOAl203-6H20)[l,2] (equation 2). Because tricalcium hydroaluminate is stable in sodium hydroxide solution[3], a lot of A1203 entering into the leaching process is lost. The CaO of the reduction slag would react with sodium carbonate prior to sodium hydroxide to form CaC03 which has lower solubility and is more stable than Ca(OH)2 . The formation of CaC03 inhibits the formation of tricalcium hydroaluminate and reduces the loss of Al203[4], The leaching principle of CaO· 2A1203 and CaO A1203 are listed as follows[5, 6]:
The filtrate without the wash liquid was a sodium aluminate solution, and it was decomposed by carbonation in the water bath at 80 °C. The finish of decomposition was determined by the decomposed rate. When the decomposion reached 90%, it was stopped, filtered and aluminum hydroxide could be obtained. Results and discussions
In the reduction slag, the aluminium is present in two main species, one is the alumina which exists in CaO-2Al203 and CaOAl203, the other is elemental Al (unreacted aluminum 3Ca(OH)2 + 2NaAl(OH)4 = 3CaO · Al203 · 6H20 + 2NaOH (2powder). ) The elemental Al can enter into the leaching solution CaO · Al203 + Na2C03 + 2H20 = 2NaAl(OH)4 + CaC03 O) completely and alumina can not. To distinguish the two parts, the leaching rate was described as the leaching rate of CaO · 2Al203 + Na2C03 + 2NaOH + H20 = NaAl(OH)4 + CaC03alumina (4) alumina and the aluminum leaching rate was described the leaching rate of element aluminum. The conditions effecting the In the leaching process, elemental Al ( unreacted aluminum alumina leaching rate are mainly leaching temperature, leaching powder ) in the reduction slag can react with sodium hydroxide time, the content of sodium hydroxide and sodium carbonate. and also with sodium carbonate, but the reaction rate of elemental Al with sodium hydroxide is much faster than with sodium carbonate, so most of the elemental Al react with sodium hydroxide. Elemental Al is fully solution in the leaching sloution. 2AI + 2NaOH +6H20= 2NaAl(OH)4 + 3H2
Effect of leaching temperature on alumina leaching rate The results of effect of leaching temperature on alumina leaching rate are shown in Figure 3, and the experiments were carried out under the conditions of leaching L/S 5, leaching time
(5)
Experimental Method
206
2h, content of sodium hydroxide 100g/L, and content of sodium carbonate 120g/L.
temperature 95°C, leaching time 2h, content of sodium hydroxide 100g/L
-
90
|
80
•
■—'"'
M
2
^ ^
i_
·
·
*
I The alumina leaching rate I | — · — The aluminum leaching rate |
h
1 —"—The alumina leaching rate 1 1 — · — The aluminum leaching rate 1 60
65
70
75
80
85
90
100
95
The content of sodium carbonate/g/L
Temperature/°C
Figure 5 Effect of content of sodium carbonate on the alumina leaching rate
Figure 3 Effect of leaching temperature on the alumina leaching rate
It can be see from Figure 5 that the content of sodium carbonate have little effect on the alumina leaching rate when the content of sodium carbonate is between 80 and 130 g/L especially over 100 g/L. In the leaching process of the reduction slag, there are at least five reactions in the leaching solution, they are the formation of sodium aluminate (equations 3, 4 and 5) , the formation of tricalcium hydroaluminate (equation 2) and the decomposition of tricalcium hydroaluminate (equation 6). There are two main role of sodium carbonate, one is reaction with CaO to form CaC03, the other is decomposing tricalcium hydroaluminate| 7,8 j.
The effect of leaching temperature on the alumina leaching rate is very notable, the leaching rate of alumina in the reduction slag increases with increasing leaching temperature and levels off when the leaching rate reaches 90°C. Effect of leaching time on alumina leaching rate The results of effect of leaching time on alumina leaching rate are shown in Figure 4, and the experiments were carried out under the conditions of leaching L/S 5, leaching temperature 95°C, the content of sodium hydroxide 100g/L, and content of sodium carbonate 120g/L.
The content of sodium carbonate depends on two factors, one is that the content of sodium carbonate must exceed the theoretical amount to avoiding the forming of tricalcium hydroaluminate and the theoretical amount is about 80g/l when the L/S is 5, the other aspect is that since the decomposition of tricalcium hydroaluminate is a reversible reaction, thus the higher the content of sodium the better in order to decompose tricalcium hydroaluminate. According to the results shown in Fig.4, the optimal content is 100 g/L.
'■ •
\s 20
40
60
80
The alumina leaching rate The aluminum leaching rate
100
120
140
160
180
3CaO · Al203 · 6H20 + 3Na2C03 :
200
ì 2NaAl(OH)4 + ANaOH + 3CaC03
(6)
Leaching time/min
Effect of the content of sodium hydroxide on alumina leaching rate
Figure 4 Effect of leaching time on the alumina leaching rate As shown in Fig.4, the leaching rate of alumina in reduction slag increases with the extension of leaching time and levels off when the leaching time extends 2h.
The effect of the content of sodium hydroxide on alumina leaching rate is showed in Figure 6. The experiments were carried out under the conditions of leaching L/S 5, leaching temperature 95°C, leaching time 2h, content of sodium carbonate 100g/L.
Effect of the content of sodium carbonate on alumina leaching rate The results of effect of the content of sodium carbonate on alumina leaching rate are shown in Figure 5, and the experiments were carried out under the conditions of leaching L/S 5, leaching
207
As shown in Figure 6, the content of sodium hydroxide has great effect on the alumina leaching rate. The alumina leaching rate increases at first, and then decreases with the increase of sodium hydroxide concentration, and there is a maximum at a content of 75g/L. Phase analysis of Leaching slag
- The alumina leaching rate - The aluminum leaching rate [
60
65
70
75
80
85
90
95
100
The main chemical composition of the leaching slag is listed in table 2, and the XRD pattern of the leaching slag was analysed in Figure 7.
105
The content of sodium hydroxide/g/L
Figure 6 Effect of content of sodium hydroxide on the alumina leaching rate Table 2 The main chemical composition of the leaching slag A1203
MgO
CaC0 3
Si0 2
Fe 2 0 3
others
17.88
6.38
70.53
2.81
0.79
1.61
24000 r
Li
22000 [
A-CaC03 o - MgOAl203
20000 [ 18000 [
V-MgO 0 - 3CaOAl203.6H20 - 2CaOSi02
;
16000 [ 14000 [ 12000 1 [ 10000 [
0
8000 [
2000 [
ΔΔ
V
6000 [ 4000 [
Δ
CO
Δ
°0
of 1
-2000
i
1
2θ/(°)
Figure 7 The XRD of the leaching slag As shown in Figure 7 and Table 2, the main phase of the leaching slag is calcium carbonate, and there are some other phases such as magnesia-alumina spinel (MgOAl203), tricalcium hydroaluminate (3CaOAl203-6H20), and calcium silicate (2CaOSi02). It is also found that the Ca02Al 2 0 3 and CaOAl203 are decomposed completely.
In the sodium aluminate solution which result from leaching of the reduction slag, the alumina concentration is 120-150g/L, and Si0 2 concentration is 0.2-0.3 g/L, and the ratio of A1203 and Si0 2 is more than 500. Since the A/S is high, the sodium aluminate solution can be decomposed by carbonation without desilication process. The composition of the aluminum hydroxide obtained is shown in Table 3.
Table 3 The composition of aluminum hydroxide[9] A1 2 0 3 Fe2Os Si0 2 0.02 AH-1 64.5 0.02 the national AH-2 0.04 0.04 standards of 64.0 china 0.06 AH-3 63.5 0.06 Al(OH)3 of this project 64.3 0.03 0.01
The chemical composition of the aluminum hydroxide products used reduction slag of magnesium smelting as materials can meet the quality standard of the alumina plant. In addition, the
Na 2 0 0.4 0.5 0.6 0.43
sodium aluminate solution is suitable to producing high added value non-metallurgical alumina.
208
Conclusions (1) The main phase of the reduction slag from magnesium smelting by vacuum aluminothermic reduction is Ca02Al 2 0 3 , and the Ca02Al 2 0 3 can be decomposed completely by leaching with alkali solution to form CaC03 and sodium aluminate solution. (2) The leaching rate of alumina from the slag is more than 86% and the leaching rate of element aluminum is more than 90% when the reduction slag leached at 95 °C for 2h. The magnesia-alumina spinel which is generated in the reduction process and 3CaOAl203»6H20 which is generated in the leaching process were the two main pathways which caused loss of alumina. (3) Aluminum hydroxide conforming to the national standards of China could be produced after carbonation precipitation. References 1. M. Gawlicki, W. Nocu-Wczelik, M. Pyzalski. Studies on the hvdration of calcium aluminates, Journal of Thermal Analysis. 1984, 29 :1005-1008. 2. Xiao Wei , Liu Wei, Zhang Nianbing. Leaching kinetics of calcium aluminate sinter. Light Metals.2009, (8):23-26. 3. Yuan Jiongliang, Zhang Yi. Study on desilication process of sodium aluminate solution by adding tricalcium hvdroaluminate. Nonferrous metals. 2003, 55 (l):60-65. 4. C. Ostrowski, J. Zelazny. Solid solutions of calcium aluminates C^A, C12A7 and CA with sodium oxide. Journal of thermal analysis and calorimetry. 2004, 75: 867-885. 5. He Runde. The extracting of alumina from powdered calcium aluminate. Light metals. 1997, (6):20-25 6 Fan Juan , Ruan Fuchang. Review on the extracting methods of aluminum from powdered calcium aluminate. Applied chemical industry. 2002, 31(1): 1-5. 7. Liu Guihua, Li Xiaobin , Peng Zhihong. Reaction behavior between calcium oxide or calcium hydroxide and aluminate solution with heavy caustic soda. The Chinese journal of nonferrous metals. 2000, 10(2):266-269. 8. Qiu Zhenzhuo. Extracting calcium aluminate by sodium hydroxide. Light Metals.1999, (4):18-22 9. The national standards of china. Aluminum hydroxide. GB/T 4294-1997.
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
APPLICATION OF THERMOGRAVIMETRIC ANALYSIS FOR ESTIMATION OF TRI-HYDRATE ALUMINA IN CENTRAL INDIAN BAUXITES - AN ALTERNATIVE FOR CLASSICAL TECHNIQUES Y. V. Ramana and Rajesh C. Patnaik Vedanta Aluminium Limited, Lanjigarh, Kalahandi, Orissa-766 027, India Keywords: Bauxite, tri-hydrate alumina, gibbsite, thermogravimetry. Trihydrate alumina content in bauxite is generally estimated by a bomb digestion system in the plants of the alumina industry using the Bayer process. Though this analytical method provides the required levels of accuracy, the large number of steps involved in the process, starting from digestion, filtration and titration to complete one analysis, becomes tedious and time consuming. If there are a large number of samples to be processed within a stipulated time to meet the process requirement, completion of the targeted number of samples in time is a challenge. This necessitates looking into quick alternative methods with equal levels of accuracy and dependability.
Abstract Tri-hydrate alumina or gibbsite (A1203.3H20) content in bauxite is a fundamental quality parameter in the Bayer alumina process using low temperature digestion. Classical techniques available for its estimation are time consuming and prone to standard and non-standard sources of error. Extensive studies on analysis of samples from Central Indian bauxite sources using thermogravimetry over varied temperature ranges and comparison with that of the data obtained from classical techniques, have revealed that loss of molecular water at different temperatures in thermo-gravimetry provides a meaningful tool to correlate with tri-hydrate alumina content by applying relevant correction factors derived from the experimental data at different concentrations of tri-hydrate alumina. The studies also have found that the thermogravimetric analysis can be used as a very fast and dependable technique with higher levels of accuracy over classical methods and is free from other interferences. The accuracy levels of the method developed were checked using reference standards.
Since the above conditions are always prevalent at Vedanta, alternative instrumental techniques which are very fast, with competitive levels of accuracy to wet chemical analysis, particularly thermal decomposition techniques, have been tried. This is a reasonable approach as the thermal techniques to study the dehydration and dehydroxylation behaviours of bauxites have been widely documented [1]. It is well known that thermal decomposition of aluminium minerals in bauxite is a factor of time and temperature. Dehydroxylation of gibbsite occurs in the temperature range of 260-380 °C with peak losses at 320-340 °C.
Introduction Central Indian bauxite sources contain three principal hydrates of aluminium, i.e., gibbsite or trihydrate alumina (A1203.3H20 or Al(OH)3)), boehmite and diaspore or monohydrate alumina (A1203.H20 or AlOOOH) with gibbsite often being the predominant mineral. The other major minerals are aluminogoethite, hematite, kaolinite, anatase, rutile and quartz. Alumina extraction using low temperature digestion in the Bayer process results in dissolution of the gibbsitic form of aluminium hydroxide where other iron and titania minerals remain undigested. Kaolinite (Al4[Si4Oi0](OH)8) shows significant dissolution in sodium aluminate liquors, subsequently forming a desilication product, whereas the effect on quartz is less significant. All the minerals present in the bauxite, and their chemical-mineralogical characteristics, will have some influence either on process control, or on the methods of estimation of primary pararmeters.
The established general decomposition pattern of gibbsite, A1203.3H20 is as below: A1203. 3H20—>A1203· x H 2 0 (as residual hydroxide) + y H2Oî Using thermogravimetry, this property of losing hydroxyls as water molecules at elevated temperatures from the gibbsite mineral present in the bauxite, and converting the weight loss into an equivalent alumina content, is accomplished. The values obtained are then compared and equated with the data previously generated using classical techniques of analysis for the same samples. From the proportions of released and residual hydroxides lost as water molecules, a relationship is established at a particular temperature by comparing the thermal decomposition data with wet chemical data and by applying a correction factor if required.
Determination of the trihydrate alumina content in feed bauxite is of paramount importance, as the total process control is directly or indirectly affected by its proportion in the bauxite and also the accuracy levels in its estimation. Vedanta Aluminium Ltd., at Lanjigarh is presently using varied sources of bauxites for production of alumina and a Central Indian source is one of them. Determination of the chemical and mineralogical composition of these sources with acceptable limits of accuracy, and with the necessary speed, is of pivotal importance to facilitate required blending and homogenising so as to feed a uniform quality of bauxite to the plant.
Materials and Methods 1. Bauxite sample collection and preparation Bauxite samples were collected from the mines of M/s. CR Mittal from Central Indian region. A total of 29 samples have been collected and prepared using Indian Standard [2]. Apart from the mine samples, a total of 10 international reference standards of bauxite have been used for comparison.
211
a combination of released and unreleased (bound hydroxyls) water molecules from the bauxite. The instrumental settings followed for the final analysis of bauxite samples are detailed in Table 1.
2. Chemical analysis of bauxite LOI of the sample is determined by subjecting 1 g of sample to 1000 °C for one hour and from the weight difference between before and after heating [3]. The major elements; alumina, silica, ferric oxide and titania in bauxite samples are also determined using the respective Indian standard test methods of [4-7]. Reactive silica in bauxite is obtained by deducting the percentage of nonreactive silica from the percentage of total silica [8].
Since the percent loss of mass corresponds to the water molecules associated with trihydrate alumina in bauxite, the weight of trihydrate as alumina is calculated as below: ,, .
3. Trihydrate alumina determination using wet chemical techniques Using an M/s. Intronics bomb digester, a known quantity (1.3 + 0.0001g and 0.65 + 0.0001g) of bauxite are heated under pressure, in a solution of sodium hydroxide (165g/l as Na2C03) respectively at 145 °C and 240 °C for gibbsite and boehmite; the insoluble matter is separated by filtration. Alumina present in the filtrate is given by titrating the equivalent of the hydroxide ions (OH) released by fluorine in a chelating solution containing sodium gluconate [9]. Mass percentage of A1203 is calculated as gibbsite for the solution at 145 °C and as gibbsite and boehmite combined at 240 °C. The difference in the values between the two determinations is expressed as boehmite.
Covers
Unit
Moisture
355 °C LOI
-
No
No
Start Temp
°C
25
107
End Temp
°C
107
355
Ramp Rate
°C/Min.
10
20
Ramp Time
Min.
00:08
00:12
Hold Time
Min.
00:20
00:20
Total Time
Min.
00:28
00:32
Max Time
Min.
00:00
00:00
-
None
None
Flow Rate
ml/Min.
Off
Off
Comparator
%
0.1
0.1
Final weight
-
at constancy
at constancy
Atmosphere
(kmàMass-MoisteeMassi* 100 initial Mass
Gioiste Mass-355 °C I d Mass)»100 — Moisture Mass
mxuÂ%*
ö
MoistufeMas£-J55^LOIMassFl^i..^* litt Moisture Mass
Molecular ratio of A1203 : H 2 0 Factor for residual hydroxidefractionof 3H20
4. Trihydrate alumina determination using thermogravimetry A thermogravimetric analyser, model TGA-701 of M/s. Leco Corporation is used for the determination of loss of mass in bauxite. Step Name
„
Released fraction of 3H 2 0 Unreleased fraction of 3H20 %of unreleased 3H20 %of released 3 H20 Molecular mass of released 3 H20 Factor for unreleased | 3H 2 0
= 1.889 = 1.06
2.89
2.87
2.85
2.83
2.81
2.79
2.77
0.11
0.13
0.15
0.17
0.19
0.21
0.23
3.67
4.33
5.0
a
6.33
7
7.67
96.3
95.7
95
943
93.7
93
92.3
52.0
51.7
51.3
50.6
50.2
49.9
1.04
1.05
1.05
1.07
1.08
1.08
■ ■
Table 2. Basis of gibbsite thermal dehydroxylation The initial step in the thermal decomposition of gibbsite is the diffusion of protons and the reaction with hydroxyl ions to form water. From the comparative analysis of thermal decomposition data with wet chemical data of trihydrate, it is observed that there is a certain fraction of unreleased water molecules characteristic of either the source of bauxite or the concentration levels of trihydrate present in the bauxite. The factor for the unreleased fraction of water molecule is calculated with different combinations of released and unreleased proportions of three water molecules available with each alumina molecule for completion of the total quantity of water. Out of several combinations, as displayed in Table 2, a fraction of 0.17
Table 1. Test conditions for bauxite observed during thermogravimetric analysis A number of samples are analyzed using different instrumental conditions in the range from 100 to 420 °C to arrive at the most common and nearest temperature where the wet chemical values are comparable with the scanned data derived by applying the necessary correction factors. These are calculated on the basis of
212
method is free from such interferences. Other variables like heating rate, external pressure, water vapour pressure, sample particle size etc., do not have a significant influence on these determinations [10].
unreleased molecules and its corresponding factor (1.06), is considered to be most appropriate for the calculation of actual trihydrate content, when compared with the values of trihydrate determined by classical techniques, for the Central Indian bauxite sources.
is.*». U,fe
Results and discussion
!
The data obtained by conducting wet chemical analysis of the major elements and trihydrate alumina content in 29 samples collected from a bauxite source of Central India (M/s. CR Mittal) has been furnished in Table 4. Data on trihydrate alumina obtained from thermogravimetric analysis and boehmitic monohydrate data obtained from wet chemical techniques using a 240 °C bomb digestion of the same samples, is also furnished for comparison. The data on the other possible interfering parameters in wet chemical analysis, like reactive silica and boehmitic alumina, are also furnished in the table to study their effect Interpretation of the above data reveals several interesting factors related to the thermogravimetric analysis and its application to the determination of trihydrate content in bauxites. Data on the wet chemical analysis of trihydrate alumina content is furnished in column 10 (ranging from 24.10 to 48.27%) and the respective values obtained by using thermogravimetric analysis are furnished in column 12 (ranging from 24.18 to 47.22%) in Table-3. Out of the total 29 samples analysed, only four samples are highlighted in the Table- 3 as showing a difference of > 1% THA as under column 13, wherever the actual THA content is more than or equal to 45 % in the bauxite. It is interesting to note that except the above four samples, all other samples have a THA of < 40% and variation is < 1%. The data also indicates that as the THA content is increasing from 40% and above, the variation between the two methods is showing an increasing trend indicating that the factorial value needs to be corrected with increasing THA content. It is also clear from the table that the variation is very insignificant if THA value is < 35% as reflected in majority of the samples studied. These observations reveal that the thermogravimetric data obtained for bauxites having < 35% trihydrate alumina for the source studied can be a perfect match with the wet chemical data and it can be used as a direct alternative technique. Similarly, up to 45% THA content, the thermogravimetric method can be accepted with 1% variation. Above 45%, it is necessary to make corrections to the factor derived from the loss of mass.
Samp!« Legation
m
TiOì
SiOì (H)
L.OÎ
1
2
3
4
5
6
7
8
Î
61 SI
Pit-AS!
31.04
4SI
3.60
3S.65
2
6132
Pa-E/S2
9.S4
8.33
0.36
56.35
3
61 S3
Pfc-D'Sl
16.66
7.60
LÔ4
4
6184
Pit-OSS
16.76
5.48
2.90
R-SiOj VtttTBA
m
TGA-THÄ at3S5*C
»iff. Eeeiumt« (M-12)
m
u 1
9
1β
11
12
13
20*1
3.26
33.47
31.72
3363
0.16
1.04
23.83
0.7:9
35.10
32.38
34.35
0.25
19.46
48.04
25.79
0.93
45.07
41.S4
44.35
0.72
100
49.15
24.85
2.56
41.30
38.37
40.67
0.63
4.77
5
6185
Pft-OS2
6.13
6.42
0.Î1
57.72
2S.57
0.00
48.27
44.55
47.22
LOS
90S
6
6186
PÜ-A/S2
25.72
7.48
2.39
4110
22.50
2.10
37.90
35.63
37.74
0.16
0.00
7
6187
PÎI-C/S5
9 SS
6.37
1.15
5512
26.63
102
44.74
41.09
43.56
LIS
8.99
8
618S
PÎÏ-D/S3
22.16
5.3S
1.24
4537
24.73
1.12
4133
39.26
41.61
0.72
OSO
9
6189
Pit-C'Sl
24.82
5.54
0.3S
43 67
24.09
0.33
41.39
3S
Î>5
40.57
0.52
0.60
10
6190
P&-D/S2
13.1$
6.74
1.27
51.30
26.76
1.12
46.78
43.03
45.61
1.17
2.81
1!
6191
Pà-E-S2
14.97
6.Ö9
0.59
51.31
25.39
0.79
41.73
3S.63
40.95
0.78
8.55
12
6192
Pà-E-S3
19.84
5.S7
1.22
4S.23
24.08
1.09
39.93
37.4S
39.70
0.23
6.20
13
6193
Pìt-C:·^
19.04
6.60
no
49.39
23.25
0.91
37.58
35.34
37.46
0.12
9.94
14
6194
Pfc-IKÔÎ
9.5Î
6.42
063
60.53
21.34
0.56
25.73
24.07
25.51
0.22
33 52
15
6195
H t - H M Ö 20.32
5.56
4.18
49.S0
19.33
3.73
24.10
22.8Î
24.18
-0.ÖS
21.24
5.57
O.il
27.il
16
6196
»-1ΚΌ3
16.30
1?
6197
PÌHK/04
15.20
5.86
Î.93
56.33
18
6198
Pit-lK/OS
10.67
6.06
0.91
60,25
19
9134
Pit-04
40.34
3.98
5.57
31.70
20
9267
Pit-C
20.5?
6.95
1.20
21
9268
Pà-A/SS
14.92
7.84
22
9269
Pfi-E-Si
16.04
23
9270
Pìt-C'S2
24
9271
Pit-ES2
25
9272
Ptt-AS3
32.85
5.71
LSI
36.99
26
9273
PÌÌ-AS4
16.97
6.14
0.54
4S.32
27
9274
Prt-AS2
32.02
4.45
2.18
3S.50
23
9275
Pit-ASl
12.42
7.Î4
1.05
50.12
29
9276
Pìt-A'Sé
25.96
4.32
9.22
39.22
2.05
24.60
23.10
24.49
19.84
1.72
25 J 3
24.17
25.62
-0.39
2S.76
21.27
0.S2
27.10
25,73
27.27
-0.17
31.87
17.29
3.59
26.79
25.41
26.93
-0.14
0.00
4649
23.3S
1 07
38.79
36.78
3S.99
-0.20
5.10
1.12
49.13
25.53
1.00
42,44
40.69
43.13
-0.69
2.97
7.40
Î.SS
5145
21.95
1.42
32.71
31.05
32.94
-0.23
16.32
7.30
6.37
0.84
56.02
23.31
0.75
47.62
43.80
46.43
1.19
4.67
15.53
7.09
1.43
4S.95
22,77
1.2S
35.27
33.42
35.43
-0.16
10.54
20.68
1.3S
34.32
32.53
34.48
-0.16
O.OC
26.20
0.49
43.92
41.96
44.48
-0.56
0.00
21.22
1.96
34.92
33.19
35.18
-0.26
O.OO
27.54
0.93
46.69
44.60
47.28
-0.59
0.00
19.82
7.64
29.21
27.68
29.34
-0.13
» ■ * >
2.30
54.SÎ
20.09
Table 3. Data on chemical composition and trihydrate alumina determined from wet chemical and thermogravimetric techniques on Central Indian bauxite samples.
The other advantages of thermogravimetric analysis over the classical techniques, is that the former facilitates multiple sample analysis i.e., up to 20 samples at a time resulting in higher throughput, whereas the later is confined to cumbersome time consuming individual analysis.
Verification with international reference samples To validate the above test data, a set of ten international reference bauxite samples with known trihydrate content have been analyzed under the same set of experimental conditions. The data of these samples with different combinations of released (2.89, 2.87, 2.85, 2.83, 2.81 and 2.79 H20) and respective unreleased (0.11, 0.13, 0.15, 0.17, 0.19, 0.21 and 0.23 H20) water have been tabulated in the Table-4. The data indicates that out of the ten, four standards namely; BXT 05, 06, 07 and 09 show a direct match with the reference value within a 1% limit of variation. By applying a factor of 1.06 derived from the release of 2.83 H 2 0 molecules, as established in the above experimental case of
In the normal course, when wet chemical techniques are used for this analysis, parameters like boehmite and reactive silica effect the determination of trihydrate alumina. To verify this effect, the analysis of boehmite and reactive silica and other parameters were conducted and furnished in the table along with trihydrate data. Reactive silica ranged from 0.33 to 7.64% (column 9) and boehmite ranged from 0 to 33.52 (column 14) in the experimental samples. These parameters are not showing any significant influence on the respective values of trihydrate alumina determined by using thermogravimetry, confirming that the
213
Central Indian bauxite samples, five standards, i.e., BXT 02, 04, 08, 09 and 11 match with less than 1% variation. The other standards namely; BXT 01, and 03 show more than 1% variation between wet chemical and thermogravimetric determinations even after applying the factor. Since the mean value of THA is considered as the standard wet chemical reference and there is a chance of variation up to 1% in actual trihydrate value for an individual standard, these interpretations are considered as indicative but trends are close to the actual data observed for experimental samples. Sample ID
THA THA·,* value
BXT 01 54.3
BXT 02
m
BXT 03 51
BXT 04
415
BXT 05
414
BXT 06
45i
BXTÔ* 39.?
BXT 08
4L?
BXT 09 26.4
BXT 11 44.?
Rei
355 *C (Tri* eut factor)
TM
Va!« 54.3 +A05 46.5 +/-Ô.7 51.0 +/-04 415 +/-0.4 414 +/-0.6 45.5 +/-0.6 39.7 +/-0.6 41.7 +/-0.4 264 +/-1.0 44." +/-0.9
2.89 H : 0
187 H 2 0
2.85 H 2 0
183 H 2 0
181 H 2 0
The data obtained with some Central Indian bauxite sources showed very good agreement between the two methods. An excellent comparison is noticed up to 35% of trihydrate alumina in the bauxite and up to 45% it is comparable with acceptable limits of variation. Above 45%, the data needs appropriate correction based on the proportion of molecules of water released/unreleased. The test data obtained is verified by comparison with that of international reference bauxite samples. Similar behavior and trends were observed, indicating validity of the experimental data. The results also show scope for adapting the technique for other sources.
2.79 H 2 0
Dia. UM Dift 1.045 Dift 1.053 Dift 1.06 Dift 1.068 Dift 1.075 Dift
49.28 -5.02 51.15 -3.15 51.5
18 51.89 141 5124 -2.06 52.63 -1.67 5198 -1.32
4123 -3.27 44.8? •L63 45.18 IM 45.52
45.8« -5.14 47,6
Acknowledgement The authors are grateful to Dr. Mukesh Kumar, CEO and Mr. Bimalananda Senapati, Head of Refinery for their support. The authors also thankfully acknowledge the infrastructural help provided by the management of Vedanta Aluminium Ltd., Lanjigarh.
058 45.82 -Ô.6S 46.17 -0.33 46.47 -0,03
■3.4 4752 •3.08 43.29 -171 48.61 -2.39 48.98 -102 49.3
-L?
References 40.42 -108 4156 -0.54 4214 -016 42.56 0.06 4185
ÔJ5 43.17 0,6? 43.45 Ô.95
1. Lodding E., 1969. The gibbsite dehydroxylation fork. Schwenker RF and Gran, PD (eds.), Thermal Analysis, Volume-2. Inorganic materials and Physical Chemistry. New York: Academic Press pp. 1239-1250. 2. Indian Standard: Methods of sampling bauxite (First Revision), IS: 1999-1987 Page 1-15. 3. Indian Standard: Methods of chemical analysis of bauxite, Parti- Determination of loss on ignition (First Revision) IS: 2000 (Part-1) - 1985, Page 1-5 4. Indian Standard: Methods of chemical analysis of bauxite, Part 2- Determination of silica (First Revision) IS: 2000 (Part2)-1985, Page 1-6. 5. Indian Standard: Methods of chemical analysis of bauxite, Part 3- Determination of alumina (First Revision) IS: 2000 (Part-3) - 1985, Page 1-6. 6. Indian Standard: Methods of chemical analysis of bauxite, Part 4- Determination of ferric oxide (First Revision) IS: 2000 (Part-4)-1985, Page 1-6. 7. Indian Standard: Methods of chemical analysis of bauxite, Part 5- Determination of titania (First Revision) IS: 2000 (Part-5) - 1985, Page 1-5. 8. IBM Manual, 1979. Manual of procedures for chemical and instrumental analysis of ores, minerals and ore dressing products, Page 11-12, Issued by the Controller, Indian Bureau of Mines. 9. Ibrahima Sory Cisse and Jiwen Ge, 2010, Determination of bauxite's phases by the bomb digest method at Kamsar laboratory ISO 9002 (Guinea), Journal of American Science, 6 (6), 139-145, 2010. 10. Paulik F, Paulik J, Naumann R, Kohnke K and Petzold D, 1983. Mechanism and kinetics of the dehydration of hydrargillites, Part I & II. Thermochimica Acta, 64, 1-12
4199 0.59 44.62 121 4452 152 4517 187 45i? 3.1? 45.91 3il 46.21 3.81
44.9
M
40.36
ÙM 41.89 119 42.18 148 42i
46.61 1.11 4652 1.42 471«
39.08 -162 40.57 -1.13 40.84 -0.86 41.15
25.6?
4.n 26.65
4158 -112 44.2
L7I 47Ì9 109 47.95 245 4817 2.7?
18 42.78 3.08 43.1
3.4 43.39 3.69
455 41.42 -018 41.74 0.04 41Ö1 U l
125 26.83 0.43 27.03 0.63 2 ? l i &ß 27.42 1.02 27.6
11
-Oi 44.5 -01 44.84 0.14 45.13 (MS 45.48 0.78 45,77 1.07
Table 4. Thermogravimetric data of international reference Standard bauxite samples These observations show that the trends noticed with Central Indian bauxite samples are mostly applicable to the international reference standards and vice versa, hence confirming that the observations for Central Indian case can also be applied for other sources within the defined ranges of trihydrate concentration. The observations also indicate that wide variations and deviations in case of BXT 01 and 03 standards are most likely due to the very high concentrations of trihydrate present, i.e., > 50 % and they may need to have a different correction factor, depending on the release of water molecules at that concentration of trihydrate. A similar effect is also observed in case of Central Indian sources with more than 45% trihydrate. Conclusion Thermogravimetric analysis based on the loss of mass of water molecules is found to be a very fast and alternative technique for assessing trihydrate alumina in comparison to classical methods of analysis which are tedious and time consuming.
214
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
DETERMINATION OF OXALATE ION IN BAYER LIQUOR USING ELECTROCHEMICAL METHOD Seval Turhan1'2, Betül Usta1'2, Yücel Çahin1'2, Oktay Uysal2 Anadolu University, Faculty of Science, Department of Chemistry, 26470 Eski§ehir; Turkey 2 ENTEKNO Industrial, Technological and Nano Materials Ltd., Anadolu University, Yunus Emre Campus, Anadolu Teknopark No: 104,26470 Eskiçehir; Turkey 1
Keywords: Bayer liquor; Oxalate; Pencil graphite electrode, Differential pulse voltammetry Abstract chromatography [7-10], high performance liquid Chromatograph [11-14], and gas chromatography [15-17]. In many factories, the oxalate was determined by titration with a standard potassium permanganate solution. This method often resulted in low precision and accuracy. Furthermore, it takes much time and many agents. Voltammetric techniques have not found use for this purpose. Recently, there has been a considerable effort in the development of voltammetric methods for the determination of some biological and industrial samples. It is generally believed that direct redox reactions of these species at the surface of usual electrodes are irreversible and therefore require high overpotential [18]. Moreover, the direct redox reaction of these species at the unmodified electrodes takes place at very similar potentials and often suffers from a pronounced fouling effect, which results in rather poor selectivity and reproducibility. In this study, we investigated the electrochemical determination of oxalate ion at pencil graphite electrode (PGE) by using differential puls voltammetry in Bayer liquor. A linear relationship between oxalate concentration and current response was obtained with good reproducibility of the current . Also real Bayer process samples were analysed by DPV and the results obtained compared with titration by potassium permanganate.
The Bayer process can be summarized as the digestion of bauxite with caustic liquor and the subsequent precipitation of hydrated alumina [1]. Most bauxite contains organic compounds in various amounts. Depending upon the digestion conditions, 510% of the organic carbon is converted to sodiumoxalate [2]. When sodiumoxalate, if not controlled in Bayer process, builds up to a certain level of supersaturation, it precipitates out in the hydrate precipitator tank. This co-precipitation affects the quality of alumina [3]. In this study, we investigated the electrochemical determination of oxalate ion by using differential puls voltammetry in Bayer liquor. A linear relationship between oxalate concentration and current response was obtained with good reproducibility of the current. Introduction The Bayer process can be summarised as the digestion of bauxite with caustic liquor and the subsequent precipitation of hydrated alumina [1]. Bauxite often contains organic matter in various amounts. Typically digestion of the bauxite during the Bayer process more man half of this organic carbon is extracted into the liquor. With recycling of the Bayer liquor the concentration of humic substances and their degration and oxidation products builds up an equilibrium concentration depending on the amount of organic carbon present in the bauxite. Because of this, in the Bayer process, the organic matter builds up to an equilibrium level, typically to 10-30 g/L of organic carbon content, determined by the outputs and the inputs [4]. Besides the bauxite input, some organic matter comes from other sources, such as process water, red mud flocculants or antifoams. The outputs are complex through the adsorption on the red mud or adsorption on the product. The accumulated organic matter and its breakdown products are known to cause numerous process problems. As a result, the liquors darken, and notable amounts of carbonate and oxalate are formed. Depending upon the digestion conditions, typically 5-10% of the organic carbon is converted to sodium oxalate [2,3]. Sodium oxalate has been shown to be very harmful with regard to alumina productivity and particle size [3, 5]. When sodium oxalate, if not controlled, attains a certain level of supersaturation, it precipitates out as fine needles in the hydrate precipitator tank. This coprecipitation affects the alumina product quality and productivity in many ways [3]. A variety of techniques have been utilized for the identification and determination of sodium oxalate and organic substances in Bayer liquors, these have included titrimetry [6], ion
Experimental Chemical and Reagents Sodium sulfate (99.0%, Aldrich), potassium permanganate (99.0%, Aldrich), and sodium oxalate (99.5%, Aldrich) reagents are commercially available as analytical grade and used without further purification. Stock solutions of oxalate solution were prepared daily from by using ultra-pure deinoized water. Bayer liquors and alumina hydrate were obtainaed from Seydiçehir Eti Aluminyum A.§. (Turkey). Apparatus Electrochemical experiments were carried out by a conventional three-electrode system. Pencil graphite (PG) was used as working electrode. Pt wire and saturated calomel electrode (SCE) were used as an auxiliary and reference electrode, respectively. Differential pulse voltammetry (DPV) measurements were performed by Autolab PGSTAT 100 Potentiostat/ Galvanostat with GPES 4.9 Version conversion software (EcoChemie, The Netherlands).
215
Electrochemical Detection of Oxalate Ion
Real Sample Analysis
The electrochemical detection of oxalate ion was performed in aqueous solution of 0.1 M Na2S04 using potential puls between 0.0 V and +1.5 V with a scan rate of 50 mVs"1 for three scans. All electroanalytical measurement were made at room temperature.
In order to verify the reliability of the electrode, it was applied for the determination of oxalate in Bayer liquors (pregnant liquor and spent liquor) and alumina hydrate samples. The oxalate concentrations in the samples were determined using the standard addition method and direct interpolation in the linear regression. Obtained results for oxalate ion concentrations from different Bayer liquors and alumina hydrate samples are given in Table I. Also oxalate in Bayer real samples determinated by using titration method. Obtained results are given Table 2
Result and Discussion Preparation of Calibration Curve for Oxalate Ion A series of DPVs was recorded at various concentrations of oxalate to determine its calibration curve. The response of the PG electrode to oxalate was found to increase with increasing oxalate concentration. Fig. 1 shows DPV curves recorded on the PG electrode in the presence of various oxalate concentrations in the range of 0.10-lOmM. Differential pulse voltammograms of oxalate had shown only a single peak (oxalate). The figure 2 shows the calibration curve obtained by measuring the DPV peak current intensity (in μΑ) vs. the concentration of oxalate (in mM) with a correlation coefficient of 0.9984.
0.250
0.500
0.750 E/V
1.000
1.250
Table I. Experimental Results of Oxalate Ion Samples [Oxalate] [Oxalate] with DPV with titration Pregnant liquor 33792 mg/L 48000 mg/L Spent liquor 21472 mg/L 26000 mg/L Hydrate alumina 0.122 mg/gsamDle 0.20mg/g samDle 1 Conclusion Electrochemistry has many advantages for making it an appealing choice for industrial oxalate ion analysis. Electrochemical method is a high specific and analitical method which is likely to get preference in sampling small volumes and complex sample matrices. The determination of sodium oxalate in several Bayer samples was undertaken and the results have been summarised in Table 1. As can be seen compared to results, titration method is higher than DPV method. Because of titration method often resulted in low precision and accuracy. In this study, PG electrodes was performed for the first time in different Bayer liquors and alumina hydrate. This method provide a simple and rapid technique to determine oxalate ion in high alkali media. Also, this method takes less time when compared with other methods. Further studies on the variations of electrode, electrolyte, linearity of response and quantitative properties of DPV method, etc. are underway. The developed method is straightforward and suitable for the determination of the oxalate ion in Bayer samples.
1.500
Fig. 1. .Differential pulse voltammograms of oxalate at various concentrations in 0.1 M Na2S04 at the PGE. Scan rate was 50mVs_1.
Acknowledgements The authors are grateful to Seydiçehir Eti Aluminyum A.C. (Turkey) for providing Bayer samples. References
10
1. T.G. Pearson, The Chemical Background of The Aluminium Industry (London, UK, Monograph No:5, Royal Institute of Chemistry, 1955), 137-143. 2. G. Lever, Travaux (1983), 13, 335 3. R. Calalo and T. Tran, "Effects of sodium oxalate on the precipitation of alumina trihydrate from synthetic sodium aluminate liquors " Light Metals, (1993), 125-133. 4. B. Gnyra and G. Lever, "Review of Bayer organics-oxalate control processes,"L/g/tf Metals, ( 1979), 151. 5. N Brown, and T.J Cole, "The behaviour of sodium oxalate in a bayer alumina plant," Light Metals, (1980), (15)105-117. 6 A.I. Vogel,, and G.H. Jeffery, Vogel's Textbook of Quantitative Inorganic Analysis (Harlow: Longman, 1986), 4th ed. 7. Grocott, S.C., Jefferies, L.P., Bowser, T., Carnevale, J., and Jackson, P.E., J. Chromatogr. A, 602 (1-2) (1992), 257.
12
[Oxalate]/mM Fig. 2. The analytical curve is linear in the range 0.10lO.OOmM of oxalate concentration with a correlation coefficient (tf2) of 0.9984.
