Hydrometallurgia a Presion

ISBN: 978-1-926872-10-0

PRESSURE HYDROMETALLURGY 2012

42nd Annual Hydrometallurgy Meeting

EditorsM.J. Collins, D. Filippou, J.R. Harlamovs, E. Peek

Organized by

Pressure Hydrom

etallurgy 2012

2012

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FRIENDS & SUPPORTERS

EVENT SUPPORTERS

GOLD SPONSORS

THANK YOU TO OUR SPONSORS

SILVER SPONSORS

BRONZE SPONSORS

Pressure Hydrometallurgy 2012

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PROCEEDINGS OF THE 51ST CONFERENCE OF METALLURGISTS, HELD WITH PRESSURE HYDROMETALLURGY 2012,

42ND ANNUAL HYDROMETALLURGY MEETING, SEPTEMBER 30 TO OCTOBER 3, 2012, NIAGARA FALLS, ONTARIO, CANADA


Editors

M.J. Collins Sherritt International Fort Saskatchewan,

Canada

D. Filippou Rio Tinto Iron & Titanium

Sorel-Tracy, Canada

J.R. Harlamovs Teck Resources Trail, Canada

E. Peek Molycorp

Peterborough, Canada

Symposium Organized by the Hydrometallurgy Section of the Metallurgy and Materials Society of the Canadian Institute of Mining, Metallurgy and Petroleum

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A Publication of The Canadian Institute of Mining, Metallurgy and Petroleum

1250 - 3500 de Maisonneuve Blvd. West Westmount, Qubec, Canada H3Z 3C1

http://www.metsoc.org

ISBN: 978-1-926872-10-0 Printed in Canada Copyright 2012

All rights reserved. This publication may not be reproduced in whole or in part,

stored in a retrieval system or transmitted in any form or by any means, without permission from the publisher.

If you are interested in purchasing a copy of this book, or if you would like to receive the latest MetSoc Publication Catalog, please call: (514) 939-2710, ext. 1327

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Foreword

The Second International Symposium on Pressure Hydrometallurgy is organized by the Hydrometallurgy Section of the Metallurgy and Materials Society of the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) and constitutes its 42nd Annual Hydrometallurgy Meeting. A primary goal of these meetings is to bring together scientists and engineers representing metal producers, equipment suppliers, engineering firms, test laboratories and academia from around the world to share knowledge and exchange ideas. Pressure Hydrometallurgy 2012 builds on the foundations established by the first Pressure Hydrometallurgy conference, held in Banff in 2004.

Although pressure hydrometallurgy is becoming an increasingly mature and conventional

technology, it remains an exciting area of both study and application. Processes operating at high temperature and under pressure enjoy fast kinetics, enhanced solubilities of reagent gases, and improved liquid-solid separation characteristics, purity of product streams and stability of residues. Pressure vessels have been commercially used for the extraction of aluminum, nickel, cobalt, zinc, uranium and copper, for the oxidation of refractory gold feeds, and for the purification and recovery of various metals and metal by-products. New construction materials and autoclave design improvements continue to extend the application of these vessels, in an effort to keep up with innovative new processes being developed at the laboratory and pilot plant stage. The scope of the present symposium embraces all of these aspects. The manuscripts provided in these proceedings describe a wide range of commercial applications covering the above topics, as well as test programs and scoping studies in areas that have yet to be realized on a commercial scale. The future of this field remains bright.

Niagara Falls, Ontario, Canada October 2012

M.J. Collins D. Filippou

J.R. Harlamovs E. Peek Editors

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Editors Biographies

Mike Collins received a B.Sc. degree in Chemistry from the University of Victoria in 1977. After employment with the Defense Research Establishment Pacific and Royal Roads Military College in Victoria, B.C., he attended McMaster University in Hamilton, Ontario, where he studied the polyatomic cations of sulphur, selenium and tellurium in super-acid systems and received a Ph.D. in Chemistry in 1984. A visiting fellowship at the National Research Council in Ottawa followed, for studies in solid-state NMR spectroscopy of main group elements. In 1985, Mike joined Sherritt Gordon Mines Ltd. in Fort Saskatchewan, Alberta, becoming manager of the Sherritt Research group in 1997. Shortly thereafter, he became Manager, Research for the newly formed Metallurgical Technologies Division of Dynatec Corporation. Additional roles with Dynatec included Senior Consultant, and Manager, Research and Analytical Services, prior to the acquisition of Dynatec by Sherritt International Corp. in 2007. Mike is currently Director, Process Development within the Sherritt Technologies Division. Work at Sherritt and Dynatec has primarily involved laboratory and pilot plant testing in the field of pressure hydrometallurgy. Author of more than 75 technical publications and holder of eight patents, Mike is a former chair of the Hydrometallurgy Section of the Metallurgy and Materials Society of CIM and was co-chair of Pressure Hydrometallurgy 2004. He received the Sherritt Hydrometallurgy Award in 2005.

Dimitrios Filippou obtained a Dipl.Eng. degree in Mining and Metallurgical Engineering from the

National Technical University of Athens (Greece) in 1988, and a Ph.D. degree in Metallurgical Engineering from McGill University (Canada) in 1994. He has extensive experience in zinc, iron, arsenic and titanium hydrometallurgy in sulphate and chloride media. Since 2003, he works for Rio Tinto Iron and Titanium at the companys Technology Centre in Sorel-Tracy, Quebec, where he currently holds the position of Metallurgist, TiO2 Products. His fields of interest include R&D project management, process scale-up and costing, and Life Cycle Assessment. He has authored several papers and in 2010, he was awarded a TMS Best Paper Award for a publication in the JOM journal. Since 2008, he is the Secretary of the Hydrometallurgy Section of the Metallurgy and Materials Society of CIM. He is also a licensed engineer.

Juris Harlamovs was brought up and educated in England, obtaining a B.Sc. in chemistry and

going on to do a Ph.D. in zinc solvent extraction at the University of Birmingham. In 1980, he joined Anglo American Research Laboratories as a Research Metallurgist. Juris then joined Cominco Research, (now Applied Research and Technology, Teck Metals Ltd.) as a Senior Research Metallurgist and has been a Group Leader, Chemistry and Environment and a Principal Hydrometallurgist. Juris has broad interests in hydrometallurgy especially in copper, zinc, nickel, specialty metals and their associated impurities; but his first love still is solvent extraction. He is currently Section Leader, Copper and Zinc. This section provides technical metallurgical back-up for the various Teck Metals mines and smelter.

Edgar Peek received a M.Sc. (1991) and Ph.D. (1996) degree in Raw Materials Technology from

the Department of Mining and Petroleum Engineering of the Delft University of Technology in The Netherlands. Raw Materials Technology is the treatment of primary and secondary raw materials using a combination of Extractive Metallurgy, Mineral Processing and Recycling skills. In 1995 he joined the Research and Development group of Falconbridge Limited in Sudbury, Ontario to work on the laboratory and pilot plant aspects of nickel laterite flow sheet development. Key areas of focus were high pressure pumping, solid-liquid separation, heat exchange and autoclave operation. In 1999 Edgar joined the Projects & Engineering group located at the Kvaerner/Hatch offices in Toronto to work on the joint hydrometallurgical and pyrometallurgical scoping study for the Koniambo laterite ore deposit. Areas of responsibility were the overall site water balance and operating cost model. From 2002 to 2006 he was seconded to the Technical Services group of the Falconbridge Nikkelverk AS refinery in Kristiansand, Norway. He worked on medium-term plant support or Six Sigma stage gate projects in the areas of tailings disposal, solvent extraction, cementation, filtration and precious metal processing. In 2006 Xstrata Plc. acquired Falconbridge Ltd. and Edgar moved back to Sudbury, Ontario to become the Manager of Extractive Metallurgy at the newly created Xstrata Process Support (XPS). XPS provides commercial consulting and metallurgical test work services to the mining and minerals industry. While Edgar focused his time on topics such as nickel pig iron, process engineering support (Aker Solutions/Jacobs Engineering) and business development, the Extractive Metallurgy team centred on primarily pyrometallurgical services.