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8. Jackson, P.E., J. Chromatogr. A, 693, (1) (1995) 155. 9. T Bowser,, and S.C, Grocott, Ion Chromatography in the Alumina and Aluminium Industries, Proc. 2nd Int. Alumina Quality Workshop, Perth, 1990. 10. Whelan, T.J., Shalliker, R.A., Mclntyre, C, and Wilson, M.A., Ind. Eng. Chem. Res.,, 44, (9), (2005) 3229. 11. Whelan, T.J., Kannangara, G.S.K., and Wilson, M.A., Ind. Eng. Chem. Res., 42 (26) (2003), 6673. 12. Whelan, T.J., Wilson, M.A., and Kannangara, K. (Proc. Aust. Organic Geochemistry Conference, 2002), 77. 13. Susie, M. and Armstrong, L.G., J. Chromatogr. A, 502 (1990), 443. 14. Xiao, J.B., Chen, X.Q., Jiang, X.Y., and Wu, S.D., Ann. Chim., 96 (5-6) (2006), 347. 15. Baker, A.R., Greenaway, A.M., and Ingram, C.W., "A Microwave Digestion-Based Determination of Low Molecular Weight Organic Acids in Bayer Process Liquor." Talanta, 42. (10) (1995), 1355. 16. Smeulders, D.E., Wilson, M.A., and Armstrong, L., "Insoluble organic compounds in the Bayer process,". Ind. Eng. Chem. Res., 40 (10) (2001), 2243-2251.. 17. Wilson, M.A., Ellis, A.V., Lee, G.S.H., Rose, H.R., Lu, X.Q., and Young, B.R., "Low-temperature products" Ind. Eng. Chem. Res., 38 (12) (1999), 4663. 18. R.N. Adams, "Probing brain chemistry with electroanalytical techniques,"^««/. Chem. 48 (14) (1976) 1126A-1138A.
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
Light Metals 2011 ALUMINA and BAUXITE
Alternative Alumina Sources Poster Session SESSION CHAIR
James Metson University of Auckland Auckland, New Zealand
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
The effect of ultrasonic treatment on alumina leaching from calcium aluminate slag Sun Hui-lan1, Wang Bo1, Guo Dong1, Zhang Xue-zheng1, Bi Shi-wen2 ^ebei University of Science and Technology; 70 Yuhua East Rd; Shijiazhuang, Hebei, 050018, China; 2 Northeastern University; No. 11, Lane 3, Wenhua Road; Shenyang, Liaoning, 110004, China Keywords: calcium aluminate slag, alumina, leaching, ultrasonic 2.1. Materials and apparatus
Abstract
The material used in the leaching process is a calcium aluminate slag which is obtained by treating iron-bearing bauxite with a blast furnace. Table 1 shows the chemical composition of the slag, while Fig.l shows the XRD results from the slag. The main phases in the slag are 12Ca07Al 2 0 3 and y-2CaOSi02. After grinding, the particle size of the slag is below 74um.
The effect of ultrasonic treatment on the alumina leaching properties of calcium aluminate slags was investigated. The effects of the main leaching parameters such as ultrasonic power, leaching time, leaching temperature, and the concentration of sodium carbonate on alumina leaching rate were determined. The mechanism of ultrasonic effects on leaching is also discussed. It was found that the increase in leaching rate was slight, but leaching conditions were improved significantly when ultrasonic treatment was used in the leaching process. Leaching temperature and sodium carbonate concentration decreased by 30°C and 60g#L_1 respectively with the addition of ultrasonic treatment. Ultrasonics would not promote the decomposition of γ2CaOSi02. The agglomeration of leaching residue was prevented by the cavitation and mechanical effect of ultrasonic treatment.
Element Content (Wt.%)
Table 1 Chemical analysis of slag A12Q3 SiQ2 CaO TiQ2 22.24
16.05
54.89
1.23
MgO 2.12
l-Y -2CaO SÌO2
Introduction
2
With the rapid development of the steel and aluminum industries, the shortage of suitable resources of iron ore and bauxite has attracted considerable attention [1]. Therefore, the utilization of low grade ores becomes imperative. Leaching alumina from resources which are low grade and refractory, such as ironbearing bauxite, red mud and fly ash, has become a research focus in alumina production [2-4]. Iron-bearing bauxite is treated by the process of "Blast furnace Sintering-". After smelting, a calcium aluminate slag is obtained, and then it is used to leach alumina [57].
2r-12CaO 7A120^
1,2
3-CaTi03 4-MgO
1,2
1 1
1,3 2,3
22
r^M wj 20
1 ill
40
34 2 60
2 80
2Theta
Fig.l XRD spectra of calcium aluminate slag
However, there are many problems in alumina leaching of calcium aluminate slag. The required concentration of sodium carbonate in the leaching solution is high, leaching time is long, and the alumina leaching rate is low [8-9]. Solving these problems effectively could improve the comprehensive utilization value of iron-bearing bauxite.
The ultrasonic cleaner (KQ-100) used in the experiments is produced by Kunshan ultrasonic instrument Ltd. The frequency of the ultrasonic wave is 40 kHz, and the power, which is adjustable, is between 40 and 100W. Stirring of the solution is achieved using a high speed mixer. The tri-mouth glass flask is fitted with a thermometer, stirrer and a reflux condenser. Scanning electron microscopy (SHIMADZU SSX-550) is used to observe the surface morphology of the slag.
Ultrasonic wave treatment has advantages in leaching processes of ores which are difficult to treat [10-12]. Both the leaching rate of refractory ores and the leaching conditions are improved with the cavitation effect of the ultrasonic wave [13, 14]. The application of ultrasonics in hydrometallurgy is concentrated in Heavy metal and precious metal metallurgy, but little research has been carried out in alumina production, especially in treating iron-bearing bauxite using ultrasonics.
2.2. Procedure The sodium aluminate solution obtained from the slag is treated with the carbonization precipitation process, and circulating mother liquid is used to leach new calcium aluminate slag. The conditions of the leaching solution for alumina digestion are listed below: caustic alkali concentration Nk=7g»L"1, a^l.6 (molecular ratio between Na20 and A1203), liquid-solid ratio L/S=5. Other conditions are variable in different experiments such as sodium carbonate concentration, leaching time, leaching temperature, and ultrasonic power.
Therefore, in order to find an effective method to improve the leaching rate and to decrease the concentration of sodium carbonate and leaching temperature enhancing leaching of calcium aluminate slag with ultrasonic treatment was studied. Experimental
221
The leaching solution is preheated to the required temperature and then poured into the tri-neck glass flask in which the weight of calcium aluminate slag is 10 g. The leaching process begins when the speed of mixer is up to 300r*min"1.
treatment could improve the leaching conditions, but could not increase the maximum leaching rate. In addition, the increase of the useless bubble and the formation of a sound barrier, were the reason why the leaching rate changed slightly when the ultrasonic power was higher under both conditions.
After leaching and dry filtration, the filtrate is used to analyze the composition concentration of solution, and the filter residue is washed and dried for analysis.
3.2 Effect of leaching time on leaching rate of slag
Results and discussion
Leaching rates with different leaching times were investigated under the following conditions: Nc=120g»L"\ T=75°C and 45°C. Ultrasonic power, fixed according to section 3.1 was 120W in the following experiments. The results are shown in Fig.3.
3.1 Effect of ultrasonic power on leaching rate of slag Variations in ultrasonic power will cause large changes in the sound intensity and the maximum radius of cavitation bubbles. These changes will affect the leaching process of the slag.
-
The effect of ultrasonic power on leaching rate of slag was investigated under the following conditions: 1), sodium carbonate concentration (Nc^^Og'L"1, leaching temperature (T)=75°C; 2), Nc= óOg'L"1, T=45°C. Leaching time (t) was 80min, and other conditions were as shown in section 2.2. The results are shown in Fig.2.
-
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-
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/
/
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t t t t
h ul t r asoni e ho ut ul t rasoni c h ul t rasoni c hout ul t rasoni c
1
1
1
60
80
100
Leaching time /min
The leaching rate of slag with ultrasonic and without ultrasonic treatment increased slightly with increasing leaching time when the leaching temperature was 75°C. That was because the reaction speed was fast at this temperature.
/ i
1
40
wi wi wi wi
Fig. 3 Effect of leaching time on leaching rate
-*-75°C, 120g.L_1
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< 72
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^^*^^^^
i
1
1
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80
120
160
200
With ultrasonic assistance, alumina leaching rate was improved visibly with increasing leaching time when leaching temperature was 45 °C. Extended leaching time had little effect on leaching rate. That was because the temperature and the initial leaching rate were low, so the cavitation effect was significant when time was less than 80min. When leaching time was over 80min most of the alumina was leached from the slag, so leaching rate increased slightly. The optimal leaching time was 80min, under this optimal condition the leaching rate with ultrasonic was close to the leaching rate which was obtained with or without the existence of ultrasonic when the temperature was 75 °C.
Ultrasonic power /w
Fig. 2 Effect of ultrasonic power on alumina leaching rate The alumina leaching rate of the slag was improved slightly with ultrasonic treatment, compared with that of slag without ultrasonics, when ultrasonic power was below 180w under condition 1. But if ultrasonic power was increased to 200W, the leaching rate could be improved by 1.5%. That was because the viscosity of the solution decreased and the movement of grains became violent when the power was high. Therefore, the leaching reaction was easy to carry out. For ultrasonic treatment at high temperature, cavitation occurred readily, but the cavitation intensity decreased with increasing temperature. In other words, because the leaching rate was high without ultrasonic and the cavitation effect decreased at high temperature, the effect of ultrasonic on leaching rate was slight.
Therefore, although ultrasonic could not increase the maximum leaching rate when leaching temperature was 45 °C and leaching time was 80min, it has decreased the required leaching temperature by 30°C.
The leaching rate increased significantly with increasing ultrasonic power under condition 2. The leaching rate changed little when ultrasonic power was up to 80W. The alumina leaching rate was improved by 5% with ultrasonic treatment. The optimal leaching rate obtained under condition 1 was close to that obtained under condition 2. This suggested that the ultrasonic
3.3 Effect of Nc on leaching rate of slag Nc(defined in section 3.1, passage 2) was high in the leaching process without ultrasonics, and this would increase the consumption of sodium carbonate. Experiments with different Nc were carried out under the following conditions: t=80min, L/S=5. The ultrasonic power was 120w. The results are shown in Fig. 4.
222
Leaching rate of slag without ultrasonic improved with increasing leaching temperature. Leaching rate with ultrasonic assistance was better than that without ultrasonics. When leaching temperature was 45°C and 75°C leaching rate of slag with ultrasonic increased by 10% and 4%. Effect of leaching temperature on leaching rate of slag with ultrasonic was not obvious. This meant that the addition of ultrasonic treatment weakened the effect of leaching temperature on the leaching rate of the slag.
94 92 90 SS ""■ 88 Φ
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—■—75°C —·—75°C —A—45°C —▼—45°C
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Generally, the increase of temperature will decrease the viscosity and promote molecular motion and Mass Transfer of liquid systems. So leaching rate without ultrasonic increased obviously with the increase of leaching temperature. When there was ultrasonic treatment, the vibration and cavitation effect also accelerated the speed of molecular motion and mass transfer in the liquid system. And under this condition the increase of temperature could not promote the leaching process unceasingly, so the effect of temperature was weakened.
with u l t r a s o n i c without ultrasonic with u l t r a s o n i c without ultrasonic
_i_
_l_
_L_
_L_
_JL_
80
90
100
110
120
Sodium carbonate concentration /g.L"1
Fig 4 Effect of Nc on leaching rate Fig.4 shows that with the increase of Nc, alumina leaching rate increased. When leaching temperature was 75 °C and Nc increased from 60 g^L"1 to 120 g'L"1 the alumina leaching rate without ultrasonic increased by 3.3%, but leaching rate with ultrasonic increased by 1.3%. The results indicated that Nc had little effect on leaching rate and the existence of ultrasonic weakened the slight effect when leaching temperature was high.
3,5 Effect of ultrasonic on stability of v^CaOSiO? in slag Although y-2CaOSi02 was stable, some of it in slag was still decomposed in sodium carbonate solution. The cavitation effect of ultrasonic treatment would cause instantaneous high local temperatures and pressures. Shock waves and micro jets would be generated which increase the reaction activity of solids [14]. Meanwhile, the shock wave could intensify the relative motion and collision probability of reactants. If the decomposition of γ2CaOSi0 2 was intensified with the existence of ultrasonic, the loss of alumina would increase, and these results were not expected [15]. Therefore, the content of silica in leaching solution was investigated under different ultrasonic power. Leaching time was 80min and L/S of slurry was 5. Other conditions and results are shown in Fig.6.
Nc had a great effect on alumina leaching rate when leaching temperature was 45 °C and ultrasonic was not used. Leaching rate increased obviously with the increase of Nc. But with the addition of ultrasonics, Nc had a little effect on leaching rate. The use of ultrasonic not only increases the leaching rate but also weakens the effect of Nc on leaching rate. The leaching rate of slag with ultrasonic under 45 °C and 60 g^L"1 was close to that of slag without ultrasonic under 75°C and 120 g·!/ 1 . In another word, under the circumstances that the extraction rate is guaranteed not to fall, the existence of ultrasonic could not only decrease leaching temperature by 30°C but also decrease Nc by 60 g#L"1.
1.4
3,4 Effect of leaching temperature on leaching rate of slag
1.2
The effect of leaching temperature on leaching rate of slag was studied when Nc was 60g#L"\ leaching time was 80min, and L/S was 5. The results were shown in Fig.5.
ί
1.0
φ 0.8 c o ° 0.6 O"
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-■-75°C,120g.U 1
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80
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160
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Fig.6 Effect of ultrasonic power on silica content
S 84
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<
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-
Silica content decreased with the increase of ultrasonic power when leaching temperature was 45 °C and Nc was 60geL"1. The silica content changed little when temperature was 75 °C and Nc was 120g»L"1. This suggests that ultrasonic treatment does not promote the decomposition of y-2CaOSi02. Silica content under the former condition was much lower than that under the latter condition. So the influencing factors of decomposition of γ2CaOSi0 2 were leaching temperature and sodium carbonate concentration Ultrasonics did not appear to affect the
—A—wi t h ul t r asoni e
-
—T—VA t hout
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Fig 5 Effect of leaching temperature on leaching rate
223
decomposition of y-2CaOSi02. On the contrary, the improved leaching conditions decreased the decomposition of y-2CaOSi02.
sodium carbonate concentration decreased from 120 g»L to 60 g-L"1. (2) Ultrasonic didn't promote the decomposition of y-2CaOSi02. On the contrary, the improved leaching condition decreased the decomposition of y-2CaO#Si02.
3.5 Effect mechanism of ultrasonic on alumina leaching property of slag
(3) The effect mechanism of ultrasonic was that ultrasonic enhanced the dispersion of solid particle and prevented the agglomeration of leaching residue.
In addition to the cavitation effect, a mechanical effect would be generated by ultrasonics. The shock wave whose speed could reach lOOOmms"1 damaged and sheared the surface of solid particles, this helps to sustain a high reaction activity. In order to study the mechanism of ultrasonic treatment, micro morphology of the leaching residue was observed by SEM. The results were shown in Fig.7 and Fig. 8.
Acknowledgements The authors greatly acknowledge the financial support of the National Nature Science Foundation of China (Project No: 50674028), and the foundation of Hebei University of Science and Technology (Project No: XL200923). The authors express their profound gratitude to the editor and reviewers of TMS. References [1]
GUO Jing-jian, "Who will ensure the supply of bauxite in China/' China Metal Bulletin, 42 (2006), 2.
[2]
J.Grzymek, A.Derdacka and Z.Konik, "Method for obtaining aluminum oxide," U.S Patent : 4149898 , 19782-21.
[3]
J.Grzymek, "Complex Production of aluminium oxide and Iron from laterite raw materials applying the calcium aluminates polymorphism," TMS Annual Meeting and Exhibition, Light Metals, 1985, 87-99.
[4]
ZHANG Jing-dong, LI Yin-tai, BI Shi-wen and YANG Yihong, "Research on integrated utilization of high-ferrum bauxite in Guigang Guangxi," Light Metals, (8)(1992), 1618.
[5]
BI Shi-wen, YANG Yi-hong, LI Yin-tai, ZHANG Jingdong and DUAN Zhen-ying, "Study of alumina leaching from calcium aluminate slag," Light Metals, (6)(1992), 1015.
[6]
SUN Hui-lan, YU Hai-yan, WANG Bo, MIAO Yu, TU Gan-feng and BI Shi-wen, "Leaching dynamics of 12CaO#7Al203 ," The Chinese Journal of Nonferrous Metals, 18(10)(2008), 1920-1925.
[7]
TONG Zhi-fang, BI Shi-wen, YU Hai-yan and WU Yusheng, "Leaching kinetics of non-constant temperature process of calcium aluminate slag under microwave radiation," The Chinese Journal of Nonferrous Metals, 16(2)(2006), 357-362.
[8]
D.J.Connor, Aluminium extraction from non bauxitic materials, (Sydney: Aluminium-Verlag Gmbh, 1988), 230250.
[9]
WANG Bo, YU Hai-yan, SUN Hui-lan and BI Shi-wen, "Effect of material ratio on leaching and self-disintegrating property of calcium aluminate slag," Journal of Northeastern University: Natural Science, 29(11)(2008), 1593-1596.
[10]
L.B.Sukla, K.M.Swamy and K.L.Narayana, "Bioleaching
P ^ ^ ^ ^ ^ ^ ^ ^ ^ ^ ^ ^ ^ ^ ■ t e M M M M M M
Fig.7 SEM image of slag with ultrasonic wave ( X 300 ]
Fig.8 SEM image of slag without ultrasonic wave ( X300 ) The dispersion of leaching residue was better when ultrasonic was used in leaching process. The particle size of most of the residue was small. But the dispersion of leaching residue was worse when ultrasonic was not used in the leaching process. The agglomeration phenomenon was serious under these conditions. In addition, other changes such as cavities caused by shock waves and micro jets were not found in the images. Therefore, the effect mechanism of ultrasonic which improved the leaching process was that ultrasonic enhanced the dispersion of solid particle and prevented the agglomeration of leaching residue. Conclusions (1) Ultrasonic could not increase the alumina leaching rate greatly, but it could improve the leaching condition obviously. The leaching temperature decreased from 75 °C to 45 °C, and
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of sukinda latente using ultrasonics," Hydrometallurgy, 37(3)(1995), 387-391. [11]
K.M.Swamy, L.B.Sukla and K.L.Narayana, "Use of ultrasound in microbial leaching of nickel from latérites," Ultrasonics Sonochemistry, 2(1)(1995), 5-9.
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E.Sayan and M.Ayramoglu, "Statistical modeling and optimization of ultrasonic-assisted sulfuric acid leaching of Ti0 2 from red mud," Hydrometallurgy, 71(3)(2004), 397401.
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A.S.Slaczka, "Effect of ultrasound on ammonium leaching of zinc from Galmei Ore," Ultrasonics Sonochemistry, 24(1)(1986), 53-55.
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LI Xiao-bin, ZHAO Zhuo, LIU Gui-hua, ZHOU Qiu-sheng and PENG Zhi-hong, "Behavior of calcium silicate hydrate in aluminate solution," Transactions of Nonferrous Metals Society of China, 15(5)(2005), 1145-1149.
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
THEORY AND EXPERIMENT ON COOLING STRATEGY DURING SEEDED PRECIPITATION Zhanwei Liu 12, Wenmi Chen \ Wangxing Li 1 2 School of Metallurgical Science and Engineering, Central South University, Changsha, Hunan, 410083, China 2 Zhengzhou Research Institute of Chalco, Zhengzhou, Henan, 450041, China Keywords: Seeded precipitation, Cooling way, Sandy alumina temperature of precipitation very low. So the a k (molecular ratio of Na2Ok to A1203) of the pregnant precipitation liquor is low and the supersaturation is high. Secondary nucleation occurs and the fining of the product particle size is caused, making industrial production fluctuate. This situation is more likely to happen when the caustic concentration is high and the seed coefficient is not very high.
Abstract The η-t curve is derived from the dynamic equation of seeded precipitation. The curve has the characteristic of a hyperbolic function and the experimental results conform to the characteristics of the curve. The η-t curve indicates that the precipitation ratio rises rapidly at the preliminary stage of precipitation, the precipitation ratio ascends slowly at middle stage of precipitation and the precipitation ratio increases more slowly at the final stage of precipitation. On the basis of this character of the η-t curve and the technology features of seeded precipitation in producing sandy alumina in Chinese alumina plants, various parameters are determined in the precipitation process as a result of intermediate lowering of temperature.
Some researchers have performed the experiment such that the caustic concentration of sodium aluminate liquor is 125g/l, the a k is 1.71, the seed coefficient is 2.30, the initial temperature of precipitation is 55°C, the final temperature of precipitation is 40 °C and the liquor was precipitated for 72h. They found the product particle size didn't become fine. Given closer analysis we know that because the concentration of the pregnant precipitation liquor is lower, a k is higher and the seed coefficient is relatively high, thus the fining of the product particle size is avoided.
Introduction The precipitation temperature is one of the most important factors influencing the precipitation ratio in the process of seeded precipitation. Based on the η-Τ curve, the precipitation ratio is highest when the temperature is somewhat lower than 40 °C. In industrial production, the temperature of the pregnant precipitation liquor after leaf filtering, is higher (90~95°C), so the liquor must be cooled down and then precipitated. For cooling, the textbook and many papers have shown that we should cool down quickly at first and then slowly. There are two main cooling strategies in industrial production. One way is that the pregnant precipitation liquor is cooled to around 60 °C directly and then precipitated. In general, this route is always taken by plants producing powder alumina. But this way has also been adopted by Pechiney and Pingguo Aluminium Company producing sandy alumina.
Things would be different if the precipitation conditions are slightly more demanding. When the caustic concentration of sodium aluminate liquor is 150~160g/l, the a k is 1.55. This sodium aluminate liquor was precipitated at constant temperature with a different seed coefficient. The product particle sizes with different amounts of seed are shown in Figure 1. 50 < 45 ,
Έ 40 Z2. 35
^
1
30 25
They must add in a great deal of seed, and so the solid content in the first precipitator reaches 750~850g/l. In order to prevent the product from becoming fine, an advanced automatic monitoring system is fitted. The other approach is an intermediate lowering of temperature, that is to say, firstly the pregnant precipitation liquor was cooled down to a temperature range between 70-^75 °C for agglomeration for some time, and then the intermediate lowering temperature is adopted to cool down the liquor for further precipitation. This way is used by plants producing sandy alumina with two-stage processes that is represented by the Aluminum Company of Switzerland.
ω 20 o
£
15 10 r 5 ! 50
55
60
65
70
75
Tenperature( °C )
Figure 1. The product particle sizes with different amount of seed
Using The Strategy Of Single Stage Cooling Or Intermediate Lowering Of Temperature
It can be seen from the Figure 1 that when the seed addition is high (300g/l and 500g/l, for example), the product particle sizes don't become fine at a precipitation temperature 60 °C, but the product particle sizes begin to become fine at precipitation temperature 55 °C; when the seed addition is 150g/l, the product
In general, if the strategy of single stage cooling is adopted, the initial temperature of precipitation must be cooled to a low temperature, such as 60 °C or under, in order to make the final
227
particle sizes are significantly fine at precipitation temperature 55 °C . When seed addition is low, such as 50g/l, even if the precipitation temperature is 70 °C, the product particle sizes also become fine, with the fining of the product particle sizes more serious at precipitation temperature 65"Cor lower.
A=\ +
1
at
C3t
η = ^ί + ^
(Λ-Α,Υ
Λ c, =· '' A.-4
(i)
C3=^
= KJ + C
The Character Of The η-t Curve In Laboratory Tests
When t=0, A=Ao, we can solve for C as follows:
c=-
We have done the precipitation testing in the laboratory to simulate plant conditions. The caustic concentration of the sodium aluminate liquor is 155g/l, the a k is 1.55. This sodium aluminate liquor was precipitated at constant temperature with different seed coefficients. The seed additions are 300g/l and 500g/l respectively with the seed obtained from the Henan Branch China Aluminium. The η-t curves are shown in Figures 2 and 3 respectively.
1
Λ-Λ
Where AQ is the concentration (g/1) of A1203 in sodium aluminate liquor. So we can solve for A as follows:
- A
K0(A0-Ae)
We will further explain the character of the η-t curve by the laboratory test.
(A-Ae)
1 A-Ae
=
The η-t curve equation (4) is a hyperbolic function, that is to say, when the seed coefficient is fitted (if the seed coefficient is too low, the induction period will be prolonged). The characteristic of the η-t curve is that the precipitation ratio rises rapidly at the preliminary stage of precipitation, the precipitation ratio ascends slowly at the middle stage of precipitation and the precipitation ratio increases more slowly at the final stage of precipitation.
(2)
Solving the differential equation (2) as follows:
-
(4)
Where:
Where A is the instant concentration (g/1) of A1203 in the sodium aluminate liquor, Ae is the equilibrium concentration (g/1) of AI2O3 in sodium aluminate liquor, t is the precipitation time (h), and Ko is precipitation rate constant. Equation (1) is transformed as follows:
-'<*-*>-**
Α,-Λ
c2t
Researchers in China and abroad have carried out many studies about the dynamics of seeded precipitation in sodium aluminate liquors, and similar dynamic equations are put forward. Next, the η-t curve is derived from the dynamic equation of seeded precipitation. Take following precipitation rate equation for example to deduce the η-t curve.
K0(A-Ae)2
A
= l-±-
The Theory Of Intermediate Lowering Temperature
=
(3)
It then substitutes into the precipitation ratio formula, equation (3), and we can find the n as follows:
The major bauxite resource is diaspore ore in China. The digestion conditions of this bauxite ore are more exigent, and the concentration of the pregnant precipitation liquor is high, that is always between 140g/l and 160g/l. If the strategy of one-stage cooling is adopted to make the initial temperature of precipitation drop to 60 °C or even lower, the fining of the product particle size is difficult to avoid.
~
Λ-Λ
i+K0t(A0-Ae)
1
-A- K t
A ι+* 0 ι(Λ-Λ)
-.
228
the precipitation ratio is about as high as that in the temperature range of 55-70 °C when the seed addition is 500g/l, the precipitation ratio is much higher than that in the temperature range of 55-70 °C after precipitating 12 hours when the seed addition is 300g/l. The Determination Of Various Parameters Of Intermediate Lowering Temperature By Using The Change Rules Of The η-t Curve The Determination Of The Initial Temperature Of Precipitation
0
10
20
30 Tira* h)
40
50
Because the precipitation ratios are about the same for the first 12 hours in the temperature range of 55-70 °C , it is proper to determine the initial temperature of precipitation in the temperature range of 65-70 °C which is a bit higher within the temperature range of 55-70°C. By doing this, the precipitation ratios are not affected and the fine particles of aluminium trihydroxide can be agglomerated very well.
60
Figure 2. The η-t curve with 300g/l seed addition
The Determination Of The Time Of Intermediate Lowering Temperature
0
10
20
30 Tìne(h)
40
50
Because the precipitation ratios are about same in the abovementioned temperature ranges, and the gaps become too big for different temperatures with the longer times of precipitation, we should cool down after precipitation for 12-14 hours. In this way the liquor can be precipitated at a low temperature in time, and the precipitation ratio is improved. If the time of intermediate lowering of temperature is too late, the liquor is still precipitated at high temperature, so the precipitation ratio is influenced. The later it is cooled down, the greater the precipitation ratio is influenced. It is not appropriate to cool down too early, for example after precipitating 7-8 hours, if so, the product particle size may become fine and scale is readily produced on the interstage cooling equipment.
60
Figure 3. The η-t curve with 500g/l seed addition
The Modes Of The Precipitator Before The Intermediate Lowering Of Temperature
The η-t curves in Figures 2 and 3 are completely in conformity with the change rules of the above-mentioned η-t curve derived from the dynamic equation. The precipitation ratio rises rapidly at the preliminary stage of precipitation. It can be seen from Figures 2 and 3 that in the temperature range of 50-70 °C, the precipitation ratios reach about 30% for the first 12 hours which is the 65 per cent of the total precipitation ratios for 60 hours when the seed addition is 300g/l; the precipitation ratios reach about 35% for the first 12 hours, which is 70 per cent of the total precipitation ratio for 60 hours when the seed addition is 500g/l. But the precipitation ratio ascends slowly over the middle stage of precipitation and the precipitation ratio increases slower and slower at final stage of precipitation. The precipitation ratios are about same at different times for the first 8 hours in the temperature range of 50-70 °C; the precipitation ratios are also about same for the first 12 hours in the temperature range of 55-70 °C ; the precipitation ratios are different at different temperatures after precipitation for 24 hours, the gap increases with longer time of precipitation, and the precipitation ratios are higher with decreasing temperature. If the precipitation temperature is too high (75 °C), the precipitation ratios are much lower than that in the temperature range of 50-70 °C at the preliminary, middle and final stage of precipitation; if the precipitation temperature is too low (50°C),
Seeded precipitation is performed in batch operation or continuous operation. Currently, continuous operation is adopted in most plants. The disadvantage of continuous operation is that the pregnant precipitation liquor with the lower a k flows continuously into the precipitator to rapidly mix with the precipitated liquor ( a k is higher) in this precipitator, so the a kis increased immediately and the precipitation ratio is decreased. Of course, the precipitation ratio of the liquor flowing from this precipitator is also decreased. From the characteristics of the above-mentioned η-t curve, we know that the precipitation ratios for the first ten hours or more, are 65-70 per cent of the total precipitation ratios for the full precipitation time. So it is at the initial stage of precipitation (more specially, it is at the first precipitator) that the seeded precipitation performed in continuous operation influences the precipitation ratio remarkably. But there are no obvious influences on the middle and final stages of precipitation. To a certain extent, the disadvantage of continuous operation can be overcome by adopting the strategy of intermediate lowering of temperature. That is to say, the pregnant precipitation liquor flows into the precipitator from the top which is stirred interlayer, so the a k of the pregnant precipitation liquor is not significantly increased,
229
then the precipitated slurry flows out from the bottom of the precipitator and was pumped into interstage cooling equipment.
These include the time of intermediate lowering of temperature, the initial temperature of precipitation and the cooling extent of intermediate lowering of temperature and so on. This has certain direct significance to the determination of the temperature decreasing strategy in the process of seeded precipitation.
The Technical Projects On The Intermediate Lowering Of Temperature Cooling The Pregnant Precipitation Liquor By Plate Heat Exchangers
Reference 1. Wangxing Li, "The Impact of Temperature on Precipitation Ratio in the Process of Seeded Precipitation," Light Metals, (5) (1998), 14-18.
The pregnant precipitation liquor was cooled down to the temperature ranges between 65/~~70°C from the temperature 95 °C by using plate heat exchangers. It is calculated that it will be possible to save 200,000 tons of steam every year, generating 10 million yuan every year in economic returns, if the spent seed precipitation liquor was preheated by the way of cooling down the pregnant precipitation liquor.
2. Yuling Wang, Huaizhong Zhu and Liwei Zhou, "Application of Intermediate Lowering Temperature in the Process of Seeded precipitation," Advanced Ceramics, (2) (2005), 27-28. 3. Junzhao Fan, "The Technology of Producing Sandy Alumina Using Sodium Aluminate Solution with High Solid Content," Non-ferrous Smelting, (3) (2002), 50-54.
The Location Of Interstage Cooling Equipments In The Series Of Precipitators
4. Jinqing Chen et al., "Depth Analyzing of Seeded Precipitation and Discussion on Intensify Method" Non-ferrous Metals, (6) (2003), 30-34.
It has been mentioned above that it is appropriate to cool down the liquor after precipitating 12-14 hours. So the location of interstage cooling equipment will depend on the retention time of the liquor in each precipitator. If the retention time of the liquor in each precipitator is range of 10~12h, we should begin to cool down the liquor after it flows out from the first precipitator. Of course, we should begin to cool down the liquor after it flows out from the second precipitator, if the retention time of the liquor in each precipitator is in the range of 5~8h.
5. Chongyu Yang, The Process Technology of Alumina, (China, NY: Metallurgical Industry Press, 1993).
The Tube Heat Exchangers As The Interstage Cooling Equipment Because there are a lot of Al(OH)3 solids in the liquor, the plate heat exchangers are easily blocked, so the plate heat exchangers are not suitable for this application. Tube heat exchangers and vacuum heat exchangers are preferred equipment. The tube heat exchanger is simple and practical, which was developed in Hungary and used successfully in industrial production. The temperature of the slurry is 65~70°C as mentioned above, and it needs to be cooled to around 50 °C, thus the cooling requirement is considerable. So if it is not enough to cool down by using one-stage cooling equipment, we can adopt a multi-stage series to cool down the liquor to the set temperature. To conclude, the temperature of the pregnant precipitation liquor after leaf filtering is 90 ~ 95 °C . It was cooled down to temperature ranges between 65 ~ 70 °C by the plate heat exchangers and flows into the first precipitator from the top, then the precipitated slurry flows out from the bottom of the precipitator and was pumped into multi-stage tube heat exchangers to cool down to temperature ranges between 50~55 °Cand then was feed into follow-up precipitator series to continue precipitation. The temperature of the precipitated slurry in the last precipitator is 40~45°C. Conclusion The characters of the η-t curve are studied by the way of the basic theory of precipitation, combined with experiments. This characteristic curve elucidates the rationale for adopting the strategy of intermediate lowering of temperature in the process of seeded precipitation and the technical parameters are determined.
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Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
EXTRACTION OF ALUMINA FROM RED MUD BY DIVALENT ALKALINE EARTH METAL SODA ASH SINTER PROCESS S. N. Mener1, A.K. Rout1, B. K. Padhi2 1
Department of Chemistry, Kiit University, Bhubaneswar, Orissa, India, Pin-751024.
2
R & D Department, Nalco, Damanjodi, Dist.-Koraput, Orissa, India, Pin -763008. Email Id: [email protected] Key Words: Red Mud, Sintering, Leaching and Divalent alkaline earth metal oxide. soda and alumina. Recoveries of 95% Na20 and 70% A1203 were obtained [5]. Alyanak et al. [6] studied on the recovery of aluminium from the red mud of the Seydisehir Aluminium Company Limited. They prepared a mixture composed of 26.6 wt% red mud, 37.2 wt% CaC03 and 36.2 wt% Na2C03. The mixture was sintered at 900°C for 2 hours then leached in water. Increasing the sintering temperature as high as 1150°C decreased the alumina extraction efficiency.
Abstract A novel method for extraction of alumina from red mud using a divalent alkaline earth metal oxide soda ash sinter process followed by thermo chemical leaching in suitable caustic concentration at 105°C for different times is presented in this paper. During sintering Sodalite/Cancrinite, Gibbsite and Boehmite in red mud is converted to sodium aluminate as well as stable phases of divalent alkaline earth metal ferrite, titanate and silicate. In the above novel economic process an extraction of alumina from red mud of 97.64%, 98.7% and 99.5% was achieved using CaO, MgO and BaO respectively. The morphological study of sinter products and leached products was studied using SEM and XRD.
In another article, a gallium recovery method was applied to bauxite ores, during alumina production by the Bayer process. Gallium movement and possible locations of gallium recovery in the Etibank-Sedisehir Aluminium Plant and some results for gallium concentrate production by precipitation were presented [7]. S. Uday [8] had worked on the recovery of iron, titanium and aluminium values from red mud by calcining it, preferably between 800°C to 1350°C and smelting the material with a carbonaceous reducing agent in an electric furnace, thus obtaining molten iron and a slag containing substantially all the titania, alumina and silica.
1. Introduction Red mud is a by-product in the manufacture of alumina that contains mainly iron oxide (54-65%) with significant amounts of silica, alumina, calcium oxide and titania, dispersed in a highly alkaline and caustic liquor [1,2]. The treatment and disposal of bauxite residue is a major operation and may account for 30-50% of operations in an alumina refinery [1]. The red mud has been accumulating at a rate of 70 million tonnes annually throughout the world [3]. For a given aluminum production rate, the quantity of red mud generated during the alumina extraction process varies significantly depending on the original properties of the bauxite and the operating conditions of the Bayer process, and in particular, the process temperature [4]. Depending on the factors, between 1 and 2.5 tonnes of bauxite residue is generated per tonne of alumina produced [1]. The disposal of the red mud is associated with pollution problems. Disposal of any solid waste is associated with space or real estate near the industry, and the cost of disposal and pollution, which are now crucial factors. Obviously, these three factors are also associated with the red mud disposal. As the red mud contains a large amount of valuable chemicals, there is a need for developing a technology for the recovery of at least some of the important chemicals [3].
The red mud has also been investigated for the making of artistic glazes in ceramic. For this purpose, no additives and some fluxing materials such as read lead, ulexide, cryolite, borax, soda ash, etc. were added to the red mud in different proportions. Finally, the optimum conditions forming novel glazes were determined [9]. The red mud was also investigated to find out whether it can be used as a construction material or not. For this purpose, the red mud, a waste material of Aluminum Plant in Seydisehir, was mixed with clay, which was taken from the same region. This mixture was pressed and sintered at between 1000 and 1150°C. It was stated that construction materials produced from the mixture had higher compressive strength and lower water absorption ratio compared with the related standards [10]. In another study, the structure of the red mud deposits of Seydisehir, Turkey was investigated by simple statistical analysis. It was shown that the ratio of the iron oxide % (D i) in the red mud to the mean iron oxide % (D2) was found to fît a linear function represented by plotting D\ID2 versus K λΙΚ 2 in a double logarithmic plot. The relation in the form of a power function was D 1= 1.0092D2/r10021/j:2-0022[ll].
Sulfided red mud is active as a hydrogénation catalyst, due to its iron sulfide content, and has been used for the hydrogénation of organic compounds, and the liquefaction of coal and biomass [2].
The thermal behaviour of the red mud was investigated in the literature. An extensive characterization was performed by thermal and X-ray diffraction analyses. Dried red mud was substantially inert up to 900°C, the loss of H 2 0 from aluminium hydroxides and of C0 2 from silico-alumino-carbonates being the only detectable effects [12]. Several methods for extracting alumina as well as for recovery of other metal values from alumina-bearing materials
In the literature, a lot of different articles that explain the properties and the investigation of the red mud tailings can be found. One such reference investigated the red mud under caustic concentrations of 10-30% to determine the extraction of soda and alumina as a function of caustic ratio, CaO: Si0 2 ratio, water content, temperature and reaction time. Hydrothermal treatment of the red mud with lime at high temperatures of 300°C was found to be an effective method for the recovery of
231
adding divalent alkaline earth metal oxide (i.e. CaO, MgO and BaO) and Na2C03 to the red mud and sintering at temperatures of 900-1100°C [14]. The silica in the red mud reacts with divalent alkaline earth metal oxide to form the relatively inert divalent alkaline earth metal silicate as shown in equation: 1. The sodium, which is in the red mud via soda ash or sodium carbonate, reacts with the alumina and forms the soluble sodium aluminate as given in equation:2 [15]. The Fe203 contents of red mud reacts with sodium carbonate to form sodium ferrite as given in equation: 3.The Ti0 2 which is in the red mud reacts with sodium carbonate forms sodium titanate as given in equation: 4. The sodium titanate reacts with divalent alkaline earth metal oxide forms divalent alkaline earth metal titanate and free Na20 as given in equation: 5. The sodium ferrite reacts with divalent alkaline earth metal oxide forms divalent alkaline earth metal ferrite and free Na20 as given in equation: 6.
have been studied [13]. The United States Bureau of Mines has been evaluating possible alternatives to the use of bauxite for several years. These include acid leaching processes and sintering processes applied to clays and anorthosites. Fly ash from burning regular coals has also been investigated extensively as a potential source for alumina. The processes studied for the extraction of alumina from fly ash include variations of existing sintering and leaching processes. There are now a few millions of tonnes red mud stored in Red Mud ponds in India. These wastes have not been investigated yet in any industrial process or used for material production as an additive. The aim of this study was to investigate these muds for maximum alumina extraction by a divalent alkaline earth metal oxide soda ash sinter process followed by thermo-chemical leaching. 2. Materials and Methods
Si0 2 +MO -> M2Si04 + 2H20 A1203 + Na2C03 -► 2NaA102 +C0 2 Fe 2 0 3 + Na2C03 -♦ 2NaFe02 +C0 2 Ti0 2 + Na2C03 -> Na2Ti03 +C0 2 Na2Ti03+MO -»MTi0 3 +Na 2 0 2NaFe02 + 2MO -> M2Fe205 +Na20
Red mud and crushed Lime (CaO) used in experimental studies were supplied from, Damanjodi, Dist-Koraput. The MgO, BaO and Na2C03 used in experimental studies were supplied from Qualigens fine Chemicals, Bombay. The major elements were analysed by classical methods as well as X-Ray Fluorescence Spectroscopy (Model PW 2400). The minor elements were analysed by AAS. The chemical composition of the red mud is shown in Table 1. X-Ray Diffraction (XRD) analysis of sintered red mud and leached red mud were carried out with a Diffractometer using CuKa (Ka= 1.54186 A) Rigaku Dmax 2200, Japan. The morphological behaviours of sinter red mud and leached red mud were studied by SEM (Model:430, Electron Microscopy, United Kingdom).
The product is then leached in an alkaline solution or water as given in equation: 7 and the sodium aluminate solution is directed to the precipitation stage of the Bayer process. NaA102(s) + M2Si04(s) + 2NaFe02(s) + 3H20 -» NaAl(OH)4(aq) + 2NaOH(aq) + M2Si04(s) + Fe203(s) (7) To produce high aluminium and sodium extractions the decomposition of divalent alkaline earth metal silicates must be as low as possible. In practice the amount of silica extracted in the leaching step is too high, as the M2Si04 is not sufficiently stable under the leaching conditions [16] and silica extractions of around 15-20% [17] occur during leaching. This ties up sodium and aluminium with the formation of Mg3Al2(OH)12, Ca3Al2(OH)12 or Ba3Al2(OH)12, Hydro garnet and DSP. The opportunity exists for a process to treat red mud to separate aluminium and sodium from silica, producing an adequately stable insoluble phase for leaching and therefore yielding greater extractions.
An amount of 100 g of red mud, 10 g of divalent alkaline earth metal oxide (i.e. CaO, MgO and BaO) and 15 g / 20 g/25 g of Na2C03 was mixed thoroughly and pulverised in a pulveriser, then the above mixture was sintered at temperatures of 900, 1000, 1100°C for 1 to 4 hrs in a ceramic crucible in a high temperature muffle furnace. An amount of 100 g of red mud, 15 g of divalent alkaline earth metal oxide (i.e. CaO, MgO and BaO) and 10 g/20 g /25 g of Na2C03 was mixed thoroughly and pulverised in a pulveriser, then the above mixture was sintered at temperatures of 900, 1000, 1100°C for 1 to 4 hours in a ceramic crucible in a high temperature Muffle furnace. An amount of 100 g of red mud, 20 g of divalent alkaline earth metal oxide (i.e. CaO, MgO and BaO) and 10 g/20 g/25 g of Na2C03 was mixed thoroughly and pulverised in a pulveriser, then the above mixture was sintered at temperatures of 900, 1000, 1100°C for 1 to 4 hrs in a ceramic crucible in high temperature muffle furnace. After sintering the sinter products were leached with 80 gpl caustic at 105°C for 1 hr in an autoclave of 2.5-litre capacity. The leached red mud was filtered after the leaching. The amount of A1203 in the leach solutions was determined by using a titrimetric method.
An alternative method of divalent alkaline earth metal oxide Na2C03 sinter is presented in this paper to produce a M : Si ratio of one in the sinter product[16]. The potential benefits are to halve raw material costs and to give a more stable insoluble phase. Also the extra sodium needed could be provided by spent liquor creating the potential for organics removal through liquor burning in the sinter step. The removal of organic impurities in sintering followed by leaching at 105°C is due to formation of TCA or Layered Double Hydroxide [Hydrotalcite: Mg6Al2(OH)16 C03.4H20)].
Table 1. Chemical Analysis of Red Mud €&C9BÉCtf
AbO>
FesQs
THfe
SiOr
N*0
CaO
VaO*
IfgQ
<*0
iS-IS
5Ï-57
3-5
S-Î2
4-6
8.8823®
Ü14Ù.2Ì
0.13-
UaO
loia*
02»
11-13
(l) (2) (3) (4) (5) (6)
However, thermodynamic calculations and laboratory tests show that MSi03 does not form under sinter conditions [18]. From the present experiments it was thought that the following divalent alkaline earth metal oxide sinter reaction was feasible due to the presence of sodalite in the red mud formed by the reaction of kaolin and sodium aluminate liquor.
3. Results and Discussion Methods of aluminium and sodium recovery from red mud have been around for many years due to the requirement of some refineries to process very high silica bauxites. The most common method to process red mud is the "divalent alkaline earth metal oxide Na2C03 sinter process". The process involves
Na2O.Al203.2Si02 ( Sodalite ) +2 MO + 4Na2C03 -* 2Na2MSi04+2NaA102+4C02+2H20 (8)
232
A similar lime sinter process is referred to in a paper discussing the integration of coal combustion with lime sintering [19]. Reaction (8) shows that soluble sodium aluminate and sodium alkaline earth metal silicates are produced. The sodium associated with sodium alkaline earth metal silicate must also be recovered in the leaching step. The objective of this project was to investigate the use of a M : Si ratio of one, in a divalent alkaline earth metal oxide, Na2C03 sinter process for use as a red mud treatment for the recovery of aluminium and sodium and to compare this with the current lime soda ash sinter system of M:Si = 2.
RM+MgO+Na2C03(1 :.1 :.25)-900-4 4000-
LO
g- 3000 -
'S
1 2000 e
1 2 3 4 5 6 7 8 9
Temperature in°C
Time in hours
900 1000 1100 900 1000 1100 900 1000 1100
1 3 4 4 4 3 1 4 1
2.50
I? 3.50
4.50
5.50
d-value
Figure 1: MgO Soda Ash Sinter Process with highest Extraction Efficiency 98.70% at 900°C.
Table 2. Extraction Efficiency of divalent alkaline earth metal oxide soda ash sinter process at different condition. Condition RM+MO (CaO, MgO, BaO)+ Na2C03 1:0.10:0.20 1:0.15:0.25 1:0.20:0.25 1:0.10:0.25 1:0.15:0.20 1:0.20:0.25 1:0.10:0.25 1:0.15:0.25 1:0.20:0.25
i 1 Irl
1.50
The extraction efficiency of divalent alkaline earth metal oxide soda ash sinter process is shown in Table 2.
SI. no.
co
Extraction 1 Efficiency (%)
RM+CaO+Na2C03(1 :.2:.25)-1100-1
Θ
61.79 66.66 97.64 98.70 97.22 96.50 99.14 99.50 95.57
Figure 2: Lime Soda Ash Sinter Process with highest Efficiency 97.64% at 1100°C.
*Sl.No.l to 3: RM + CaO + Na2C03 sinter process, Sl.No.4 to 6: RM + MgO + Na2C03 sinter process, and Sl.No.7 to 9: RM + BaO + Na2C03 sinter process. In the synthetic MgO Na2C03 sinter stage, calcination at 900°C for 4 hrs with red mud, MgO and Na2C03 in a ratio of 1:0.1:0.25 was found to produce the most stable Sodium Magnesium Silicates, Di-Magnesium Silicates, Magnesium Ferrite (SEM Fig.6) and Magnesium Titanate (SEM Fig.7). This was confirmed by XRD (as shown in Fig .1 and Table 3) and maximized the extraction of alumina at 98.7%. The lowest extraction efficiency of 47.84% obtained in the MgO Na2C03 sinter process with red mud MgO Na2C03 ratio (1:0.15:0.10) sintered at 1000°C for 1 hr is due to the formation of insoluble magnesium aluminium silicate hydrate and sodium aluminium magnesium silico titanate [NaAlMgSiTiOx].
In the synthetic BaO Na2C03 sinter stage, calcination at 1000°C for 4 hrs with red mud, BaO and Na2C03 in a ratio of 1:0.15:0.25 was found to produce the most stable Sodium Barium Silicates, Di-Barium Silicates, Barium Ferrite (SEM Fig. 8) and Barium Titanate (SEM Fig.9) confirmed by XRD (as shown in Fig. 3 and Table 3) which maximized at 99.5% the extraction of aluminium and sodium. The lowest extraction efficiency of 32.56% obtained in the BaO Na2C03 sinter process with red mud BaO Na2C03 ratio (1:0.2:0.1) sintered at 1100°C for 3 hrs is due to the formation of insoluble Barium Aluminium Silicate Hydrate and Sodium Aluminium Silicate. RM+BaO+Na2C03(1 :.15:.25)-1000-4
In the synthetic CaO Na2C03 sinter stage, calcination at 1100°C for 4 hrs with red mud, CaO and Na2C03 in a ratio of 1:0.2:0.25 was found to produce the most stable Sodium meta aluminate, Sodium Calcium Silicates (Fig.4.a) Di-Calcium Silicates (Fig.4.a), Calcium Titanate (Fig.4.b) and Calcium Ferrite (Fig.5.a and 5.b) confirmed by XRD (as shown in XRD Fig. 2 and Table 3). This procedure maximized the extraction of alumina at 97.64%. The lowest extraction efficiency of 38.35% obtained in the CaO Na2C03 sinter process with red mud CaO Na2C03 in a ratio (1:0.1:0.15) sintered at 900°C for 2 hrs is due to the formation of insoluble calcium aluminium silicate.
B
1800
5.50
Figure 3: BaO Soda Ash Sinter Process with highest Extraction Efficiency 99.50% at 1000°C.
233
The existing CaO Na2C03 sinter product ties some calcium up with iron due to the formation of Ca2Fe205 (d value= 2.70,2.67,2.10,1.84, XRD Fig. 2 and Table 3), which have Whisker like structures as shown in the SEM images in Fig.5.a and 5.b. This meant that extra calcium was added to ensure the maximum amount of silica was reacted with the calcium and not with sodium or aluminium. At 1100°C, instead of sodium aluminate and sodium ferrite forming, Na2AlFe04 formed in products with iron present. For the synthetic new lime soda ash sinter process without iron it was found that a two stage caustic or water leach (105°C/60 min., 105°C/240 min.) was required to gain maximum Na and Al extractions due to the decrease in Al in solution over time as shown in Figure 10. It was found that in an MgO Na2C03 sinter process sintering at 900°C for 4 hrs with red mud MgO Na2C03 in a ratio of 1: 0.1:0.25 followed by leaching, formation of Sodium aluminate, MgTi03 and Mg2Si04 maximise alumina extraction up to 98.7%. In a BaO Na2C03 sinter process sintering at 1000°C for 4 hrs with a red mud BaO Na2C03 ratio of 1:0.15:0.25 followed by leaching and formation of Sodium Aluminate, Ba2Fe205 (SEM Fig.8) and BaTi03 (SEM Fig.9) give maximum alumina extraction efficiency of 99.5%.
When Iron is introduced to the system there are effectively two soluble phases (NaA102 + NaFe02) in which the ratio of soluble sodium to soluble aluminium increases. The results in Figures 12 and 13 show the benefit of the new alkaline earth metal Na2C03 sinter process in terms of aluminium extraction. Much of the sodium and all of the aluminium is extracted from the existing alkaline earth metal sinter process in the first stage. The second stage merely completes the leaching of sodium. In the new red mud alkaline earth metal sinter extraction the resultant solids have a sodium content of 0.36 % compared to the 6.00 % sodium content of the red mud. The aluminium levels have been reduced from 17.8% to 0.42%. A more stable sodium alkaline earth metal silicate would increase these extractions and could be achieved with the optimisation of sinter temperature and leach conditions.
This allowed the removal of the majority of the aluminium before the decomposition of the sodium divalent alkaline earth metal silicate and subsequent reaction to form DSP, Hydro garnet or Mg3Al2(OH)12, Ca3Al2(OH)12 or Ba3Al2(OH)12. To ensure leaching conditions were the same for both systems the two-stage process was also employed for the existing lime sinter products. The results of extractions for the sinter products without iron are displayed in Figure 11.
Fig.12.New vs E x i s t i n g S y n t h e t i c CaO N a 2 C 0 3 Sinter
Fig.13 New Vs E x i s t i n g Red Mud CaO N a 2 C 0 3 Sinter 10
20 Tim e(hrs)
Fig.10 Single Stage Extraction over time
100
<
Mag-
2.00 KX
E H T * 15.00 kV
Fig.11.New vs Existing S y n t h e t i c CaO N a 2 C 0 3 S i n t e r w i t h o u t Iron
f°»m'
Detector = SE 1 Date :11 Jul 2007
Figure 4a: Formation of Sodium Calcium Silicate and DiCalcium Silicate (2.00 KX)
234
Figure 4b: Formation Calcium Titanate (4KX) J
*c
i "XL' % MM
Figure 5a: Formation of Calcium Ferrite (4KX)
Figure 7: Formation of Magnesium Titanate (4.05KX)
Figure 5b: Formation of Calcium Ferrite 10K
Figure 8: Formation of Barium Ferrite (4.69KX)
235
2. In the MgO-Na2C03 sinter process alumina extraction achieved was 98.7%. 3. In the BaO-Na2C03 sinter process, the extraction efficiency achieved was 99.5%. But from an economic point of view, the lime-soda sinter process is best and suitable method for extracting the remaining alumina from Bayer's process waste residues (Red Mud). References
Figure 9: Formation of Barium Titanate (3.88KX) Table 3. Formation of major phases confirmed by XRD during sintering of red mud followed by leaching Sinter Process Red Mud + CaO +Na2C03
Red Mud + MgO+Na2C03
RedMud + BaO + Na2C03
Formation of Major phases Dicalcium Silicate Calcium Titanate Calcium Ferrite NaA102 CaSi03, CaV206 CaMn03 Calcium Aluminum Silicate Dimagnesium Silicate Magnesium Titanate Magnesium Ferrite NaA102 Dibarium Silicate
Characteristic d spacing (A) 2.21
Barium Titanate Barium Ferrite NaA102
2.13,2.24 1.63,2.43,2.63,2.78 2.13,2.23
1
2.70,2.52,2.87 2.70,2.67,2.10,1.84 2.13,2.59,2.64,2.95 2.70,2.67,2.21 2.67 2.64 2.53 2.70,2.53,2.09 2.96 2.96 2.78,3.15
4. Conclusion 1. By adoption of the sintering method followed by alkaline leaching, the alumina extraction achieved was 97.64% in LimeSoda Sinter process and impurities like Si02, V 2 0 5 and Mn0 2 are removed from the liquor. These impurities are removed from the crystal lattice of Hematite as well as Sodalite and Cancrinite to form insoluble materials like CaSi03 (d value= 2.70, 2.21), CaV206 (d value= 2.70, 2.67) and CaMn03 (d value= 2.67) confirmed by XRD.
1. Q. D. Nguyen and D. V. Boger, Int J Miner Procès, (54) (1998), 217-233. 2. J. Alvarez, S. Ordonez, R. Rosai, H. Sastre and F.V. Diez, Appi CatalA: Gen, (180) (1999), 399-409. 3. B. K. Mohapatra, M. B. S. Rao, R. Bhima Rao and A. K. Paul, "Characteristics of Red Mud generated at Nalco Refinery, Damanjodi, India", (Light Metals 2000 as held at 129th TMS Annual Meeting; Nashville, TN; USA, 2000), 161-165. 4. L.Y. Li, Waste Management, (21) (2001), 525-534. 5. P. J. Creswell and D. J. Milne, (Proceedings of the Annual Meeting on the Light Metals AIME, TX, USA,1982), 227-238. 6. C. Alyanak, C. Oktaybas, H. Sesigur and E. Acma, (Proceedings of the International Metallurgy and Materials Congress, Istanbul, Turkey, 1995), 91-97. 7. H. Yuzer, E. Avci, O. Emrullahoglu, E. Gencer and T. Haser, (Proceedings of the International Metallurgy and Materials Congress, Ankara, Turkey ,1993), 215-223. 8. S. Uday, Metallurgical and Chemical Process Limited (Hamilton, Canada), U.K.Patent GB 843607(1960). 9. Z. Mete and A. Cam, (Proceedings of the Chemistry and Chemical Engineering Congress, Istanbul, Turkey, 1992), 1924. 10. Kara M., Ekerim A. and Emrullahoglu O. F., (Proceedings of the International Metallurgy and Materials Congress, Istanbul, Turkey, 1995), 1435-1440. 11. Sahin S, Hydrometallurgy, (47) (1998), 371-376. 12. V.M. Sglavo, R. Campostrini, S. Maurina, G. Carturan, M Monagheddu, G. Budroni and G. J. Cocco, Euro Ceramic Soc, (20) (2000), 235-244. 13. A.K. Mehrotra , P.K. Bashnor and W.Y. Svrcek, Cart Jour ChemEng.,(57) (1979), 225-32 14. W. R. King, "High Caustic Mud-Sinter process for high silica weips Bauxite", Non-Metallic Materials Research, Kaiser Aluminium and Chemical Corporation, (RR 80-71,Project 7991-60310,1980). 15. P. Leci and A. Guidi, "Pyrogenic Attack of Bauxite" extractive Metallurgy of Al, (1) (1962), 231-249. 16. A. Hartsborn, "Lime Sinter with Calcium: Silica ratio: :l",Comalco, (RTS Technical Note 9392, 2000). 17. Q. Likuan, "The Recovery of Aluminium by Soda-Lime Sinter Process", (Zheng Zhon Conference Proceedings 1993). 18. Osborne J., "Thermodynamic Analysis of Lime Sintering of De-Silication Products ", (Rio Tinto Technical Note, 27731/1997). 19. V. L. Rayzman and I. K. Filipovich, "Integrating coal combustion and red mud sintering at an alumina refinery," 70M,51(8)(1999),16-18.
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
DISSOLUTION KINETICS OF SILICON FROM RED MUD IN PURE WATER Xiaohui Li, Kai Huang, Hongmin Zhu* School of Metallurgical and Ecological Engineering, University of Science and Technology Beijing, 100083, China (* Corresponding author: Tel: +86-10-62332267; Email: [email protected] ) Keywords: red mud; dissolution kinetics; liquid layer diffusion; Noyes-Whitney equation atmosphere and so decreasing the greenhouse effect.
Abstract
Haraguchi et al proposed that the utility of steelmaking
In this study, the dissolution kinetics of silicon from sintering
slag as the nutritive elements to promote the growth of algae
red mud in pure water was investigated. Experimental factors
[8,9]. Now in this study, we have tried to check the feasibility of
such as particle size and stirring speeds were studied to test their
using red mud as a nutrient source. Our experiments indicated
effect on the extraction fraction of silicon from red mud. It was
that it is also quite viable to use red mud for the significant
found that the silicon extraction increases with stirring speed
promotion of the growth of algae and the results will be
and decreasing in the particle size of red mud. Based on the
published later. In the process of the production of alumina, a
experimental results, the Noyes-Whitney equation was adopted
huge amount of red mud is produced causing serious
to describe and explain the dissolution behavior of the red mud
alkalization of the soil and polluting the underground water after
in pure water, and the experimental data fit the model quite well
long term storage, and the treatment and comprehensive
in the whole process of dissolution.
utilization of red mud is a troublesome problem. Though a large number of studies have been conducted on the comprehensive
Introduction
utilization of red mud, the research progress is not significant.
Thousands of phytoplankton live under the surface water
However, taking accountant of the fact that red muds are also
of the ocean and constitute the primary productivity of the ocean.
rich in Fe and Si, we tried to investigate its dissolution behavior
The phytoplankton breeds rapidly under appropriate conditions,
preceding its real application in promoting the growth of algae
and the growth of phytoplankton would consume huge amounts
of the ocean.
of carbon dioxide through photosynthesis [1]. Of the global emissions of carbon dioxide, about 50% of those from human
Experimental
activities are captured the oceans [2]. Phytoplankton growth
The red mud sample used in this study was donated by
needs not only C, N, P nutrient sources, but also Fe, Si and other
Shandong Aluminum Company. Samples were dried, crushed
trace elements. Martin et al found that deficiency of iron can
into smaller particles by ball milling and sieved to the different
inhibit the growth of algae [3,4]. However, there are many
particle size fractions. Particles with the size range of
oceans that lack trace elements such as Fe, Si. It is reported that
39-185μπι were sent for analysis of chemical composition by
about 20% of the ocean area has excessive N, P nutrients, but
X-ray fluorescence spectroscopy.
the phytoplankton biomass is very low, and this area is known
The dissolution experiments were carried out in the plastic
as high nutrient-low chlorophyll (HNLC) [5]. The researchers
containers. The suspension was agitated with a Teflon stirrer,
found that the iron limits phytoplankton productivity in the
driven by a magnetic stirring apparatus. A water-bath heater was
HNLC, thus the deficiency of iron affected the carbon dioxide
employed to provide the stable heating conditions at the
transportation from the upper sea water to the deeper [6,7]. If
predetermined temperature. When the temperature reaches the
iron was added in the HNLC, the phytoplankton growth could
required value, a certain amount of red mud particles was added
be promoted, consuming excessive N, P nutrients and
and the dissolution reaction was started. After the different
accelerating the carbon transporting from the ocean surface to
reaction time intervals, the solution was sampled, filtered and
depth and ultimately reducing the carbon dioxide level of
the clear supernatant solution was sent for analysis of the silicon content by UV visible spectrophotometer (UV2000). Taking
237
into account the different factors to test the extraction efficiency
interface between the particles with the solution. It was shown
of silicon from the red mud, dissolution time intervals from 0 to
that the decreasing of particle size can provide greater diffusion
120 min were examined.
area, which made the extraction of silicon increase.
In order to investigate the dissolution behavior of red mud,
4
we have carried out long time dissolution trials up to 100 hours.
3.5
An SEM instrument was also used to observe the morphology variation of the solid particles.
t
Results and discussion Chemical composition
3 2.5
O 1.5
From the X-ray fluorescence spectrum, the composition of
1
red mud can be determined as shown in Table 1. It was shown
0.5
that the main metal elements include iron, silicon, aluminum
0
and calcium etc, in which these elements mainly existed in the form of their oxide. According to our knowledge, all these
— · — 80rpm —♦— 160rpm
0
20
as silicon from dissolution in the leach liquor would be quite
4
Table 1.Chemical composition of the red mud sample (wt.%) Si0 2
MgO
P2O5
MnO
Γ1Ο2
A1203
Cont
36.01
29.61
21.44
6.49
1.71
1.7
1.31
1.73
120
0.5 g/L, T=20°C)
low.
Fe 2 0 3
100
concentration in the leach liquor (particle size range: 46~74μπι,
predicted that the concentration of the interesting elements such
CaO
60 80 time /min
Fig.l. Effect of stirring speed and time on the silicon
oxides belong to the sparingly soluble phases, so it can be
Cmp
40
3.5 3
Effect of stirring speed Fig.l shows the variation of extracted silicon concentration
E a a
2.5
U
1.5
2 —♦— 39~46μ m
1
in the leach liquor from the red mud with different dissolution times, for the samples under different stirring speeds of 80rpm,
0.5
and 160rpm. It can be seen that the silicon in the red mud could
0
be extracted rapidly into water in the first 60 minutes, then the dissolution rate would gradually decrease, after 120 minutes it
0
—♦— 74~185μ m
20
40
60 80 time / min
100 120
nearly reached an equilibrium state. It was found that increasing
Fig.2. Effect of particle size and time on the silicon
of the stirring speed could promote the extraction of silicon
concentration in the leach liquor (T=20°C, 0.5 g/L, 300 r/min)
from the red mud.
Kinetics analysis
Effect of particle size
In order to study the changes of surface morphology before
Fig.2 shows the variation of extracted silicon concentration
and after long term dissolution, samples were observed by SEM
in the leach liquor from the red mud with dissolution time for
after dissolution for lOOh. As shown in the Fig.3, there seems
the samples with different particle sizes of 74~185μπι and
almost no difference in the microstructure for the samples after
39~46μπι. It was found that the decrease of the particle size
long term dissolution in pure water. Considering that the
could enhance the extraction of silicon from the red mud, i.e.,
extraction of silicon in pure water is quite low as shown in Fig.2,
the smaller the particle size, the higher the extraction rate of
the maximum amount of silicon extraction from the red mud
silicon. With decrease of particle size, the corresponding
reached only 3.57ppm. It can be concluded that the dissolution
particle surface area will increase, leading to a larger contact
238
of silicon from red mud in the pure water was a very slow and is
particles, the variation of the silicon concentration extracted into
a nearly homogeneous dissolution process.
the water may follow the liquid layer diffusion control model as shown in equation (1) which was derived from the classic Noyes-Whitney equation [10], and at t=to, C=0 , Csurface=Cs, the
«
corresponding initial conditions of equation (1) can be derived.
,n<
(1)
FTF ) = T' + c ·
In which, D—diffusion efficiency of silicon through in the bulk liquid layer; C—silicon concentration in the leach liquor at the moment of t; Cs— water solubility of silicon compound under the experimental conditions; A—contact interface area between the red mud particles and water; δ—liquid layer thickness; Cerror—the error of experimental data.
3.5 3 ~
2
·5
"
2
g
1.5 1
Fig.3. Scanning Electron micrograph photos of red mud sample
0.5
before and after dissolution lOOhours, (a) before, (b) after.
0
0
20
40
When red mud particles were put into the pure water, the
60 80 time / min
100
120
particle surface reached saturation solubility rapidly and a liquid
Fig.4. Fitting curves based on the liquid layer diffusion equation
layer will be established around the particle in the dissolution
for different particle sizes of red mud
process. Between the two sides of the liquid layer, a
(T=20°C,300r/min,0.5g/L)
concentration gradient will appear between the solid particle surface and bulk solution, which makes the elements diffuse
Ψ
from the particle surface through the liquid layer into the bulk
3
solution. The experimental results suggest that the dissolution is
2.5
quite fast at the beginning, then the extraction of silicon will gradually decrease and after
— · — 80rpm ♦ 160rpm
•
? 2
120min it nearly reaches
equilibrium. It was found that the extraction of silicon from the red mud is increased with faster stirring. Because in this case,
^y^m
^
the thickness of liquid layer will gradually become thinner and the extraction of silicon will consequently increase. So the
^/+
1 0.5
observed dissolution seems follow a process of dissolution of
n
silicon through a saturated solution layer to the bulk solution, and it can be defined as limited by liquid layer diffusion. Under the different conditions of the dissolution of red mud
239
20
40
60 80 time /min
100
120
1
Fig.5. Fitting curves based on the liquid layer diffusion equation
References
at different stirring speeds (particle size ranged ΐη:46~74μπι,
[1] Liu Q, Environmental Chemistry (Beijing, Chemical Engineering Press, 2004), 127-133.
T=20°C, 0.5 g/L)
[2] Zhang ZB, Liu LS, Advance of Ocean Chemistry (Beijing , Chemical Engineering Press, 2005), 65-67.
Based on the dissolution experimental data of silicon from
[3] Martin JH, Broenkow WW, Fitzwater SE, et al., "Yes, it does:
red mud in pure water at different stirring speed and particles
A reply to the comment by Banse," Limnol. Oceanogr, 35 (3)
sizes, the fitting results were calculated and depicted in Fig.4
(1990), 775-777.
and Fig.5, respectively. All the curves fitted quite well according
[4] Martin JH, "Glacial-interglacial C0 2 change: the iron
to the equation (2). It was found that the whole process of
hypothesis," Paleoceanography, (1990), no.5:l-13.
dissolution was in agreement with the results of our proposed
[5] Leblanc K, Hare CE, Boyd PW et al., "Fe and Zn effects on
assumption, suggesting that the whole dissolution process of
the Si cycle and diatom community structure in two contrasting
silicon was controlled by liquid layer diffusion step. Moreover, this simple diffusion model can be used to describe this kind of
high and low silicate HNLC areas," Deep Sea Res .1, (2005), 52.
dissolution process where the extraction rate of substances are
[6] Martin JH, Coale KH, Johnson KS, et al., "Testing the iron
quite low, there is negligible change of the particle size, and the
hypo thesis in eco- systems of the equatorial Ocean," Nature,
particle surface reaches saturation solubility in a very short
(1994), no. 371: 123-129.
period of time.
[7] Coale KH, Johnson KS, Fitzwater SE, et al., "A massive
Conclusions
phytoplankton bloom induced by an ecosystem scale iron
The dissolution process and kinetics of silicon dissolution
fertilization experiment in the equatorial Pacific Ocean," Nature,
from red mud in pure water was studied. It was found that the
(1996), no.383: 495-501.
silicon extraction was a quite slow and is a nearly homogeneous
[8] Haraguchi K, Taniguchi A, "Effect of simultaneous
dissolution process with the change of particle size negligible
enrichment of dephosphorization steelmaking slag and treated
after the long dissolution times. In this kind of dissolution
municipal sewage on growth of coastal phytoplankton
system, the concentration of silicon would reach a saturation
assemblage," ISU Intern., 8(4) (2003), 430-437.
equilibrium state around the particle surface in a very short time,
[9] Haraguchi K, Suzuki K, Taniguchi A, "Effects of
that is, a saturation layer was quickly established close to the
steelmaking slag addition on growth of marine phytoplankton,"
phase interface, and its thickness was affected by stirring speed
ISU Intern., 43(9)(2003), 1461-1468.
and so on. Based on the dissolution experimental data for silicon
[10] Noyes, A.A., Whitney, W.R., "The rate of solution of
from red mud in pure water, the models are fitted very well
solid substances in their own solutions," J. Am. Chem. Soc,
according to the equation, suggesting that the whole dissolution
19(1897), 930-934.
process of silicon follows the liquid layer diffusion model.
240
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
The effect of cooling rate on theteachabilityof calcium aluminate slags Wang Bo 1 , Sun Hui-lan1, Zhang Xue-zheng1, Bi Shi-wen2 ^ e b e i University of Science and Technology; 70 Yuhua East Rd; Shijiazhuang, Hebei, 050018, China; 2 Northeastern University; No.ll, Lane 3, Wenhua Road; Shenyang, Liaoning, 110004, China Keywords: calcium aluminate slag, cooling rate, self-disintegrating, alumina leaching The cooling rate will not only affect the components and structure of the slag, but also affect the self-disintegration rate of the slag[3]. If the cooling rate is fast, the glass phase will be formed in the between the crystal grains. The glass phase first will prevent the crystal transformation of 2 C a O S i 0 2 (from ß to γ), and second will cover the surface of 2 C a 0 7 A l 2 0 3 . Therefore, the alumina leaching rate of the slag decreases. Besides, a slag which does not self-disintegrate is hard to crush and grind, and this process will also consume a lot of energy.
Abstract A calcium aluminate slag was synthesized by melting a CaC0 3 , Si0 2 and A1 2 0 3 mixture at 1500°C for lh. The effect of cooling rate on alumina leaching and self-disintegration of the calcium aluminate slag were studied. The mechanism was examined using the methods XRD, SEM, and laser particle size analysis. The results indicated that the phase components, alumina leaching property and self-disintegrating property of the calcium aluminate slag were better when cooling rate was lower than 20°Cmin"1. The formation of 2CaOAl 2 0 3 «Si0 2 and 3CaOAl 2 0 3 decreases alumina leaching rate and self-disintegrating rate when the cooling rate was higher than 20°Cmin"1. The conversion rate from ß-2CaOSi0 2 to y-2CaOSi0 2 was still high when cooling rate was between 20°Cmin"1 and SO'Cmin"1. Therefore, the main reason of the decrease in self-disintegrating rate was the formation of2CaOAl 2 0 3 *Si0 2 .
With regard to the optimal cooling rate of calcium aluminate slags Eremin [4] considered that the proper value was S'Cnnin 1 ~ 6T>min_1. But Zhang Jing-dong suggested that if the cooling rate of slag was slower than lO'Omin" 1 when the initial temperature of the slag was between 1400°C and 1250°C, the alumina leaching rate would be higher. Therefore, studies of the effect of cooling rate of calcium aluminate slags have been carried out by previous researchers, based on actual slags, and the results have been useful [5-7]. However, the optimal cooling system was not determined, and the effective mechanism was unclear according to these studies. Therefore, in order to confirm the phase transformation law and the effect of cooling rate on calcium aluminate slags, a study of cooling rate through XRD, SEM and particle size analysis has been carried out and is reported in this paper. In order to exclude the effect of impurities and other factors, the materials which were used to synthesis the slags were analytically pure.
Introduction The cooling rate of calcium aluminate slags is one of the important parameters in the "sintering - blast furnace smelting alumina leaching" process, which is used to utilize iron-bearing bauxite. During the process of blast furnace smelting, iron and a calcium aluminate slag are separated, and then the slag is subjected to a cooling and self-disintegrating procedure[l, 2]. Calcium aluminate slag is in a liquid state when the temperature is 1500°C. The slag completes the transformation from liquid phase to solid phase during the cooling process. There are many pathways as the melt of calcium aluminate slag cools down. The first is that the melt, which has no time to crystallize when the slag is quenched, exists in the form of a glass phase. The second is that the liquid has no time to react with the solid phase which is generated during cooling process, and then crystallizes independently. The third is equilibrium crystallization according to the phase diagram. Therefore, the cooling rate is the key to optimizing the alumina leaching properties and self-disintegration of the slag.
Experimental 2.1. Materials and apparatus Calcium aluminate slag was synthesized by mixing reagents A1 2 0 3 , Si0 2 and CaC0 3 . The A/S (the weight ration of A1 2 0 3 to Si0 2 ) of slag was 1.3, and the C/A (CaO/Al 2 0 3 , molar ratio, except CaO of 2CaOSi0 2 ) of slag was 1.7. The components of materials were shown in Table 1. Table 1 Components of materials Components CaO A12Q3 SiQ 2 W% 57.25 24.16 18.59
Calcium aluminate slag system is polysystem, so the components of solid, which is crystallized from the liquid phase, are different from that of liquid. If the cooling rate is very fast, it will affect the diffusion of liquid phase and then change the solid composition and crystallization path. And besides, it is important that there is a peritectic reaction between gehlenite and liquid phase, and this reaction is very slow. If the cooling rate is fast, the liquid has no time to react with gehlenite, it will form a new solid/liquid mixture and the effect will be different than if gehlenite had formed. That is, the new liquid system will deviate from the best operating region, and does not crystallize the phase which is optimal for leaching.
2.2 Smelting of calcium aluminate slag The samples with different cooling rates were smelted in a vacuum furnace, and the container was a graphite crucible. A Vacuum furnace was used in the experiments largely due to its thin insulating layer, and thus low thermal mass. It could thus realize rapid heating and cooling. The smelting temperature was 1500°C and the holding time was one hour. The sample was taken out of the vacuum furnace at
241
400 °C. Cooling rates of the calcium aluminate slag between 2°C min"1 and 40 °C min"1 were studied in this paper.
7000
* - Y -2GO S02
2.3 Leaching of calcium aluminate slag
• - 12QO 7AL203 v - GO
£
O Ü
4000
3000
1000
ujy^ 20
40
60
80
2Theta
2.4 Analysis
Fig. 2 XRD spectra of slag with cooling rate=10°C -min"1
The contents of A1203 and Si0 2 of samples and filtrate were analyzed by a chemical method. The Phase components of calcium aluminate slag were identified by X-ray diffraction analysis (PANalytical PW3040/60). A Malvern laser particle analyzer was used to analyze the particle size of the slag.
7000 » - Y - 2CaO Si 0 2 6000
* - CaO Al O
4000
£
3.1 Phase transformation law of slag with different cooling rate
c 8
3000
In order to study the phase transformation law of calcium aluminate slags with different cooling rates, XRD analysis was used in the experiments. The results are shown in Fig. 1 to Fig. 4.
2000
^UJIM^
7000 - Y -2QO Si02 6000
• - 12CaO 7AI 2C>
5000
Results and discussion
2Theta
- 12GO 7ALj03 - QO ALP3
Fig. 3 XRD spectra of slag with cooling rate=20°C min"1
5000
w
Αψ3
5000
The sodium aluminate solution obtained from slag leaching should then be treated by a carbonization precipitation process, and the circulating mother liquor used to leach new calcium aluminate slag. Therefore, the feasible conditions for alumina digestion were: leaching temperature was 75 °C, leaching time was 2h, L/S ratio was 4.5(by weight), caustic alkali concentration was 7g/L and sodium carbonate concentration was 120 g/L. The leaching experiments were carried out in a magnetically stirred, constant temperature water bath. After leaching and dry filtration, the filtrate was used to analyze the composition of the solution, and the filter residue was washed and dried to preserve for analysis.
4000 - γ -2CaO
Si0 2
- 12CaO 7AI 2 0 3 - CaO Al 2 0 3 - 2CaO ΑΙ 2 0 3 · S 0 2
20
IwJJi^^ 40
60
5
3000h
80
2Tteta Fig.l XRD spectra of slag with cooling rate=2°C min"1
^HwllW 2Theta
Fig. 4 XRD spectra of slag with cooling rate>40°C min"1 Fig.l shows that when the cooling rate was 2°C-min"1, the calcium aluminate slag had a proper phase component. The main phases were 12Ca07Al 2 0 3 (alumina is easy to be leached out) and γ-
242
2CaOSi0 2 (slag could self-disintegrate because of its phase transformation). Little CaOAl 2 0 3 (the alumina teachability is lower than 12Ca07Al203) was formed in the slag. When the cooling rate increased to lO'Cmin"1 and 20°Cmin"1 the phase components of slag were similar to the slag whose cooling rate was 2°C-min"1(Fig.2 and Fig.3). 2CaOAl203»Si02 which was difficult to leach was not found under the three conditions. But 2CaOAl203»Si02 could be found in slag when cooling rate was above 40°C-min ^ . 4 ) . ß-2CaOSi02(2.78 , 2.61 , 2.19Â) was not found in slag either when cooling rate was not above 20°Cmin"1. It could be found when cooling rate was above 40°C-min"1, but its amount was limited. Therefore, its characteristic peaks were very low and were not marked in Fig.4.
Therefore, a certain amount of 3CaOAl 2 0 3 and 2CaOAl203»Si02 would appear when the cooling rate was very rapid. This would affect the alumina leaching property and selfdisintegrating property of calcium aluminate slag.
According to the former results when the cooling rate was under 20°C-min"1 the calcium aluminate slag would have an ideal phase composition. If the cooling rate was accelerated continuously, 2CaOAl203»Si02 would appear and effect the phase of slag. Trace quantities of ß-2CaOSi02 was formed when cooling rate was over 40°Cmin"1. And this indicated that 2CaOSi0 2 had a rapid crystal transformation speed, so cooling rate had a little effect on crystal transformation of 2CaOSi0 2 under the conditions studied in this paper. These results accorded with Zhang Xiong [8], who deemed that 2CaOSi0 2 had a higher crystal transformation ratio when cooling rate was between 2°Cmin"1 and 500 °C min"1. He further noted that the transformation rates of y-2CaOSi02 were 100%, 95% and 90% when the conditions were slow cooling (about 2°Cmin"1), natural cooling (about 60 °C min"1) and wind cooling (about 500 °C min"1).
The slag with different cooling rate was well-mixed in order to analyze the granularity of slag. The particle size results analyzed by Malvern 2000 are shown in Fig.6.
3.2 Effect of cooling rate on self-disintegrating property of slag The self-disintegration of calcium aluminate slag can reduce the energy consumption during production, and it is a very important characteristic of calcium aluminate slags. The percent content of granularity which is lower than 74um in samples is defined as the self-disintegrating rate.
In addition, when the cooling rate of calcium aluminate slag was over lO'C-min"1 trace amount of particles (particle size bigger than 1mm) which didn't self-disintegrate completely appeared. XRD analysis was carried out on these grinded particles, and the result is shown in Fig.5.
15
• - 12CaQ
7Ai,0 3
* -CaO
6000
4000
2000
n
'.
30
Fig.6 indicates that the self-disintegrating rate of slag decreased with acceleration of cooling rate. However, the self-disintegrating rate of slag was still higher than 90% when cooling rate was not higher than 20°Cmin"1, this suggested that under this condition the effect of cooling rate on pulverization was little. The results could be verified by XRD (Fig.l to Fig.3). ß-2CaOSi02 and 2CaOAl203»Si02 were not found in slag, so the crystal transformation of y-2CaOSi02 was complete and the particle size was relative stability.
* - 3CàO AL,03 8000
25
Fig.6 Self-disintegrating rate of slag with different cooling rate
12000
10000
20
Cooling rate / eC · min"1
•
ILjJL L L ^ .
If cooling rate increased continually, the self-disintegrating property of slag decreased obviously. When cooling rate was 40°Cmin"1 the self-disintegrating rate was only about 80%. There were two points for the decrease according to XRD analysis. The first was the formation of 2CaOAl203#Si02 (Fig.4) which decreased the content of 2CaOSi0 2 of slag, thus the selfdisintegrating rate decreased. The second was the existence of the trace of ß-2CaOSi02 which decreased the crystal transformation rateof2CaOSi0 2 .
^^LJ
2Theta
Fig.5 XRD spectra of particles without self-disintegrating Fig.5 shows that the major phase of particles which didn't selfdisintegrate was 3CaOAl203, and the minor phases were f-CaO and 12Ca07Al203.
Further, according to report [8] the crystal transformation rate was still higher than 95% even under natural cooling condition (about 60 "C-min"1). So the effect of the existence of ß-2CaOSi02 on pulverization of slag was little.
243
and 3CaOAl203 which were difficult to leach would be formed if the cooling rate increased continually. And because of that the self-disintegrating property and alumina leaching property of slag would decrease obviously.
Therefore, the major reason of the decrease of self-disintegrating rate was the formation of 2CaOAl203#Si02; the minor reason was the existence of ß-2CaOSi02.
(2) The crystal transformation rate was still high when cooling rate was between 20°C-min"1 and SiTC-min"1. Therefore, the main reason of the decrease of self-disintegrating rate was the formation of 2CaOAl203»Si02.
4,3.4 Effect of cooling rate on alumina leaching property of slag The alumina leaching experiment of slag was carried out in order to study the effect of cooling rate on the alumina leaching properties. The leaching conditions and process were as in section 2.3, and the results are shown in Fig.7. 88
84
2
The authors greatly acknowledge the financial support of the National Nature Science Foundation of China (Project No: 50674028), and the foundation of Hebei University of Science and Technology (Project No: XL200921). The authors express their profound gratitude to the editors and reviewers of TMS.
■
86
* 82 "53
Acknowledgements
-
References
■V
80
«50
Ì78
ä
-
[1]
J.Grzymek, A.Derdacka and Z.Konik, "Method for obtaining aluminum oxide," U.S Patent : 4149898 , 19782-21.
[2]
ZHANG Jing-dong, LI Yin-tai, BI Shi-wen and YANG Yihong, "Research on integrated utilization of high-ferrum bauxite in Guigang Guangxi," Light Metals, (8)(1992), 1618.
[3]
BI Shi-wen, YANG Yi-hong, LI Yin-tai, ZHANG Jingdong and DUAN Zhen-ying, "Study of alumina leaching from calcium aluminate slag," Light Metals, (6)(1992), 1015. N.I.Eremin, "Investgations on the complex processing of bauxites," (Bauxite-Alumina-Aluminum, Symposium of ICSOBA . Budapest: Research Institute for Non-Ferrous Metals, 1971), 329-335.
3 76 ■\.
74 72
-
7n
l_
0
5
i
i
1
1
1
1
1
10
15
20
25
30
35
40
45
Cooling rate/ °C · rri n 1
Fig.7 Leaching property of slag with different cooling rate The alumina leaching rate of slag showed a decreasing tendency with the increase of cooling rate (Fig.7). When the cooling rate was in slow cooling stage (<5"C-min-1) the slag had a better phase components and particle size, so the alumina leaching rate was higher (>85%).