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Since 2011 Edgar is Manager of Process Development, Rare Metals Division, Neo Performance Materials (recently merged into Molycorp Inc.) based in Peterborough, Ontario. Author of more than 20 technical publications and holder of one patent, Edgar is the former Treasurer and current Chair of the Hydrometallurgy Section of the Metallurgy and Materials Society of CIM and was the co-chair of Chloride Metallurgy 2002.

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Table of Contents

Pressure Hydrometallurgy 2012 Foreword ......................................................................................................................................................... v Editors Biographies ..................................................................................................................................... vii

REFRACTORY GOLD Design of the AGA Brasil Refractory Gold Pressure Oxidation Plant ............................................................ 3

M. Collins, K. Buban, M. Faris, I. Masters and M. Antonio The Mobilization and Potential In-autoclave Recovery of Gold During Pressure Oxidation and Leaching of Auriferous Sulfide Materials ............................................................................................... 15

G.P. Demopoulos, J.-C. Parisien-La Salle and D. Blais Kinetics of the CESL Gold Process ............................................................................................................... 31

E. Asselin, P. Sauve and H. Salomon-de-Friedberg Enhancing CIL Gold Extraction by Hard Paraffin Wax Blanking of Double Refractory Ore Pre- Or Post-pressure Oxidation (POX) ........................................................................................................ 41

G. Van Weert, J. Jiang, O. Wang and Y. Choi Alkaline Pressure Oxidation of Pyrite in the Presence of Silica: A Surface Study ........................................ 53

A. Dani and V.G. Papangelakis Gold Extraction from Pyritic Ores Through Simultaneous Pressure Leaching/Oxidation ............................. 63

J.L. Valenzuela G., L.S. Quiroz C., J.R. Parga T., P.J. Valenzuela G. and P. Guerrero G.

MODELING Simulating the Iron(III) Precipitation Rate During Medium Temperature Chalcopyrite Oxidation in Batch and Continuous Autoclaves ................................................................................................................. 75

J.D.T. Steyl Use of Numerical Methods, Scale-up and Lab Tests in the Design of HPAL Autoclaves ............................ 91

J. Jung, W. Keller and A. Zucht Chemical Modeling in Pressure Hydrometallurgy Using OLI ..................................................................... 105

V.G. Papangelakis, A. Anderko, G. Moldoveanu, M. Carlos and G. Azimi

PROCESS DEVELOPMENT Influence of Iron on the Complexation and Oxidation of Quebracho: An Investigation on the Stability of Sulfur Dispersing Agent ...................................................................................................... 121

L. Tong, D. Dreisinger, B. Klein and J. Li Advances in the Use of Polytetrafluoroethylene During Medium Temperature Pressure Oxidation of Chalcopyrite ............................................................................................................................................ 137

E. Guerra

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Hydrothermal Production of Gallium Oxide ............................................................................................... 143

L. Becze, E. Peek, C.J. Ferron, G. Evans and K. Reading Synthesisof Li2TiO3 Powders by a Hydrothermal Process .......................................................................... 153

Y.J. Li, L. Li, C. Xu, V.G. Papangelakis and L. Kong Effects of Hydrothermal Processing on the Crystallinity and Morphology of the Precursor of In2O3 ....................................................................................................................................................... 163

Y. Li, Z. Liu, Q. Li, C. Wu and Z. Liu Pressure Leaching of Alkali-pretreated Limonitic Laterite Ore in Nitric Acid ............................................ 175

C. Wang, Y. Zhang and Z. Xu Low-temperature Pressure Leaching Kinetics of Zinc Sulphide Concentrate ............................................. 187

Z. Xu, F. Guo and C. Wang Treatment of Thompson Copper/Arsenic Residue ....................................................................................... 199

T. Xue, A. Singhal, I. Mihaylov and D. Wrana The Co-processing of Nickel Sulphide and Laterite Materials Using Low Oxygen Pressures .................... 211

R. McDonald, M. Rodriguez, J. Li, D. Robinson, M. Jackson and T. Hosken

COPPER The Mechanism of Chalcopyrite Leaching in Acidic Ferric/Ferrous Sulfate Media in the Presence of Silver-enhanced Pyrite ............................................................................................................................ 229

G. Nazari, D.G. Dixon and D.B. Dreisinger Reduction Kinetics of Fe(III) on Chalcopyrite Up to 110 C ...................................................................... 241

G.K. Yue and E. Asselin Hydrothermal Purification and Enrichment of Chilean Copper Concentrates. The Behavior of Bornite, Covellite, Pyrite and Enargite ........................................................................................................ 257

G. Fuentes and J. Vials Uniform Corrosion of Titanium Under Medium Temperature Chalcopyrite Concentrate Leaching Conditions .................................................................................................................................... 269

J. Liu, E. Asselin and A. Alfantazi Arsenic Stability and Characterization of CESL Process Residues ............................................................. 281

R. Bruce, K. Mayhew, G.P. Demopoulos and A. Heidel Chalcopyrite: Bioleaching Versus Pressure Hydrometallurgy ..................................................................... 295

F. Habashi

ZINC Leach Residue Filtration at the Hudbay Zinc Pressure Leach Plant ............................................................ 305

S. Shairp

Copper and Cadmium Removal from Zinc Sulphate Solution ..................................................................... 313 S.M. Rebelo and R.G. Helberg

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EQUIPMENT DESIGN Stress Analysis of Brick Lined Autoclaves ................................................................................................. 327

K. Brooks Effect of Pressure in Leaching of Low Grade Sulphide Ore at Ambient Temperature Development of Hydrostatic Pressure Reactor .................................................................................................................. 335

M.J. Latva-Kokko and T.J. Riihimki Direct Contact Heat Transfer Between Steam and Aqueous Slurries .......................................................... 343

D.T. White Refractory Lining Design for Pressure Oxidation Autoclaves ..................................................................... 367

I. Donohue, E. Barrette and M. Pearson Advanced Control of Pressure Hydrometallurgical Plants .......................................................................... 379

R.K. Jonas The Flash Recycle System for the Cooling of Autoclaves Process Design and Operation .......................... 391

F. Crundwell and N. Steenekamp Special Design Considerations for Pressure Hydrometallurgy Pilot Plants ................................................. 401

P. Martin Application of Tube Digestion Technology to the Non Ferrous Metals Industry ........................................ 413

J.A. Gorst, B. Haneman and W. Slabbert

PLATINUM GROUP METALS Development of the Twin Metals Minnesota Flowsheet Incorporating the Platsol Process .................... 427