[4]
When the cooling rate was in medium cooling stage (ö'C-min"1 ~ 20°C-min"1) the alumina leaching rate decreased but was still above 80%. That was because the main phases were still 12Ca07Al203 and y-2CaOSi02. The decrease of leaching rate was because of the particles which didn't self-disintegrate. The formation of 3CaOAl203 and the existence of f-CaO which didn't react with A1203 or Si0 2 in these particles decreased the content of 12Ca07Al203.
[5]
J.Grzymek, "Complex Production of aluminium oxide and Iron from laterite raw materials applying the calcium aluminates polymorphism," TMS Annual Meeting and Exhibition, Light Metals, 1985, 87-99.
[6]
WANG Bo, YU Hai-yan, SUN Hui-lan and BI Shi-wen, "Effect of material ratio on leaching and self-disintegrating property of calcium aluminate slag," Journal of Northeastern University: Natural Science, 29(11)(2008), 1593-1596.
[7]
D.J.Connor, Aluminium extraction from non bauxitic materials, (Sydney: Aluminium-Verlag Gmbh, 1988), 230250.
[8]
Zhang Xiong, "Quantative control of C2S transformation", Journal of The Chinese Ceramic Society, 23(6)(1995), 680-684.
When cooling rate was in fast cooling stage (20°Cmin1 ~ öCTC-min"1) the alumina leaching rate decreased obviously. The main reason was the formation of 2CaOAl203#Si02 and 3CaOAl203; the minor reason was the decrease of the selfdisintegrating rate. The two reasons interacted with each other, and decreased the alumina leaching property obviously under this condition. Conclusions (1) The phase component, self-disintegrating property and alumina leaching property of calcium aluminate slag was good when cooling rate was slower than 20°Cmin"1. 2CaOAl203»Si02
244
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
PREPARING POLYMERIZED ALUMINUM-IRON CHLORIDE WITH RED MUD LUGuilin1'2, YUHaiyan1, BIShiwen1 School of Materials and Metallurgy; Northeastern University ,Shenyang, Liaoning, 110004, China 2 School of sciences; Shenyang University of Technology, Shenyang, Liaoning, 110870, China Keywords: Red mud, Flocculant, Polymerized aluminum-ferric chloride, Basicity, Hydrolyzing
Abstract A new method is proposed for the preparation of polyaluminium ferric chloride (PAFC) with red mud. The dependence of pH of the PAFC on basicity, is discussed and the difference in the products of the process of red mud hydrolysis from products in the process of sodium hydroxide hydrolysis is studied. The results show that basicity increases with increasing pH, the preferred pH value is 3.5, and basicity is up to 85.55%. Comparing the appearance of PAFC with that of crystalline aluminium chloride, polyaluminium chloride by SEM, can indicate that the distribution of materials is not uniform on the surface of PAFC. Energy Dispersive analysis indicates that the content of aluminum, iron, and oxygen accounts for 60.73% in the white areas, 43.95% in the black areas in the process of red mud hydrolysis; whereas the content of aluminum, iron, and elements in the white areas is 49.74%, 23.45% in black areas in the product from sodium hydroxide hydrolysis.
Experimental Materials and Equipment Waste red mud, hydrochloric acid and sodium hydroxide were used as starting materials. The waste red mud was obtained from Shandong aluminum factory in China. Compositions and content of red mud are shown in Table 1. The equipment is shown in Table 2. Table 1 Compositions and content of red mud Composition
Alumina
Iron oxide
Silica
Titania
Others
Content(%)
20.54
26.73
25.21
1.85
25.67
Table 2 names and types of equipment
Introduction Red mud is a waste byproduct generated during the production of aluminafrombauxite. Amounts of red mud vary with the different quality and processing of ores containing aluminum. If bauxite is used as starting raw materials, the product of per ton alumina will result in approximately 1 ton of dry red mud, and will produce about 7 tons dry red mud if nephelite is used as the starting raw material. Today, worldwide production of red mud is estimated to be about 150 million tons each year, and production of red mud is more than 4 million tons in Chinail]. So red mud occupies largeland areas, and pollutes the environment12"51. In order to solve the above problems, various means of utilization of red mud have been suggested by worldwide metallurgy workers, for example, red mud was used as catalyst16"91, plastics padding110"111, cement[12], production of castings1131, glass-ceramics1 415] and so on. Red mud is predominantly a mixture of oxides of iron, aluminium and titanium with relatively smaller amounts of silica and magnesia. Utilizing the components of aluminium and iron in red mud to prepare flocculant-polymerized aluminum-ferric chloride (PAFC) has been reported[161. The processes is that red mud reacts with Hydrochloric acid, the mixture is filtered, the filtrate is neutralized and hydrolyzed with sodium hydroxide. This process will not only consume a lot of sodium hydroxide, but also the utilization rate of hydrochloric acid is very low and content of byproducts (mostly NaCl) is quite high. Because red mud contains iron oxide, alumina, sodium aluminate and other alkaline substances which can be used to react with aluminum chloride, iron chloride and free hydrochloric acid, replacing sodium hydroxide with red mud to neutralize and hydrolyze the filtrate will save a lot of sodium hydroxide, increase the utilization rate of hydrochloric acid and purity of the products, and cut the cost of production.
Names
types
scanning electron microscope
Hitachi S-3400N
Synthetic Principles of PAFC Alumina(Al203) and Iron oxide(Fe203) in waste red mud react with hydrochloric acid to form [A1(H20)6]C13, and [Fe(H20)6]Cl3. Fe203+6HC1+9H20 ►2[Fe(H20)6]Cl3 (1) A1203+6HC1+9H20 ► 2[A1(H20)6]C13 (2) The addition of red mud (or alkali) to the filtrate will increase the pH value of system, when the pH value reachs some limit, [A1(H20)6]3+ and [Fe(H20)6]3+ will hydrolyze to form H+and a series of coordinate ions. The equations are as follow: [A1(H20)6]3+ - 2*L~ _
[Al(OH)(H20)5]2+ + H+
2+
[Al(OH)(H 2 0) 5 ] -^~- [Al(OH)2(H20)4]++H+
(3) (4)
[Al(OH)2(H20)4]+ - 2ST- [Al(OH)3(H20)3] + H+
(5)
[Fe(H 2 0) 6 ] 3 + -^~- [Fe(OH)(H20)5]2++H+ [Fe(OH)(H20)5]2+-™\- [Fe(OH)2(H20)4]+ + H+ [Fe(OH)2(H20)4]+ -22:- [Fe(OH)3(H20)3] + H+
(6) (7) (8)
With the increase of concentration of OH", coordinated water hydrolyzes and substances formed will polymerize. Bridge will form between OH" and OH", [A1(H20)6]C13 and [Fe(H20)6]Cl3 gradually are converted into dipolymer, tripolymer, finally high molecular weight species PAFC [Al2(OH)nCl6-n]m, [Fe2(OH)nCl6-n]m. Synthesis Processes of PAFC Hydrochloric acid was added to waste red mud. It reacted with waste red mud for a period of time, then the mixture of liquid and
245
solid was filtered and washed, filter liquor and filter cake were obtained. The filter cake then reacted with hydrochloric acid, and was filtered and washed once again. Both filter liquors were mixed, and neutralized and hydrolyzed with either waste red mud or sodium hydroxide. If waste red mud is used to neutralize and hydrolyze the filter liquors, the mixture of liquid and solid will be filtered and filter liquor obtained will be concentrated and dried, and PAFC will be obtained; If sodium hydroxide is used to neutralize and hydrolyze the filter liquors, the solution obtained will be concentrated and dried, and PAFC will be obtained. The production process of red mud neutralization and hydrolysis are shown in figure 1.
«—H
red mud
CD 3—content of effective composition (%) ( calculated as it was AI2O3) 0.01699—Mass (g) of A1 2 0 3 equivalent with 1.0 ml NaOH standard solution [C (NaOH)=1.000mol/L] Results and Discussion Effect of P H Value of Solution on Basicity of PAFC Different volumes of sodium hydroxide solution were added to the filter liquors obtained when red mud reacted with hydrochloric acid. The pH value of the filter liquors was adjusted to 1.5, 2.0, 2.5, 3.0, 3.5 respectively. The alumina and iron oxide content was determined for each sample, and then basicity determined. The results of the experiments are shown in Table 3. Table 3 Effect of pH on basicity of production
hydrochloric acid
i—^ 1 filter and washing iv filter liquor
1r
+—\
filter cake
hydrochloric acid
filter arid was]ling
Ψ
^r
_2 r filter cake
—► red mud
filter liquor
1 filter
x— 1
1f filter cake
i
r
filter liquor
drying
1 products
Fig. 1 Production process of neutralized red mud Determining the Basicity of PAFC One third of the mol ratio of OH" and Al3+ in a sample is called the basicity. Generally, the effects of flocculation increase with the increase of basicity. Firstly, potassium dichromate titration was used to determine content of iron oxide ω i, and EDTA titration was used to determine content of alumina ω 2 . Alumina was used to calculate effective content ω 3, ω 3 = ω 2 +0.6384ω i, 0.6384 is a factor in which iron oxide was converted into alumina. Then quantitative hydrochloric acid solution was added to the sample of PAFC, sodium fluoride was used to shelter Al3+, sodium hydroxide standard solution was used to titrate above solution. Use equivalent distilled water(no carbon oxide ) and hydrochloric acid as blank sample and make the same experiment. The formula of basicity W4 is as follow:
(Vo- V)xCx0.01699 mco^
X 100%
basicity (%)
1.5
5.82
2.0
72.27
2.5
76.35
3.0
80.89
3.5
85.55
When pH value of solution is between 1.5 and 3.5, the basicity of products increases with the increase in pH of the solution, but for pH value >3.5, the solution will change into a gel. If the pH value of the solution is low, most of the base will be used to neutralize free hydrochloric acid, and the hydrolyzing degree of A1C13 and FeClß should be low, as the contents of hydroxyls are low. Thus basicity of the product is low when the pH value of solution is less than 1.5. When the pH value of solution is more than 2.0, then some of base reacts with A1C13 and FeCl3, and some of chloride in AICI3 and FeCl3 is changed into hydroxyls. So basicity of the products increases with the increase of alkalinity. When the pH of the solution is more than 3.5, A1C13 and FeCl3 will be changed to deposits of Al(OH)3 and Fe(OH)3.
w
W4=
pH
Comparison of PAFC with Aluminum Chloride and Polymeric Aluminum Chloride In order to determine whether the synthetic products are the target PAFC or not and, simultaneously, determine the composition of the products, crystalline aluminum chloride and polymeric aluminum chloride were chosen to compare with the synthetic products. Their appearance and composition were investigated using scanning electron microscope and energy dispersive analysis. The experimental results are shown in figure 2, figure 3 and figure 4. Conditions for the preparation of the synthetic products were that the liquid-solid ratio of hydrochloric acid to red mud is 4:1, the concentration of hydrochloric acid is 6 mol/L, reaction time is 60 minutes, reaction temperature is about 109°C, the extraction is the second extracting (see fig. 1, pH value of the filter liquors was adjusted to 3.5 with sodium hydroxide solution. The synthetic product- PAFC was obtained after the solution was dried.
(9)
Vo—volume (ml)of sodium hydroxide standard solution consumed by blank sample V— volume (ml)of sodium hydroxide standard solution consumed by sample of PAFC C—concentration of sodium hydroxide standard solution(mol/L) m— mass of sample PAFC(g)
246
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*
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1%
4
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►
#* *, * ^ * >
/^^^^B m 1
* m
I
:
,, ■·■
S3400n 15 OkV 9 7mm x2 00k SE 6 /18/20C 8 12:3B'
(..
;
'
m
20.0um
(a) SEM of PAFC (a) SEM of crystal aluminium chloride «.s -
39.62 32.42 27.96
OJT AIE CX
-
S.1
OK NaK AUL OK CaK FeK
Al
55.44 26.90 17.66
19.56 10.31 11.74 38.87 01.0S 18.44
34.34 12.60 12.22 30 80 00.76 09.28
0
a
KG*
3.4
-
1.1
-
0.«
-
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j
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~
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(b) Energy spectrum of crystal aluminium chloride Fig.2 Crystal aluminium chloride
(b) Energy spectrum of white areas of PAFC OK NaK ΛΙΚ OK CaK FeK
11.99 25.06 03.96 51.04 0046 07.50
20981 30.52 04.11 40.31 0032 03.76
(c) Energy spectrum of black areas of PAFC Fig.4. PAFC of pH=3.5 with sodium hydroxide neutralisation. (a) SEM of polyaluminium chloride
According to figure 2(a), figure 3(a) and figure 4(a), the appearance of the synthetic product is much different from that of crystal aluminum chloride and similar to that of polymeric aluminum chloride. The surface of the crystalline aluminum chloride is smooth, and the distribution of materials is uniform. Distribution of materials is not uniform on the surface of PAC and PAFC. The white and black areas occur on their surface. According to the energy spectrum of white areas in figure 2(b), figure 3(b )and figure 4(b), crystalline aluminum chloride contains only Al, Cl, and O; polymeric aluminum chloride contains traces of Fe, and Na in addition to Al, Cl, and O; PAFC contains high levels of Fe, Na in addition to Al, Cl, O. The appearances and compositions of white areas show that the white areas should be PAFC. Comparing compositions of white areas in figure 4(b) with that of black areas in figure 4(c), the content of Fe, and Al in the white areas are more than that of Fe and Al in the black particles. Fe and Al in
(b) Energy spectrum of polyaluminium chloride Fig.3 Polyaluminium chloride
247
According to figure 5(a), the distribution of materials of PAFC obtained in process of red mud neutralizing is more uniform than that in process of sodium hydroxide solution in the figure 4(a). The white areas of materials are quite large in figure 5(a). Comparing figure 5(b) with figure 4(b), content of Al, Fe and O in PAFC obtained in process of red mud neutralizing is higher than that in process of sodium hydroxide neutralizing. The content of Al, Fe and O is 60.73% in the former, and 49.74% in the latter. The content of Na is 0.95% in the former, and 10.31% in the latter. The results indicate that the content of PAFC obtained in process of red mud neutralization is higher than that in the process of sodium hydroxide neutralization in the chosen white areas. The content of by-product NaCl is lower in the former than in the latter in the white areas. Comparing figure 5(c) with figure 4(c), the content of Al, Fe and O in PAFC obtained in the process of red mud neutralization is higher than that in process of sodium hydroxide neutralization in the black areas.The content of Al, Fe and O is 43.95% in the former, and 23.45% in the latter. The content of Na is 3.78% in the area analysed in figure 5(c), and 25.06% in figure 4(c), which also indicates that content of PAFC obtained in process of red mud neutralization is higher than that in process of sodium hydroxide neutralization in the chosen black areas, and the content of by-product NaCl is lower in the former than in the latter in the darker areas.
the white particles account for 30.18%, but is only 11.46% in black particles. The effect of Neutralizing approach and Hydrolysis on the Appearance and Composition of PAFC Red mud was leached with Hydrochloric acid, the filter liquor was neutralized and hydrolyzed to pH=3.5 with red mud. The results of the experiment are shown infigure5.
III,. "Ill &:, 'sa%.
::
* ;ii :
:
:"■'
;
'
Π5 ] (a) SEM of PAFC
The composition of the material in figure 5(c) can indicate whether compositions of the black areas in figure 5(a) are PAFC or not. The proportion of Fe and Al is 29.6%. Subtracting the proportion of the atoms of Na from that of atoms of Cl gives 38.28%, which indicates each Al (or Fe ) can bond with 1.3 Cl atoms. The residual atoms (Cl ) were hydrolyzed into hydroxyl (-OH). The analysis above shows that compositions of the black areas infigure5(c) are PAFC, in addition to a little by-product NaCl.
6.2 -
I2S
3.f
OK NaK AIK CIK FeK
-
C»
28.68 00.95 Î5.67 38.31 16.38
47.32 01.09 15.33 28.52 07.74
KCi« 2.5 -
1.2 -
Al
II IH 1.00
Conclusions
I
(1) Red mud can be used as a starting material to prepare PAFC. (2) The basicity of PAFC increases with increasing pH, and the preferred pH value is 3.5. (3) Comparing the appearance of PAFC with that of crystalline aluminium chloride and polyaluminium chloride by SEM can indicate that the distribution of materials is not uniform on the surface of PAFC. The Energy Dispersive spectrum can be used to analyze the composition of PAFC. (4) The Energy spectrum indicates that the content of PAFC obtained in the process of red mud neutralization is higher than that in the process of sodium hydroxide neutralization; the content of by-product NaCl is lower in the former than in the latter. Process of red mud neutralizing is better than process of sodium hydroxide neutralizing.
Fe 2.00
3.00
4.00
S.00
6.00
7.00
8.00
9.00
10.00
11.00
12.0«
(b) Energy spectrum of white areas of PAFC
4.2 H
l*g?g?»i] OK 3233 03.78 NaK 23.35 AIK 52.27 CIK FeK 082?
22.52 04.SÖ 25.2S 43.08 04.32
KCf*
References 1. JIANG Yijiao, NING Ping, "An overview of comprehensive utilization of red mud from aluminum production", Environmental Science and Technology, (2003), 26(1), 40—42. 2. R. A. A. Blackman, K. W. Wilson "Effects of red mud on marine animals", Marine Pollution Bulletin, (1973) 4(11) 169-171. 3. S. Baseden, D. Grey, "Environmental study of the disposal of red mud waste", Marine Pollution Bulletin, (1976), 7(1), 4-7.
(c) Energy spectrum of black areas of PAFC Fig.5 PAFC with pH value 3.5 in process of red mud neutralizing
248
4. Shaobin Wang, H.M. Ang, M.O. Tadé, "Novel applications of red mud as coagulant, adsorbent and catalyst for environmentally benign processes",_Chemosphere, (2008), 72(11), 16211635. 5. Ekrem Kalkan, "Utilization of red mud as a stabilization material for the preparation of clay liners", Engineering Geology, (2006), 87( 3-4), 220-229. 6. Cakici A I, Yanik J, Uçar S, Karayildirim T, Anil H, "Utilization of red mud as catalyst in conversion of waste oil and waste plastics to fuel", Journal of Material Cycles and Waste Management, (2004), 6(1), 20—26. 7. Juan J. Llano, Roberto Rosai, Herminio S astre, Fernando V. Diez, "Catalytic hydrogénation of anthracene oil with red mud", Fuel, (1994) 73(5), 688-694. 8. Jale Yanik, Md. Azhar Uddin, Kazuo Ikeuchi, Yusaku Sakata, "The catalytic effect of Red Mud on the degradation of poly (vinyl chloride) containing polymer mixture into fuel oil", Polymer Degradation and Stability, (2001), 73( 2), 335-346. 9. Jorge Alvarez, Salvador Ordonez, Roberto Rosai, Herminio Sastre, Fernando V. Diez, "A new method for enhancing the performance of red mud as a hydrogénation catalyst", Applied Catalysis A: General, (1999), 180( 1-2), 399-409. 10. Singh B, Gupta M. "Surface treatment of red mud and its influence on the properties of particulate-filled polyester composites", Bulletin of Materials Science, 1995, 18 (5) : 603-621. 11. Soo-Jin Park, Dong-Il Seo, ChangwoonNah, "Effect of Acidic Surface Treatment of Red Mud on Mechanical Interfacial Properties of Epoxy/Red Mud Nanocomposites",_Journal of Colloid and Interface Science, (2002), 251(1), 225-229. 12. Maneesh Singh, S. N. Upadhayay, P. M. Prasad, "Preparation of special cementsfromred mud", Waste Management, (1996), 16(8), 665-670. 13. Nilza Justiz-Smith, Vernon E. Buchanan, Gossett Olive. "The potential application of red mud in the production of castings", Materials Science and Engineering A, (2006) 420(1-2), 250253. 14. Jiakuan Yang, Dudu Zhang, Jian Hou, Baoping He, Bo Xiao, "Preparation of glass-ceramics from red mud in the aluminium industries", Ceramics International, (2008), 34(1), 125-130. 15. Vincenzo M. Sglavo, Stefano Maurina, Alexia Conci, Antonio Salviati, Giovanni Carturan, Giorgio Cocco, "Bauxite 'red mud' in the ceramic industry. Part 2: production of clay-based ceramics", Journal of the European Ceramic Society, (2000), 20(3), 245-252. 16. ZHANG Lingfen, HUANG Jianfang, MO Yazh, "A study on the preparation of polymerized aluminum ferrum chloride from red mud"._Water Purification Technology 1998,(03):2~4.
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
Adsorption of Polyethylene Glycol at the Interface of Dicalcium Silicate - Sodium aluminate solution Haiyan YU1, Xiaolin PAN1, Zhongke LU2, Tingting DING1 1
School of Materials Metallurgy, Northeastern University; Shenyang, Liaoning, 110004, China 2
Shandong Branch, Chalco; Zibo, Shandong, 255065, China
Keywords: Polyethylene glycol; ß-CaOSi02; Sodium aluminate solution; Adsorption thermodynamics [3-5]. However, the basic theories, such as the
Abstract
models of adsorption, the laws at the interface of sodium Surfactants, such as polyethylene glycol (PEG), are widely used
aluminate solution and ß-CaOSi02, and the effect of adsorption
to inhibit the secondary reaction of dicalcium silicate (ß-
on the solid-liquid interface were seldom studied. Consequently,
CaOSi02) in sodium aluminate solution. The adsorption isotherm
according to the development of Gibbs adsorption isothermal
of PEG on ß-CaOSi02 in sodium aluminate solution at 80 °C
formula on the solid-liquid interface [6-10], the adsorption laws of
was researched in this paper. It was found that the type of surface
PEG in clinker-sodium aluminate solution systems and the
adsorption of PEG is saturated adsorption, and the adsorption
mechanisms of PEG inhibiting the secondary reaction were first
behavior belongs to "S"-type, according with the multi-molecular
studied in this article.
layer adsorption model of the Freundlich Equation. The surface coverage of PEG close to saturated adsorption derived from the
Experimental procedures
2
specific surface area of ß-CaOSi02 is 0.723 mol/μπι . The relation curves between the surface pressure and the molecular
Dicalcium silicate was prepared by pure calcium oxide and silica.
area of adsorption film were obtained according to the adsorption
Calcium oxide and silica were first mixed with the molar ratio of
results at the solid-liquid interface by Gibbs formula. The
2 to 1, grinded 2 hours in a mortar, and then calcined at 1350 °C
variation of interfacial energy caused by adsorption, as well as the
followed by quenching in the air.
relationship between the relation curves and the type of adsorption
Sodium aluminate solution with caustic ratio (i.e. molar ratio of
was also discussed.
Na20 to A1203) of 1.30 was prepared by the red mud lotion from the Shanxi Branch of the China Aluminum Company, industrial sodium hydroxide and aluminum hydroxide, and analytically pure
Introduction
sodium carbonate. The concentrations of Na20 and Na2C03 of sodium aluminate solution were 36 g/L and 15 g/L, respectively.
In the sintering process of alumina production, there is about 30%
PEG was added to sodium aluminate solution with the
dicalcium silicate (ß-CaOSi02) in the clinker. During the
concentration of ß-CaOSi02 75 g/L. The concentrations of PEG
leaching of clinker, as ß-CaOSi02 reacts with sodium aluminate
in sodium aluminate solution were 0, 80 mg/L, 120 mg/L, 160
solution, both A1203 and Na20 re-enter into the red mud, resulting
mg/L, 200 mg/L, 240mg/L, respectively. The mixed solutions
in the loss of A1203 and Na20 [1]. It was found [2] that additions
were stirred magnetically (600 r/min) for 1 hours in 80 °C water
of surfactants, such as polyethylene glycol (PEG), can inhibit the
bath followed by centrifugal separation. 0.5 ml supernatant fluid
secondary reaction of ß-CaOSi02 with sodium aluminate
was taken out to determine the content of PEG by the method of
solution, which increases the leaching rate of A1203.
spectrophotometry using a 722S Spectrophotometer.
Much of work was done on the additives for inhibiting the
The adsorbed amount per unit weight of adsorbent at equilibrium,
secondary reaction, which was mainly concerned on the selection
Qe, can be expressed as follows:
and application of additives as well as the adsorption kinetics and
251
(1)
Qe=(C0-Ce)/W0
The value
of jirdin C with
different concentrations can be
obtained by graphic integration, and then the value of π with
where C 0 and Ce is the initial concentration and the residual
different adsorbed amounts can be calculated. At very low
concentration of additives (mg/L) in the solution, respectively,
concentration, the relationship between Γ and In c is considered as
and W 0 is the concentration of adsorbent (g/L) in the solution.
a line. Accordingly, the area of each adsorbed molecule (σ) at different
Gibbs Adsorption Isotheral Formula
amounts of Γ is given by the Avogadro constant (NA) as: According to the interface Gibbs formula in the ideal solution
σ = 1/(ΠνΑ)
[11], the basic form of surface tension, γ, under constant
(6)
temperature and pressure is given by: Results and Discussion (2) i=1 A .
Adsorption isotherm
i=1
The adsorption isotherm of PEG on ß-CaOSi02 in sodium
where ^ is the surface molar quantity of i component, As is the
aluminate solution is shown in Fig. 1. The type of adsorption
surface area, μϊ is the chemical potential of i component, T[ is the
isotherm belongs to "S"-type, which indicates that the adsorption
excess amount of i component per unit surface (i.e. the surface
of PEG on 2CaOSi0 2 accords with the Freundlich multi-layer
adsorbed amount or surface concentration (mol/m2)).
model, and follows the Freundlich Equation as expressed in Eq.
As the interfacial tension numerically equals to the interfacial
(7):
pressure (π), the relationship is d π = -d γ. lnß,=lnk +
Thus, άπ = Υ^Γ,RT dinCi
InC.
(7)
(3) where k and n are constants.
For binary systems, if Group 1 is identified as the major component, Γι (1) =0. The adsorbed amount of Group 2 is Γ2(1). Then, Eq. (3) can be written as:
-ày=RTYf
dlnc2
(4)
For the adsorption in the dilute solution, it is reasonable to consider the surface excess amount Γ2(1) as the actually measured amount (apparent adsorbed amount), so Γ 2 (1) is replaced by Γ and c 2 is replaced by c in Eq. (3). For the solid adsorption from liquid, lô
Eq. (3) is integrated on both sides:
π=
γ80-γ$ι=1ΐτ\ΐτά\ηα
2Ô 30 40 50 Equilibmim concentration (nig/L)
Fig. 1 Adsorption isotherm of PEG on ß-CaOSi0 2 at 80 °C
(5)
where γ8ο and YSL is the solid-liquid interface free energy before
The relationship between the residual concentration (Ce) and the
and after adsorbing the solute, respectively.
value of Q/Qe is shown in Fig. 2. The curve can be fitted by the Freundlich Equation as Eq. (8):
252
l n ß , =1.242 In Ce -4.184
interface [14, 15]. Therefore, the adsorption of polyethylene
(8)
glycol is mainly physical adsorption, which is also confirmed by IR detection. In addition, the curve in Fig. 3 changes at the point of 1.09 mg/g, which is approximately considered as the adsorption 1.0
capacity of the covered monolayer.
♦ JS
0.5
3 c
1
R 2 =0.963
y
0.0
r
H
♦
-
yS
-0.5 -1.0 .1 R
2.0
S
1
1
2.5
3.0
i
3.5
·
4.0
4.5
InCe
Fig. 2 Freundlich isotherm of PEG on ß-CaOSi02 0
As shown in Fig. 2, the experimental data can be well simulated
5 σ öuii2/raole) 10
15
Fig. 3 Relationship between the interfacial pressure and
by the adsorption isotherm. The adsorption of PEG at the interface
molecular area
of dicalcium silicate-sodium aluminate solution belongs to "S"type, having the characteristic of multi-layer adsorption.
When the adsorption capacity is 2.45mg/g, the surface coverage is
Adsorption film on B-CaO-SiO?
calculated to be 2.24, which is larger than 1, demonstrating that
The π-σ curve of the interfacial film on ß-CaOSi02 can be
the adsorption of PEG is a multi-molecular layer. Meanwhile, the
obtained from the adsorption results by using Eq. (5) and (6). As
single molecular area is 1.383 μηι2/ιηο1, and the corresponding
shown in Fig. 3, the shape of π-σ curve is similar to the adsorption
coverage is about 0.723 mol/ μπι2.
of insolubles in the liquid surface [12]. Phase transitions take place at the interface of dicalcium silicate-sodium aluminate
Conclusions
solution, which is in accordance with the formation of surface micelles. It is known that polyol compounds are typically non-
It was found in the research that the adsorption of PEG at the
ionic surfactants, with a structural formula of H-(0-CH2-CH2)n-
interface of dicalcium silicate-sodium aluminate solution is a
OH. When the concentration of polymer is lower, the molecules
saturated multi-layer adsorption and the type of adsorption is "S"-
of PEG are adsorbed at the interface of the ß-CaOSi02 by -OH
type in accordance with the Freundlich Equation. The adsorption
groups [13], and no significant changes of interfacial tension
of PEG on ß-CaOSi02 in sodium aluminate solution belongs to
happen.
physical adsorption. The π-σ curve can be drawn according to the
As the concentration of polymer increases, the surface micelles
Gibbs formula, which is available to obtain useful information
(or called semi-micelles) are formed between PEG molecules and
from the shape of the π-σ curve.
adsorbed molecules by the hydrophobic effects of C-H bonds, resulting in a sharp increase of adsorption. At the same time, the
References
interface pressure changes rapidly, while the correspondingly molecular area only changes in a cell range, which is correlative
[1] BI S W, YU H Y. Production technology of alumina [M].
to the formation of surface micelles. The reason of the sudden rise
Beijing: Chemical Industry Press, 2006: 250-251.
in the adsorption capacity is the formation of hemi-micelles at the
253
[2] ZHANG C Z. Research on additive for inhibiting secondary
Foundation item: Project (50974036) supported by the National
reaction in clinker leaching process [D]. Northeastern
Natural Science Foundation of China: Project supported by
University. 2008: 49-50.
Aluminium Corporation of China
[3] REN G K. Technical research of secondary reaction inhibitor and sweetening process [J]. Light Metals, 2008(5): 16-18. [4] LI T C. Technical Research of Secondary Reaction Inhibitor and it's Sweetening Process [J]. Nonferrous Metals, 2002(1): 26-28. [5] SHEU E Y, STORM D A, SHIELDS M B. Adsorption kinetics of asphaltenes at toluene/acid solution interface [J]. Fuel, 1995,74:1475-1479. [6] OLOF S, TOBIAS H, TORGNY S, THOMAS A. Adsorption of delmopinol at the solid/liquid interface - The role of the acid-base equilibrium [J]. Journal of Colloid and Interface Science, 2010, 350: 275-281. [7] PAUL J, JOHN R. The adsorption of a polysaccharide at the talc-aqueous solution interface [J]. Colloids and Surfaces A: Physicochemical and Engineering Aspects, 1998, 139: 27-40. [8] WU Z S, LI C. Kinetics and thermodynamics of b-carotene and chlorophyll adsorption onto acid-activated bentonite from Xinjiang in xylene solution [J]. Journal of Hazardous Materials, 2009, 171: 582-587. [9] SOMASUNDARAN P, KRISHNAKUMAR S. Adsorption of surfactants and polymers at the solid-liquid interface [J]. Colloids and Surfaces A: Physicochemical and Engineering Aspects, 1997, 123-124: 491-513. [10] WANG Y J, SONG Z F. Spectroscopy and Chromatography [M]. Beijing: Peking University Press, 1995: 225-239. [11] HU Y. Modern Chemical and Engineering Thermodynamics [M]. Shanghai: Shanghai Science and Technology Press, 1994: 239-240. [12] HARKINS W D. The Physical Chemistry of Surface Films[M]. New York Reinhold, 1952: Chap2. [13] CHEN H W, ZHOU Z K. The Analyses on the Impact of the clinker
stripping
conditions
on
the
Secondary
Reaction[J].Light Metal, 2001(8):21-24. [14] SCAMEHORN J F, SCHECHTER R S, WADE W H. Journal of Colloid Interface Science, 1982, 85:463-467. [15] ZHU B Y, ZHAO X L, GU T R. Journal of Chemistry Society Faraday Transactions, 1988, 84 (ll):3951-3958.
254
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
PRODUCTION OF HEMATITE ORE FROM RED MUD P.K.N.Raghavan1, N K. Kshatriya1 and Katarzyna Wawrynink2 bharat Aluminium Company Limited, Korba (CG) India, 495-684 2
Warsaw University of Technology, Poland
Keywords: red mud, zerosite, hematite, thenardite universally as solid residue. The Jarosite generated in a Zinc refinery is normally associated with free sulfuric acid and ferric iron. The Jarosite mainly contains iron,sulfur,zinc, lead and calcium. Jarosite released from industrial process is complex and its quality and quantity make the of safe disposal complex. The most widely used method for disposal of Jarosite is the Jarofix Process.In this Jarofix process the Jarosite is treated with hydrated lime and then stabilized with ordinary Portland cement before dumping in a pond. At present the long term stability of Jarofix material remains unsatisfactory and when the material does break down it is likely that the trace metals will be released into the surrounding environment with potentially adverse consequences. It would be advantageous if red mud and Jarosite could be transformed into other forms less harmful to the environment. In the present work an attempt is made to evaluate the utilization potential in developing marketable commodities using red mud and Jarosite for safe and effective waste management.
Abstract Red Mud generated as a waste material in the Bayer process is a potential source of metal values. Red Mud is highly caustic residue (pH > 13).Approximately 1-2 tonns of red mud are being generated per MT of alumina produced. Disposal and storage problems will prevail due to the leachable soda content and the volume of material to be handled. Jarosite is a hazardous waste of Zinc refinery and is both difficult and costly to manage and to dispose safely. Jarosite is normally associated with sulfuric acid and Ferric iron in the waste produces additional acid on hydrolysis. Red Mud is source of hydroxyl and sodium ions whereas Jarosite is source of sulphate and hydrogen ions. Reaction between red mud and Jarosite in water will therefore produce a liquid phase containing sodium sulfate and a solid phase dominated by iron hydroxide. The liquid phase was evaporatively concentrated to precipitate salt and the solid phase was calcined to produce a ferruginous solid that could be used as an iron ore. The reaction of red mud and Jarosite produces two marketable commodities (Gypsum) & (Thenardite). In this paper an attempt is made to evaluate procedures for the preparation of iron ore and marketable byproduct salts by reacting the Balco's Red Mud and Jarosite from Hindustan Zinc of Vedanta Group.
Materials and Method Sample collection: The solid residue, Jarosite was obtained from our group company Hindustan Zinc Limited ( HZL) and Red Mud from Bharat Aluminium Co.Ltd.,Korba (India).The samples were oven dried (105-1100C) separately, sieved through 150 micron sieve. For characterization sampling was done from the oven dried sample adopting the conning and quartering method.
Introduction For every tonne of alumina produced in the Bayer processing of bauxite, 1 to 2 tonnes of red mud generated depending upon the quality and grade of the bauxite and, to some extent, the processing conditions. Red mud consists of almost all the elements present in the parent bauxite from which it is derived, many of them in enriched form, in addition to sodium compounds formed by reacting with caustic soda used for digestion, and calcium compounds if lime is added during the processing stages. It is infact a complex mixture of various oxides like Fe203, Ti02, A1203, Si02, Na20 occurring in major amounts. Disposal and utilization of red mud are common problems for all alumina refineries in the world. Industries located far the coastal region (as is the case with alumina refineries in India) face significant disposal problems. To dispose off a huge volume requires enormous civil, mechanical and process efforts causing huge expenditure. All alumina producers are, therefore keen to make effective utilization of red mud. During metallic Zinc extraction from Zinc concentrates huge quantity of Jarosite are released
Characterization of chemical & Mineralogical properties: The major constitutes of red mud were analyzed by XRF PW 2440 Phlips, Netherland. Heavy metals were detected by Atomic Absorption Spectrophotometer (AAS), SOLAAR S 4AA, and Thermo Elemental, USA with SOLAAR software package. The mineralogical analysis were carried out using PANlytical X'Pert Cubix Pro series diffractometer equipped with copper target tube,X'celerator detector and operated at 40kV and 30 mA. Diffraction data were analyzed using PANlytical X'Pert High Score plus Version 2.1.
255
The process for recovery of Hematite and Thenardite by reacting with Red mud and Jarosite is summarized in Fig. 1. Results
W&FER
SOLUTION
SOLID
EVAPORATE
CALCINE
The various major constituents present in Jarosite are given in Table 2 while that of the heavy metals such as Cr, Ni, Cu and Co in Jarosite are shown in Table 3. The % composition of major constituents of Red Mud and Calcined material are shown in Table 4. X- Ray diffractograms of Jarosite, Red Mud, Calcined solid (Hematite Iron) and Thenardite are shown in Fig. 2, Fig.3, Fig. 4 and Fig. 5 respectively.
800 \
The major mineral phase of Jarosite is potassium iron sulfate hydroxide (KFe3(S04)2(OH)6) and iron sulfate hydrate (2Fe203S03.5H20) (Fig.2).The Main crystalline phase of Red Mud are : Hematite (Fe203), Goethite (FeOOH), Gibbsite (Al(OH)3), Boehmite (AlOOH), Cancrinite (3NaAlSi04.NaOH), Quartz (Si02), Anatase (Ti02) etc (Fig.3).
HEMATITE SOLUTION
SOLID
FeA
ZYPSUM CHILLED
THENARDITE Fig.l. Schematic representation of recovery of i»matite from red mud
Table2.Chemical Properties of Jarosite (in %)
Experimental The process for recovery of Hematite from red mud by reacting with Jarosite is briefly outlined below: (i) A mixture of Jarosite along with red mud and water were mixed at different ratio at ambient temperature (280C) with magnetic stirring. The ratio of Jarosite, Red Mud and Water are given in the following Table 1. Table. 1: Various Ratios of Jarosite to Red Mud Used in the Experiment SN
Jarosite: Red
Mud
Water
pH
(weight ratio)
(ml)
1
1:1
300
6.5
2
2:1
300
6.0
3
4:1
300
6.0
To investigate the possible use of these neutralization reactions two parts of Jarosite and one part of red mud were reacted for 24 hours at ambient temperature. (ii)The neutralized slurry was filtered and washed with water. (iii) The filtered solution was concentrated by evaporation and two minerals (Gypsum and Thenardite) could be easily separated Gypsum is much less soluble than Thenardite and can be precipitated first and removed. (iv) The solid phase was calcined at 8000C for 2 hours to produce a ferruginous feed stock for blast furnaces producing pig iron.
256
Si
2.45
Fe
24.54
Al
2.12
Zn
4.47
Ca
4.13
S
11.68
Pb
0.52
K
0.30
Na
0.35
Mg
0.28
Mg
0.16
um
s asm
Cr
165
Ni
81
Cu
980
Co
42
Table3· Heavy Metals in Jarosite (in ppm) The Chemical Composition of BALCO's Red Mud is given in Table-4. Constituents %
Red Mud
Calcined Material
m]
I . .11
u;y wUäJi.... ...
T T TT.riTTTT r
10
20
Ί âû
40
5
" ' T 0Û
0611KM" [*21iieB3
taras
Fig.2. X-Ray diffractogram of Jarosite
Fe203
38.38
53.13
Si02
7.21
7.10
Ti02
17.37
6.87
A1203
20.61
Trace
Na20
6.39
0.26
P205
0.35
0.15
LOI
8.03
2.19
Table4. Chemical Properties of Red Mud and
L-.n ;
Calcined Material (in %) Couts
30
40 Posi Don [ïïheta] Couts .
QMZ
Fig.3. X-Ray diffractogram of Red Mud
Fig.4. X-Ray diffractogram of Solid Calcined Material (Hematite)
257
Chief Executive Officer and Whole Time Director, Mr. Gujan Gupta during the progress of this work. The authors would like to thank the Management of Vedanta Resources for allowing us to publish this paper.
Counts ■i prepgrd
References 1. Gaines et al, Dana's New Mineralogy Eighth Edition Wiley, 1997. 2. K L Bhat, K A Natarajan,T Ramchandran, Electroleaching of Zinc leach residue, Hydrometallurgy, 1987,18,pp.287-303
jy^uhi'mmekßJ
Π-tüs^i-— 23
30
40
■""!