J.A. Brown, C.A. Fleming and G. Barr The Chemistry and Mineralogy of a Nickel Copper Matte Leach ............................................................... 443

G.C. Summerton, D.C. Craig, P. Dinham, N. McCulloch and S. Dowling The Lonmin Platinum Base Metal Refinery Operations and Continual Improvements 1985 to 2012 ................................................................................................................................................ 457

N. Steenekamp and M. Turner-Jones Author Index ................................................................................................................................................ 469

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Refractory Gold

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DESIGN OF THE AGA BRASIL REFRACTORY GOLD PRESSURE OXIDATION PLANT

*M. Collins1, K. Buban1, M. Faris1, I. Masters1, M. Antonio2

1Sherritt Technologies 8301 113 Street

Fort Saskatchewan, AB, Canada T8L 4K7 (*Corresponding author: [email protected])

2 AngloGold Ashanti Brasil Fazenda Cristina s/n, Zona Rural, Brumal, Santa Brbara

Minas Gerais, Brazil

ABSTRACT

AngloGold Ashanti Brasil, with assistance from Sherritt International Corporation, has recently commissioned a new autoclave for the pressure oxidation of refractory gold feeds at their Corrego do Stio operation in Brazil. Design of the new plant, including feed acidulation, pressure oxidation, oxidized solids washing and solution neutralization circuits, was based on the results of test work conducted by Sherritt at their pilot plant in Fort Saskatchewan, Alberta. This paper describes the results of the pilot plant study as well as highlights of the design exercise.


Proceedings of the 42nd Annual Hydrometallurgy Meeting held in conjunction with the 51st Annual Conference of Metallurgists of CIM (COM 2012)

Niagara, ON, Canada Edited by

M.J. Collins, D. Filippou, J.R. Harlamovs, E. Peek

3

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INTRODUCTION

AngloGold Ashanti Brasil (AGA Brasil) has recently commissioned a refractory gold pressure oxidation plant at the site of the original So Bento operation in Brazil. The design of the So Bento pressure oxidation facility was based on Sherritt technology; Sherritt carried out the original process development test work in the mid-1980s, provided a process design package and detailed specifications for critical pressure oxidation equipment and participated in the training of operators and commissioning of the plant. The So Bento plant was operated between 1986 and 2007, when the original ore body was exhausted (Silva et al., 2004). AGA Brasil purchased the facility from Eldorado Gold in 2008, with the intent of refurbishing the plant to process refractory gold concentrates produced from the nearby Crrego do Stio mining area.

Sherritt International Corporation has a long history in the development and application of commercial pressure oxidation and pressure leaching processes. Sherritts development of technology for the pressure oxidation of refractory gold ores and concentrates commenced in the 1980s in collaboration with Homestake for application at the McLaughlin project in California. This was followed by several commercial applications in Canada, Papua New Guinea and Brazil (Berezowsky et al., 1991). More than thirty pilot plant campaigns investigating the pressure oxidation of refractory gold ores or concentrates have been conducted by Sherritt since the 1980s, and projects with Sherritt contributions to the process development studies and/or engineering design account for about half of the commercial plants that have operated to date with this type of process (Collins et al., 2011a; 2011b).

AGA Brasil contacted Sherritt Technologies, a division of Sherritt International Corporation, in August 2008 and requested proposals for test work and engineering activities to support the development of capital and operating cost estimates for the Crrego do Stio Project. The proposed scope of work included:

A review of preceding mineralogical and batch pressure oxidation studies, conducted by others, with the Crrego do Stio concentrates, and provision of commentary on the employed methods and the veracity of the data;

Process development studies, comprising batch test work, to identify the optimum operating conditions for treatment of concentrates produced from the Crrego do Stio area, including the Cachorro Bravo and Laranjeiras ore bodies, and a blend of ore samples recovered from both ore bodies;

Up to 120 hours of continuous pilot plant test work, including integrated operation of acidulation, pressure oxidation, countercurrent decantation wash and solution neutralization circuits;

Commensurate cyanide leach test work on selected samples from batch and continuous testing to demonstrate gold extraction from the pressure oxidation solids;

Complementary liquid-solids separation testing to provide data for sizing the commercial thickeners and filters;

A trade-off study, comprising limited batch testing and preliminary engineering activities, to compare the treatment of the Crrego do Stio concentrates under operating conditions that could be applied at the existing So Bento facility with the optimum operating conditions that could be accommodated using new equipment; and

Preparation of a process design package for the pressure oxidation, solids washing, solution neutralization and effluent treatment areas.

There was significant incentive to evaluate the treatment of the Crrego do Stio concentrates in the existing So Bento equipment. Of particular interest was the possibility of using the existing autoclaves at the So Bento site. The original autoclaves are limited to a maximum operating temperature of 190C and a pressure of 1 700 kPa(g), but are of sufficient size to allow extended pressure oxidation retention times for treatment of available Crrego do Stio concentrates.

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The original scope of process test work for the Crrego do Stio Project included a limited program of batch pressure oxidation (POX) tests under So Bento conditions followed by cyanide leach tests on the POX solids to evaluate gold recovery. All of these tests were to be conducted under the same methodology as the process development tests which examined higher temperature and pressure POX conditions, the so called Preferred Conditions case, which would require a new autoclave. However, following evaluation of the batch test results, Sherritt recommended that the continuous pilot plant campaign should include a period of operation to test the So Bento conditions and generate the required data for process comparison. Accordingly, a 48 h period of operations under So Bento POX conditions was also conducted during the pilot plant test.

The Sherritt process development test program commenced on January 5, 2009, and detailed engineering activities commenced in March 2010, culminating in start-up of the new commercial pressure oxidation circuit in January 2012. A chronology of the major events during execution of the project is provided as Table 1.

Table 1 Chronology of the Crrego do Stio project

Activity Date Receipt of feed materials for process development work December 22, 2008 Start of Sherritt batch tests January 5, 2009 Complete review of previous test work January 15, 2009 Issue plan for continuous pilot plant March 9, 2009 AGA Brasil personnel on site in Fort Saskatchewan March 23, 2009 Continuous pilot plant POX autoclave test March 23 to March 28, 2009 Issue pilot plant results update April 3, 2009 Issue trade-off study summary of process test work May 15, 2009 Issue final process development test report June 12, 2009 Basic engineering April to August 2009 Detailed engineering March 2010 to March 2011 Issue autoclave shell specification April 1, 2010 Autoclave shell delivered to site January 13, 2011 Feed on to autoclave January 12, 2012

This paper describes the results of batch and pilot plant test work carried out by Sherritt for AGA Brasil to confirm the amenability of the Crrego do Stio concentrates to treatment by pressure oxidation for extraction of the contained refractory gold and to determine the technical feasibility of treating the Crrego do Stio concentrates in the existing, or in a slightly modified, So Bento pressure oxidation facility. Highlights of the engineering activities, leading up to start-up of the new commercial facility, are also provided.

PROCESS DESCRIPTION

A block flow diagram of the flowsheet evaluated in the process development studies is shown in Figure 1.

Gold flotation concentrate slurry is fed to acidulation, along with recycled first wash thickener overflow solution as a source of acid for decomposition of the carbonates contained in the feed. Decomposition of the carbonates prior to pressure oxidation limits the quantity of carbon dioxide that may be evolved in the autoclave. In the commercial plant, CO2 evolution in pressure oxidation would require an increased vent rate to maintain the required oxygen partial pressure and thereby would reduce oxygen utilization.