5C
" T " ' §0
"T1 70
PcsHion f2~hets]
Fig.5. X-Ray diffractogram of Thenardite Conclusions Every Bayer process alumina refineries produce very high volumes of Red Mud, disposal of which in open fields has the potential to cause great environmental damage. It not only causes harm to vegetative sources, but also extends great health risks to both humans and animals. Jarosite wastes generated from the hydrometallurgical processes contain significant quantities of compounds of sulfur, lead, zinc, iron etc. The presence of toxic substances in Red Mud and Jarosite make these waste hazardous and causes serious problems in for their disposal. Due to weathering there is a release of toxic elements in soluble form which can ultimately contaminate the soil, ground water and aquatic life due to improper management of such hazardous wastes. The present work shows that it is possible to efficiently and cost effectively neutralize red mud from BALCO Alumina Refinery by mixing with Jarosite waste from Hindustan Zinc refinery (HZL). Neutralization of red mud using Jarosite will reduce both the costs and the risks associated with the long term storage unneutralized red mud. This neutralized material can be stored without liability. Neutralization process will provide a variety of cost and operational benefits to BALCO and HZL including (1) Helping conserve fresh water (2) Creating a soil medium for revegetation (3) Reducing future and stock piled caustic red mud and Jarosite (4) Potentially resulting in the production of marketable sodium salts (5) Creating a raw material hematite iron ore. Acknowledgement The authors acknowledge the constant encouragement from our Chief Operating Officer, Mr. Bibhu Prasad Mishra and
258
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
Light Metals 2011 ALUMINUM REDUCTION TECHNOLOGY
ORGANIZERS
Mohd Mahmood Aluminum Bahrain- Alba Manama, Bahrain Abdulla Ahmed Aluminum Bahrain- Alba Manama, Bahrain Charles Mark Read Bechtel Corporation Montréal, Canada
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
Light Metals 2011 ALUMINUM REDUCTION TECHNOLOGY
Enviroment- Emissions / Anode Effect I SESSION CHAIR
Robert Baxter Bechtel Corporation Montréal, Canada
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
HF MEASUREMENTS INSIDE AN ALUMINIUM ELECTROLYSIS CELL Karen Sende Ösen1, Thor Anders Aarhaug1, Asbj0rn Solheim1, Egil Skybakmoen1, and Camilla Sommerseth2 2
^INTEF Materials and Chemistry, NO-7465 Trondheim, Norway Department of Materials Science and Engineering, NTNU, NO-7491 Trondheim, Norway Keywords: HF levels, aluminium electrolysis cell
the relative contributions to HF emission, and to locate the main HF generation sites. The industrial measurements have been carried out in the gas channel between individual cells and the main duct system. From such measurements, essential information regarding the importance of certain parameters such as crust integrity has been obtained [4]. Still, the fluoride concentration in the duct represents the average of the conditions at the top of the cell.
Abstract HF emissions to the working atmosphere may still be a problem for the aluminium industry. The objective in the present work was to study how the HF evolution is distributed between feeder holes, other openings in the crust, gases diffusing through the crust, fumes from the secondary alumina residing on top of the crust etc. A movable "gas sniffer" connected to a Tunable Diode Laser was used to measure the HF concentrations at the above mentioned locations. The stationary HF level in an open flaming feeder hole was approximately 9000 ppm, when measured a few cm above the bath surface. In comparison, when the probe was positioned 5-10 cm above a crust area with good integrity, the HF concentration was in the range 5-10 ppm. The results support the notion that most of the HF evolves from open feeder holes and the tapping hole.
Useful and more direct information can be obtained by doing measurements locally on top of the cell. The main purpose of the present work was to gain knowledge about how the hydrogen fluoride evolution is distributed locally between the feeder holes, secondary alumina, openings in the crust, and gases diffusing through the crust as illustrated in Figure 1. To make measurements locally at the top of the cell, a portable sniffer based on online Tunable Diode Laser monitoring was constructed.
Introduction HF emissions to the working atmosphere continues to be a problem, and new challenges concerning emissions to both internal and external atmosphere emerge as a consequence of the ongoing cell technology development. In this situation, it is important to know precisely which factors affect the formation of hydrogen fluoride, and where it evolves. Hydrogen fluoride is generated when fluorides present in the bath or in the vapour phase react with moisture, f AlF,(diss) + H 2 0(g) = 2HF(g) + lAl 2 0 3 (diss) NaAlF4(g) + H 2 0(g) = 2HF(g) + 1 Al 2 0 3 (s) + -|Na3AlF6(s)
Figure 1. Schematic representation of water sources and possible HF-evolution sites (represented by the red arrows) in an aluminium electrolysis cell.
(l)
Design Criteria for "HF-sniffer"
(2)
Based on the data from the literature [5], as well as calculations using the fluoride evolution model by Haupin and Kvande [2], the total fluoride evolution from a rather large modern cell, such as the cells in Hydro's plant at Sunndals0ra (SU4), will be in the order of 30-40 kg F/t Al. Between half and two thirds of this is due to HF. Depending on the gas suction rate, the concentration of HF in the duct will be around 300 ppm (volume), or about 250 mgF/Nm3. This value represents the average situation on top of the cell. The maximum possible gas concentration can be estimated by assuming that all HF is formed in the bath and follows the cell gases out from below the crust. 20 kg HF/t AI then corresponds to about 3.4 vol% HF (24 000 mgF/Nm3) in the
It has long been established that two of the main sources of water are the structural hydroxyl in the primary alumina and the moisture content in the air [1, 2, 3]. A certain amount of HF is also generated from electrochemical oxidation of hydrogen in the anodes. It is also a well known fact that the HF formation rate increases with increasing temperature, decreasing bath ratio, and decreasing alumina concentration [1,2]. Quite a lot of industrial measurements and a few laboratory studies have been carried out during the recent years to quantify
263
in this case, since it turned out that a certain minimum suction rate was necessary to get a good reading.
cell gas coming up from, e.g., an open feeder hole. This can be regarded as the maximum concentration that can be expected. Formation of 20 kg HF/tAl is equivalent with 0.48 wt% water in the alumina, if all water reacts to HF. The total amount of water in the alumina is normally much higher, which indicates that most of the water is not available for reaction to HF. Furthermore, the equilibrium in Eq. (1) is not completely displaced towards the right hand side. Figure 2 shows the concentration of HF and H 2 0 in the cell gases at 960 °C ("normal" bath composition) as a function of the content of water in the alumina. In these calculations, it was assumed that all water is available for reaction with the bath. 1
1
1
1
Figure 3 shows a principal sketch of the set-up. The instruments were placed on a table close to the wall by the emergency exit. It was then out of the way for all motorized traffic, and the magnetic field was moderate. The tunable diode laser (TDL) was a HF and H 2 0 Lasergas II Single Gas monitor from Neo. The PFA tubings (diameter 1/4") between the laser and the probe were approximately 20 m long and flexible. The probe could then be transferred manually from one measuring site to the next. Gas was sucked continuously through the probe and into the laser measuring cell with the help of a pump placed at the end of the line at a rate between 5 and 15 1/min. N2 (g) was purged through the path of the laser beam between the emitter and the receiver outside the actual measuring cell to eliminate the effect of humidity in the ambient air that otherwise would be present. The laser was connected to a computer, and the values recorded every 10th second. During some of these measurements, a thermocouple (Pt-PtRh 10%, type S) was placed at the probe tip. The purpose of this was to document the vertical position of the probe. Figure 4 a shows a picture of the probe tip with the thermocouple, and Figure 4 b shows how the probe was inserted through a hole in the cover.
1
HFy/^ .o
Η^θχ^ "suction probe" with thermocouple
c: o O
.....
0.5
1.0
1
1
1.5
2.0
%" PFA-tubings (appr. 20 m)
sample gas
Inert gas into outer circuit
2.5
Tunable diode laser (measuring cell 10 cm)
Inert gas out from outer circuit
"Gas watch" counter for the gas volume that
Water in alumina / wt%
Figure 2. Equilibrium concentration (Eq. (1)) of H 2 0 and HF in the cell gases as a function of the H20content in the alumina, assuming that all water in the alumina is available for reaction with the bath. 11.5 wt% A1F3, 4.5 wt% CaF2, 3.5 wt% A1203, 960 °C, 93 % current efficiency.
Vacuum pump
S
Figure 3. Gas sniffer, as it was used in the 2nd campaign. Improvements from 1st Campaign: New laser (higher upper limit), shorter measuring cell (shorter response time and higher upper limit), inside coated with PTFE, sapphire glass instead of fused silica, pressure probe installed, and thermocouple at the probe tip.
Experimental Set Up and Procedure The measurements were performed at an end cell at SU4 Hydro Sunndal in Norway during two campaigns, one in October 2008 and one in May 2010. The laser used in the 1st campaign had a more sensitive measuring range and was therefore suitable for measurements only at sites where the HF concentration was below 750 ppm, such as on top of the crust far away from any openings. The second campaign was performed with modified equipment. The measuring cell was coverered with PTFE (Teflon ®) to provide better protection towards HF than the steel used earlier. The cell was also made shorter. Furthermore, a sapphire glass window was used instead of fused silica. The shorter measuring cell resulted in a quicker dynamic response and contributed to a higher upper concentration limit since the physical principle utilized in TDL technology involves a correlation between concentration and optical path length. A pressure probe was placed in the measuring cell. This improved the instrument's capability to handle underpressure, something which was essential
Figure 4. a) Probe tip with thermocouple. b) Probe inserted through a cell cover.
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Above Crust and in Open Tap Hole and Feeder Hole. When the probe was placed as shown in Figure 5, the HF stationary HF concentration varied between 5 and 10 ppm; see Figure 7. When the probe was positioned above the feeder hole as shown in Figure 6, the concentration was in the range 100-200 ppm.
Figure 5. Probe positioned on top of the crust.
100
150 time/minutes
Figure 7. HF and H 2 0 levels above crust (probe positioned as shown in Figure 5). Figure 6. Probe positioned above feeder hole.
Figure 8 shows how the concentration varied as the probe was moved from above a blocked feeder hole to an open tap hole close to the bath surface. The probe was then flushed with N2, then placed in an open feeder hole close to the bath surface. Keeping the probe in the open tap hole or the open feeder hole made the concentration rise abruptly to above the measuring range of the instrument. It should be noted that the probe was positioned only a few centimeters above the bath surface, and not as shown in Figure 6. The flat signal when placed in the open tapping hole is simply a signal overload. The reason for the slightly stepwise increase when held above the open feeder hole is that the probe had to be yanked away every time the crust breaker came down before feeding. This resulted in a small dilution effect.
To verify the system, measurements were performed in the duct (the measuring site was located at a place close to the cell where only the off gas from one particular end cell is passing) where the expected concentration range could be calculated to be around 300 ppm as mentioned earlier. During one of these periods, a NaOH containing filter was placed at the end of the line. This was done to confirm that the online measurements were in consistency with the well established Sintalyzer fluoride analysis method. Then the probe was placed on top of the crust as shown in Figure 5, and further hand held above an open tapping hole and above and inside an open flaming feeder hole. Figure 6 shows the probe placed above an open flaming feeder hole. Results 1st Campaign Duct Measurements. A more convenient short measuring probe was used during the first duct measurements. This probe was kept in place over night for approximately 16 hours and 40 minutes while recording the HF and H 2 0 concentrations continuously. To trap the HF passing through the laser, several filters soaked with NaOH were placed at the end of the line and kept there for about 16 hours and 13 minutes, that is, during almost the entire time period. The subsequent Syntalizer analysis performed on the filters showed an average concentration of 213 mg/Nm3 HF in the gas passing the laser during the measurement The average HF concentration calculated from the laser measurements done at the same time interval was 253 ppm, which coreesponds to 210 mg/Nm3.
0.0
20.0
40.0
60.0
80.0
100.0
time/minutes
Figure 8. HF and H 2 0 levels at various locations inside the cell. 2nd Campaign
The values recorded by the laser varied from slightly below 200 ppm to slightly above 400 ppm, clearly correlating with the feeding cycles as would be expected.
The main purpose of this campaign was to obtain a quantitative value of the HF levels in the open feeder holes and tap holes.
265
Duct Measurements. As in the first campaign, to confirm that the system was without large leaks, and that the values obtained were reliable, various measurements were done in the duct. Figure 9 shows HF and H 2 0 levels when a short probe and the long probe were placed in the duct with and without dust filter. As can be obeserved, when used with a dust filter, the long probe exhibited a larger response time than when used without the dust filter (after 4 minutes the measured value is still far below the expected stationary value). This was probably caused by the fact that a larger pressure drop is built up across the dust filter. When the dust filter was removed, it took about 4 minutes to obtain a "reasonable" value (approx. 300 ppm). The rest of the measurements were therefore conducted without using the dust filter.
As shown in Figure 10, the highest temperature (approximately 876°C) was measured inside the open feeder hole, a few centimetres above the bath surface.The bath temperature at 14:00 was measured to be 954 °C. When held in the open tap hole, the bath surface could not actually be observed visually. The temperature being "only" between 600 and 700 °C, might indicate that the probe was further away from the bath surface than when held in the open feeder hole, or it might indicate that the temperature above the bath is higher in a flaming hole due to the oxidation of CO. 1.8 1.5 1.2
3
O
N2 in outer circuit
in the duct with "long probe" - without dust filter
2.5
3? 0.9
2
0.6
%,5
0.3 1
14:15
14:25
14:35
14:45
14:55
15:05
time 0.5
Figure 11. HF and H 2 0 levels in an open tap hole and in an open feeder hole plotted together with the course of events, a) 14:20-14:30: in open tap hole, b) 14:3014:34: break, c) 14:34-14:43: in open flaming feeder hole d) 14:43-14:44: break e) 14:44-14:52: back in open feeder hole, f) 14:52: break
Figure 9. Measurements of HF and H 2 0 in the duct. Measurements in Open Holes After the duct measurements, a thermocouple was attached to the tip of the probe. The probe was transferred to an open tap hole, and then to a flaming open feeder hole. The measurements are shown in Figures 10 and 11. The figures represent the same time period, i.e, the same events, but presented somewhat differently. Experience from the last campaign had shown that there was a big concentration gradient above the open holes. Therefore, to get as high a reading as possible, the probe was held only a few centimetres above the bath surface. However, since the crust breaker regularly came down (every 2-3 minutes), the probe had to be pulled up and away from the path of the crust breaker. This to a certain extent contributed to dilution of the gas sucked into the probe.
The highest HF value measured was more than 9000 ppm (0.9%), and this was obtained when the probe was held in an open feeder hole (periods c and e in Figure 11). During the same periods, the water vapour level reached a low value when the HF values were high. This suggests that a large part of the moisture is converted to HF according to Eq. (2) at the prevailing conditions; high content of fluoride vapour and high temperature. When the probe was removed from the open feeder hole, the concentration of water vapour rose to a higher level. This is especially pronounced in period f in Figure 11. Why the concentration exceeds its original background level is uncertain.
time
Dynamic Response Time. The dynamic response time is here defined as the time between an imposed change on the system and the first response recorded. The thermocouple has a very fast dynamic response compared to the HF and H 2 0 signal (and is also logged every two seconds), so its dynamic response could be used as a time basis for calculating the response in the other recorded items. In this way, the dynamic response time for the HF and H 2 0 signals are estimated from Figure 12 to be 21-33 seconds and 10-11 seconds respectively; the probe is taken out from an open feeder hole at 13:36:18 and the measured HF concentration starts to decline between 13:36:40 and 13:36:51. H 2 0 starts to increase between 13:36:18 and 13:36:51. The probe is put back in at 13:38:46 and the HF concentration starts to increase between 13:39:05 and 13:39:17. The onset of water level decrease is between 13:38:44 and 13:38:54.
Figure 10. HF and H 2 0 levels in an open tap hole (14:20-14:30) and in an open feeder hole (14:34-14:52) plotted together with the temperature measured.
Stationary Response Time. The stationary response time is here defined as the time it takes for the system to obtain a new stationary value (where this is expected) after a change has been
14:35
14:45
266
be around 4.5 %, more than four times as much as the actually recorded amount. This may indicate that even though there, in principle, would be overpressure and a gas flow directed upwards in the open flaming feeder holes due to the gas evolved underneath the anodes, air may be sucked into the holes and possibly also underneath the crust. Possibly, this can be related to waves and splashing at the bath surface. The observation made in this work, that there is a large difference in HF concentration when measured "above" the hole as opposed to "inside" the hole also supports this. A follow up on the present work was later done at Alcoa Mosj0en, where levels up to 2.6 % HF were detected in crust openings; this is reported elsewhere [7]. The feeder holes at Alcoa Mosj0e had a smaller diameter and were more cylindrically shaped than the ones observed at Hydro Sunndal.
imposed. For HF, this was estimated from the duct measurements to be approximately 4 minutes for increasing concentrations and approximately 2 minutes for decreasing concentrations. For H 2 0, the current data do not provide enough information to estimate the stationary response time. The response time of the laser itself is believed to be around one second, so the discussion above refers to the whole system.
Conclusions The stationary HF level in an open flaming feeder hole was approximately 9000 ppm, when measured a few cm above the bath surface. In comparison, when the probe was positioned 5-10 cm above a crust area with good integrity, the HF concentration was in the range 5-10 ppm. The results support the notion that most of the HF evolution emits at the open flaming feeder holes and tapping hole. The gas sniffer equipment worked according to the purpose, and this method has proven a useful tool for these types of measurements. Future campaigns will be carried out, both at other smelters but also with respect to other kinds
Figure 12. Close up of HF and H 2 0 levels as they change when the probe is taken out from an open feeder hole and put back in. Summary and Discussion
Acknowledgement
Table I shows a summary of the HF levels measured at the various sites. It seems rather clear that the HF gas generated exits almost entirely through the crust openings. Keeping the probe above a blocked feeder hole gave no indication of HF desorption from secondary alumina. The fact that surface adsorbed fluoride does not re-evolve during ore feeds to the pot, has also been reported by others [4,6].
The present work was funded by The Norwegian Research Council and the companies involved in the project Resource Optimation and Recovery in the Materials Industry (ROMA): Hydro Primary Metal Technology, Alcoa Norway, S0r-Al, Elkem, Eramet, Vale Manganese Norway, Fesil, Finnjord, and Alstom. Permission to publish the results is gratefully acknowledged. We would also like to express our gratitude towards Jan Olav Polden at Hydro Sunndal for providing vital equipment and for his assistance and help whenever needed. Thanks are also due to Chris Torjussen and his colleges at NEO for valuable advice in connection with the laser measurements.
Table I. Summary of HF levels. Location
Range
Comments
Above crust ("far" |awayfromfeeder hole) Above open feeder hole Duct
5-10 ppm
October 2008
5-200 ppm
October 2008
200-400 ppm, fluctuate according to feeding cycle 50 ppm Above blocked feeder hole Just above bath, open 3000-5000 ppm tap hole 1 In open feeder hole 0.90-1.0% 1 just above bath
References 1.
W. E. Wahnsiedler, R. S. Danchik, W. E. Haupin, D. L. Brackenstose and J. W. Colpitts, "Factors Affecting Fluoride Evolution from Hall-Heroult Smelting Cells", Light Metals 1978. pp 407-424
2.
W.E. Haupin og H. Kvande, "Mathematical Model of Fluoride Evolution from Hall-Heroult Cells", Light Metals 1993. pp 407-411
3.
Margaret Hyland, Edwin Patterson and Barry Welch, "Alumina Structural Hydroxyl as a Continuous Source of HF", Light Metals 2004. pp 361-365
4.
Michael L. Slaugenhaupt, Jay N. Bruggeman, Gary P. Tarcy and Neal R. Dando: "Effect of open holes in the crust on gaseous fluoride evolution from pots" Light Metals 2003. ppl99-206
October 2008 October 2008 May 2010 May 2010
If the suction rate during the duct measurements is approximately 7000 Nm3/h (it was not measured accurately for the actual cell during the campaign), the HF generation rate can be calculated to be around 27 kg HF/ton AI. If the raw gas that exits the feeder hole were to be completely undiluted, the HF concentration would
267
5.
Edwin Patterson: " Hydrogen fluoride emissions from aluminium electrolysis cells" PhD-thesis, The University of Auckland, 2002
6.
Neil R. Dando and Robert Tang: "Fluoride evolution/emission from aluminium smelting pots impact of ore feeding and cover practices" Light Metals 2005. pp 363-366
7.
Camilla Sommerseth, Karen Sende Ösen, Thor Anders Aarhaug, Asbj0rn Solheim, Egil Skybakmoen, Christian Rosenkilde and Arne Petter Ratvik, Light Metals 2011 (current issue)
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
LASIR™-R - THE NEW GENERATION RoHS-COMPLIANT GAS ANALYZERS BASED ON TUNABLE DIODE LASERS Alak Chanda1, Gervase I. Mackay1, Keith L. Mackay1, John T. Pisano2, Jean-Pierre Gagne3, Pierre Bouchard' unisearch Associates Inc., 96 Bradwick Drive, Concord, Ontario, Canada L4K 1K8 university of California at Riverside, 1084 Columbia Avenue, Riverside, CA, USA 92507 3 STAS, 1846 Rue Outarde, Chicoutimi, Qc, Canada G7K 1H1 Key Words: Tunable Diode Laser Spectroscopy, Hydrogen Fluoride, HF, emission monitoring, process control, aluminum smelter, gas analyzer the energy is absorbed at that wavelength. Since the amount that a molecule vibrates or rotates at a specific wavelength will be dependent on its capability to absorb energy, it will behave differently at different temperatures and pressures. This alludes to the fact that both temperature and pressure compensations for line intensities are required for any spectroscopic measurement. Such compensations can be made either by measuring the property changes of the molecules directly (for simpler molecules) and applying a mathematical correction or by preparing a compensation database empirically (for complex molecules) with laboratory experiments performed under controlled operating environments.
Abstract The laser-based optical gas sensor using Tunable Diode Laser Absorption Spectroscopy (TDLAS) is rapidly gaining favor wherever high sensitivity, real-time measurement and freedom from interferences are required [1,2], specifically for the measurement of HF in primary aluminum smelters and other process gases at various metal smelters. It eliminates the problems associated with extractive gas sampling techniques. A first generation of equipment designed by Unisearch Associates and based on the near-IR TDLAS appeared on the market in the mid-90s. It has since been improved significantly by employing fast scan measurement techniques. Now, the system is more compact, robust, easier to operate and calibrate, and it can be simply audited. The utilization of fast scan measurement techniques with the state-of-the-art high speed electronics and sophisticated software optimization algorithms not only provides enhanced stability and sensitivity but expands the dynamic range of the measurements to five orders of magnitude. This new generation RoHS-compliant instrument (Restriction of Hazardous Substances) is also very inexpensive compared to other gas analysis instruments. The technology has found applications in many industries that require fast, accurate measurement of trace emissions for both process control and environmental regulatory requirements. These include, but not limited to, HF emissions in aluminum smelters, NH3 slip measurements for DeNOx scrubbers in the Power Industry, HC1 and HF emissions monitoring in the Cement Industry and in Incinerators and for process control in Steel Smelting. This paper describes the new generation TDLAS-based gas analyzer and, as an example, its use in primary aluminum smelters.
The advantage of TDLAS over other techniques for concentration measurement is its ability to achieve fast, very low detection limits of the order of parts per billion with virtually no interference from the presence of other gas molecules. The LasIR™ is also widely being used by various aluminum smelters worldwide to measure gaseous hydrogen fluoride [3, 4]. Hydrogen fluoride (HF) is emitted during the aluminum smelting process. To avoid HF leakages in the work area, the electrolytic cells (pots) are hooded and the emitted gases are captured and vented through scrubbers. Hydrogen fluoride is a very toxic gas and represents a major health and safety hazard both for workers in the pot rooms and for neighboring environments. In addition, there is also an economic incentive to recovering this fluoride. The Tunable Diode Laser technology is presently used in over a hundred smelters in the world as an emissions monitoring device used to adjust emission control practices and equipment. Description of Technology
Introduction
The gas analyzing system consists of a tunable diode laser based analyzer and a small optical head that launches the laser light through a medium (that may contain the species of interest to be measured) and receives the transmitted laser light at the other end of the measurement medium. The analyzer itself can be located in a control room, as far as 1 km away from the measurement location. Afiber-opticcable and a coaxial cable link the analyzer to the optical head. Data are logged on to a compact flash disk built inside the TDLAS-based gas analyzer and/or an external PC via the Ethernet or RS232 outputs. Data is also transmitted to the plant database.
A basic TDLAS setup consists of a tunable diode laser light source, light transmitting optics, an optically accessible absorbing medium through which the light beam can be transmitted, receiving optics and detectors. The emission wavelength of the tunable diode laser is tuned over the characteristic absorption lines of a species in the gas in the path of the laser beam. This causes a reduction of the measured signal intensity, which can be measured by a detector and then used to determine the gas concentration. Most molecules are unsymmetrical and polyatomic. When the molecule becomes excited by absorbing near infrared energy, it increases its vibrational and/or rotational frequency. It is this tendency of the molecule to absorb energy that leads to spectroscopic identification of the molecule by the wavelength of absorption and the concentration of the molecule by how much of
The quantification is based on Beer-Lambert Law that relates the absorption of light to the properties of the material through which the light is traveling. It states that when a radiation of
269
frequency I(v) passes through an absorbing medium, the intensity variation along the path of the beam is given by: I(v)=I0(v)e-a(v)cl
and aligned probe model is also available that can be mounted at any stack/duct for short-term or long-term measurements.
[1]
where,
i-600m
Œ-
is the transmitted intensity of radiation after it has traversed a distance / through the medium, (in milliwatts) I0 (v) is the initial intensity of the radiation, (in milliwatts) o(v) is the absorption cross-section of the absorbing species, (cm2/# of molecules) C is the number density (concentration) of the absorbing species, and (in # of molecules/cm3) / is the path length or the traverse distance through the medium (in cm). I(v)
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Remote measurement of gases using TOLAS provides the concentration that actually represents a value averaged across the entire path. This gives in situ measurement an advantage over point sampling in which a probe is inserted to draw the gas out of the duct, especially when concentration gradients or heterogeneities exist. The path averaged value obtained from in situ sampling is better representative of the overall concentration within the process.
Probe Measurements
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Portable wand tor leak detectìon «sing a*sg&e Stoefßpöc oabte up to 1 km long
Figure 1. Schematic of TDLAS-based gas analyzer and the optical head configurations that can be connected to the analyzer byfiber-opticand coaxial cables. Portable stand-alone models are also shown. The TDLAS-based gas analyzer consists of appropriate hardware and software that require no field calibration. This eliminates the requirement of having calibration gas cylinders available on site. For regulatory purposes, compact hand-held audit modules are provided that allow the user to periodically audit the system as a whole (analyzer and optical heads) when desired or use the periodic automated audit feature. Automated audit results can be stored and displayed on screen. Alternatively, where auditing is to be performed by flowing a known level of gas on the path of the optical beam, an optional add-on module is available that can be configured as an inline or a stand-alone offline audit assembly. Calibration, although not required, can also be performed with this add-on module where regulatory requirements have to be fulfilled. With the use of a TDLAS-based analyzer, fast, real-time, in-situ measurements of gas concentrations linear to five orders of magnitude (sub-ppm to percent levels) can be made under various industrial environments.
The use of a laser source coupled with a fiber-optic cable allows the light to be optically multiplexed to make simultaneous measurements of the species of interest at different locations. With the TDLAS-based gas analyzer, measurements can be made from 1 to up to 16 different locations with a single analyzer. In addition, multi-species can also be measured with the same analyzer where the absorption signals of the species lie relatively close to each other and within the tunable range of the laser. Measurement of gases in explosion-proof environment can easily be done by using optical heads that require no electrical components (with configurations that transmit and receive the optical signal viafiber-opticcables). Unisearch has recently introduced a new generation of TDLAS-based gas analyzers that are RoHS-compliant and available in either rack-mount or table top models as standards. With the advent of high speed electronic data processing and measurement techniques, the costs of these types of TDLAS based analyzers are significantly reduced. The gas analyzers come in a compact form that makes them suitable for aluminum smelter environments. They are available in configurations from simple portable single-channel analyzers to systems capable of measuring up to 16 channels that can monitor both multiple locations or different gas species. All the measurement channels are independently controlled, which permits a single multi-channel analyzer to simultaneously measure gas levels in duct/stack, long-path ambient air, extractive sampling and any such combination. Each channel can handle very different gas levels with no interference with other channels. The possible measurement configurations are shown in Figure 1. The system can also be configured as a stand-alone portable gas analyzer with the optical head integrated with the electronics. This light-weight (< 5 kg), energy-efficient (consumes less than 20 W) model can be mounted on a tripod stand and, with the use of a retro-reflector array, allows measurement of various gases in an open-path mode to up to several hundreds of meters. A stand-alone, pre-configured
HF Gas Measurements as Process Control in Primary Aluminum Smelters One of the primary requirements of a gas analyzer, especially when used as a Continuous Emissions Monitoring system, is the accuracy and linearity of the measurements. This data is concurrently important when a system is used for process control applications such as maintaining scrubber efficiency and improving work practices, as well as reducing emissions. Various test and certification agencies exist that are able to verify the functionality of such analyzers. One such certification agency is the well-recognized German TÜV. Approval requires the use of two essentially identical analyzers measuring the same gas. Laboratory and field (industrial plant) tests are conducted in order to verify the manufacturer-stated specifications. These include, among many other parameters, the linearity, accuracy, precision and the operation of the analyzer under various environmental conditions. Figure 2 shows the linearity and accuracy of two TDLAS-based gas analyzers for low levels (0 to 2.5 ppm-m) of HF gas
270
measurement. As can be seen from the plot, the measured HF gas levels were almost identical between the two analyzers. AH Alumina Feeds Returned To 100%
HF measurements with two LasIR analyzers - TUV testing
— a —LaslR#2 y=1.002x +0.011 R* = 0.999
10:25
12:25
14:25
16:25
18:25
2025
Time (hh:mm)
Figure 3. Variation in HF emissions as measured by a TDLAS-based gas analyzer as a function of alumina feed to the dry scrubber.
Actual HF Concentration {mg/m3)
Figure 2. HF measurements with two TDLAS-based LasIR™ during TÜV certification.
A number of installations use TDLAS-based systems to monitor scrubber efficiency. In a typical installation, a single controller is used to monitor the HF concentrations at both the inlet and the outlet of the scrubbers. Figure 4 shows a plot of such measurements at a newly installed scrubber. Measurements were made in an in-situ mode. This plot shows that the new scrubber was able to remove more than 99.8% of the H. This information also allows the scrubber operator to maintain the desired reduction of fluoride emissions from aluminum smelting by adjusting the alumina feed rate (and the secondary alumina recycle rate) to the dry scrubber to optimize the capture efficiency of the gaseous fluoride byproducts. The measurement also indicates to the operator the surface adsorption inefficiencies that arise from losses of scrubber material such as bags either breaking or leaking, or saturation of the reactive material.
In aluminum smelting process, aluminum metal is extracted from its oxide alumina, generally by an electrolytic process called the Hall-H6roult process. As such, aluminum smelters use enormous amounts of electricity and are often located very close to large power generating stations. Primary aluminum is produced by the electrolytic reduction of alumina (A1203), dissolved in a mixture of molten cryolite (Na3AlF6) and A1F3. Since HF is emitted during the process, the electrolytic cells (pots) are hooded and the emitted gases are captured and vented through scrubbers. Hydrogen fluoride is a very toxic gas and represents a major health and safety hazard both for workers in the pot rooms and for neighboring environments. In addition, there is also an economic incentive to recovering this fluoride. Dry scrubbing technology is used at most of the world's aluminum smelters. The hydrogen fluoride is chemisorbed directly on to the alumina from the hot exhaust gas. Fresh (primary) alumina is injected directly into the raw exhaust gas stream, which mixes and reacts with the alumina. Then the reacted alumina, as well as the particulate fluorides and other particulate materials, are removed from the exhaust gas stream by bag filtration. The alumina collected (secondary alumina), which contains almost all thefluoridesand particulates emitted from the reduction process, is then fed to the pots. Thus, the entire process operates as a closed loop for the captured cell emissions. It has been found that manipulations in the feed of alumina to the dry scrubber altered the levels of HF emitted from the scrubber outlet that is usually vented to open air. Therefore, a continuous in-situ measurement is a useful tool to optimize the feed of alumina to the dry scrubbers, and/or the secondary alumina recycle rate.
Also of importance is the accuracy and reliability of the measurements when made under an industrial environment. The gas temperature, pressure, humidity and operating environmental parameters can constantly change as a function of time. In order to validate the measurements, a continuous 5-day comparison study of the HF gas data was made between those reported by a TDLAS-based gas analyzer and the standard wet chemical method. Results are shown in Figure 5. Apart from the response time where the gas analyzer shows sharper peaks, the results were virtually the same.
Average = 218 mgini3
Scrubber Inlet
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E
An example of the control of alumina feed is shown in Figure 3, obtained from tests conducted at an installation site. The higher HF levels were produced by reducing the primary alumina feed to the scrubber. In just a few minutes, the levels rapidly rose from 0.47 mg/Nm3 to 0.8 mg/Nm3. A few minutes later, when the HF levels appeared to reach stable values, the recycle rate for injected secondary alumina was reduced. The concentration rose to almost 1.8 mg/Nm3. To evaluate the scrubbing response, all parameters were set to normal. Once all the alumina feeds were turned on, an exponential decay in HF emissions was observed.
E
o
I
1 Scrubber Outlet
03 =
Average = 0.42 mg An3 1
i
1 12:00
TIME
Figure 4. HF measurements at the inlet and the outlet of a dry scrubber; the scrubber efficiency is determined to be better than 99.8%.
271
increased beyond 50.5%. The results are shown in Figure 8 and indicate that the measurements are linear over the entire measured range. Such systems become increasingly useful in industries such as nuclear material processing plants.
Wet Chemical Method
HF measurement at the top of a roof line in an aluminum smelter
3.0
I 2.5 E
2 3 No. of days
^t
Figure 5. HF measurements at the inlet and the outlet of a dry scrubber. The scrubber efficiency is determined to be better than 99.8%.
2.0
Mj
Figure 6 shows the measurement of HF gas in the exhaust manifold of a single pre-baked anode electrolytic cell. The measurements were made in a path length of 0.5 meter and approximately 1 meter away from the end of the pot. Periodic reproducibility as a result of automated alumina feed and other related processes can easily be seen on the expanded plot.
L>*W^
20:00
.JAAJAU^
1:00 Time (hh:mm)
6:00
Figure 7. Measurement of HF at the top of a pot room of an aluminum smelter.
Single Pot {Pre-baked Anode} Duct HF Concentration Measurement
Online Linearity Measurement of HF Test Gas at a Metal Fluoride Processing Plant 58 5% 45.5 %
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18:15
18:30
16:45
17:00
Time ihr.min)
Figure 6. HF off-gas measurement from a single pot 1 meter away from the pot end.
Figure 8. Linearity measurement of HFfrom0% to 50%. Conclusion
Modern cells fitted with hooding to capture the emissions are usually operated at a slight pressure. Process interventions such as the manual alumina feed, changing of anodes or metal tapping and suppression of anode effects, all require opening of the hoods. Such interventions and any leakage in the pot hooding itself result in the emission of HF into the pot room, which is ultimately vented through the roof of the pot room. By measuring the HF emissions in the pot room and correlating their occurrence with specific pot tending activities, improved work practices can be developed and better control of current efficiency can be achieved. Figure 7 shows the levels in a pot room measured with a gas analyzing system. The sampling time was 5 seconds and the path length was 400 meters.
The Tunable Diode Laser based gas analyzer is being used as a tool for making in-situ measurements of emissions and flue gases in various industries. An example of its use specific to, and not limited to, aluminum smelters has been provided. Tunable Diode Laser Absorption Spectroscopy affords the most interference-free analytic method for measuring gas concentrations where fast and accurate response is desired. Such systems are particularly well suited to process control applications. These are small in size, can be designed as calibration-free, rugged and virtually maintenance-free, making it easy to install and operate. The use of fiber-optic cables permits the instrument to be located in an environmentally suitable area, remote from the actual measurement location. In addition, optical multiplexing permits measurements at a number of locations with a single instrument, making the systems very cost effective. Measurement of gases in explosion-proof areas can easily be done. The data is used for the control and optimization of various process related parameters. This leads to energy efficiency, reduction in production time and improvement of the quality of the products.
There are several processes where the gas level changes over a wide range of magnitude. This can be as a result of a specific step in the process or the process as a whole. In order to verify the dynamic linearity of the analyzer, tests were made in an industrial environment where provisions of controlled variations in the injected HF levels existed. The HF gas level was varied from 0% to up to 50%. In the set up used, the levels of HF could not be
272
The technology has found applications in many industries that require fast, accurate measurement of trace emissions for both process control and environmental regulatory requirements. These include NH3 slip measurements for DeNOx scrubbers in the power industry, HCl and HF emissions monitoring in the cement industry and in incinerators, off gas remediation of H2S in wastewater treatment and nickel smelters, deuterated water leakages in nuclear power plants and for process control in steel smelting, to mention just a few. References 1. John T. Pisano, Claudia Sauer, Tom Durbin and Gervase Mackay, "Measurement of Low Concentration NH3 in Diesel Exhaust using Tunable Diode Laser Adsorption Spectroscopy (TDLAS)", Society of Automotive Engineers Conference, Paper 2009-01-1519, Detroit, MI, February (2009). 2. H.A. Gamble, G. I. Mackay, J. T. Pisano and R. Himes. "Post Combustion NOx Control Process Monitoring using near-IR Tunable Diode Laser (TDL) Spectroscopy to Measure Slip Ammonia", 20th Annual International Forum Process Analytical Technology, Arlington, VA (Washington D.C.), February (2006). 3. Harold I. Schiff, Alak Chanda, John Pisano and Gervase Mackay, "The LasIR™-A Tunable Diode Laser System for Environmental and Industrial applications", PIE International Symposium 3535 on Industrial and Environmental Monitors and Biosensors, Boston, MA, November (1998). 4. Harold Schiff, John Pisano, Alak Chanda, Dave Karecki and Gervase Mackay, "Measurements of Fluoride Emissions in Aluminum Smelters by Tunable Diode Laser Spectroscopy", AWMA Annual Meeting, St. Louis, MO, June 1999, paper 115.
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
USE OF SPENT POTLINING (SPL) IN FERRO SILICO MANGANESE SMELTING TMS1, Paulo von Krüger2 *TMS (The Minerals, Metals & Materials Society); 184 Thorn Hill Rd.; Warrendale, PA 15086, USA 2 Universidade Federal de Ouro Preto Keywords: Spent Potlining, Ferro Silico Manganese Abstract period that goes from 2000 to 3000 days, the pot cathode lining fails and must be rebuilt. At this time, the old lining is dismantled. This material is called SPL (Spent Pot Lining) and is constituted of the starting lining material and the components absorbed during the operating pot life. Figure 1 shows the steps of the SPL formation [1]
In this work an evaluation of the possibilities to employ Spent Potlining (SPL) as a component of the burden of a Submerged Arc Furnace (SAF) producing Ferro Silico Manganese Alloy is investigated. On this subject, a characterization of the SPL's most probable components as well as their interaction with the existing species in the ferroalloy furnace were carried out. Additionally relevant features of the ferroalloy smelting were identified and characterized. Those figures were introduced in a thermo chemical program where the SAF operational conditions were simulated in order to check the technical feasibility of that use.
SPL can be easily separated in two well defined fractions: The carbon fraction is constituted of the old cathode blocks and non carbon fraction, constituted of old refractory and insulating bricks, ramming paste, electrolytic bath and a series of components formed by the interaction between those components and metal and air.
The simulation results showed that, on the technical point of view, the SPL carbonaceous fraction is a suitable component of the SAF burden, producing Ferro Silico Manganese alloys.
From a general point of view, the reductant is the chief cost item in carbothermic reduction process. Among these processes are the manganese alloys smelting. Ferro Silico Manganese smelting is an electrometallurgical carbothermic process. In such a process, besides its reducing characteristics, carbon is the main element responsible for the electrical characteristics of the burden.
Although some of the figures generated by simulation were confirmed in several industrial exploratory tests, a more detailed test program is advisable. Introduction The electrolytic cells employed in aluminum winning are lined with carbon that is in contact with the bath and metal and insulating material placed between the former and metallic shell. Through its operating life, the cathode lining changes chemically and physically, that leads to its expansion and deterioration. As a consequence a splitting process grows up and after a
Last, but not least, the ash content and composition have an influence on slag volume and characteristics and, depending on the nature of other components, on alloy quality and furnace operation.
New lining
,JOL
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Lining deterioration
Dismantled k
1 st cut (example) 2nd cut (example)
Figure 1 - SPL formation scheme [1]
275
In this paper the possibilities for employing the SPL's carbon fraction as a burden component on ferroalloys smelting is evaluated. Specifically, the selected ferroalloy is from Ferro Silico Manganese type.
Spent Pot Lining (SPL) Essentials [2] The aluminum pot lining is constituted of a series of components, each with its own purpose.
In a first approach, aiming to check the technical feasibility of that alternative, the main physical and chemical characteristics of the SPL are established and the resultant data will be "introduced" in the Ferro Silico Manganese furnace burden and the furnace behavior with this new component was analyzed. Although this first evaluation is a theoretical approach, it indicates if there are constraints that would impair this application. The next step, industrial scale trials would be essential.
Figure 2 shows the scheme of a typical lining. [5] As can be seen, the bottom lining is constituted of three kinds of non carbon materials and a fourth that are the carbon blocks. Among the non carbon materials, the berth is an alumina layer, the refractory are silico-aluminous bricks and the insulating layer is made of low density and low thermal conductivity bricks. Not characterized in the figure, the side lining is constituted of carbon bearing paste, silicon carbide and ceramic materials
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Figure 2 - Pot lining scheme (typical) Carbon blocks, constituting the top layer of the bottom lining, are by far the most important component of this system. The carbon block assemblies form the cathode of the electrolytic system.
Chemical Characterization SPL material comprises carbon and refractory materials, penetrated metal and bath components and minor amounts of aluminum nitride, aluminum carbide and cyanide salts. There is not an exact composition of SPL; it varies over a wide range, depending on several factors, such as lining design, pot age, etc.
Carbon blocks must meet several requirements: among them are high electrical conductivity, low reactivity and high mechanical strength.
The whole of SPL can be divided in two "cuts"; the first cut corresponds to the carbonaceous part (carbon blocks) and the part of the lining beneath the blocks assembly correspond to the second cut.
Depending on the type of the block, those properties can vary among a relatively ample range, in a new lining. However, during their operating life, the blocks absorb bath components and interact with it and the surrounding environment. Within this process, both the physical and chemical characteristics change. For the purpose of this paper, the properties to be considered are those existing at the end of the lining life.
Component Carbon Total Fluorides Free Alumina Total Aluminum Total Sodium Calcium Quartz P Sulfur
Table I shows typical composition of SPL, as well as those from the first and second cuts. According to what was mentioned before this is just a reference composition
Table I. SPL's typical com ponents and concentrations First Cut (carbonaceous) Total (56%) (100%) 54-64 6-20 0-15 5-15 5-12 0.5-4.0 0-6 0-650 g/t 0.1
33.1 15.7 22.3 15.1 14.2 1.8 2.7 0.3 0.1
Second cut (non carbonaceous) (44%) 18.2 4-10 10-50 12.6 12.0 1-6 10-50 0-300 g/t 0.1
and insulating material, whose properties are extremely variable, both geometrically, physically and chemically. Both cuts are impregnated with bath components and metal.
As suggested in the third drawing of figure 1, the first cut is constituted of big pieces of high resistance carbonaceous materials, from the former cathodes blocks. The second cut is a mix of paste, refractory
276
Although the whole of the spent potlining are together from the pot cathode, fortunately the two cuts are quite different and can be easily separated each other.
As will be seen later, the material that meets the most suitable properties, considering the burden characteristics of a Ferro Silico Manganese furnace, is the first cut, both from the chemical, physical and metallurgical point of view.
The photo of Figure 3 shows pot being dismantled. As can be seen, there is a sharp difference between the carbon blocks (1 st cut) and the other components (2nd cut)
On the other hand, a greater part of the second cut is constituted of loose or powdery material that would seriously impair furnace operation. This suggests that only the carbonaceous fraction (first cut) can be considered for the proposed application.
Figure 3. Section of the lining of an aluminum pot, during its dismantling (light line enhances carbon blocks) The number of reaction is extensive; for the sake of simplicity this paper only considers the most relevant.
In order to evaluate the metallurgical behavior the figures shown in Table I must be detailed, identifying the actual compounds present in the material, specifically in the first cut.
Figure 4 shows the evolution of the main compounds during the pot lifetime. For the purpose of the present work, the reference figures are those at the time of pot shutdown.
For that, the possible reactions between carbon lining an surrounding components, as well as their relative preferences must be known and analyzed.
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Besides these components there are the carbon and quartz. Both the former and the later are diluted during lining life. Based on these aspects, the SPL Carbonaceous fractions were calculated with the resultant values shown in the following list. Naturally, it must be pointed out that these figures can vary from plant to plant. Another remark is that the criolite and sodium aluminate were decomposed in its elementary compounds with the purpose of the metallurgical operation evaluation. SPL reference composition: C 55.98% A1203
7.18% Na20 22.70% Si0 2 1.55%
1.22% 5.00%
Electrical Properties Electrical resistivity is an important parameter both in aluminum and in ferro silico manganese smelting. In the first instance, the resistivity must be as low as possible. For the ferroalloy smelting the operation is carried out in a submerged arc furnace; the high temperature zone depends on the electrode tip
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277
position that depends on the voltage and burden resistance.
graphitization and impregnation. The first one improves the conductivity and the later increases the resistivity. Figure 5 shows this variation both in ambient and operating temperatures.
During its operating life, the aluminium smelting cathode block is submitted to two kinds of processes:
Cathodic lining - resistivity x time
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Figure 5. Cathode block - change of the resistivity during operating life [3] The negative slope in the first section of the curves suggests that during initial cathode operation that graphitization prevails over impregnation of bath species. The positive slope in the final section indicates that the block is already fully graphitized
and only impregnation occurs. Based on that it can be said that at the end of its useful life, the resistivity of a cathode block are nearly the same, doesn't matter if it was originally of amorphous carbon or fully graphitized types [3]
Reduction temperatures - selected oxides
300
600
900
1200
1500
1800
Temperature K
Figure 6. FeSiMn smelting scheme (easily reduced iron is not shown) and reduction temperatures
Production of FeSiMn is carried out by the simultaneous reduction of manganese and silicon, by carbon, in submerged arc furnaces (SAF), as shown schematically in figure 6.
MnO + C = Mn + CO, depends on its activity in slag; the higher the concentration of free MnO, the lower the reduction temperature. On the contrary, the higher the concentration of MnSi04, the higher its reduction temperature (fig 6).
Burden is constituted by manganese ore, manganese bearing slag, quartz (or quartzite) carbon source and fluxing agents. The manganese reduction temperature from its lower oxide, MnO:
Reduction of silicon from silica (Si02): Si0 2 + 2C= Si + 2CO, requires higher temperatures, as shown in figure 6. Therefore, in order to get a more intense reduction of Silicon, the Manganese reduction temperature must
Ferro Silico Manganese (FeSiMn) Essentials[6]
278
be higher, that means higher concentrations of silica in the slag. On the other hand, the more silica in slag, the more viscous is it. As far as high viscosities impairing the mobility of reagents, inhibiting the reduction reactions, one can conclude that the higher the concentration of Silicon in FeSiMn alloy, the more difficult the process (indeed, there is a limit above it that the process is not feasible).
the SPL characteristics presented above, two simulations [5] of the SAF operation with and without SPL in the burden were carried out. The purpose is to evaluate by comparison, the technical feasibility of the former. All other components were the same in both cases. The simulation was carried out by means of a thermochemical program previously developed for the evaluation of ferroalloys smelting procedures. Its validation was made by the comparison with actual operations [5,7].
There are several ways to determine slags viscosities. This can be done experimentally or applying empirical expressions or diagrams. In this paper the hereafter expression [4] was adopted because it considers species encountered in SPL and so its influence can be evaluated. This is a theoretical approach that should be confirmed experimentally, but for the purpose of this work, it is considered acceptable.
Side Reactions Γ3/71 Before going on with the simulation, a thermodynamic analysis was made, aiming to identify the most probable reactions and the resultant compounds. The behavior of the oxides is well known and is the same for both alternatives. The Na20 behavior can be compared to that of K 2 0 that is a very common component in manganese ores.
Ιημ = -12,0 + 0,G525(%Al2O3) + 0,088 l(%Si02)* 0,0409(%CaO)* - 0,042l(%Na20)* - 0,0194(%F)* + 14941/T where (%CaO)* = (%CaO) + (%MgO) + (%MnO) + (%FeO) + (%B203) (%Si02)* = (%Si02) + (%P205) + (%Ti02) + (%Zr02) + (%Fe203) (%Na20)* = (%Na20) + (%K20) + (%Li20) (%F)* = Z(% Fluorides)
On the fluoride side, NaF and CaF2 are quite stable and are incorporated in the slag. Aluminum fluoride (A1F3) may react with silica, silicon or lime, according to the following reactions 4A1F3 + 3Si02 = 3SiF4 + 2A1203 4AlF3 + 3Si = 3SiF4 + 4Al 4A1F3 + 3CaO = 3CaF2 + A1203
Though not being necessarily precise, the qualitative influence of the components on slag viscosity is well posed.
Considering the furnace temperature and the concentrations of the species, it was assumed that all A1F3 would react with Si0 2 , forming volatile SiF4 (first reaction) The formation of CF4 is not likely to occur.
Process Evaluation Based on the features of the FeSiMn smelting and on
Table II. FeSiMn smelting - Results of simulation - conventional procedure Chemical composition Inputs kg/1 alloy Manganese ore 1668 Alloy Slag 22.31% Mn rich slag 910 Si Si0 2 41.61% 334 Quartz 13.45% Fe A1203 23.20% Mn Limestone 230 63.55% CaO 18.19% Al Coke 369 0.00% MnO 16.62% CaF2 Petcoke 155 C 0.70% 0.00% SPL P 0.02% NaF 0 0.00% Outputs kg/1 alloy Electrical Energy Viscosity Alloy 1000 3762 kWh/t alloy 0.58 Slag 929 1
1
In the simulation with SPL it was considered that it was introduced substituting petroleum coke. As far as carbon content of petcoke (>90%) is much greater than that of the first cut of SPL (-54%), the substitution on a weight by weight basis implies in the charging of additional metallurgical coke, to achieve the required carbon. The mass balance of carbon was completed by means of metallurgical coke in order to keep the burden resistivity constant. Table III shows the results of this simulation.
Process Simulation Γ5Ί In the reference process, without SPL, it was considered a conventional operation, using manganese ore and manganese rich slag (from HCFeMn production), quartz and limestone. The reducing agent was a mix of metallurgical coke (70%) and petroleum coke (30%), which is a blend in manganese alloys smelting operations. Table II shows the results of this simulation
279
Table III. FeSiMn smelting - Results of simulation - procedure with SPL Chemical composition Inputs kg/1 alloy Slag Alloy Manganese ore 1678 Mn rich slag 22.31% 39.30% Si 910 Si0 2 Quartz 334 13.45% 23.61% Fe A1203 Mn 63.57% CaO 17.18% Limestone 230 Al 0.00% MnO 15.70% Coke 428 CaF2 0.66% 0.25% Petcoke C 0 P 0.02% NaF SPL 3.60% 155 Viscosity Electrical Energy Outputs kg/1 alloy 3793 kWh/t alloy 0.47 Alloy 1000 984 1 Slag 1 i
1
From the electrical point of view the burden must have suitable resistivity by assuring optimum positioning of electrode tips. For that it would be desirable that the SPL resistivity be similar to that of the substituted reducing agent. Figure 7 shows the change of the resistivity with temperature for charcoal, coke and SPL[2,3 and 6]
Comparing the results from the two tables it can be seen that all but the reducing agent, components proportions are practically the same in both cases. The slag weight is greater when SPL is employed due to its ash content being greater than that of petcoke; this also implies greater energy consumption. On the other hand the sodium and fluoride content of SPL lead to a lower slag viscosity, what enhances the reaction efficiency.
Resistivity x Temperature 350
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Figure 7. Resistivity dependence with temperature - Selected reducing agents According to the above figure, SPL resistivity is the lowest when compared to coke and charcoal. However, is must be pointed out that the figure from SPL corresponds to one resistivity whereas those from coke and charcoal are bulk resistivity's. This suggests that the SPL bulk resistivity, considering the same particle size of that as coke, would be very close to that of this later.
References 1. P. von Krüger, P. Possibilidades de valorizagδo e inertizagδo do Revestimento Gasto de Cubas de aluminio (RGC) . l°Congresso Internacional da Industria do Aluminio - ABAL S. Paulo 2000 2. M. S0rlie, and H. A. 0ye, Cathodes in Aluminium Electrolysis 2nd Edition. Düsseldorf: AluminiumVerlag Gmbh, 1994. 3. M.S0rlie, and H. Grau, Properties Changes of Cathode Lining Materials During cell Operation. Las Vegas Light Metals 1995 TMS - pgs 497 - 506. 4. LA Silva.et al, .Some Empirical Correlations for Evaluating Viscosity of Molten Mold Flux for Continuous Casting of Steel. 1st Process Metallurgy and Structural Engineering of the Int'l Materials Research Congress, Cancun, Mexico - 2000 5. P. von Krüger, Metalurgia - Fundamentos Teoricos e suas Aplicagöes na Redugδo do Aluminio. Apostila de Curso. (internal publicashion- ALUMAR, ALBRΔS e VALESUL, 1995,1997,1998. 6. A. Lucio et al Metalurgia dos Ferro-ligas. Internal Publication - Depto de Engenharia Metalurgica EE/UFMG-1978 7. P. von Krüger, Internal Reports - not published
Concluding remarks The results of the simulations and other features carried out in this work suggest that, in principle, the employ of SPL's First Cut as an auxiliary reducing agent (with additional fluxing characteristics), is a technically feasible possibility. This is a preliminary conclusion based on simulations that must be confirmed by industrial tests and that impact on other parameters such as environment and economics must be investigated.. Initial exploratory industrial tests were made and the results, although essentially qualitative, are promising.
280
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
REDUCTION OF PFC
EMISSIONS AT POT LINE 70 KA OF COMPANHIA BRASILEIRA DE ALUMINIO
Henrique Santos1, Danilo Melo1, Jocimar Calixto1, Jefferson Santos1, Joäo Miranda1 Companhia Brasileira de Aluminio; 347 Moraes do Rego; Aluminio,Sao Paulo, 18125-000, Brazil Keywords: PFC, Anode Effect, Energy Consumption Abstract Following world tendency the Primary Aluminum Industry has been committed to the reduction of PFC emissions in order to reduce the greenhouse effect. Anode Effects is a significant source for such emissions. The Companhia Brasileira de Aluminio (CBA) pot line is equipped with 70 kA VSS Montecatini technology from the end of 60's with side feed pots, has been making efforts and developing actions to improve work practices and the operating process with the aim of reducing both the duration and frequency of anode effects. The anode effect, which was used for decades to assess the process behavior, ensuring their performance and preventing future problems, has been replaced by indicators supported by technological developments and the process computer improvement. Projects were developed and implemented as follow: Liquids control, Resistance control, Feed control, Anode/Cathode control and mainly by team work practices through operating training.
Figure 1. The Companhia Brasileira de Aluminio This paper was based on investigations and testing conducted at the 70 kA Pot Room that has 174 electrolytic cells with side feeding and went through the process of increasing current from 65 kA to 70 kA, from 2002 to 2005.
This paper shows achievements for projects cited, that caused significant impacts by reducing the duration and frequency of anode effects inline with world best practices for VSS technology. A reduction of 66 % of PFC emissions was achieved, from 1.2 t C02e per t Al in 1998 to 0.41 t COs e per t Al in 2009 calculated by Slope Method.
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Introduction The Companhia Brasileira de Aluminio, a company of Votorantim Metais, has an installed capacity to produce 475,000 tons of primary aluminum per year, with nine reduction lines totaling 1,508 electrolytic cells of VSS technology operating at 70 kA, 90 kA and 127 kA.
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Figure 2. Pot Room 70 kA form CBA The anode effect, previously considered essential for evaluating the operational performance of the pot line, has been losing its importance over the years as a best practice for pot operation. The systematic suppression of the anode effect was modified in order to reduce the duration from 1.86 min per cell day to 0.63 min per cell day (Figure 3), and thefrequencyof anode effect from 1.2 AE per cell day to 0.3 AE per cell day (Figure 4). The logic of feeding, which is held at intervals of four hours, has been changed
281
to reach 39 consecutive feeds. But for these reductions were possible, actions were necessary to reduce the number of anode problems, improve the liquid level control and operating procedures.
Figure 5. The Resistance Control System
The next steps for achieving further reductions targeted r improvement of routine operational and process control, because it was necessary to reduce the number of operational problems associated with anode quality.
Figure 3. Average Anode Effect Minutes per Cell Day
Reducing the number of anode problems
Development
The high number of operations related to problems with the anode was a hindrance to continue reducing the number of AEs to the levels proposed. In this, actions were taken in order to obtain an anode quality and stable operation. (See Table I and Figure 6)
Deployment of the reduction control system Up to 1998 the voltage control system was manual hampering an accentuated reduction of anode effects. The implementation of the resistance control system was the first step that led to a significant reduction of anode effect going from 1.1 AE/cell day to 0.7 AE/cell day in 2 years as shown in figure 4.
Table I. Table of anode improvement actions Problem Action Result Excessive anode Reduction of the Reduction of exposure below the size of the gas exposed anode gas collector duct. collector; from 25 cm to 18 cm Reduction of the depth of the pot shell. High number of Improvement of Reduction of fissure in the anode system of studs fissures in the replacement; anode;
Anode Effect Frequency
Increased frequency of stud recovery; Year
Anode quality
Figure 4. Evolution of anode effect frequency at Pot Room 70 kA This result is mainly due to the development of control parameters that became effective in the prediction of anode effects, providing an effective method that would allow reducing the number of anode effects to the levels achieved.
282
Changing composition of anode paste.
Reduced leakage of paste; Improved of the efficiency of stud replacement. Improvement of paste and anode quality.
to kill the AE.
Number of anode problems versus anode height 144
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The results achieved in reducing the emission of PFCs were significant, reducing emissions of 1.20 tonnesC02-e/tonneAl in 1998 to 0.41 tonnesC02-e/tonneAl in 2009 close to the estimated reduction 93% based on 1990 which provides for 0.34 tonnesC02-e/tonne Al in 2020[3] (Figure 7).
Figure 6. Reduction of anode problems Liquids control To achieve an extended period of time without the occurrence of an anode effect and to avoid occurrence of operational problems such as sludge, delay, and others, it was necessary to ensure a stable level of bath as well as an efficient bath chemistry control.
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Table II. Table of improvements of liquids control Problem Action Result Bath level control Implementation of Stability of bath daily measurements of level bath's level after metal tapping Bath chemistry Weekly sampling of Improvement of 1 control bath to measure superheat chemistry. control
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Figure 7. Trends in PFC Emissions per Ton Primary Aluminium Production. Source: 2008 Global Anode Effect Survey Results [3]
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To ensure the proposed results were achieved, actions were implemented with aim of improve operational practices such as to meet the operational standards, which were important to reduce the kill time of the AE. However, the most effective action was the operational training, which allowed the dissemination of knowledge of the automation tool, such as interpreting graphs of resistance and noise behavior, which allowed more rapid detection and efficient operator response, as well as standardization of the actions and awareness of the importance of each task in building the final result. (See Table III)
2 0 0 8 PFC E m i s s i o n s B e n c h m a r k i n g ( V S S )
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Table III. Table of actions of operational practices Problem Action Result Excessive time to Information via radio to Reduction of kill the AE operator of the crust time to kill the break vehicle when an AE. AE occurs;
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PFC Emission· (tonnes C02-*q/tonn« AI)
Figure 8. PFC Emissions Benchmarking (VSS) - Source: International Aluminium Institute - Anode Effect Benchmarking 2008 m The result is due to the reduction in the number of anode effects per year in the order of 75,000 AE/ year in 1990 to 19,000 AE / year in 2009, a reduction of 65000 AE/year which equates to a reduction of about 3.0 GWh/year in energy consumption taking
Installation of visual and audible alarms that provide the operators the information of time
283
into account only the anode effect frequency and a reduction of 66 % of PFC emissions from 1998 to 2009.
Conclusion The biggest challenge for succeeding in reducing PFC emissions was the preparation and awareness of staff to use new tools for operational analysis, to predict and deal with problems without the occurrence of anode effect. The result achieved has motivated the team to strive to improve even more, with the assurance that the problems occur, but can be predicted and studied.
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References
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1. Guimaraes, C. T., et al, Redugδo no Indice de Efeito Anodico Nδo Programado na Redugδo 4 da Albras, (Paper presented at I Seminärio de Reducäo do Aluminio, Säo Paulo, 2003).
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2. Anode Effect Benchmarking 2008, International Aluminium Institute
Anode Effect Frequency (ano<$e effects per
Figure 8. Anode Effect Frequency - Source: International Aluminium Institute - Anode Effect Benchmarking 2008I2]
3. Marks J., Bayliss C, 2008 Global Anode Effect Survey Results (Light Metals 2010), p259. 4. Results of the 2008 Anode Effect Survey, International Aluminium Institute
2008 AEM CD Benchmarkinj{VSS)
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Figure 10. Average Anode Effect - Source: International Aluminium Institute - Anode Effect Benchmarking 2008[2]
Erreision of C02 eq (AE); Anode Effect Min; Anode Effect Frequency 2,00-r
Varfafata - · - t C02 eq (AE)^Al - * - Anode Effect Minutes per Cell ··# Anode Effect Fr»qu«ncy
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1999 2000
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Figure 10. Evolution in PFC emissions, AEM and AEF
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Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
Towards Redefining the Alumina Specifications Sheet - The Case of HF Emissions 1
Linus M. Perander, 2Marco A. Stam, Margaret M. Hyland, Barnes B. Metson ^ight Metals Research Centre, The University of Auckland, New Zealand 2 Aluminium Delfzijl B.V., Delfzijl, The Netherlands
Alumina quality, HF emissions, Dry-scrubbing, Gibbsite, Residual hydroxide, Pore size distribution available for adsorption and the adsorption capacity, however it could be expected that the pore size distribution also plays an important role in providing access to internal reaction sites. The porosity and surface area arises from the structure of the transition aluminas which dominate smelter grade aluminas and reflect the incomplete conversion of gibbsite to alpha alumina in the calcination stage [5].
Abstract For smelting applications, alumina quality is typically defined in terms of chemical and physical properties, with emphasis on impurity elements, surface area, moisture content, particle size distribution and attrition index. However, these properties fail in prediction of the true HF generation potential, as well as the real capacity for HF removal in the dry scrubbers. Using plant measurements and additional laboratory characterization of a number of alumina samples a broadening of how alumina quality is specified is argued for. Measurements of the residual gibbsite/boehmite content and the pore size distribution, coupled with characterization of the alumina microstructure, can be used to predict and understand the generation of HF during feeding and dissolution as well as the ability to capture HF in the dry scrubbers.
Residual hydroxyls are an integral part of the transition alumina structures (and often reported as LOI). The less calcined the alumina, the higher surface area, and correspondingly, the higher the level of residual hydroxides in the structure. Together with other OH sources (such as gibbsite), these represent the main source of HF formation in the electrolyte [6]. In the cell HF is formed in the electrolytic process when the released water (or OH) reacts with the electrolyte (NaAlF3). It is then easy to see that there is a conflicting relationship between generation of HF from residual hydroxide and capturing the HF by increasing the surface area which inevitably results in more structural hydroxyls and increased HF formation in the first place.
Introduction For both regulatory and process stability reasons, reductions in fluoride emissions from aluminium smelters are an essential part of process improvement. The modern, injection type, dry scrubbing system comprises of a reactor (where alumina is injected and comes in contact with the collected cell off gases) and a separate filtration system (where the reacted alumina and other particulates are separated from the gas stream). These reactors are designed to enable good mixing between gas and solids and can be operated at very high efficiencies with over 99.5 % of the gaseous and particulate fluorides typically being captured [1]. With good pot hood collection efficiency, and the dry scrubber recovering and recycling most of the particulate and gaseous fluorides, the operation may almost be regarded as a closed loop between the cells and the scrubber [2].
The surface adsorbed water (represented by the MOI values) on the other hand are rapidly flashed off when the alumina is fed into the electrolytic cell [7]. This may actually be beneficial for the dissolution process as it helps with the dispersion of the material [8]. In this paper four alumina samples with very similar specifications, but which demonstrate very different HF emissions levels, are considered. The measurements currently reported on the alumina specification sheet fail to explain this HF generation potential or accurately predict dry-scrubber performance. By using low-cost, and often already existing, laboratory equipment measurements that better account for this may readily be obtained and used to better predict and control the operation, particularly with regards to emissions. This argues for some redefinition of the alumina specifications sheet, particularly when such specifications are used for example in prediction of emissions.
The chemistry and mechanism of the dry scrubbing process (i.e. the adsorption and reaction of HF with alumina) was explored in depth by Gillespie [3, 4]. Gillespie argued that since the fluoride adsorption capacity is related to both the specific surface area of the alumina and the relative humidity during the adsorption, the reactions must occur in a surface process. The author presented a mechanism which involved several steps, starting with the adsorption of water on the surface of the alumina followed by HF adsorption and acidification of this surface layer which would dissolve the alumina surface and form A102" and AlO2" species and finally the precipitation of oxy- and hydroxyfluorides. Most importantly the reaction was shown to be irreversible under dry scrubbing conditions (temperature and atmospheric), an important finding which enabled new developments of the dry scrubbing process and control strategies to be made.
Experimental The samples examined in this study were obtained during the unloading of the respective alumina shipments and directly from the conveyor belt going to the alumina storage silo. There is only one alumina silo on site which rules out alumina blending as a source of variation. The GTC (Gas Treatment Centre) inlet and outlet HF concentrations are monitored continuously using NEO laser-based gas monitors. This is an infrared single-line absorption spectroscopy measurement which provides a continuous reading of the HF content in the gas stream across the laser path. One laser is mounted in the stack directly after the GTC, and one in each inlet channel to the GTC. The inlet values reported here are averages between these two inlet channels.
Although the mechanisms for the reaction between HF and alumina (in the dry scrubbing process) are relatively well understood, the role of alumina microstructure and porosity is less clear. There is a direct relationship between the surface area
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TGA experiments were conducted to determine MOI and LOI values, using a Shimadzu TGA-50 apparatus. Approximately 35 mg of each sample was heated, in an open platinum crucible in air, with a constant heating rate of 20 °C min"1 from ambient temperature to 1100°C. A baseline curve was also produced, and subtracted from each sample curve, by recording the weight change when an empty platinum crucible was subjected to the same heating conditions as above.
Table 1. Alumina properties as reported on the specifications sheet for the 4 alumina samples used in this study.
Values for MOI and LOI were obtained based on thermo gravimetric data. As per the ISO standard, the ranges for MOI and LOI were taken as 50-300°C and 300-1000°C, respectively [9]. It should however be pointed out that according to the standard measurement for MOI and LOI values the experiments are performed with a retention time of 2 hours at 300 and 1000°C. The results presented here may therefore not correlate directly to measurements performed according to standard methods but can be used to compare the samples and for calculating the water and hydroxyl content.
Alumina A
Alumina B
Alumina C
Alumina D
LOI (300-1000
0.59
0.51
0.79
0.91
68.0
66.3
75.6
73.0
N/A
N/A
4.0
5.0
1.6
1.2
1.9
1.6
12.0
13.0
4.8
10.87
Bulk Density kgdm-1
1.04
N/A
1.00
0.98
Note that the parameters omitted from the table were not reported on the specifications sheets Table 1 indicates that aluminas A and B have relatively similar parameters (in terms of MOI, LOI, BET-SA and fines content) which is also to be expected as these samples represent different shipments from the same alumina refinery. Aluminas C and D are different compared to A and B, as they both have higher specific surface areas and LOI values. Alumina C also has a significantly lower attrition index than the rest of the samples.
Nitrogen porosimetry experiments were conducted to evaluate the specific surface area and pore volume, as well as pore size and pore size distribution. The measurements were performed on a TriStar 3000 apparatus. Approximately 1 gram of sample was used for the analysis. For the measurements the sample is initially degassed under a nitrogen purge at 120°C for 12 hours. Note that in standard analysis different degassing temperatures may be used. The temperature of 120°C was chosen to avoid the reaction of any gibbsite or boehmite present in the sample. The results obtained here thus represent the whole sample as received (without changing its phase composition). The measurements were conducted under liquid nitrogen temperatures and using nitrogen as the adsorbed gas. The BET method [10] was used to evaluate surface area and the BJH method [11] was used for the pore volume and pore size and distribution evaluations.
Based on the parameters reported in table 1 one could expect aluminas A and B to result in similar HF emissions and C and D to perform equally well but perhaps with a higher background HF emission level due to the higher LOI values. Some of the additional HF generated due to the larger amount of structural hydroxyls present in aluminas C and D could be expected to be offset by the significantly higher BET surface areas. However, as Table 2 indicates, these aluminas did not perform as expected. When alumina A was used the fluoride emissions increased by approximately 60 %, compared to alumina B. Similarly, alumina D resulted in twice the emissions compared to alumina C; far more than what could be expected based on the relative LOI values.
X-ray diffractometry was performed on a Rigaku apparatus using Cu K-alpha X-rays. The analysis was performed using a 2Θ angle of 10-80° and a step width of 0.02°. Rietveld refinement [12, 13] of the diffractograms was then performed using the FuUProf software [14]. The fine sub 40 micron size fraction was separated from the bulk by sieving and analysed separately.
Table 2. Associated hydrogen fluoride emissions (total gaseous per hour andfluorideper cubic meter) and scrubber efficiencies. Fluoride Emission mgm"3 Total Gaseous HF Emission kg h'1 Scrubber Efficiency (%)
X-ray diffractometry with Rietveld refinement allows the alumina phases present in smelter grade aluminas to be quantified. However this technique does not allow the amorphous phases such as rho- or chi-alumina that may affect the performance of the alumina to be measured directly [15, 16]. The highly crystalline phases such as alpha alumina and gibbsite produce sharp, welldefined peaks whereas the less crystalline transition alumina phases produce broader and more diffuse peaks. Several studies report on how the alumina phase composition influences the smelter performance (see for example [17] and references therein).
Alumina A
Alumina B
Alumina C
0.86
0.61
0.56
Alumina D 1.13
1.29
0.92
0.60
1.21
N/A
N/A
99.70
99.48
Since HF is generated in an electrochemical reaction between the bath and any hydroxyl species (H 2 0, OH", NaOH...), a reasonable measure of the HF generation potential is a quantification of the water and hydroxyl content of the aluminas. This is typically done in so called Loss on Ignition and Moisture on Ignition measurements [9], but may also be obtained using thermogravimetric analysis techniques. As will be shown the LOI measurement (300-1000°C) does not account for all the OH associated with the alumina.
Results and Discussion Table 1 displays a selection of properties from the alumina specification sheets for the 4 alumina samples discussed here. It is recognized that other properties are of importance for understanding alumina performance; however, these properties have been chosen as they are most often associated with dryscrubbing and HF generation, which is the topic of this paper.
Figure 1 shows the obtained thermograms for aluminas A and B as well as the thermogram for a standard industrial Bayer gibbsite. It can be seen that the relative weight loss is larger for alumina A, which indicates higher MOI and/or LOI values, and as can be seen from Table 1 the reported LOI value was slightly higher for
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sample A. The higher LOI value indicates more structural hydroxyls (OH groups retained after the calcination in the Bayer process) and thus a greater HF formation potential. What may also be seen is a sharp weight loss at around 270°C for both aluminas; this indicates the presence of residual gibbsite (or un-calcined material) in the alumina. When heated, gibbsite starts to transform to boehmite and/or transition alumina at around 250°C (depending on the reaction conditions). In the thermogram below (figure 1, blue line) this can be seen as a rapid loss of mass (starting at around 250°C) as the OH-groups in the gibbsite reacts to form H 2 0. The residual gibbsite in the alumina has the potential to react with the electrolyte to form HF, but is obviously not accounted for by the LOI (300-1000°C) value. The MOI (50-300°C) value is typically associated with, and dominated by, the surface water/moisture, although, when gibbsite is present it will be included in the MOI value.
1.830% MO! (50-250C) (0.6880mg)
U
2.061% MOI (50-3G0C) (0.7748mg) 0.9901% LOI (250-1100C) (0.3723ing)
400
600 800 T e m p e r a t u r e (°C)
1000
|-0.01£g
I
1200
Figure 2. Thermogram for alumina A with the MOI and LOI limits indicated. The blue line is the derivative weight loss (wt-% °C"1) which is a way to better identify specific phase transitions. To estimate the HF generation potential of alumina, the LOI value provides a reasonable estimate of the amount of reactive OH, particularly when the extended range (250-1100°C) is considered. However, when gibbsite and/or boehmite are present in the alumina these components should also be quantified and reported. For this a number of methods have been suggested, such as: DSC [18] Near Infrared Spectroscopy [19] and quantitative XRD, employed in this study and elsewhere [15, 20]. Although readily available, these methods have not been widely adopted in the industry for routine characterisation. Table 3 provides this comparative data for the samples considered in this study.
400
600 Temperature (eC)
Table 3. Additional characterisation data for the four aluminas: MOI (to 300°C) and LOI (300-1000°C and 250-1100°C) values (by TGA), surface area, average pore size and total pore volume (by nitrogen porosimetry), and gibbsite and alpha alumina content (by XRD with Rietveld refinement).
Figure 1. Thermograms for the two of the alumina samples: alumina A (red dashed line, left y-axis) and alumina B (green line, left y-axis) as well as a reference gibbsite sample (blue solid line, right y-axis). Thus, in order to account for the gibbsite as a potential source for HF formation, the temperature limits for the MOI and LOI values need to be re-evaluated. When gibbsite is present in the alumina a more appropriate measurement of the surface water/moisture (less likely to be transported into the electrolyte and react to form HF) is perhaps the weight loss between 50 and 250°C. An LOI value from 250 to 1100°C would then account for both the residual OH and the gibbsite, the primary HF sources. The 'new' limits can be seen in Figure 2. Note, however that although this is perhaps a more useful way to specify the MOI/LOI limits when gibbsite is present, the 'exact' limits for when surface water and/or residual OH is removed cannot be specified as these overlap to some extent. Also, as will be discussed later, not all H 2 0 released from the gibbsite is likely to react with the bath to form HF (the gibbsite dehydroxlation reaction at high temperatures is very rapid and some water will flash off without coming in contact with the bath).
MOI (to 300 °C) wt-% LOI (300-1000
Alumina D I
Alumina A
Alumina B
Alumina C
2.061
1.954
2.478
2.645
0.722
0.682
1.014
1.052
0.990
0.917
1.523
1.448
68.1
69.0
70.7
76.1
5.78
5.70
8.78
11.20
0.210
0.215
0.193
0.207
0.93
0.56
1.98
2.65
8.1
7.6
3.9
6.5
The inevitable effect of residual gibbsite is that a large amount of water (as OH) enters the cell with the alumina. As can be seen from the thermogram in figure 1, approximately 35 wt-% of gibbsite will react to form H 2 0 when heated. This will occur very rapidly when it comes in contact with the electrolyte, inducing a phase transition into the transition aluminas. The phase transition, will be competitive with dissolution but the very rapid formation of large amounts of water vapour will create a volcano effect, carrying fine particles out of the cell and potentially increasing dusting.
287
Assuming that all OH in the gibbsite reacts to form H 2 0 and A1203 (according to reaction 1), 2AI(OH)3(s) + Ei -> Al203(s) +3H20(g)
IM
where ¸χ is the reaction energy, a theoretical increase in the energy demand can be calculated. Based on the amount of water that is evolved the additional (apart from the structural hydroxyl, measured by the LOI tests) HF generated due to the gibbsite in the SGA can also be calculated. To calculate the increased HF generation capacity due to gibbsite in the alumina it is assumed that all H 2 0 will react according to reaction 2: 3H20(g) + 2AIF3(g/s) -> 6HF(g) + Al203(diss)
For comparison two measured HF emission values for aluminas with known MOI and LOI values are included. Not surprisingly the measured values are slightly lower than the calculated ones, but should still provide a reasonable estimate of the HF generation capacity. The measured values were sourced from the literature [7]. In the study dry and 'hydrated' alumina samples were fed to a plant cell and the resulting HF emissions measured. It can be seen that the MOI, or moisture content, has little or no contribution to the HF emissions. Figure 4 displays HF measurements (GTC inlet and outlet levels) when aluminas C and D were used. The higher inlet GTC HF concentration when alumina D was used may be attributed to the slightly higher gibbsite content. However, the gibbsite content alone could not explain the significantly poorer performance of alumina D in terms of HF removal. To understand the performance of these aluminas the porosity and microstructure needs to be evaluated as well. A direct relationship between specific surface area and HF adsorption capacity has been reported [3]. High specific surface area is often considered advantageous in terms of optimal dry scrubber performance, and can often be the only way to deal with increased fluoride emissions. There is also a relationship between surface area and residual hydroxyl content. The residual hydroxyls are retained in the crystal lattice of the transition alumina phases due to incomplete calcination of gibbsite. In other words, higher surface area stems from more of the low order transition aluminas (gamma, chi, rho alumina) which also have more residual hydroxyls retained in the lattice and hence also have a higher HF generation potential.
121
In practice it is not likely that all H 2 0 will react to form HF (as some will obviously exit the cell with the off gases). It should be noted that the HF lost to the environment (due to the incomplete capture) constitutes a direct loss of bath, which will have to be supplemented. Depending on the gibbsite content this may have a significant economic impact if particulate and gaseous emissions are not fully captured. XRD analyses revealed that 0.93 and 0.56 wt-% of alumina A and B, respectively, is gibbsite. Assuming that 1.92 kg of alumina is needed to produce 1 kg of aluminium metal it is then calculated that 18 and 11kg of gibbsite is introduced, for alumina A and B, respectively, to the cell per metric ton of Al produced. Based on the reactions above we can then also calculate that the 0.93 and 0.56 wt-% of gibbsite in the aluminas results in an additional 8.2 and 5.0 kg of HF being formed per ton of Al metal, for alumina A and B, respectively, on top of the 'background' HF from the structural hydroxyls (reflected in the LOI value). The additional HF due to the presence of (0.93 wt-%) gibbsite is about 20% of the total (theoretical) HF. Although these are theoretical maximum HF levels (based on complete reactions) it can be seen that even small amounts of gibbsite in the alumina have the potential to significantly increase the total HF burden.
300
GTC Inlet HF Concentrations and HF Emissions using Alummas C and D (5 days data)
Calculated and Measured HF emissions from structural hydroxyl (LOI) in kg(HF) per T(AI) 50 45 40
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Calculated HF Alumina A + OH from 0.93% Gibbsite
♦
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Calculated HF Alumina A
0.8
,
0.9
,
2000
Figure 4. Dry scrubber (GTC) inlet HF concentrations and corresponding HF emissions when using aluminas C and D (5 days data).
,
The pore size distribution and mean pore size, as well as surface area are a measure of the extent of the calcination reactions [5]. As the calcination reactions progress, the pore size grows, but pores coalesce, causing the surface area to decrease as a result. A narrow pore size distribution in a range that allows HF molecules to be easily transported to internal surfaces and reactive sites is desired and results from a well controlled calcination process. Pore size distribution is also an effective way to identify if the
LOI (wt-%)
Figure 3. Theoretical HF generation capacity (calculated based on equations 1 and 2) and measured HF generated by structural hydroxyls in the alumina (numerical values sourced from ref. [7]). In figure 3 calculated (solid line), based on reaction 2, HF generation capacities are plotted as a function of their LOI values.
288
alumina is blended. A mixture of two differently calcined aluminas, or an alumina containing under- or over calcined material will result in a broad pore size distribution or a distinctly bimodal distribution. For a well-controlled calcination process, a narrow pore size distribution centred around 6-8 nm is typically observed, and desired (see figure 5). Higher calcination temperatures or longer calcination times usually result in an increased pore size and highly over-calcined material (more thetaalpha alumina) shows up as a peak at >10 nm pore sizes. Material with a lower degree of calcination results in a lower average pore size and generally shows as a peak in the <2 nm pore size range on the distribution plots. Poor control of the calcination process results in a broader peak in the 4-10 nm pore size range. +
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Figure 6. More than one way to achieve a surface area target. Alumina D displayed a distinctly bi-modal pore size distribution, indicating the presence of under-calcined components as well as over calcined material. From an operational viewpoint the blend of over- and under calcined material is detrimental for a number of reasons. As mentioned previously the low order transition aluminas contain more residual hydroxyls (confirmed by the LOI measurements) resulting in more HF generation upon dissolution. At the same time, the fine pore size is likely detrimental for HF removal. The narrow pores restrict access to internal porosity and readily become blocked (when HF reacts to form oxy-fluorides). This further restricts access to internal sites, thus reducing the capacity and rate of HF absorption. Moreover, the over-calcined components in the sample are likely to have a poor dissolution rate. Thus, the currently used surface area specification, derived under equilibrium conditions is not ideal for predicting scrubbing efficiency and it would be helpful to extended this to include information about the pore size distribution. This information is readily available using most of the commercial surface area instruments by examining the nitrogen adsorption isotherms using the BJH method.
10.0 Pore Diameter (nm)
Figure 5. BJH pore size evaluations for alumina C reveals a 'normal' pore size distribution. As reported in table 3, the surface area measurements revealed that alumina D had a slightly higher surface area than alumina C (76.1 and 70.7 m2/g, respectively) which could be assumed to improve the dry-scrubbing efficiency and offset some of the additional HF that is likely to be generated by the higher LOI content. However, when the pore size distribution plots are examined (figures 5 and 6) some interesting differences between the two samples can be seen. Alumina C shows one peak at a pore size of approximately 8 nm whereas alumina D clearly has a bimodal pore size distribution with one peak centred below the 1.0 nm pore size limit of this technique, and the other centred at a much larger pore size of approximately 12 nm. This would indicate that alumina D contains a large quantity of undercalcined material and that the rest of the sample is over-calcined (rich in theta and alpha alumina). As the small pores contribute more to the total (or average) surface area the alumina D therefore has a larger nitrogen specific surface area. Note that by controlling the calcination conditions, surface areas in excess of 350 mV 1 may be achieved.
Conclusions Plant measurements and additional laboratory characterization of a number of alumina samples was used to highlight some of the shortcomings of the alumina specifications sheet when it comes to understanding alumina performance, particularly with regards to HF generation and emissions as well as feeding and dissolution characteristics. It was demonstrated that residual gibbsite in the alumina results in increased HF levels (which coupled with poor hooding/collection efficiency results in an additional loss of electrolyte) and feeding perturbations. Therefore, knowing the amount of gibbsite in the alumina will help to predict and control the process response to the alumina quality. Several low cost options for determining the gibbsite content in alumina exist (XRD, NIR, DSC or TGA for example) and should be adopted more widely.
289
Nitrogen porosimetry and investigations of alumina microstructure revealed that the low order transition aluminas (gamma, chi, rho alumina) have a limited pore size and plant measurements indicate that this is detrimental for the dryscrubbing process. It seems that the narrow pores, although resulting in large average surface areas, may become blocked thus restricting access to internal surfaces which effects the HF adsorption capacity adversely (at the same conditions). These observations indicate that a pore size distribution specification would be a better requirement for alumina quality (than a single surface area target) as this ensures good scrubbing and may even be energetically favorable in the alumina refinery. The pore size distribution may readily be measured through nitrogen adsorption techniques using existing equipment both in the smelters and alumina refineries.
12.
Acknowledgements
16.
The authors would like to thank Anita Folkers and Karin Bouwmeester for their assistance with the alumina samples and HF measurements, the interpretation of the results and their contributions to this manuscript.
13. 14.
15.
17.
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3. 4. 5. 6. 7.
8. 9. 10. 11.