The acidulated slurry is thickened and the underflow slurry is fed to the POX autoclave. Oxygen is added to each compartment of the autoclave to oxidize the sulphide minerals and liberate the gold. The

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flashed POX slurry is fed continuously to the first of a series of CCD wash thickeners. The underflow slurry from the last wash thickener is the feed to cyanide leaching.

The acidulation thickener overflow solution is forwarded to the solution neutralization circuit, where the free sulphuric acid and the acid generated by hydrolysis of metals in this step are neutralized with either limestone or flotation tailings. Metals precipitation is completed by addition of lime to further increase the pH. The neutralized slurry is thickened, and the overflow solution is forwarded to the CCD wash circuit to meet the wash water requirements.

Figure 1 Process flow diagram for treatment of Crrego do Stio concentrate

FEED MATERIALS

Two different concentrate samples, Cachorro Bravo and Laranjeiras, were received in Fort Saskatchewan from AGA Brasil on December 22, 2008, as well as a sample of flotation tailings for solution neutralization tests. Each sample was well mixed to produce uniform feed for testing. In keeping with the mine plan, the concentrates were blended in a 1:1 weight ratio for the bulk of the batch tests and for the feed to the pilot plant campaign. The chemical compositions of the individual concentrates and the 1:1 blend are summarized in Table 2. The concentrates were relatively fine, at 88 to 89% passing 27 m. The blend was prepared as a slurry of 56.0% solids for feed to the pilot plant campaign.

Table 2 Composition of Crrego do Stio concentrates tested during Sherritt pilot plant campaign

Feed Au Ag Al Sb As C CNAL* Fe Si S g/t g/t % % % % % % % % Cachorro Bravo 66.2 1.8 6.3 0.4 4.4 2.0 0.8 16.2 17.6 9.0 Laranjeiras 28.4 4.9 5.9 3.6 1.1 2.4 1.3 8.3 22.2 7.5 Blend 47.5 3.5 6.2 2.1 2.8 2.3 1.1 12.2 19.8 8.3

* CNAL = non-acid leachable carbon, equivalent to organic carbon

The flotation tailings contained 6.1% Al, 2.4% Ca, 2.1% C (most of which was carbonate), 5.0% Fe, 2.4% Mg and 27.8% Si. The flotation tailings solids were coarser than the concentrates, at 40% passing 27 m. The neutralizing capacity of the flotation tailings was 0.10 g H2SO4 per gram of solids, or about one-tenth that of the limestone used. For use in the pilot plant, the flotation tailings were prepared as an homogeneous slurry of 65.8% solids.

PROCESS DEVELOPMENT STUDIES

The amenability of the Crrego do Stio concentrates to the pressure oxidation process was evaluated in batch tests, followed by a continuous pilot plant campaign to generate design data.

Concentrate Quench Water Lime; Cyanide

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Acidulation, CCD washing and solution neutralization circuits were operated in an integrated fashion with the pressure oxidation circuit in the pilot plant campaign. Extensive cyanide leach tests were also conducted on the pressure oxidation product solids and intermediate solids from the individual autoclave compartments during the campaign.

Batch Test Program

Acidulation

Four batch acidulation rate tests were conducted on the separate concentrates and ten additional tests were conducted on the 1:1 blend. The effects of acid addition, temperature and retention time on the rate and extent of carbonate decomposition were examined. Between 75 and 90% carbonate decomposition was achieved.

Pressure Oxidation and Cyanide Leaching

Four batch POX rate tests were conducted at 220C with the separate concentrate samples. Several rate samples were collected during each POX test so that gold extraction vs. sulphide oxidation curves could be produced. The primary goal of the tests was to evaluate the effect on gold extraction of chloride removal from the acidulation solids by washing prior to pressure oxidation. It was found that washing the acidulation solids prior to POX was not necessary, since sulphide oxidation in all cases was rapid, exceeding 99%, and gold extraction in cyanide leaching typically exceeded 96% for the 40 min POX solids and 97% for the 60 min POX solids.

Sixteen additional batch POX rate tests were conducted with the 1:1 blend. Test variables were temperature, pressure and concentrate particle size. For temperatures between 210 and 230C sulphide sulphur contents of the POX solids were between 0.03 and 0.04 wt% after 40 min retention time, and gold extractions in cyanide leaching were between 95.7% and 98.5%. The extent of gold extraction increased with increasing temperature and retention time. Regrinding the feed had no measurable effect on the ultimate extent of sulphide oxidation or gold extraction. POX conditions of 220C and 60 min retention time were selected as the base case for the preferred conditions in the pilot plant test, although it was proposed to also evaluate 230C, and 220C at 90 min retention time.

At 190C (So Bento conditions), sulphide oxidation in the batch tests was slower and gold extraction in the range of 94 to 95% required a POX retention time of up to 180 min. Similar results were obtained with solids recycle or addition of calcium lignosulphonate. (The purpose of either solids recycle or calcium lignosulphonate addition is to disperse elemental sulphur, which is expected to be formed in a greater proportion at the lower POX temperature.) Regrinding the feed resulted in increased sulphide oxidation rate but gold extraction following POX was not increased. Based on these results, it was proposed to operate the So Bento case in the pilot plant with as-received concentrate, i.e., without additional grinding, with solids recycle, at 190C and with 180 min continuous retention time.

The POX discharge slurry was conditioned at 95C. The purpose of the conditioning step is to decompose basic iron and/or iron-arsenic sulphate precipitates, which would otherwise increase lime consumption in the cyanide leach step. Generally, the conditioning step decreased the iron, arsenic and sulphur contents of the POX solids, confirming the efficacy of this step for treatment of the Crrego do Stio concentrates, especially for POX at 230C.


Two solution neutralization batch rate tests were conducted with selected solutions that were produced in the acidulation and POX batch tests. The retention time was 240 min. Neutralization was conducted in two stages, with addition of either flotation tailings or limestone in the first stage, followed by addition of lime in the second stage.


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As expected based on the acid neutralization capacities of the two materials, limestone was much more effective than flotation tailings for neutralization in the first stage. The pH achievable with flotation tailings (pH 2.6) was considerably lower than that with limestone (pH 6.8). As a result, considerably less lime was required when used with limestone versus the flotation tailings. In both tests, aluminum, arsenic, iron, manganese and silicon were precipitated to below the detection limits (

Figure 2 Extent of carbonate decomposition during operation of acidulation pilot plant circuit; variations in operating conditions are indicated by vertical dashed lines, separating operating periods 1 to 5

The acidulated slurry settled very well in the continuous circuit. The underflow slurry solids content ranged from 45 to 55%, with an average of 49%.

Pressure Oxidation

The pressure oxidation circuit consisted of feed systems for acidulation thickener underflow slurry and quench water, the pressure oxidation autoclave, a flashing system, and conditioning tanks to dissolve basic iron sulphates in the POX solids. The autoclave was a horizontal, multi-compartment vessel, operated with a double-sized first compartment (C1/C2) and four additional compartments (C3, C4, C5 and C6).

The pressure oxidation circuit was operated for 120 hours with 99.8% on stream time. The circuit processed about 1 500 kg of thickened acidulated slurry containing about 750 kg of solids and about 1 100 L of quench water.