Well wood, G.A., The practice of dry scrubbing. Light Metals (Warrendale, PA, United States), 2001: p. 371377. Zhang, W., X. Liu, P. McMaster, and M. Taylor, Modeling of impurity balance for an aluminum smelter. Light Metals (Warrendale, PA, United States), 1996: p. 405-11. Gillespie, A.R., Mechanistic Studies of HF Adsorption on Alumina. PhD Thesis, The University of Auckland, New Zealand, 1997. Gillespie, A.R., M.M. Hyland, and J.B. Metson, Irreversible HF adsorption in the dry-scrubbing process. Journal of Metals, 1999. 51(5): p. 30-32. Perander, L., Evolution of Nano- and Microstructure During the Calcination of Bayer Gibbsite. PhD Thesis, The University of Auckland, New Zealand, 2010. Patterson, E.C., Hydrogen Fluoride Emissions From Aluminium Electrolysis Cells. PhD Thesis, The University of Auckland, New Zealand, 2002. Hyland, M., E. Patterson, and B. Welch, Alumina structural hydroxyl as a continuous source ofHF. Light Metals (Warrendale, PA, United States), 2004: p. 361366. Haverkamp, R.G., Surface studies and dissolution studies of fluorinated alumina. PhD Thesis, The University of Auckland, New Zealand, 1992. Aluminium oxide primarily used for the production of aluminium - Determination of loss of mass at 300° C and 1000°C. ISO 806:2004. Brunauer, S., P.H. Emmett, and E. Teller, Adsorption of gases in multimolecular layers. Journal of the American Chemical Society, 1938. 60: p. 309-19. Barrett, E.P., L.G. Joyner, and P.P. Halenda, The determination of pore volume and area distributions in porous substances. I. Computations from nitrogen isotherms. Journal of the American Chemical Society, 1951. 73: p. 373-80.
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Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
DESIGN OF EXPERIMENT TO MINIMIZE FLUORIDE AND PARTICULATE EMISSIONS AT ALUMAR Eliezer Batista1, Edson Montoro2, Paulo Miotto2, Luciano J. P. Souza2 Department of Potrooms and Technology 1 Alcoa Technical Center, 100 Technical Drive, Alcoa Center, Pa 15069 2 Consorico de Aluminio do Maranhäo - Alumar, BR 135, Km 18, Säo Luis, Brazil, Zip code 65065-604 Keywords: Fluoride Emissions, Particulate Emissions, Design of Experiment
was decided to make a full factorial experiment 2k at Alumar to find out the main factors and the size of their impacts. The objectives of this work are to develop statistical models to predict the impacts of operational activities on fluoride and particulate emissions and recommend actions to minimize them.
Abstract Most of Aluminum plants have been struggling to minimize the fluoride and particulate emissions at the main source, which is the pot rooms, in order to reduce the environmental impacts. Nowadays, this challenge has been more difficult to be accomplished because of some reasons, such as: amperage increase, alumina quality deterioration and pot room expansions. Alumar, one of Alcoa's units, following a corporate vision, is continuously searching for alternatives to eliminate or reduce the environmental impact caused by operations. Many actions and studies are in place currently. This full factorial experiment 2k was done with the aim of identifying the main factors and their impacts on fluoride and particulate emissions. The statistical model is showing that the fluoride emission has been affected mainly by Pot Draft, Pot Dressing, and the Usage of Compressed air for Housekeeping with R2 at 82%, and for particulate at 58%. Based on the models, certain actions were recommended to minimize both of these emissions. In addition, this paper describes, step by step, how this kind of experiment can be applied to the Aluminum industry.
Experimental This study was started up at Alumar with a brainstorming of the possible contributors of fluoride and particulate emissions in the potroom. It was done by a very experienced team with people from several areas such as EHS, bath handling area, laboratory and potroom (operators, technicians and engineers). By using statistics and subject matter knowledge, the team hypothesized that the potential causes would be the following: o o o o o o o o
Introduction Alcoa has been recognized by governmental and nongovernmental agencies around the world for their accomplishments related to environmental protection and sustainable growth. Because of this, the understanding and minimization of fluoride and particulate emissions is always a subject of intense interest. Alumar, one of Alcoa's units located in the northeastern region of Brazil, following this corporate vision, is continuously researching the root causes of fluoride and particulate in order to minimize them. Previous works were published about it, such as a study developed to map and determine the impact of each operation (anode setting, tapping, among others) in the fugitive fluoride emissions in the pot rooms [1]. It is already known that the fluoride and particulate emissions are affected by several factors inside the potroom such as bath temperature and ratio (bath chemistry), alumina quality, amperage level, operational activities, pot draft, etc. For example, it was shown the importance of the anode cover integrity to prevent gaseous and particulate fluoride evolution in another Alcoa research [2], which says that the crust integrity was capable of changing the amount of fluoride evolution by more than 1000% while the bath chemistry changes could only change the amount of evolution by 20%. Since fluoride and particulate emissions are impacted by several variables and their effects depend on the interactions of them, it
o o o o o o o o o o
Amperage - load creep Usage of compressed air for housekeeping Increase the number of days between pot dressing Analysis error - Lack of reliability at the Lab Sampling and preparation error - Lack of reliability at the Lab Stuck material and open hole in the ducts of gas Lack of Pot Draft Delays to remove the butts from potroom after anode settings Inefficient cleaning machines for housekeeping High alumina LOI High bath temperature Low bath ratio Excess of Burn-off (Anode failure requiring an extra set) Excess of anode effects Deficient balance of pot draft Pot superstructure cover is not maintained at the right position High amount of fines in the anode cover High percentage of alumina in the anode cover
Variable Definitions The most likely causes were selected from this list, based on experience and previous work of the team. The decision was to study the effect of Pot Dressing, Usage of Compressed Air for Housekeeping and Pot Draft. The pots at Alumar are currently dressed every 6 days by the Pot Tender Operator. This activity is really important to keep the anode cover integrity and consequently to minimize the emissions.
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Housekeeping is made every shift in a group of pots at Alumar. Compressed air has been used to clean some components of the pots to guarantee the minimal TPM (Total Preventive Maintenance) requirements related to cleanliness. Alumar uses A-398 reactors for gas treatment. Pot draft normally ranges from around 3500 to 4000 ACFM. This reactor has a very good efficiency, but the fugitive emission increases in the potroom if the pot draft is below a minimal level.
Results The experiment was set up and run one time for each combination. Total fluoride and particulate emissions were collected for a period of 8 hours using an isokinetic test, based on US EPA Methods 13 (Determination of total fluoride emissions from stationary sources) and 14 (Determination of fluoride emissions from potroom roof monitors for primary aluminum plants). Every 8 hours the reacted solution of NaOH 0.1N was analyzed, using the ion selective electrode method and the experiment was moved to new settings. Table 3 is showing the results for fluoride emissions and Table 4 for particulate emissions. The experiment was run one time for each combination.
Statistical Experiment The interaction among the variables and their effects on the fluoride and particulate emissions needed to be known for better understanding of the causes. It was decided to make a full factorial experiment 2 3 , three factors with two levels for each one, as can been seen in Figure 1. By definition Factor A is the variable Usage of Compressed Air for Housekeeping, Factor B is Pot Dressing Frequency and Factor C is Pot Draft.
Table 3- Fluoride emission results for each combination 23
(+)— Factor (C)
cC
+
B-
B+
b
a
ab
be
ac
abc
(1) c
0.51
0.622
0.554
0.675
c+
0.188
0.272
0.271
0.652
c
Table 2 presents the level for each variable. The experiment was run with these levels for both response variables (fluoride and particulate emissions).
Low level
High Level
A Compressed Air for Housekeeping
Don't use compressed air
Use compressed air as usual
B Pot Dressing Frequency
Pol dressing every 2 days
Pot dressing every 5 days
C Pot Draft
3000 ACFM
43O0ACPM
B
B+
B-
B+
1.043
1.592
2.282
1.620
0.404
0.434
0.916
2.454
Y=0.468+0.07A+0.087B-0.122C
Table 2- Level definition for each variable Factor
A+
A-
The statistical analysis of experimental data was done using the software DOE-KISS (®-Air Academy Associates, Colorado Springs, CO, USA). The following table (Table 5) is presenting the results for fluoride emissions. The interaction among the studied factors doesn't impact the fluoride emission. F value is above the Sig F (for Alpha = 5%), which means it is a significant model. The factors Pot Draft (C) and Pot Dressing Frequency (B) are really relevant for Fluoride Emissions, while the Usage of Compressed air for Housekeeping (A) is in a "gray zone". Nevertheless, the latter was confirmed as a significant factor using others techniques such as Yates. It is recommended to maintain it as part of the final model. The model has a high strength with R2 at 0.8 and the prediction equation for it can be seen in equation 2 (coded effects).
A+ B
C
+
Table 1- design matrix for experiment 2 3 with all runs B-
B+
C
Xijki = μ + Ai + Bj+ ABy + Ck + AC* +BCjk + ABCijk + Z,(ijk) (1)
+
B
( ♦ )
Each combination was tested once in a period of 8 hours. Table 1 summarizes and gives a better understanding of how the data will be collected. The equation 1 describes the expected response model for this kind of experiment. Table 1 shows a design matrix for experiment 2 3 with all runs.
A-
B+
23
I I Factor (B) (-) (+) Figure 1 - Graphical depiction of full factorial DOE 2
23
B
Table 4- Particulate emission results for each combination
— (+) Factor (A) —
A+
A-
292
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Table 6- Statistical analysis for the variable particulate emission
T a b l e 5- Statistical analysis f o r t h e v a r i a b l e f l u o r i d e e m i s s i o n Multiple Regression Analysts
Multiple R e g r e s s i o n Analysis
'Tltr^T 1 1
1
Factsr Const A B C
Name
Coeflf i 0,46800 Jateamento 0,07000 Rastelamenlo 0,08725 Pot Draft -0,12225
Rsq
0,8283
Adj Rsq
0,6995
StdError
0.1067
F
6.4306
SigF
0,0520
1
Pf? Tail) i Tol ! < 0.0002 0,1371 1 : X 0.0818 1 ; X 0.831'/ 1 : X
Õ-ËÌ Model 1
I Factor I Name
1 I I j
A B
C
Low High Exper
i Jateamento ! Rastelamento
i Pot Draft
-1 -1
- 1 : 1
1 1
0 °
0
I \
1
| j
Source
Nam«
(:*«« Pt2TMt;T«4 1.3431? OJÖOli
-imwma Qi7m
A C
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-03113
0.4205
. · , _ _ ·_
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< 1 [Factor
1 K
0,2231
A B C
1
Prediction Ar$Rs<j I
| Y-hat !Std Error
SS
MS
0,2
3
0,1
Error
0.0
4
0,0
Total
0.3
7
Lower Bound Upper Bound
0,1478809 0,7881191 I
I
S.gF
I
&<►««#
♦•if-
MS
$
0.4
Low High Exper 0 0 0 .· Ã
Prediction Y-hat StdError I
1,343125 0,5921187
99% Prediction Interval
11
I R*(jr*wjfe« '■ 1J9 ' · :
Pareto chart in Figure 2 indicates that the effects Pot Draft (C) and Pot Dressing Frequency (B) are the most important, while the usage of air for housekeeping (A) has a lower impact in the fluoride emissions.
A2
.:....,...vi...,:J
7 ' ♦ · ! " · ♦
I
I . .
Lower Bound Upper Bound
·. . L.
«0,433231 3.119481
. j .:....
j ,j
Pareto chart in Figure 3 indicates that the Usage of Air for Housekeeping (A) is really the most important effect. Pot Draft (C) has some impact but it is much lower than Factor A.
Fluoride Emission Abs. Value of the Factors
n_n Particulate Emission Abs. Value of the Effects
0.7 0.6
|o.5-
JKKHKM
| O
0.4-
J«
|
0-20.1
Comp. Air fur HouseUeping
0
^
l_
Comp. Air for Housekteping
Figure 2 - Pareto chart for absolute values for each effect for fluoride emissions
Draft
Lr
Effect Name
Figure 3 - Pareto chart for absolute values for each effect for particulate emissions.
The results for particulate emissions are as shown in Table 6. The interaction among the studied factors doesn't impact the particulate emissions. As in the fluoride model, the F value is above the Sig F (for Alpha = 5%), which means it is a significant model. The factor Usage of Compressed Air for Housekeeping (A) is the most relevant for this response variable. Pot Draft (C) is in a "gray zone", however, it was confirmed as a significant factor using others techniques such as Yates. Pot Dressing Frequency (B) is not a significant factor for particulate emissions. The model has a reasonable strength with R2 at 0.5, regarding it is an industrial experiment, and the prediction equation for it can be seen in equation 3 (coded effects). Y=l. 34+0.47 A-0.29C
Name
: Miwfmto-ú S 1 Rast*r3«r«nto A 1 Draft Ë-.Á ·| Ë i
I
B ^ ^ - ^ w i l ^ i
T*Wl
«Dr··»«
_'_ ♦:
0468 0,1067064
99% Prediction Interval 1 Recession
factor COM
Conclusions and Recommendations According to the statistical model, fluoride emission can be significantly reduced increasing the Pot Draft (main factor) and improve Pot Dressing as much as possible. The worst condition would be to set the Pot Draft to -1 and the Usage of Compressed Air and Pot Dressing to +1. Considering this condition, the fluoride emission would be 0.74 kg/tAl; and, if the settings were the opposite values mentioned above, the result would be 0.18 kg/tAl. Particulate emission can be significantly reduced if the Usage of Compressed air for Housekeeping is minimized or avoided. Setting the effects to maximize the particulate emission (Usage of Compressed Air to +1 and Pot Draft to -1), will make the particulate emission 2.1 kg/tAl and 0.57 kg/tAl to the opposite settings. The team looked for alternatives to move the factors toward lower emissions.
(3)
293
Fluoride Emission Applying Recommendations
Pot Draft By removing the stuck material and closing the holes in the external ducts of gas, the pot draft can be improved by 500 ACMF. Figure 4 shows the current condition of the duct of gas after more than 20 years of operation and the state of it after the cleaning process.
m A
fl
rfl
^ψ
Λ :;j9B
:
Before Cle*ining
Comet
Condrfion (I>pical)
R*commend»non_2nd ssnphng
Figure 6 - Fluoride emission results under recommendation conditions for a period of 48 hours. Although the model shows 0.23kg/tAl after applying the recommendations for a sampling period of 48 hours, the measured results were 0.34 and 0.37 kg/tAl. It is clear that the implementation of those actions will result in lower fluoride emission.
Afte r Cleaning
Figure 4 - External gas ducts before and after the cleaning process Pot Dressing
Particulate Emission Applying Recommendations
It is possible to improve the pot dressing by reducing the frequency from every 5.3 to 3 days, without increasing the head count. Currently, there are four Pot Tenders by room and they are responsible for the dressing and other operational activities. Their production hours can be optimized if the dressing is done by two operators, and the additional activities are done by the others, as can been in the following figure (Figure 5). Current Condttan - 8 pots dressed by shift by room - Every 5,3 days
JED
Current Condition (Tipical)
Picommendauon. 1 it samphng
Recommendaaon_2nd sampling
Figure 7 - Particulate emission results under recommendation conditions for a period of 48 hours.
Target Condition - 1 4 pots dressed by shift by room - Every 3 days
There is a good agreement between the measured particulate emission and model predictions under the recommendation conditions for a sampling period of 48 hours. As mentioned for fluoride emissions, it is clear that the implementation of those actions will also result in lower particulate emission.
in tuip«$&·* ««nodiijäes
Figure 5 - Pot Dressing Frequency to reduce the emissions.
Acknowledgements Colleagues Joel Cämara, Elisio Bessa and Ciro Kato are acknowledged for their contribution, support and valuable discussion about this subject.
Usage of Compressed Air for Housekeeping It can be reduced by at least 50% through the usage of cleaning machines and changes on the TPM requirements.
References
It was decided to set the conditions according to those recommendations and measure the emissions to check the benefits of those recommendations and to compare the results with the model's predictions. Figure 6 shows the results for fluoride emission and Figure 7 displays the results for particulate emission. The sampling time was 48 hours for both samples.
294
1.
Nagem, N., Batista, E., Silva, A., Gomes, V., Venäncio, L., Souza, L., Understand the Fugitive Emissions at Alumar, TMS, Light Metals 2005, pp. 289-292
2.
Tarcy, G.P., The Effect of Pot Operation and Work Practices on Gaseous and Particulate Fluoride Evolution, TMS, Light Metals 2003, p 193-198.
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
INNOVATIVE DISTMBUTED MULTI-POLLUTANT POT GAS TREATMENT SYSTEM Geir Wedde, Anders S0rhuus, Odd Bjarn0 Alstom Norway AS, Drammensveien 165,0277 Oslo, Norway Pot Gas Treatment, Emissions, Energy Recovery It is obvious that the GTC size and footprint must fit within a smelter requirements and practical arrangement and it seems that this limit has been reached. Alstom has developed a solution that divides the traditional GTC into smaller sub-centers that are distributed along the pot rooms. These small plants are called Decentralized Distributed Scrubbers (DDS - patented).
Abstract Gas Treatment Centers (GTC) are traditionally arranged in the courtyard between pot rooms and handle vast quantities of pot gas (3- 4 million m3/h) in a large number of filter compartments (1530) with demanding space requirements and challenging control of operation.
Decentralized distributed scrubber (DDS)
Arranging the gas treatment in decentralized installations as Decentralized Distributed Scrubbers (DDS) would obviously save on duct work, alumina handling, storage and civil work. In addition Alstom's DDS integrates silos, heat exchangers, scrubbers, fans and stack into an extremely compact and efficient multi-pollutant control and recovery technology with incomparable footprint. The close integration with pots improves pot gas collection and simplifies the alumina distribution to pots. Module based design allows for short delivery time and early start-up. This paper discusses and reviews the benefits and mitigation of the technical challenges offered by the DDS solution.
Gas treatment plants for pot lines must satisfy the stringent requirements that are forced on them through the continuous production process that governs the aluminium electrolysis. This means a 24/7 operation year round including high equipment redundancies. DDS lav-out and arrangement The difference between the GTC and DDS lay-out in the courtyard between two pot rooms is illustrated in figure 2. The DDS has been laid out close to the pot room walls and connected to a section of pots. A number (12) of identical DDS consisting of equal pot ducting systems, filters, fans and alumina (fresh and enriched) handling systems are shown.
Introduction New greenfield aluminium smelters and brownfield smelter amperage increase or expansions have increasingly demanded larger gas flows to be handled by gas treatment systems leading to an economical challenge as well as space (footprint) issue.
WfM. Ìί^ÌÌίÌψÌ
Gas flows from sections of pots are traditionally handled by Gas Treatment Centers (GTC) arranged in the courtyard between pot rooms. The size of GTCs has dimensions of a large office buildings being up to 150 m long and 30 m wide and rise to levels of 30 m. For operational flexibility and gas and alumina flow considerations the GTC is made up of a number of identical compartments that operate in parallel. The number of compartments could be as many as 30 or above which may be challenging to tune, operate and to detect failures or faults among the individual units.
Figure 2: GTC lay-out (1) and corresponding DDS lay-out (2) The illustrated GTC handles pot gas from a total of 180 pots arranged in two pot rooms with 45 pots in each half of the pot room on both sides of the cross-over passage way. The GTC is designed to operate with a total pot gas flow of 2.5 million m3 per hour. A typical arrangement of DDS would be to handle gas from 15 pots (-210,000 mTh) in each DDS which means a total of 12 small gas treatments plants. This includes allowance for running 2 pots simultaneously on forced draft (~24,000m3/h) for each section. Arrangement and design of a DDS The smallest gas treatment center should include a minimum of 2 reactor/filters and 2 exhaust fans with sufficient redundant equipment for independent operation. Any other combination of
Figure 1: Gas Treatment Center (GTC) for modern pot lines
295
number of pots and compartments is readily adaptable with DDS small compartment sizes and adjustable bag lengths. An arrangement of a DDS is shown in figure 3. Each of the two units is a fully equipped Abart [1] dry scrubber including fresh and enriched alumina storage, fresh alumina metering and transport, reactor/filter, alumina recycling feeder, fan and stack.
the pots through an Alstom Alfeed pot feed system [2] or other feed systems on the market. This saves intermediate silos and alumina transfer points that tend to increase segregation possibilities of the alumina and disturbances of pot operations. The hopper capacity for stored enriched alumina is about 3 tons per compartment and a direct by-pass from the fresh alumina storage to the pot feed system is available ensuring alumina feed to pots at all times. Recycling of enriched alumina within one compartment is achieved with individual feeders with adjustment control. Performance enhancements Each compartment has individual control of gas flow by a dedicated exhaust fan; fresh alumina feed and enriched alumina recycling through frequency controlled drives. Each compartment should run with the same flows and feeds and with the above features DDS has the possibility to individually fine-tune each compartment. This allows the DDS system to easily and accurately achieve equal gas flow with equal alumina feed through each compartment. Each DDS compartment utilizes independent bag cleaning control which allows simultaneous tuning of pressure drop, flow and particulate emission (particulate emission is directly related to pressure drop of a filter).
Figure 3: Arrangement and design of a DDS plant
The direct fresh alumina feed from silo into the gas stream ensures constant feed without any surges and reduces stops for cleaning of airslide systems.
The adaption of fans on top of the fresh alumina storage requires additional stiffening and vibration pads. The mini-stacks are equipped with sound absorbers both at fan inlets and in ministacks.
Overall the DDS has the potential of reduced emissions through optimized adsorbent utilization and enhanced plant stability level.
Benefits of DDS
The DDS is designed to utilize multiple bag lengths, but 8 m bag length is preferred for improved cost efficiency.
The benefits of gas treatment plants arranged as DDS are many but also challenges and some disadvantages are obvious. These are analyzed below with the assumption that fresh alumina is stored in one central silo and the fresh alumina requirement for the DDS is achieved through dense phase transport up to the individual DDS.
Power consumption DDS represent a significant saving in power consumption of the gas treatment installations for a smelter. Pot room ducting, crossover ducts, filter to stack ducts and stack are additional ducting with additional resistance of a traditional GTC. The reduced pot gas pressure drop for a DDS equals about 15 % of the total GTC gas pressure requirement which is a 15 % reduction of the power consumptions of the main fans.
Foot print There is a substantial reduction in foot print area by implementing 12 DDS plants rather than one GTC (figure 2). One DDS unit of 4 compartments takes up approx 90 m2 of area in the courtyard and 12 DDS will then occupy nearly 1100 m2 while a GTC would cover about 2300 m2 which is more than double of the DDS. This saving in space or foot print is of significant value for the smelter.
Steel weight It is easily seen that the duct system has been substantially reduced as the DDS are located close to the pots and use smaller and identical pot duct sections. Ducts from GTC filters to the exhaust fans and the large stack are eliminated. As DDS integrate the required silo capacities, both the fresh and enriched alumina silos can be eliminated. One may expect an increased weight of the DDS filters due to the increased total number of compartments; however, DDS with 8 m long bags has the same weight per unit of gas flow as the Alstom standard GTC equipped with 6 m bag length. With the arrangement as seen in figure 2 the reduction in steel weight of DDS is in the order of 30 % of a comparable GTC.
In addition the flexibility in the location of each DDS in the courtyard allows for an increased optimization of space and pot room and utility building lay-outs Alumina storage and handling Integrated with each Abart filter compartment in a DDS is fresh alumina storage of 25 tons. This capacity reflects close to 2 days consumption of alumina of the pots. From the storage of fluoride enriched alumina in the filter compartment hoppers the enriched alumina may be fed directly to
296
Equipment scope There is a significant saving on alumina transport and silo fluidization, but an increase of alumina feeders. Although GTC has only 6 large fans compared to 48 fans for DDS, the GTC fan cost is about 4 times that of DDS. Filter bags and cages are reduced by the DDS by approx. 20 % because of extended bag length. Electrical items and instruments As each DDS will have to be electrified from a central source, number of power and instrument cables will increase substantially for a DDS arrangement. Motor starters, instruments, plant switches and signals, are proportional to the number of filter compartments which may amounts up to 20 % extra for the DDS system. Although the power tension required is LV - low voltage (< 460 V) and that the cabinet etc for each DDS is located in a common cubicle that can be mounted and furnished in a work-shop and shipped to site, the electrical scope may be more than twice that of the GTC.
B2JI I
I cmntm.cotmx».wiomr%cM}Amßunr\
L.
Figure 5: Installation of work-shop erected DDS modules Erection of several DDS units can be achieved simultaneously and independently which is not the case for a GTC where many items are dependent on each other as for erection of filter compartments, alumina handling to and from these compartments, stairs and platforms, and electrical and instrumentation. Further benefits
TO na«
w i n
Civil work for a GTC installation is a significant part of total cost. The foundations for the heavy silos, filter compartment support, main fans and stack are extensive compared with the simpler arrangements for the DDS.
Figure 4: Basic DDS E&I topology
For smelters with pot line expansions DDS represent a simple and unique opportunity for stage wise implementation as the system connects to a smaller number of pots. The bolted design eases the installation in an operating smelter as magnetic field strength complicate welded construction.
However, DDS simplifies total electrical lay-out and is exampled in figure 4. A GTC will have substations for both MV and LV, as the main fans normally require MV. MV can be removed, and a number of identical cubicles for each DDS installed.
There are projects where the GTC falls on the critical path of the smelter erection sequence. The DDS will allow quick and independent erection and potentially early start-up of metal production.
It is reason to believe that the extra electrical scope of the DDS can be significantly reduced by working close with the smelters utilizing and incorporating the existing electrical system lay-out. Manufacture, transport and erection The size of a DDS compartment is fulfilling the requirement on regulation of transport on roads. DDS is suitable for offsite and offshore pre-assembled, with manufacture in work-shops and transport to site. Cost efficient production of filter panels, hoppers and tops utilizing automatic, robotic manufacturing systems has been demonstrated with success for other products as Alfeed. This allows fast erection in lay-down areas or directly on site as seen figure 5.
Challenges of DDS There are challenges to the DDS solution that sometimes would become opportunities but also a re-think of the smelter operation may facilitate all the benefits that have been addressed above. Pot gas temperatures As the DDS is closer to pots it should experience higher gas temperatures (15-20 °C) than the GTC. It is well established that pot gas temperatures affect the performance of the dry scrubber and in order to meet current requirements the gas temperature at scrubber inlet should be 115-120 °C [3]. This requirement means that the DDS may need additional cooling upstream the filter compartments. If dilution air was to be used a significant quantity of air has to be added and the cost efficiency of the DDS would suffer.
Fluidization of filter hopper and mini-airlifts and pot feed system is generated by one blower/fan only using air from the pot room plant air lines as stand-by. It is also preferable to use plant air instead of installing dedicated air compressor and dryer equipment for bag cleaning air needs.
Integration of heat exchanger Alstom has developed heat exchangers operating on the pot gas upstream the filters [4]. Improvements on this technology are ongoing and a heat exchanger integrated (IHEX) into the inlet duct/reactor of the Abart filter compartment has now been realized. This allows significant savings on the heat exchanger system as support steel, stairs and platforms, and the damper arrangement is combined with the DDS design as seen in figure 6.
Feeders, recycling units and main fans are easily accessible in the DDS and spare units can be stored at the smelter for fast replacement of broken units. During situations of "sick" pots where pot gas temperatures increases and fume evolutions are enhanced the HEX will achieve improved performance and cooling as the temperature differences of pot gas and cooling water increases. Short term high HF evolutions can be controlled with increased fresh alumina feed. Cover material for pots Many smelters and pot technologies use enriched alumina and mixed with bath for covering and dressing pots. For DDS this can be facilitated by inter-connecting several pot feed sections from several DDS units would allow these DDS to run with slightly higher feed equal to the cover material needs. The remaining DDS will have to be slightly starved of fresh alumina feed but the overall emission level from DDS will not be affected.
Integrated c heat exchanger
The additional flow of enriched alumina could be extracted from the pot feed system at an appropriate location into an airlift feeding a cover material silo. Optionally one can feed a tanker with cover material and feed the silo in batches which would then overcome the high capacity discharge needs from silo to pot tending cranes.
Figure 6: The integrated heat exchanger (IHEX) for DDS This innovative solution reduces the pot gas temperature to any desired practical levels (90-120 °C) depending on the purpose of the heat exchanger whether it should only cool the pot gas ahead of the gas treatment plant and /or to recover waste heat for use inside or outside the smelter. The additional 15 (- 20) °C temperature level of the pot gas entering an DDS means a significant contribution to higher value heat (exergy) of about 50 kW/pot. For a 180 pot smelter this amounts to 9 MW in addition to a traditional (GTC) heat recovery potential with IHEX of 15-20 MW.
Emission monitoring The DDS is prepared for the installation of HF and dust measuring equipment, easy accessed from the filter top. To reduce the number of monitors required a system for extraction of gas from each mini-stack into a laser based analyzer has been developed. Emission dispersion
Operation redundancy
Discharge of treated pot gas from GTC stacks would require stack height level above pot room roof louvers to meet stringent emission requirement through emission dispersion modeling. For DDS it would have been detrimental to its design simplicity to have enforced same stack heights as on the GTC.
The DDS is designed to utilize multiple bag lengths, but 8 m bag length is preferred for improved cost efficiency. For special purposes and increased flexibility as when the N-l redundancy principle applies the DDS can be equipped with shorter bag length (5-7 m) which will increase the number of compartments. Connecting pair-wise DDS will add further flexibility and with frequency operated motors of main fans one compartment can be taken off-line with only a slight reduction in total gas flow from pots.
As DDS introduce multiple stacks at locations across the pot line the dispersion of the emissions is significantly improved and DDS emissions can be released at lower levels. The emission through the pot room roof is of similar concentration levels as from the DDS mini-stacks and it would be natural to use similar discharge height level for the mini-stacks.
Even more gas flow capacity is achieved when the HEX is designed with some over-capacity that allows further cooling of the gas.
Magnetic fields The DDS is located closer than a GTC to the pot room and is therefore exposed generally to higher magnetic field strength. This has to be considered especially regarding electronic equipment.
Therefore the combination of size and number of compartments and with a heat exchanger the DDS solution has sufficient gas flow capacity to satisfy traditional redundancy needs of a smelter.
298
The performance of "standard" polyester needle felt filter bags is demonstrating consistent low particulate emission at about 0.9 1.1 mg/Nm3, and the particular test campaign below (figure 8) shows emissions during pulse air cleaning of the filter bags reaching about 12 mg/Nm3 as a peak value and reducing to a value close to zero between each cleaning action. The total emission over several cleaning cycles is 1.04 mg/Nm3
The main fans, however, are located on top of the DDS which increase the distance to the source (electrical bus bars) of the magnetic fields. It has been calculated that the field strength at the DDS main fans is similar to the one at the GTC main fans. The DDS is normally made as a bolted solution, which means considerably less welding and ease of erection if the construction has to take place in the magnetic field of an existing operating smelter.
Filter3 L D 1 2 8 2 S t a n d a r d
Multi-pollutant gas treatment More often S0 2 scrubbing is required for aluminium smelters, in particular the large mega-smelters where S0 2 emission may have a detrimental impact on the local ecology. Wet S0 2 scrubbers are used which absorb S02 into an alkali solution and smelters operating at coastal areas would benefit from using seawater as the absorbent [5]. To accommodate the DDS principle of simplicity and modularization a S0 2 scrubber design has been developed and ongoing tests will confirm the performance which are targeting levels that exceed current S0 2 scrubbers for GTC
Figure 8: Particulate emission peaks during pulse cleaning of "standard"filterbags.
Operational experiences
Filter bags with enhanced micro-denier fibers embedded in the outer filtration surface achieve an exceptionally low peak emission consistently lower than 3.0 mg/Nm3 as seen in figure 9. The total particulate emission is reduced to only 25 % of the "standard" bag emission to as low as 0.25 mg/Nm3.
Experiences with DDS have been obtained since 2007 where 2 DDS plants (figure 7) were installed at an aluminium smelter. Each DDS consists of 2 compartments and handles 115.000 m3/h. The two DDS are interconnected to achieve highest flexibility. Fresh alumina is fed to the four integrated silos in the DDS by an airslide system running from an elevated existing silo. The enriched alumina is fed to 2 pot sections.
Filter! L D 1 2 8 4 M i c r o D e n i e r _ Pulse is avg of 3 pulse cycles I.e. 3 x (
Figure 9: Particulate emission of "micro-denier"filterbags Two test campaigns using manual test equipment, of emission of gaseous fluorides (HF) has been recorded at an average level of 0.26 mgF/Nm3 with F-content of the enriched alumina at 1.80 %. For low HF concentrated pot gas with 1.15 % F-content in enriched alumina the HF emission level was as low as 0.04 mgF/Nm3. Figure 7: Operating DDS installations
Conclusion
Emission performance
The principles of centralization do not always reduce costs and ease the operation of a system. The DDS solution as distributed gas treatment plants located close to the pot room and handling gas from a limited number of pots (10- 30) has many advantages
Particulate emissions has been followed up and monitored regularly as testing of filter bag materials is part of an agreed cooperation with the smelter.
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and benefits that should be explored further by the aluminium industry. The reduced cost and power consumption in addition to the reduced foot-print of 50 % and the increased potential of waste heat recovery would obviously make this industry more competitive. DDS has flexibility and operational performance that will meet future emission requirements. Particulate emission levels at only 1/10 of traditional level is achieved by applying filter fabric made with micro-denier fibers. Incorporating a SO2 scrubber in the DDS solution makes it a cost efficient multi-pollutant gas treatment technology. References 1. 2. 3. 4. 5.
G. Wedde, P Henriksen "Experiences and developments through 10 years of operation with the Abart dry scrubber" Light Metals 2007 S. Ose, A. Sorhuus, O. Bjarno, G. Wedde " Alfeed, a new alumina feeding system to aluminium pots" Light Metals 2010 A. Heiberg, G.Wedde, O Bockman, S. Strommen " Pot gas fume as a source of HF emission from aluminium smelterslaboratory and field investigations" Light Metals 1999 A. Sorhuus, G. Wedde, K. Rye, G. Nyland „ Increased energy efficiency and reduced HF emission with new heat exchanger", Light Metals 2010 S Broek, B Rogers "The origin and abatement of S02 emissions from primary aluminium smelters" Light Metals 2009
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
FLUORIDE EMISSIONS MANAGEMENT GUIDE (FEMG) FOR ALUMINIUM SMELTERS Nursiani Tjahyono, Yashuang Gao, David Wong, Wei Zhang, Mark P. Taylor Light Metals Research Centre, The University of Auckland, Private Bag 92019, New Zealand Keywords: Fluoride emission, Control, Management, Guidelines, Work practices Abstract
These legislative limits, the health and safety of the employees and the wellbeing of the environment are major drivers for the smelters to reduce their fluoride emissions.
All smelters worldwide operate under strict fluoride emission limits and the reduction of fluoride emissions is further driven by health and environmental considerations. A Fluoride Emissions Management Guide (FEMG) has been written by the Light Metals Research Centre (LMRC) on the invitation of Australian Aluminium Council (AAC), under the Asia-Pacific Partnership (APP) on Clean Development and Climate. The aim of the FEMG is to provide a better understanding of the factors affecting and ways of reducing fluoride evolution and emissions in smelters, and further, to provide smelters with an operational guide for reducing and managingfluorideemissions.
Mechanisms of Fluoride Generation and Escape Pathways The majority of gaseous and particulate fluorides emitted from an aluminium smelter are originally generated in a pot. The amount will depend on cell technology, raw materials, process control, and ambient conditions (temperature and humidity) where the smelter is situated [10]. A typical pre-bake, point-fed pot generates about 15-30 kg total F/t Al, with >99% of these fluorides typically captured in the scrubbing system [10,11].
Reduction in fluoride emission can be achieved by improvement of operational standards in the smelter. This includes operational, control and maintenance practices in the potroom, GTC, and other areas in the smelter that generate fluoride. Along with incorporated audit guidelines and training package, implementation of this guide can, in a short time, lead to significant reduction influorideemissions in aluminium smelters.
Gaseous fluoride (HF) is produced when A1F3 in the bath or as bath vapour (NaAlF4) reacts with a hydrogen source in one or two reactions: 2A1F3 (g or diss) + 3H20 (g) => 6HF (g) + A1203 (s or di SS ) 3NaAlF4 (g) + 3H20 (g) => 6HF (g) + A1203 (s) + Na3AlF6 (s)
(1) (2)
Introduction Particulate fluorides, on the other hand, are generated when bath vapour condenses to form droplets or fine bath particles are picked up and entrained in the pot gases.
Generation of fluorides from the aluminium smelting process is unavoidable with the current technology. Hence, all smelters worldwide operate under strict limits of fluoride emissions and these limits are expected to become tighter with time. Table I provides a comparison of production-based regulatory limits for total airborne fluoride emission in primary pre-bake aluminium smelters for different regions of the world, including the current best practice.
Potrooms and Gas Treatment Centres (GTCs) are often seen as the main sources of fluoride emissions. While it is important that the different factors related to fluoride emissions from these two key areas are addressed and controlled, it is equally important to realise that a fluoride emissions problem may also be related to other plant areas. Examples are the fluorides released from transport/delivery of spent anode butts and the bath processing plant. The mission to reduce and control smelter fluoride emissions spans across all plants within the smelter and is a responsibility of every individual in the smelter.
The majority of larger capacity modern smelters today limit their fluoride emissions to 0.5-0.6 kg F/t Al [1]. These limits are practically achievable, as demonstrated by the performance of Aldel and Tomago smelters (0.46 kg F/t Al in 2005 [2] and 0.56 kg F/t Al in 2007 [3], respectively). Some of the better performing smelters in the world have reported emissions as low as 0.29 kg F/t Al (Portland in 2008) [4] and 0.25 kg F/t Al (Karmoy in 2006) [5]. Table 1. Comparison ofproduction based regulatory limits and the current best practice for total airbornefluorideemissions in the pre-bake aluminium smelting industry. Regulated Airborne Emissions Limits for Total Fluoride (kg F/t Al) World Benchmark 0.2 [6, 7] Europe - Iceland 0.35 [1] Europe / OSPAR 0.6 [8] USA-EPA 0.6-1.5* [9] * The limit of 1.5 kg F/t Al is only applicable to older potlines as specified in the National Emission Standards for Hazardous Air Pollutants for Primary Aluminium Reduction Plants. Region
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Through careful management and control of both operation and maintenance practices around the plant, further reductions in fluoride emissions can be achieved. However, the factors which affect and the ways of reducing fluoride emissions are sometimes not clearly understood, hence there is a need for a management guide to address fluoride emissions in the smelter as a whole. This paper describes the Fluoride Emissions Management Guide (FEMG) which has been prepared by the Light Metals Research Centre (LMRC), The University of Auckland, on the invitation of the Australian Aluminium Council (AAC), under the Asia-Pacific Partnership (APP) on Clean Development and Climate. The paper will focus on the content of the guide and the importance of managingfluorideemissions in the smelters.
anode setting and bath pouring operations are examples of work practices that have a direct effect on fluoride emissions. These operations are generally simpler to target compared to those with an indirect impact, such as the influence of poor anode quality on carbon dust generation in pots, which in turn increases fluoride emissions from more frequent dust skimming and open pot doors.
WhatistheFEMG? The FEMG is a comprehensive guide to managing fluoride emissions in primary aluminium smelters, with a focus on improvements to operational, control and maintenance practices to allow smelters to maximise their environmental performance with their current technologies. It aims to increase understanding of factors controlling fluoride evolution and emission, as well as providing an operational guide to smelters for reducing and managing fluoride emissions. The FEMG covers fluoride emissions from the potroom, GTC and other smelter systems that contribute to emissions, along with incorporated audit guidelines for each area. The guide is specifically prepared for pre-bake, point-fed pot technologies with dry scrubbing technologies.
HDBBBBH Operation
The FEMG is broken down into 6 main chapters, which are: • Introduction & Theory - covers the background of fluoride generation and drivers to reduce fluoride emissions. • Overall Fluoride Emission Management System covers the overall concept and approach for managing smelter fluoride emissions. • Potroom Systems for Reducing Fluoride - specifies the Key Performance Indicators (KPIs) and control points for work practices in the potroom. • Gas Treatment Centre (GTC) Systems for Reducing Fluoride - addresses KPIs and control points for work practices in the GTC. • Smelter Systems Outside the Potroom and GTC details KPIs and control points for other areas in the smelter that affect fluoride emissions. • Fluoride Emission Measurement - lists standard and recommended smelter fluoride measurement methods.
Process Control
Maintenance M
l
Smelter
HF Measurements
· Remedial Actions
Boundary
Figure 1. Overall concept behind thefluorideemission management system. Monitoring Systems Successful management of any process requires good measurements. Measurement systems are important to determine whether the operations are performing within the required specifications, otherwise actions can be taken to rectify the problem. The most effective monitoring system designs include a combination of numbers, charts, colour and alarms, which provide meaningful and relevant information to the users. For monitoring of fluoride emissions, specific equipment and methods, such as US EPA Methods 13A and 14, have been developed to measure the level offluorideemissions as required by law.
Pictures are extensively used to illustrate issues and potential improvements to lower fluoride emission through daily operations. The recommendations contained in the guide have been proven by world-class smelters and are based upon current best practices, whereby if adopted, significant improvement in fluoride emission can be achieved.
Response Systems Once information from the monitoring systems has been received and analysed, response systems can be developed. They include: • Remedial actions • Work standardisation • Staff education and training • Plant and process changes and improvement
Overall Fluoride Emission Management System As previously mentioned, reducing fluoride emissions in the smelter should be the responsibility of all smelter personnel and not be focused only at the potroom and GTC areas. The overall concept of the fluoride emission management system applied in the FEMG is a systematic-interconnected network consisting of four main themes, as illustrated in Figure 1. They are: • Control of work practices within specifications • Monitoring systems • Response systems • Audit systems. Each of these themes has a very important, specific role in reducing fluoride emissions and requires the involvement of all smelter employees. Control of Work Practices When the smelter work practices are not within process and work specification, they can result in a direct or indirect negative impact on the smelter's fluoride emissions. Exposed bath during
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The response systems can range from an immediate to long term response. They are important for keeping practices within specifications, as well as for operational and financial optimisation. Furthermore, the response systems are utilised to prevent and reduce undesirable process conditions from occurring. Smelter Audit Systems An audit is an on-site review conducted to determine the existing status of a process and whether a certain quality condition has been achieved, ensuring that issues, problematic areas and improvement opportunities are identified and addressed. The level of audits in the smelter can vary from simple (i.e. hooding tightness) to comprehensive assessment (GTC operation). It should be conducted on a regular basis by the smelter operations and management team, and cover the entire systems in the smelter to maintain the integrity and effectiveness of every part of the smelter systems.
collapse, resulting in more open holes in the crust, which again increases duct temperatures in the GTC.
Why is Fluoride Emission Management Important? All smelters worldwide are driven by the same factors to control and reduce their fluoride emissions. Some of the problem areas that smelters face are explored below. High Total Fluoride Emissions Some smelters face difficulties in keeping their fluoride emissions well below the legal limits, which may result in the possibility of heavy fines or even plant closure, should the situation continue to occur. High levels of fluoride emission will also affect the health of employees, particularly those working in areas where fluorides are generated. Skin/respiratory irritation, dental/skeletal fluorosis, and liver, kidney, or heart damage are some of the risks employees face with prolonged high exposure tofluorides[12]. A change in the operational standards is one way of making a difference in the level of fluoride emission in the smelter. By going through each step in every work practice, potential pathways of fluoride evolution can be easily identified. In metal tapping, for example, the fluoride is generated during the opening of tap end doors, as shown in Figure 2. Pot suction is lowered when there are gaps in the cell (including opening of doors and hoods) due to the loss of static pressure in the pot. The reduction of pot suction will lower the pot capture efficiency and hence, results in higher fluoride emission. Furthermore, fumes released from the crucible vent during the metal siphoning process are a major source offluorideemission during metal tapping.
# Hoods Off / Doors Open Figure 2. Schematic diagram showing howfluorideevolution and emission occur when the tap end door is opened during metal tapping procedure [Top], and how pot suction affects the pot capture efficiency and HF total emissions (not to scale) [Bottom].
When the pathways of fluoride generation have been identified, potential improvements can be recognised and change in work practices can be made. As shown in Figure 3, by comparing the smelter's current work practices against the world's best practices, poor work practices can be easily targeted and operational standards can be systematically improved. Audit checklists, like that shown in Table II, are a good tool for smelter managers or superintendents to use for conducting regular audits, ensuring that work practices are maintained within specified standards.
Inlet duct gas temperature is an important operational parameter. Above certain temperatures, the scrubbing efficiency is lowered as particulate fluorides captured on the surface of the bagfilters begin to react, forming HF which is released as a gaseous stack emission. The bagfilters in the GTC also have critical temperature limits, above which they are irreversibly damaged - in some smelters, the gas would then be bypassed from the GTC, sending all pot gases to be released into the environment without scrubbing. Both of these aspects will lead to higher total fluoride emissions and reduction in the operating safety margin of a smelter.
Increase in Amperage The aluminium industry continues to shift towards higher amperage cells to obtain higher levels of productivity and greater economic returns. This shift requires better control and management of fluoride emissions before it can proceed. Increases in line amperage will inevitably lead to higher heat generation from the pots. Poor management of pot heat balance can lead to direct increases in fluoride emissions, for example from more frequent unscheduled anode changes and carbon dusting. Furthermore, if pot heat balance is not managed properly, gas duct temperatures in the GTC can rise excessively, decreasing the efficiency of the gas scrubbing system. Greater heat generation in the pot can also cause increased rates of crust
Table II. Checklist of key areas to reducefluorideemissions during metal tapping. I Check items 1
Only open the tap end door before metal tapping
YES
NO
2
Use a hose to direct the extracted fumes back to pot
YES
NO
3
Cover the tap hole by sprinkling some crushed bath or alumina into the tap hole
YES
NO
Close the tap end door when metal tapping is finished.
YES
NO
\T"
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BAD Practice
Opening the tap end doors of a group of pots which are to be tapped.
Opening the tap end door of a single pot right before tapping.
Releasing the environment.
Connecting a hose onto the crucible vent and directing the extracted fumes back to the pot.
extracted
fumes
to
the
potroom
Figure 3. Examples of recommended best practice to minimise fluoride emissions during initial steps of metal tapping, presented as comparison of existing work practices and best practice (bad vs. good practice). By installing an online measurement and monitoring of inlet duct gas temperatures, a pre-warning system for high temperatures, as illustrated in Figure 4, can be adopted. Such a system will allow for investigation or action before a bypass from the GTC is activated. Furthermore, regular measurement of inlet duct gas fluoride level allows for detection of long term shifts/fluctuations in fluoride generation from the potroom. This will enable adjustments in scrubbing efficiency, for example, by increasing secondary alumina recycle rate to respond to higher levels of fluoride in the incoming gas to the GTC. Iniet Duct Temperature (deg C)
k/\/\^v\/\/vm Time (days)
Table III. Potential improvements that can be done to reduce fluoride emission due to shift in inlet duct gas temperature. Potential improvement to reduce fluoride emission due to increase in inlet duct gas temperature and fluoride level.
Critical Filter Temperature
120 100
With continual increases in amperage, however, existing GTCs can only cope to certain extent before some capital investment is required to accommodate the increase in the level of fluoride generation. Some potential improvements are shown in Table III.
Low
Cost
Pre
i-warning Temperature
Investigate High Cost
Figure 4. Schematic of continuous temperature measurement of GTC duct gas inlet temperature to detect and treat issue before critical temperature is reached.
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Use higher temperature bagfilter materials to sustain higher temperature operation and minimise GTC bypass on hot days. Measure and monitor inlet duct gas fluoride level using stack-type testing methods, over regular intervals (e.g. once per fortnight). If at risk of exceeding temperature limits of bagfilters, implementing a heat exchanger system to cool inlet gas ducts before the gas reaches the bagfilters at the GTC. Continuous monitoring (laser) of inlet duct gas fluoride level. Install higher scrubbing capacity (more gas flow) for higher amperage or higher line capacity.
Customisation of FEMG & Identification of Improvements
Customisation and Implementation of FEMG
Using the customisation material and photos, a tailored FEMG for specific smelter's technology and operational, control and maintenance practices can be developed. From the audit review, KPIs and improvement opportunities to reduce fluoride emissions across all areas of the smelter can be identified. This will ensure the content of a customised FEMG and any improvement to work practices is specific and relevant to all the personnel in the smelter.
As potroom and GTC technologies vary from smelter to smelter, it is important for the management guide to be adaptable to any variation in technologies and work practices. The FEMG is customisable to meet the need of any specific smelter. Ultimately, the guide achieves its purpose when the content is being implemented in the smelter. The development of an implementation plan is important to allow the smelter to distribute the relevant information to all its personnel that play a role in fluoride emissions. It will also ensure that maximum gains can be achieved by the smelter in reducing fluoride emissions.
Strategic Planning and Implementation of FEMG It is important to develop a strategic plan to address areas of improvement. The implementation process may include: • Smelter personnel training - training of smelter staff in the identified areas of improvement based on the customised FEMG and training workshops. • Implementation of changes in practices - smelter personnel to apply the necessary changes as per the guidelines in FEMG. • Regular audits - smelter managers to monitor the changes in practices by conducting regular audits. • Fluoride monitoring equipment - fluoride emission monitoring equipment could be used to monitor the changes in the smelter's fluoride emission as improvements to practices are implemented. This can be done by comparing baseline measurements (1-3 months prior to changes) against measurements 1-3 months after implementation of FEMG practices.
An example of the customisation and implementation process for the FEMG is summarised in Figure 5 and consists of several steps. Fluoride Emission Management Guide (FEIV1G) Customisation & Implementation Plan External Audit & Customisation Visit Audit:
Customisation Material:
On Operating, Control & Maintenance Practices as per FEMG
3.g. Photos of good / bad practices, diagrams, etc from Smelter
Improvements to Practices Identified
FEMG Customisation for Smelter
If adopted, smelters stand to achieve significant improvement in environmental performance and operational standards for fluoride emissions reduction. Case Study
Strategic Planning & Implementation of FEMG Training of Staff
Customised FEMG ■ Training Workshops
A smelter case study with the FEMG was carried out by the LMRC. An audit was conducted in two sections of a potroom and in a GTC area in a smelter, using the checklists provided in the FEMG. During the auditing process, some good practices and improvement opportunities were recognised. One example was the use of anode rod seals to minimise emissions from the pathway shown in Figure 6.
Adjust Practices
As per Recommended Improvements
Fluoride Measurements
Monitor Changes with Implementation
Regular Audits
Monitor Changes to Practices
Figure 5. Customisation and implementation process of FEMG. External Audit and Customisation Visit The purpose of an initial audit and customisation visit is to obtain relevant information and data on smelter technology, as well as smelter's operation, control and maintenance practices to customise the FEMG. The information obtained will ideally include photos of bad/good work practices, operational schedules, smelter layout, and other relevant smelter information. Concurrently, operational, control and maintenance practices in the smelter will be reviewed to assess the current practices. The audit will be based on the audit guidelines/checklists and recommended best practices specified in the FEMG.
Figure 6. Schematic diagram offluorideemission from the gap left between the anode rod and the skirt due to damage or missing seal
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The openings in the cell superstructure around the anode rods are designed to accommodate anode change process. Prior to the audit and training sessions, most of the operators did not understand the role of anode rod seals. Hence, no routine check on the condition of the anode rod seals was done. On average per pot, five seals were missing and seven seals were damaged.
References 1. A. Moras et al, "Modern Potline Gas Treatment Technology for High Amperage Pots - The Alcoa Fjardaal Experience," Light Metals, 2010, pp 239-242. 2. M. Stam et al, "Operational and Control Improvements in Reduction Lines at Aluminium Delfzijl," Light Metals, 2007, pp 243-247. 3. M. Meyer, G. Girault, and J. M. Bertolo, "Development of a Jet Induced Boosted Suction System to Reduce Fluoride Emissions," Light Metals, 2009, pp 287-292. 4. "Alcoa in Australia: Smelter Fluoride Emissions - Point Henry and Portland Aluminium," taken from http://www.alcoa.com/australia/en/info page/environ air.asp 5. H. P. Lange et al, "Innovative Solutions to Sustainability in Hydro," Light Metals, 2008, pp 211-216. 6. S. Lindsay, TMS Alumina Short Course (2004) 7. S. J. Lindsay, "Effective Techniques to Control Fluoride Emissions," Light Metals, 2007, pp 199-203. 8. Emission targets for 2010. OSPAR Recommendation 98/2 on Emission and Discharge Limit Values for Existing Aluminium Electrolysis Plants, www.ospar.org 9. US EPA (1997) "National Emission Standards for Hazardous Air Pollutants for Source Categories; National Emission Standards for Hazardous Air Pollutants for Primary Aluminium Reduction Plants; Final Rule," 40 CFR Parts 9, 60, and 63. 10. E. Patterson et al., "Reducing Fluoride Emissions from Aluminium Electrolysis Cells," (Proc. 7 th Australasian Aluminium Smelting Technology Conference and Workshop, 2001) 11. M. M. Hyland, B. J. Welch, and J. B. Metson, "Changing Knowledge and Practices towards Minimising Fluoride and Sulphur Emissions from Aluminium Reduction Cells," Light Metals, 2000, pp 333-338. 12. U.S. Agency for Toxic Substances and Disease Registry (2003) Toxicological Profile for Fluorides, Hydrogen Fluoride and Fluorine.
Through the training sessions, the importance of anode rod seals on maintaining pot suction and therefore, keeping the fluoride in the pot becomes clear to the operators. These anode rod seals can reduce fluoride emissions by as much as 20% [7]. After the training sessions, regular checks were scheduled on anode rod seals whereby, damaged seals were repaired and missing seals were replaced immediately. Feedback obtained from the operators was generally positive. Operators found the guide to be very practical, particularly the pictures or schematic drawings used to illustrate the problems and the recommendations. The theory and background information contained in the FEMG for each practice provides a better understanding of the importance of each operation - through this understanding, it is easier for operators to implement the FEMG's best practice recommendations. Implementation of the FEMG was also a good opportunity to test the robustness and practicality of the guide. Feedback from the operators and smelter personnel was integrated into the guide, improving the quality and usefulness of the guide. Conclusions The mission to manage and reduce fluoride emissions extends across all areas of the smelter. Every individual in the smelter is responsible and plays a vital role in achieving this goal. Although the potroom and the GTC are likely to be the main areas contributing to fluoride emissions, it is important to realise that other plant areas, such as the bath plant, may also contribute to fluoride emissions. The FEMG provides a structured approach to a better understanding of the factors affecting and the ways of reducing fluoride generation and emissions. The ultimate aim of this customisable guide is to provide practical and technical information to help smelters achieve significant improvements in their environmental performance with their current technologies. Through careful management of operation, process control and maintenance practices across the plant, significant reduction in fluoride emissions and improvement of operational standards can be achieved in the smelters. Acknowledgement The authors would like to thank the Australian Aluminium Council (AAC) - acting on behalf of the Asia Pacific Partnership (APP) on Clean Development and Climate Aluminium Task Force - for their generous sponsorship of the FEMG and Dr. Eng Fui Siew for his extensive contributions to the development of the FEMG.
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Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
Light Metals 2011 ALUMINUM REDUCTION TECHNOLOGY
Enviroment- Emissions / Anode Effect II SESSION CHAIR
Marco Stam Aluminium Delfzijl Delfzijl, Netherlands
Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
On Continuous PFC Emission Unrelated to Anode Effects Wangxing Li, Qingyun Zhao, Jianhong Yang, Shilin Qiu, Xiping Chen Zhengzhou Research Institute of CHALCO, Zhengzhou, China, 450041 China National Engineering & Technology Research Center for Aluminum, Zhengzhou, China, 450041 Jerry Marks J Marks & Associates, LLC, 312 NE Brockton Dr, Lees Summit, MO 64064 Chris Bayliss International Aluminium Institute, London SW1Y4TE, United Kingdom Keywords: Continuous PFC, reduction cells, anode effects, smelters Abstract An investigation was made into operating parameters that could result in continuous PFC emission unrelated to anode effects in two smelters. Anode current distribution was measured in QY smelter. No obvious correlation was observed between the nonhomogenous anode current and continuous PFC emission. Bath temperature and alumina concentration were synchronously measured during continuous PFC monitoring in LX smelter. No continuous PFC emissions were observed when reduction cells are operated at stable bath temperature and alumina concentration. No clear relationship emerged among these factors and further surveys need to be made in other smelters. The effect of feeding mechanism was also investigated. Metal tapping and anode exchange can disturb cells' balance, which result in cell voltage rising, and then may cause continuous PFC emission. It is found that the continuous PFC emission only occur in some particular cells and is different from cell to cell even with the same line current. The continuous PFC emission is only a small portion in total PFC emission.
Investigation Objectives in QY smelter Single-cell studies were performed on two 320 kA cells. Objectives of the study were as follows: (a) Determine the proportion of continuous emissions unrelated to anode effects to the overall PFC emissions, (b) Determine if continuous emissions occur in all cells or in only a few cells. (c)Determine if continuous emissions could take place as a result of non-homogenous current distributions among cell anodes. Investigation Objectives in LX smelter Single-cell measurement was performed on two 176 kA cells which belong to different potlines in LX smelter. The purposes are to determine if bath temperature, feeding mechanism and alumina concentration are the factors affecting continuous PFC emissions. Results and Discussions Investigation in QY Smelter
Introduction PFC survey was performed in CHINALCO smelters (1) using accepted IAI/USEPA measurement methodology. The survey was done in different running statuses such as normally-operated, power-limit, newly-started and new control-mode. PFC emission of normally-operated smelters is close to that of western ones (2). During this survey non-AE PFC emission, however, also called as continuous PFC emissions were first observed. These continuous PFC emissions occurred when no anode effects were taking place. What caused these continuous PFC emissions? Is it universal for all cells or particular for certain cells? Could it be eliminated? A research plan to study continuous PFC emission from Chinese smelters was made in China National Engineering & Technology Research Center for Aluminum/Zhengzhou Research Institute of Chalco (ZRI). Results from the first phase of the study are discussed in this paper.
Gas sample came from two different single cells. Cell numbers were QY-320-6 and QY-320-7. Monitoring duration for QY-320-6 was eleven hours. Two AE peaks were collected (see Figure la). The maximum CF4 emission rate is 0.463 g/s. And CF4 emission is almost zero when there are no anode effects (see Figure lb). Results are shown in Figure 2a from QY-320-7 during six-hour measurement, during which one AE peak was monitored. The maximum CF4 emission rate is 0.848 g/s. During most of measurement period, CF4 emission is close to zero when no anode effects occur (see Figure 2b). Anode Current Distribution Measurement Anode current of cells was usually measured with Millivolt Fork. Voltage drop is measured on every anode rod between two probes at fixed distance apart. The current in each rod is:
Research Plan and Objectives
_ AV; _ AVjA
The existence of continuous PFC emission is like a puzzle. A research plan was developed in ZRI to solve the puzzle. Two smelters were selected to perform studies on continuous PFC emission in the first phase of the plan. One is a pilot plant (code QY), and the other is a common industrial smelter (code LX). QY has two potlines, 320kA cells and 162kA cells respectively. LX has three potlines, all are 176kA cells.
' ~ R, ~ P{T, )L
Where: AVi = mV drop, Ri = resistance of the rod between the fork tines, A = cross-section of the rod, pi = resistivity of the rod (function of rod temperature Tj between the tines), L = distance between fork tines. In cell control practice it is assumed that all the rods have the same
309
temperature. Then, the resistance is the same in each rod and will fall out in the calculation. The sum of measured electric currents in individual rods is then easily normalized to the line current. Measured current distribution in anode rods is normalized to line current in the following way: _ _ < ¢
r
Ä
r
Ó( ^ Line current, kA; (AV)i = Individual rod voltage, mV; N = Number of anode rods. Anode current of two QY-320 cells was measured, calculated and plotted (see Figure 3).
-i
·—i
0
50
«—i
100
'
1
·
150
1
200
«—i
250
1—·
300
1
350
' — i
400
Monitoring duration, min
Figure 2a Original CF4 Emission Trace of QY-320-7
0.5-
0.4-
0.3-
"Φl
2 o 0.2-
0.1 -
0.00
100
200
300
400
500
LJ 600
700
Monitoring duration, min Monitoring duration, min
Figure la Original CF4 Emission Trace of QY-320-6
Figure 2b Continuous CF4 emission of QY-320-7
0.05QY-320-6 X
bC
0.04""«12. 0
i éï.ï
TJH"
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Figure lb Continuous CF4 emission of QY-320-6
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Graph lb came from graph la by cutting the AE peak which appeared after 660 minutes' monitoring, they took over the same measurement period of QY-320-6. Graph 2b came from graph 2a by cutting the AE peak which came forth after 350 minutes' monitoring, they also described the same measurement period of QY-320-7.
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Anode current distributions of QY-320-6 and QY-320-7 were measured once two hours. First current measurement was performed when PFC monitoring had gone on 120 minutes. The mean value of five measurements from QY-320-6 was plotted in Figure 3a. Three measurements were performed on QY-320-7 and the mean value was plotted in Figure 3b. Anode current distribution in QY smelter is similar to anode current distribution variability measured in Western cells (3). There is no obvious relationship between maximum anode current and continuous PFC emission.
It can be learned from investigation in QY smelter that continuous CF4 emissions are only small portion of the total PFC emissions. The amount of continuous PFC varies from cell to cell from near zero to a detectable level. The measured variability of current flow among the anodes is similar to that of Western reduction cells. There was no clear correlation of continuous PFC emission with maximum anode current. Investigation in LX smelter
Percentage of Continuous PFC Total CF4 and non-AE CF4 from two cells were calculated (see Table I). Continuous PFC emission is only small portion of the measured PFC emissions from 320 kA cells. There are different CF4 emissions from different cells even with the same line current.
Single-cell measurement was performed on two 176 kA cells which belong to different potlines in LX smelter. The purposes of this investigation were to determine if bath temperature, feeding mechanism and alumina concentration affect continuous PFC emission. Results from LX176-334 cell Exhaust gas from LX176-334 cell was sampled and monitored continuously for 47.54 hours. No AE peaks were observed. Continuous CF4 emission was almost zero during the whole measurement (see Figure 4).
Table I Proportion of Continuous CF4 Emission
Cell number
QY-320-6 QY-320-7 QY-320-7
Total CF4
188.55 443.74 443.74
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311
was 87%. No clear relationship was found between alumina concentration and continuous PFC emission.
During measurement, the change trend of feeding interval is described in Figure 5 together with cell voltage. It is obvious that feeding is under good-control. No clear correlation was found between feeding mechanism and continuous CF4 emission. As compared Figure 5 with Figure 4, it can be seen that detectable CF4 emitted soon after setting new anode. Bath temperature and alumina concentration were measured during PFC monitoring (see Figure 6 and Figure 7). Bath temperature was measured with K-thermo-couple which has autoreading function. Alumina concentration was measured with LECO-RO500C.
300
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Figure 7 Trend of alumina in bath during measurement
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Results from LX176-435 Cell Measurement duration is 45.61 hours for LX176-435 cell. No AE peaks were observed during monitoring as well. Continuous CF4 emission is near to zero during the whole measurement (see Figure 8). Change trends of cell voltage and feeding interval are plotted (see Figure 9). As compared Figure 9 with Figure 8, no clear correlation was found between feeding mechanism and continuous CF4 emission. Metal tapping and anode exchange caused rising of cell voltage. It can be seen that metal tapping could cause a little continuous CF4 emission. Bath temperature of LX176-435 was also measured. Bath temperature changed in a narrow range of 924-933 °C during the whole measurement (see Figure 10). There is no clear evidence to show continuous CF4 emission cause by unstable bath temperature.
1200 1500 1800 2100 2400 2700 3000
Running time, min
Figure 6 Trace of bath temperature during measurement The maximum bath temperature difference in LX176-334 cell was 13 °C . As compared Figure 6 with Figure 4, it seems that continuous CF4 easily emits at lower bath temperature. The influence of bath temperature on continuous PFC needs further investigation. Benefit from good feeding mechanism, alumina concentration in the cell showed stable and in good-control (see Figure 7). The percentage of alumina concentration in the range of 2.0%~3.0% D.00T
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Figure 9 Change curves of cell voltage and feeding interval during measurement exchange and metal tapping disturb the stability of cells, and could cause continuous emissions. Continuous PFC emission from LX176 cells was almost zero. Influences of bath temperature and alumina concentration on continuous PFC emission are not clear in the survey of LX176 cells. Perhaps low line current could play a role in continuous PFC emission. More surveys need to be done on higher amperage cells. Future Plan
0
300
600
900
Anode exchange and metal tapping cause instability of cells in the short-term, which may result in continuous PFC emission. More studies should be done to determine further whether anode exchange or metal tapping cause continuous PFC emission. Continuous PFC emission may take place when bath temperature goes down. More investigation should be carried out to determine the further influence of bath temperature. More detailed survey need to be done to find out if local alumina concentration influences PFC emission. More investigation is needed to determine if higher amperage (>300 kA) is easier to cause continuous CF4 emissions than lower amperage cells like <200 kA. Reduction of PFC emissions requires efforts from smelters all over the world. ZRI would like to strengthen cooperation with IAI in the future. More surveys will be done with the help of a more powerful and flexible equipment donated by Canadian Aluminum Association and Environment Canada as part of the Asia Pacific Partnership on Clean Energy and Climate (4).
1200 1500 1800 2100 2400 2700 3000
Running time, min
Figure 10 Trace of bath temperature during measurement There was a good feeding mechanism, stable bath temperature, quite homogenous alumina concentration in LX176 potlines. Therefore, very low AEF took place in LX176 potlines and the continuous PFC emissions in the measured cells were almost zero. Because anode exchange and metal tapping disturbed the stability of reduction cells, they would have some effects on continuous PFC emission. Perhaps line current in LX smelter was a bit low so that it had very low continuous PFC. No clear correlation was found among bath temperature, alumina concentration and continuous PFC emission.
Acknowledgement This work is supported by The National Natural Science Foundation of China (No. 50974127) and The National Key Technology R&D Program (No. 2009BAB45B03).
Conclusions Different cells have different CF4 emission level even with the same line current. Continuous PFC emission from QY smelter is only a small portion of total PFC emissions. The measured variability of current flow among the anodes is similar to that of Western cells. No obvious relationship was found between maximum anode current and continuous PFC emission. Anode
References 1. Protocol for Measurement of Tetrafluoromethane (CF4) and Hexafluoroethane (C2F6) Emissions from Primary Aluminum
313
Production. U.S. EPA (Washington, D.C.) and IAI (London, U.K.). April 2008 2. Results of the First Survey in CHINALCO Smelters, Zhengzhou Research Institute of Chalco, May 2009 3. Perfluorocarbon Emissions and Primary Aluminum Production, Jerry Marks, February 2010 4. Asia Pacific Partnership on Clean Energy and Climate, Available at http://www.asiapacificpartnership.org/, September 29, 2010.
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Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
Monitoring airfluorideconcentration around ALUAR smelter in Puerto Madryn (Chubut Province, Argentina) 1
Jorge Zavatti1; Claudio Lopez Moreno1; Juliana Lifschitz1; Gabriela Quiroga1 ALUAR Aluminio Argentino SAIC, Ruta Nacional N° A010 s/n; Puerto Madryn, Chubut, U9120OIA, Argentina Keywords: Fluoride inmission, air monitoring network, method validation The emission of Total Fluoride (gaseous + particulate) (F) is a significant environmental aspect of the aluminum industry. Therefore the government have established regulations aimed at controlling the environmental performance of smelters through monitoring of F" emissions and inmission.
Abstract The emissions of Total Fluoride (gaseous + particulate) (F) are one of the significant environmental aspects of the aluminum industry. Therefore the government has established regulations aimed at controlling the environmental performance of smelters through monitoring of F" emissions.
The Authority of Chubut Province (Argentina), sets emission limits at 1 kg FVton Al and following recommendations from the World Health Organization, a 16.0 μg F7m3 guideline for inmission has been established [3].
Chubut Province (Argentina), sets emission limits, 1 kg FVton Al and following recommendations from the World Health Organization, a 16.0 μg F/m3 guideline for inmission has been established.
Although measurements were started years before, since 1998 ALUAR has been operating an Air Quality Monitoring Net (RMCA stands for its initials in Spanish), with the purpose of systematically measuring the air Total F concentration in the city of Puerto Madryn and surroundings.
In this paper we present: i)the monitoring network that ALU AR deployed in Puerto Madryn, 12 air sampling stations; ii)sampling procedures and analysis adopted; iii)the results of air fluoride concentration from July 2009 to June 2010; iv)the validation of these results by three scientific institutes of Argentina.
Sampling and analysis are based on ASTM D3267 [4] and ASTM D3270 standards. However some modifications have been made on sampling in order to simplify the procedures. Two samples are taken every week from each and every one of the RCMA stations, sampling periods being of 72 and 96 hours. The results are then recalculated for and average of 24 hours.
The results show that in Puerto Madryn (100000 inhabitants; 2 km to the South of ALU AR's Plant), the median concentration in air was 1.61 μg F/m3 (N = 471 - p95 = 5.32). These values met the inmission guideline level determined by the supervisory authority.
In this context some doubts were publicly raised concerning the accuracy of the information get by ALUAR. These doubts were mainly based on the fact that the sampling was not made following in all the details an international standard. Some objections were also made to the design and location of the monitoring stations
Introduction The ALUAR smelter was originally started in 1974 with 400 SWPB and open pots, working at 150 kA. Starting at 1987 the original pots were transformed to PFPB and gas treatment centers were installed. This and other technical improvements allowed the original potlines to reach now 200 kA, with a high current efficiency and a good general performance [1]. In 1999, ALU AR started its expansion using API8 pots, later upgraded to AP22. This expansion process is still underway.
Therefore, and in order to ensure the quality of the information sent to the authorities and communicated to the residents of the city, Aluar decided to ask to three well known research centers in Argentina (CNEA [5], INQUIMAE [6], PLAPIQUI [7]), to make independent monitoring of the F in air around the smelter.
ALUAR site (42° 44' 21.35" S - 65° 02' 46.73" W) (4 potlines, 736 PFPB cells and 36,000 ton Al/month), is located at 2 km to the North of the city of Puerto Madryn, with 100000 habitants, in the Province of Chubut.
It was required to each Institute to select points close to some of the existing monitoring stations, to design and install sampling stations at these points, to take samples representing at least 1500 hours and to analyze them. The sampling and analysis procedures should be selected by each Institute. The ASTM D3267, ASTM D3268, ASTM D3269 and ASTM D3270 standards were used.
Puerto Madryn is located in an environmentally sensitive area of the Atlantic coast of Argentina, characterized by the presence of flagship species of marine mammals like the Southern Right Whale.
The monitoring by the three Institutes, which made their work in a completely independent way and without exchange of information between them, was made between July 2009 and June 2010. During that time, ALUAR made his own monitoring as usual allowing a direct comparison of the results.
Due to the presence of these species, some of which are endangered, in 1999 the UNESCO inscribed the Peninsula Valdes, very close to Puerto Madryn, on the World Heritage List [2]. The region is also a very important touristic area. The climate of Puerto Madryn is characterized by low rainfall (200 mm/year), prevailing winds from the southwest and an annual average temperature of 13.6 ° C.
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• Due to practical reasons the ALU AR sampling is made for 72 or 96 hours, while the regulations consider the average concentration for 24 hours. Therefore, the values are affected by a factor >1 to take into account the difference in the sampling time [8] Purposely, both corrections tend to increase the final result. In this way, ALU AR always takes a "safe" position with reference to the adherence to the environmental regulation. The reported concentration could be overestimated but never underestimated. Table 1: CNEA validation of results from the ALUAR-RMCA. Monitoring period from July the 28th to November the 13th of 2009. Research Center: RMCA-ALU AR CNEA Station H (3.8 km to the West of ALU AR) | Median: 1.35 F μg/m3 1 Median: 0.34 F μg/mi 3 3 Percentile 5: 0.21 F μg/m Percentile 5: 0.21 F μg/m Percentile 95: 4.10 F μ/m3 Percentile 95: 2.34 F μg/m3 N:67 N:32 Station E (0.7 km to the East of ALU AR) | Median: 1.03F^g/m3 Median: 1.79F>g/mj Percentile 5: 0.32 F μg/m3 Percentile 5: 0.84 F μg/m3 Percentile 95: 6.79 F μg/m3 Percentile 95: 3.82 F μg/m3 N:31 N:67 Station O (0.8 km to the North of ALU AR) | Median: 0.48 F μg/m3 Median: 1.12 F>g/m^ Percentile 5: 0.31 F μg/m3 Percentile 5: 0.60 F μg/m3 Percentile 95: 3.45 F μg/m3 Percentile 95: 2.60 F μg/m3 N:65 N:31 The number of samples for station differs between CNEA and the RMCA since CNEA took samples of 24 hours of duration, while RMCA Net covers periods of 72 and 96 hours.
Figure 1. ALUAR air sampling network at Puerto Madryn city. Results The results produced by the RMCA show that in Stations A, B, C, N and D located at urban areas in Puerto Madryn city (100000 inhabitants; 2 km to the South from ALU AR's Plant), the median of concentration of F" in air is 1.61 μg /m3 (N = 471 -p95 = 5.32). These values met the inmission guideline level determined by the supervisory authority (16 μg F/m3).
All the air F" concentration values shown in tables 1, 2 and 3 proves that the emission generated by ALU AR operations, do not exceed the corresponding guideline for inmission (16 μg F/m3), set by the environmental authorities of the province of Chubut (Argentina).
The average concentrations of these urban stations do not differ significantly between them and it was not possible to link the temporal evolution of these F concentrations to the wind performance.
At the same time it is worth mentioning that the three research centers that participated in the validation of data systematically report concentration values that are lower than those reported by RMCA in each instance.
Tables 1, 2 and 3 show the results of air F concentration obtained by the Research Centers and those obtained by the ALUAR RMCA, taken simultaneously and at points very close between them, for different periods of monitoring.
Conclusion • The monitoring by three independent and prestigious Institutes has confirmed the correctness of the F" inmission values reported by ALU AR
In most cases the ALU AR values are higher than those obtained by the Institutes. These differences can partially be explained by two reasons:
•The values reported by ALU AR tend to be higher than those obtained by the Institutes. This is explained by the use of corrections that ensure that the concentrations can be overestimated but cannot be underestimated.
• The ALU AR values are corrected in order to take into account the loss of absorbent liquid during the sampling. The ASTM D3267 does not consider any correction for this loss.
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• The results of this study validate the inmission results reported by ALU AR to the authorities since 1998.
References [1] P. Navarro, C. Daviou, L. Daurade. "ALUAR'S al20 PROJECT: A SUCCESSFUL WAY UP TO 200kA". IV International Aluminium Congress (May 18th, 19th and 20th, 2010-Säo Paulo): 594-609. [2] http://whc.unesco.org/en/decisions/2534 [3] World Health Organization. "Air Quality Guidelines for Europe 2nd Edition", WHO Regional Publications, European Series, No. 91 (2000): 143-145. [4] American Society for Testing and Materials (ASTM) (www.astm.org). [5] CNEA http://www.cnea.gov.ar/cac/utt nuevo/unidades 1 .htm [6] INQUIMAE - http://www.inquimae.fcen.uba.ar [7] PLAPIQUI - http://www.plapiqui.edu.ar [8] D. Bruce Turner. "Workbook of Atmospheric Dispersion Estimates". (1994): 4-8 to 4-9.
• The F" concentrations around the ALU AR smelter satisfy the regulations sets by the authorities for the environmental protection • The results confirm that the potroom operating procedures and emission control equipment, originally started in 1988, are still operative for the environmental protection, in spite of the large increase in the production of the plant. Table 2: INQUIMAE validation of results from the ALUARRMCA. Monitoring period from December the 4th of 2009 to March the 19th of 2010. Research Center: RMCA - ALU AR INQUIMAE Station O (0.8 km to the North to ALUAR) Median: 1.66 F μ^ιη Percentile 5: 0.47 F" μg/πrJ Percentile 5: 0.54 F μg/m3 3 Percentile 95: 2.31 F μg/m Percentile 95: 2.84 F μg/m3 N:24 N:26 Station C (3.5 km to the South to ALU AR) Median: 0.84 F μ^πΓ Median: 1.44 F μg/m Percentile 5: 0.46 F μg/m3 Percentile 5: 0.20 F μg/m3 3 Percentile 95: 1.58 F μg/m Percentile 95: 5.15 F μg/m3 N:27 N:30 Station D (3.5 km to the South-East to ALU AR) Median: 1.98F>g/m3 Median: 1.11 F μg/m Percentile 5: 0.73 F μg/m3 Percentile 5: 0.38 F μg/m3 3 Percentile 95: 9.82 F μg/m3 Percentile 95: 1.91 F μg/m N:26 N:27 Both INQUIMAE and RMCA samples were taken covering periods of 72 and 96 hours. Table 3: PLAPIQUI validation of results from the ALUARRMCA. Monitoring period from June the 2nd to August the 9lt of 2010. Research Center: RMCA - ALU AR PLAPIQUI Station E (0.7 km to the East to ALU AR) Median: 1.95 F μg/m3 Median: 2.69 F μg/mJ Percentile 5: 1.17 F μ^πι3 Percentile 5: 0.66 F μg/m3 3 Percentile 95: 6.90 F μg/m3 Percentile 95: 8.13 F μg/m N:20 N:29 Station C (3.5 km to the South to ALU AR) Median: l!53F]Ig/m1 Median: 0.21 F μ^πι' 3 Percentile 5: 0.52 F μ^πι3 Percentile 5: 0.04 F μ^πι 3 Percentile 95: 4.28 F με/m3 Percentile 95: 0.99 F μ^πι N:20 N:27 Station G (2.8 km to the North-East to ALU AR) Median: 0.19 F^g/m 3 Median: 2.46 F μg/πr Percentile 5: 0.09 F μg/πrs Percentile 5: 1.13 F μ^πι3 3 Percentile 95: 0.40 F μg/m Percentile 95: 8.51 F με/πι3 N:22 N:20 Both PLAPIQUI and RMCA samples were taken covering periods of 72 and 96 hours.
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Light Metals 2011 Edited by: Stephen J. Lindsay TMS (The Minerals, Metals & Materials Society), 2011
REDUCTION OF ANODE EFFECT DURATION IN 400KA PREBAKE CELLS Wei Zhang1, David Wong1, Michel Gilbert1, Yashuang Gao1, Mark Dorreen1, Mark Taylor1 Alton Tabereaux2, Melinda Soffer2, Xiaopu Sun2 Changping Hu3 Xueming Liang4, Haitang Oin4 Jihong Mao5, Xuehui Lin5 1 Light Metals Research Centre, the University of Auckland, New Zealand institute for Governance and Sustainable Development, US 3 China Nonferrous Metals Industry Association 4 Henan Zhongfu Industrial Co. Ltd, China 5 Northeastern University Engineering& Research Institute, China Keywords: aluminium electrolysis, PFC, anode effect stress to the cell lining from expansion and contraction, melting of the side ledge causing the electrolyte composition to change and increased anode consumption. Other negative impacts include increased hydrogen fluoride emissions, increased operator intervention, exposure of operators to safety risks during manual AE termination, bath spillages and floor damage. The primary aluminium industry is striving to decrease both the number and duration of anode effects [5]. Moreover, with the prospect of a future price on carbon, the emission of these greenhouse gases will impact production costs. It has been calculated that with a C0 2 tax of 15 US $ /ton C0 2 , each anode effect minute per day per cell will increase production cost by about 1.2% [7].
Abstract In order to improve energy efficiency and reduce green house gas emissions, the aluminum smelting industry has been continuously working on reducing both anode effect frequency (AEF) and duration (AED). However, there is still a long way to go to achieve zero anode effect (AE) on very high amperage, low specific power consumption cells due to the added complexity of the process. A new program to quickly terminate AEs has been developed by Light Metals Research Centre, the University of Auckland, in conjunction with the efforts of the Asia Pacific Partnership on Clean Development and Climate (APP) to facilitate investment in clean technologies and to accelerate the sharing of energy efficient best practices. A pilot project was initiated to test an automatic Anode Effect Termination (AET) program on 400kA cells in Zhongfu, China. This paper demonstrates the success of the new anode effect termination (AET) program in killing AEs on this cell technology without conflicting with normal cell operations. The resulting decrease in average anode effect duration (AED) is demonstrated.
The frequency and the duration of anode effects depend upon the technology of the cell, the operational procedures and also upon process control at each smelter. Hence, the amount of CF4 and C2F6 may vary from one plant to the other, according to these parameters. Some types of pot allow more sophisticated interventions to minimize the consequences of an anode effect, such as computer process to control the feed and voltage [8]. Several strategies for AE quenching have been tested and implemented in aluminium smelters, such as pumping or quenching of the anode system, or lowering anodes until they are short-circuited by the metal [9]. Recently, a new AE quenching method has been proposed by Pablo Navarro et al. While a pot is in an anode effect a high proportion of electrical current is conducted from the sides rather than the bottom surfaces of the anodes. This is because the insulating gas layer considerably hinders the flow of current through their bottom surface. This produces an increase in the horizontal current densities and therefore the magnetohydrodynamic (MHD) instability in a pot during the AE. With Navarro's AET strategy, the anode beam is lowered to a particular ACD, at which the MHD instability of the pot is further increased, generating a self-sustained metal wave that short-circuits the anodes, resulting in removal of the insulating gas layer and ultimately termination of the AE. An intense bath circulation is also generated, redistributing alumina within the bath. The particular ACD required to do this is characteristic of each pot technology and depends upon the MHD design. Compared to the traditional AET strategy of physically immersing anodes in the metal pad to short circuit the anodes, this new approach is more energy efficient and faster in killing AEs [9-11].
Introduction The primary aluminium production process has been identified as the largest anthropogenic source of two kinds of perfluorocarbon (PFC): tetrafluoromethane (CF4) and hexafluoroethane (C2F6). PFCs comprise one category of major greenhouse gases with long atmospheric lifetimes and significant global warming potential (GWP). PFCs have an estimated atmospheric lifetime of 10,00050,000 years, and an estimated Global Warming Potential (GWP) of 6,500 and 9,200 times that of carbon dioxide (C02) [1]. Human-made PFCs are very effective absorbers of infrared radiation, so that even small amounts of these gases contribute significantly to the radiative forcing of the climate system [2]. Therefore the aluminium smelting industry has been trying to reduce PFCs emissions [3, 4]. PFCs are generated during anode effects (AEs) in aluminium electrolysis process. Anode effects occur when the concentration of dissolved alumina in the molten bath reaches low values or anode current density is higher than the critical current density [5]. During an anode effect, an insulating layer of gas bubbles appears under the anodes, increasing pot voltage from 4-5V to more than 8V, up to 40-50V, which depend on the type and operating conditions of pots) [6]. In addition to causing environmental damage from the production of PFCs, AEs have other negative impacts. These are increased energy consumpt