The target parameters for each period of autoclave operation are summarized in Table 3. The first four periods tested the high temperature, preferred conditions case. Periods 1, 2 and 3 were operated at 220C, with the retention time and chloride levels being the target process variations. The temperature was elevated to 230C in Period 4. The So Bento conditions case was tested in Period 5, at a significantly lower temperature, 190C, and pressure, 1 650 kPa (gauge); the retention time was 2 to 3 times longer than in the higher temperature periods, and oxidized solids were recycled, by way of the acidulation circuit, to prevent agglomeration of any elemental sulphur that could be formed in the autoclave at the lower temperature.

For each operating temperature the ratio of thickened acidulation product slurry and quench water was adjusted to achieve a target slurry solids content in the autoclave as predicted by preliminary mass and heat balances performed prior to the continuous campaign. Prior to the start of Period 5, the acidulation

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circuit was manipulated to ensure that the transition from fresh concentrate to a 1:1 blend of recycled POX solids and fresh concentrate was a step change as opposed to a gradual transition.

Table 3 POX pilot plant operating conditions

Operating Condition Period 1 Period 2 Period 3 Period 4 Period 5 Run Time, h 0 to 24 24 to 42 42 to 60 60 to 72 72 to 120 Temperature, C 220 220 220 230 190 Pressure, kPa(gauge) 2 570 2 570 2 570 3 050 1 650 Retention Time, min* 90 60 60 60 180 Chloride in Quench Water, mg/L 13 13 nil nil nil Solids Recycle Ratio 0 0 0 0 1:1

*Based on single pass solids retention time in the autoclave

Sulphide oxidation profiles across the POX autoclave were developed by collecting samples from each autoclave compartment during the campaign, as summarized in Table 4. The values provided in the table are for samples that were collected after the autoclave had reached steady state in each operating period, rather than average values for all samples collected during the campaign. The autoclave retention time corresponding to each autoclave compartment could be readily calculated based on the feed rates and the known internal geometry of the autoclave.

Table 4 Sulphide oxidation profile during POX pilot plant test

Period Temperature Sulphide Oxidation (steady state values), % C C1/C2 C3 C4 C5 C6 1 220 89.9 95.9 99.2 99.4 99.4 2 220 77.7 83.0 90.7 99.0 99.5 3 220 78.0 85.8 97.4 99.3 99.4 4 230 82.5 98.3 99.2 99.3 99.3 5 190 88.6 94.4 98.0 99.1 99.1

Sulphide oxidation exceeding 99% was achieved in all 5 periods of operation. As expected, the rate of sulphide oxidation increased with temperature. The retention times required to reach 99% sulphide oxidation were about 45 and 60 min for operation at 230 and 220C, respectively. When operating at 190C, 99% sulphide oxidation was not achieved until compartment C5, corresponding to about 160 min retention time (single pass basis) in Period 5.

Cyanide amenability (CNA) tests, which are cyanide leach tests that use small quantities of solids and do not involve intermediate sampling, were performed on samples of the conditioned solids every two hours and on a suite of selected samples from the pressure oxidation circuit every six hours. In addition, carbon in leach (CIL) cyanide leach tests, conducted at higher pulp density with larger quantities of solids and with intermediate sampling, were performed on a conditioned and washed sample of the POX solids every six hours. These higher pulp density tests are more useful for defining reagent requirements for the commercial operation. Nonetheless, as shown in Figure 3, there was very good agreement between the gold extraction values obtained in the CNA and CIL tests, indicating that the more numerous CNA results provide a reliable assessment of the trends in gold extraction under the range of POX conditions tested.

Table 5 summarizes CNA gold extraction results obtained during the pilot plant campaign based on cyanide leaching of samples taken after the autoclave had reached steady state in each operating period. Samples collected from each autoclave compartment and from the discharge of the conditioning circuit were tested.


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Figure 3 Gold extraction in cyanide leaching of pilot plant POX solids following conditioning

Table 5 Gold extraction from POX solids in CNA cyanide leach tests during pilot plant campaign

Period Temp. Gold Extraction in CNA Cyanide Leach Tests (steady state values), % C C1/C2 C3 C4 C5 C6 Conditioned

1 220 90 to 93 90 to 94 93 to 95 93 to 95 94 to 95 93 to 94.5 2 220 86 to 88 90 to 92 92 to 93 92 to 94 93 to 95 93.5 to 94.5 3 220 84 to 88 88 to 89 91 92 93 92 to 93 4 230 90 to 91 94 to 95 95 95.5 95.5 95.5 5 190 88 to 91 90 to 91 91 to 92 91 90 89.5 to 91.5

Final gold extraction achieved in cyanide leaching when operating the pressure oxidation circuit at temperatures of 230, 220 and 190C (i.e., for solids collected from compartment C6 or following conditioning) was about 95, 92.5 to 94.5, and 90%, respectively. Based on the shapes of the gold extraction profiles for cyanide leaching of solids collected from the individual autoclave compartments, there was no indication that increased gold extraction could be obtained by extending the autoclave retention time beyond 60 min at 220 or 230C, or beyond 180 min at 190C.

The effect of varying chloride concentration in the pressure oxidation solution was also investigated, since it was expected that preg-robbing by organic carbon under POX conditions may affect gold extraction. Two levels of chloride concentration were tested, approximately 30 and 15 mg/L in the POX solution before flashing, corresponding to addition of quench water containing about 13 mg/L chloride or essentially no chloride, respectively. There was very little difference in the results obtained over this small range in chloride concentration. Nevertheless, there were indications during the test program that small variations in oxygen distribution and oxidation reduction potential in the autoclave were influencing gold extraction. It appeared that the preg-robbing effect was diminished by operating at 230C compared with operating at 220C, since gold extraction in Period 4 (230C) was about 95%, compared with 92.5 to 94.5% in Periods 1 to 3 (220C).

Conditioning following pressure oxidation did not have a significant impact on gold extraction. However, significant redissolution of ferric iron, arsenic and sulphate sulphur in the solids did occur in the conditioning step. The value of the conditioning step was especially evident when operating the autoclave

88

90

92

94

96

0 24 48 72 96 120Run Time, h

Gol

d Ex

tract

ion

, %

CNA

CIL

Period 1 Period 2 Period 3 Period 4 Period 5


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at 230C, as the sulphate sulphur assay of the solids changed from about 3.3% in the final autoclave compartment to about 2.4% following conditioning. This difference would lead to a decrease in lime consumption in cyanide leaching of about 15 kg/t compared with not conditioning the solids.

Countercurrent Decantation Wash

The conditioning tank discharge slurry from the pressure oxidation circuit was washed with solution neutralization thickener overflow solution in a three-stage countercurrent decantation (CCD) wash circuit to remove the bulk of the acid from the POX solids prior to cyanide leaching for gold extraction.

Overflow solution from the first wash thickener was forwarded to the acidulation circuit to serve as lixiviant for decomposition of carbonate in the concentrate feed. The underflow slurry from the third wash thickener was collected for further testing. Every six hours, samples of the third wash thickener underflow slurry were taken, further washed and subjected to cyanide leaching tests to determine gold extractions.

The countercurrent decantation (CCD) wash circuit was operated from 10.5 to 120 h run time, with an on-stream time of 99.3%. The acid concentration in the wash circuit solutions depended on the feed solution acid concentration and the wash ratio. The acid concentration across the CCD circuit decreased from about 30 g/L in the circuit feed to between 1.5 and 3 g/L in the entrained liquor with the third wash thickener (W3) underflow slurry. The product solution, W1 overflow, typically contained about 15 g/L H2SO4.

The POX solids settled well in the CCD wash circuit, to about 40 to 45% solids, and allowed from 91 to 96% of the sulphuric acid and metal salts to be washed from the conditioning tank discharge slurry during the pilot plant operation. A wash ratio of about 2.5 L of wash solution per L of W3 underflow entrained solution gave the desired extent of washing in the circuit.


In solution neutralization, acidulation product solution was reacted with limestone (or flotation tailings) and slaked lime in two stages to neutralize the free acid and precipitate metals. The solution neutralization circuit consisted of a solution feed tank, slurry feed tanks for limestone or flotation tailings and lime, five stirred tank reactors in series, and a thickener.

Overflow solution from the solution neutralization thickener was forwarded to the CCD wash circuit to serve as wash solution. The underflow from the solution neutralization thickener was collected for further testing.

In the first stage of neutralization, limestone or flotation tailings solids were added as a slurry to neutralize the sulphuric acid and to hydrolyze and precipitate metals at a target pH of 5. Limestone consumption was about 25 to 30 g per L of feed solution, or about 0.7 g limestone per g of equivalent acid. Limestone utilization was calculated at about 93 to 94%, based on assays of carbon in the solids and flow rates. When flotation tailings were added in place of limestone, the addition rate was about 7.5 g per g of equivalent acid in the feed solution, and a pH of 3.3 was achieved. Carbonate utilization from flotation tailings was calculated at about 44% based on assays of carbon in the solids and flow rates.

In the second stage of neutralization, lime was added to precipitate metals to low concentrations at a target pH of 9. Lime consumption was about 15 to 18 g CaO per L of feed solution when limestone was added in the first stage of neutralization, but was approximately doubled to 32 g CaO per L of feed solution when flotation tailings was added to the first stage.

Antimony and arsenic were both precipitated to near completion in the first stage of neutralization with either limestone or flotation tailings addition to a pH range of about 3 to 5. Significant amounts of


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aluminum, iron and silicon were also precipitated in the first stage with limestone. Nearly complete precipitation of aluminum, iron, manganese and silicon occurred in the second stage of neutralization with lime addition to the range of about pH 8.5 to pH 9. Some magnesium remained in the calcium-saturated solution after lime addition to about pH 9, and ranged from about 0.2 to 1.3 g/L Mg.

The solution neutralization slurry settled very well. The underflow slurry solids content typically ranged from 54 to 58% when limestone was added to the first stage of neutralization. The underflow pulp density increased from about 1 350 g/L (or about 45% solids) without recycle of product slurry to the first stage of neutralization, to about 1 500 g/L (or about 58% solids) after one day of slurry recycle. The underflow pulp density and solids content remained fairly similar when flotation tailings were added in place of limestone; however, the quantity of solids being handled was significantly greater when flotation tailings were used.

ENGINEERING DESIGN

A trade-off study completed by Sherritt provided preliminary economic information for the implementation of pressure oxidation technology under various conditions. Further, basic engineering activities were completed for two cases: a) installation of a new autoclave at high temperature (Period 4 in Table 3); and b) refurbishment of one existing So Bento autoclave at original conditions (Period 5 in Table 3). The cost and performance data from these studies supported AGA Brasils steering committee in their ultimate decision to install a new autoclave.

A review of the So Bento equipment showed that much of the available equipment outside the autoclave area was in good mechanical condition and could be used, with some refurbishment, to process the Crrego do Stio concentrates and oxidized slurry. Within the autoclave area, one of the autoclaves could be used to treat concentrates under the conditions tested during Period 5 of the continuous pilot plant operation. Some rearrangement of the existing equipment, piping rework and instrumentation upgrades were deemed to be required to put the plant in working order for this case. However, the higher gold recovery demonstrated during pilot plant Period 4 offered sufficient economic benefit to justify the additional cost of new equipment. The additional time to design and procure a new autoclave had minimal impact on the projected startup date, as both cases required new equipment in milling and flotation as well as refurbishment of existing equipment in the remainder of the plant.

The process design criteria for the Crrego do Stio Project were developed based upon results of the laboratory and pilot plant test work, information supplied by AGA Brasil, and Sherritts experience and expertise in similar projects. The pressure oxidation plant was designed for a nominal treatment rate of 60 000 t/a refractory gold concentrate using an operating basis of 24 hours per day, with an annual 85% on stream time.

The autoclave shell design that was developed during basic engineering was utilized without modification during the initial stages of detailed engineering, and underwent only a small number of changes recommended by the fabricator. This allowed the order to be placed for the autoclave shell within days of receiving approval to proceed, in order to ensure that autoclave delivery would not delay plant construction.

The refurbishment of feed, discharge and vent handling equipment in the So Bento autoclave circuit was considered during detailed engineering with the objective of reducing cost and delivery time. However, in most instances new equipment was specified in order to optimize the equipment layout in the available space, and to maximize environmental performance. Costs were reduced by reusing some of the original agitator components and refurbished electrical equipment.


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CONCLUSIONS

The process development studies conducted for the Crrego do Stio Project demonstrated significantly improved gold extraction from Crrego do Stio concentrates when processed at 220 to 230C prior to cyanide leaching compared with pre-treatment at 190C. Overall, the value of continuous piloting for process demonstration and sound engineering design was confirmed during the project. The following key parameters were defined, which otherwise were not clearly identified based on batch test results:

Improved gold extraction with POX treatment at 230C compared to 220C; Lower gold extraction with pre-treatment at 190C, even with solids recycle and very long solids

retention times in the autoclave; Liquid solids separation testing of pilot plant process slurries confirmed the suitability of existing

thickeners; and Continuous neutralization confirmed the suitability of the existing tanks.

As a result, following an economic evaluation, AGA Brasil made the decision to build a new pressure oxidation circuit. The results of the process development work were successfully used by Sherritt to produce a commercial design for the new pressure oxidation circuit, which was commissioned in January 2012 and is currently in operation.

ACKNOWLEDGEMENTS

The authors wish to thank the management of AGA Brasil and of Sherritt International Corporation for permission to publish these results. The significant contributions of the professional and technical staff of Sherritt Technologies are also gratefully acknowledged.

REFERENCES

Berezowsky, R.M.G.S., Collins, M.J., Kerfoot, D.G.E., & Torres, N. (1991). The commercial status of pressure leaching technology. JOM, Vol. 43, No.2, February 1991, 9-15.

Collins, M.J., Hasenbank, A., Parekh, B. & Hewitt, B. (2011). Design of the new Lihir Gold pressure oxidation autoclave. In B.R. Davis and J.P.T. Kapusta (Eds.), New Technology Implementation in Metallurgical Processes (pp. 101-110), MetSoc. of CIM, Montreal, Canada.

Collins, M.J., Yuan, D., Masters, I.M., Kalanchey, R. & Yan, L. (2011). Pilot plant pressure oxidation of refractory gold-silver concentrate. In G. Deschenes, R. Dimitrakopoulos and J. Bouchard (Eds.), World Gold 2011 (pp. 145-155), MetSoc. of CIM, Montreal, Canada.

Silva, L., Guimaraes, R. & Milbourne, J. (2004). Process modifications to the Sao Bento concentrator of Eldorado Gold. In M.J. Collins and V.G. Papangelakis (Eds.), Pressure Hydrometallurgy 2004 (pp. 781-791). CIM, Montreal, Canada.


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THE MOBILIZATION AND POTENTIAL IN-AUTOCLAVE RECOVERY OF GOLD DURING PRESSURE OXIDATION AND LEACHING OF AURIFEROUS SULFIDE MATERIALS

*G.P. Demopoulos, J.-C. Parisien-La Salle, D. Blais

Department of Mining and Materials Engineering McGill University

Montreal, Quebec, Canada H3A 2B2 (*Corresponding author: [email protected])

ABSTRACT

In this work the dissolution, precipitation and adsorption behaviour of gold in acidic iron sulphate media containing small amounts of chloride ions is investigated via batch laboratory tests and OLI thermodynamic calculations over the temperature range 25-250C. Chloride ions may be present in industrial pressure oxidation/pressure leaching (POX/PL) operations due to elevated water salinity levels or due to intentional addition as is the case of the CESL Copper Pressure Leaching Process. Thermodynamic calculations show that AuCl4- may form depending on the free Cl-/Au ratio and temperature, its stability increasing with increasing Cl- concentration and decreasing with increasing temperature. On the other hand during POX/PL, we know that massive precipitation of iron in various hydrolysis products and compounds occurs. Hence of interest is the question how AuCl4- that may form during leaching interacts with the iron precipitation products. Gold(III) chloride was found to co-precipitate with iron(III) (oxy) hydroxides and arsenates forming in the autoclave but not with natrojarosite. Upon further investigation it was discovered that activated carbon added in the autoclave (as done in CIL) acts as preferential sink for the mobilized gold, hence opening the potential for the development of a direct cyanide-free gold recovery process in acidic sulphate leach reactors.


Proceedings of the 42nd Annual Hydrometallurgy Meeting held in conjunction with the 51st Annual Conference of Metallurgists of CIM (COM 2012)

Niagara, ON, Canada Edited by

M.J. Collins, D. Filippou, J.R. Harlamovs, E. Peek

15

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INTRODUCTION

The recovery of gold from ores and concentrates has evolved over the years with processing routes becoming increasingly complex. Given the high value of gold, any reduction in losses during recovery operations constitutes an important economic benefit. One such loss that needs to be addressed is the potential solubilisation of gold during pressure oxidation (POX) of refractory sulphidic gold feedstocks. For example, Agnico Eagle Mines reported a loss in gold recovery due to the formation of a gold chloride complex (Garofalo, 2009). Gold is known to form complexes in solution, such as gold(III) chloride, if an excess of chloride ions are present. The fate of such soluble gold chloride during pressure oxidation and the subsequent neutralization of the acidic discharge slurry will determine if gold is lost or recovered in the downstream cyanidation operation. Thus it is of interest to know under what conditions gold chloride may stay soluble or co-precipitate along primary (formed in the autoclave) or secondary (formed during neutralization) precipitates. Furthermore, the mode of co-precipitated gold, such as adsorption on iron compounds, like goethite and hematite (Machesky et al., 1990; Nachayev, 1984); substitution in one of the precipitated phases; or finally chemical reduction to metallic state would have consequences in terms of gold recovery.

Another process where gold is recovered from the residue generated in an autoclave is copper pressure leaching. There are two general approaches in pressure leaching of copper: the high-temperature, or total sulphur oxidation and the medium-temperature approach, exemplified by the CESL process (Defreyne et al., 2006). Drawbacks of this technology are: in the former case excessive cost due to oxygen consumption and neutralization of the generated acid; and in the latter case low gold recoveries due to the formation of elemental sulphur which interferes with cyanidation. Cyanidation remains by far the key process used worldwide on free-milling ores and on gold recovery systems from autoclave discharge, but increasingly serious concerns are expressed about the toxicity of cyanide and the risks associated with its handling and transportation.

For these reasons the behaviour of gold(III) chloride in acidic iron(III) sulphate solutions during hydrothermal or neutralization treatment was investigated. In particular two types of experiments were carried out. The first involved atmospheric pressure precipitation of iron(III) (with and without arsenate) by neutralization to pH 4 in the presence of gold chloride (100 mg/L Au as HAuCl4). This investigation involved examination of solution composition (iron(III) concentration, Fe/As molar ratio, Cl/Au molar ratio) and temperature (20-90C). The second series involved autoclave experiments under various conditions favouring the formation of crystalline iron(III) and arsenate phases and monitoring the behaviour of gold chloride over the temperature range of 200-250C. The behaviour of gold(III) chloride in terms of stability (resistance to reduction) in solution is analyzed with the aid of the OLI thermodynamic package to set the scene for the experimental investigations undertaken. The adsorption of gold chloride on iron oxides is also covered.

After examining the behaviour of gold chloride during autoclave iron compound precipitation, it was decided to investigate a new cyanide-free and direct (one-step) concept for the recovery of gold from refractory gold pressure oxidation and copper pressure leaching reactors-autoclaves. The new process concept involves the extraction of gold during the actual pressure oxidation/leaching operation in the autoclave via the use of a small amount of chloride ion addition (if not already present) plus activated carbon. For maximizing gold recovery the process is designed in such way that favours the precipitation of iron as sodium jarosite since the latter does not interfere with gold recovery in contrast to hematite (or ferrihydrite) that does. Under these conditions gold (in the form initially of aurochloride complex) adsorbs on carbon from which it can be recovered following screening, elution and electrowinning or reductive precipitation/cementation.

It is the scope of this paper to review the dissolution/adsorption/co-precipitation of gold as chlorocomplex during iron precipitation occurring in POX autoclaves or in neutralization reactors following PX and to demonstrate the potential use of activated carbon directly in hot leaching reactors (atmospheric to pressure) to recover in-situ the mobilized gold.


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PROCESS AND CHEMISTRY REVIEW

Pressure Oxidation

The pressure oxidation treatment of refractory gold-bearing minerals involves a number of processing steps as seen in Figure 1 (left). Slurried crushed/ground ore, or concentrate, is fed into a multi compartment autoclave where it is made to react with oxygen. Oxidized iron, arsenic and sulphur (as sulphate) are released temporarily into solution where most ions later form precipitates in-situ. Depending on the operating conditions a wide range of precipitates may form.

Figure 1 (left) General flowsheet for the recovery of refractory gold via pressure oxidation; (right) General CESL copper pressure leaching process flowsheet

Pressure oxidation may be conducted in either acidic or alkaline media. Acidic media remains the prominent route used in industry, whereas alkaline oxidation may be applied industrially in special cases. For proper oxidation the autoclave operating temperature should be above 180C to ensure complete oxidation of all sulphides present to the sulphate form of sulphur. Elemental sulphur forming at lower temperatures is undesirable as it interferes with the oxidation process and the downstream recovery of gold by cyanidation (Demopoulos & Papangelakis, 1989). After pressure oxidation the solids and liquids are separated and the solids are sent to cyanidation to recover the gold, which was liberated during pressure oxidation (POX). The liquid, which still contains dissolved iron (and arsenic), is neutralized to completely precipitate all the metals from solution. As in POX, stable gold complexes can react with compounds that form during neutralization.

Pressure Leaching of Copper

The high temperature total oxidation process described earlier for the processing of copper concentrates is essentially the same (from a gold recovery standpoint) with the POX process for refractory gold ores; hence it will not be discussed further. Instead the medium-temperature (150C) type of copper pressure leaching process and in particular the CESL (developed by Teck) process (Defreyne et al., 2006) is described. The key ingredient of the process is the addition of 6-15 g/L chloride as catalyst to accelerate the leaching kinetics of chalcopyritethe most common and difficult to leach copper sulphide mineral. It is interesting to observe here that despite the presence of this quantity of chloride ions in the CESL autoclave not dissolution or losses of gold have been reported. This may suggest that if any solubilisation


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happens, it must be short-lived with gold reporting ultimately to the residue from which it is recovered. The generalized CESL process flowsheet is shown in Figure 1 (right). The feed, copper flotation concentrate, is sent to the pressure leaching autoclave. The copper sulphides are fully oxidized and the copper is solubilised. Any pyrite present reacts only partially because it is a highly noble mineral that forms a galvanic couple with chalcopyrite. The majority of oxidized iron reports in the residue as hematite and jarosite. At the same time elemental sulphur is the dominant sulphide oxidation product. A counter current S/L separation/washing circuit (CCD) separates the copper-containing solution and the gold-bearing solids. The copper-containing solution is sent to a copper recovery process (typically consisting of SX and EW). The washed solids (which may have been previously separated from the elemental sulphur by flotation) are sent to another autoclave where they are subjected to pressure cyanidation (Barr et al., 2007) in order to recover gold by dissolution as soluble gold-cyanide complexes. The products from the second autoclave stage are sent to a carbon-in-pulp (CIP) circuit to recover the leached gold with activated carbon. The pulp is screened and the gold-loaded carbon is transferred to an elution circuit while the remaining pulp is sent to a cyanide destruction plant and subsequently disposed in a tailings pond. The activated carbon is stripped exactly as in the POX process and the gold cyanide solution (eluate) is electrowon to recover gold.

In-situ Iron Precipitation

The pressure oxidation chemistry and kinetics of pyrite and arsenopyrite are described elsewhere (Papangelakis & Demopoulos, 1990a, 1999b, 1991; Long & Dixon, 2003). Upon oxidation these minerals release iron (and arsenic), which precipitate in situ. Depending on temperature, oxygen pressure and solution composition parameters such as acidity, total sulphate content and presence of co-ions for example sodium, many products can precipitate, such as hematite, basic ferric sulphate (BFS) and several types of jarosites that may be stable or metastable (Cheng & Demopoulos, 2004). If arsenic is present, then additionally ferric arsenate compounds may form as recently studied by Gomez et al. (2011).

Gold Dissolution During Pressure Oxidation

During pressure oxidation in sulphate media, generally no or very little complexing ligands (like halide ions) are present, hence the common understanding is that refractory gold remains in the elemental state if already present as such or simply converts to that as soon as it is liberated. It is conceivable that temporarily solubilised gold is reduced in-situ on oxidized mineral product surfaces via short-term adsorption or not and reports with the residue solids. Gold is subsequently recovered from these residues by cyanidation under an oxidizing environment. However in the presence of significant amounts of complexing agents like chlorides, arising from either brackish process water or the ore itself can lead to greater gold solubilization and possible losses.

A classic example of gold solubilisation, which caused serious complications, was at the Lihir POX plant in Papua, New Guinea. In this case chlorides caused the solubilisation of gold within the autoclave and the subsequent plating of gold inside the discharge nozzle (Ketcham et al., 1993). More recently gold losses due to solubilisation of gold as a result of chloride complexation was reported by Agnico-Eagle (Garofalo, 2009).

Gold has two oxidation states: aurous (Au1+) and auric (Au3+). These are stabilized only upon complexation with strong ligands. Thus the solubilisation of gold in cyanide solutions occurs via the formation of the cyano-aurous complex Au(CN)2-, while in the presence of halide ions gold is solubilised in its auric form as Au(Cl)4- (Demopoulos et al., 1989; Liu & Nicol, 2002).

The importance of complexation in rendering gold soluble can be evaluated by referring to appropriate reduction potentials and Pourbaix diagrams. Equation 1 gives the standard reduction potential corresponding the Au3+/Au(s) couple. When this potential is compared to that of oxygen/water couple (Equation 2) it becomes clear that gold in its elemental state is the most stable form.


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Au3+ + 3e- Au(s) e = 1.498 V (1)

O2 + 2H+ + 2e- H2O e = 1.23 V (2)

With the presence of chloride ions as a complexing agent, the potential required for the oxidation of gold does change and allows for the lixiviation of the metal since its standard reduction potential now (0.99 V) is less than that of oxygen (1.23 V).

AuCl4- + 3e- Au(s) + 4Cl- e = 0.99 V (3)

It is interesting to report here that with increasing temperature the reduction potential of the Au(Cl)4-/Au couple decreases making the oxidation of gold possible even by the Fe3+/Fe2+ couple as it can be deduced from the data plotted in Figure 2 (Liu & Nicol, 2002).

Figure 2 Reduction potentials of the Au(Cl)4-/Au, O2/H2O and Fe3+/Fe2+couples as a function of temperature, [Fe3+] = [Fe2+] = 0.1M (Liu & Nicol, 2002)

The possible dissolution of gold during pressure oxidation as a result of the presence of chlorides and the relative stability of gold(III)-chlorocomplexes as a function of temperature and chloride ion concentration was studied recently in our laboratory using the OLI thermodynamic software package (StreamAnalyzer Version 3.0). Some of the calculations are presented in Figure 3; the calculations were made for 5 10-4 M of AuCl3 (corresponding to approximately 100 mg/L Au). According to this data, the concentration of stable soluble gold complex decreases upon a temperature rise, but increases upon raising the excess chloride ion concentration. For example we calculated that the solubility limit of the gold (III) chloride complex before it is reduced to metallic gold for the case of 0.0715M HCl (or 2.3 g/L Cl) solution is 5 10-4 (or 100 mg/L) Au at 227C. Although in industrial systems this may be much lower due to the presence of other reducing agents such as unreacted sulphides, nevertheless these numbers serve to substantiate the notion that solubilisation (even if partial or temporary, i.e. metastable) of gold can occur hence the importance of considering its deportment and possible in-situ recovery.


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Figure 3 The effect of temperature on gold (III) chloride complex concentration for various added amounts of excess chloride (as HCl)

Adsorption of Gold (III) Chlorocomplexes

During the in-situ precipitation of iron oxides (and iron arsenates) in the POX autoclave or neutralization reactors, soluble species (cations or anions) may co-precipitate via adsorption (Demopoulos, 2009). In this section, the possible co-precipitation of AuCl4- via adsorption on iron precipitates is considered as well as reference is made to previous studies involving carbon adsorption of gold(III) chlorocomplexes.

Adsorption of Gold(III) Chloride on Iron Precipitates

Complexes of gold including gold(III) chloride may adsorb on iron (hydroxy)oxides via chemisorption. The formation of surface complexes on an oxide involving anions (as, for example, arsenate species (Demopoulos, 2009; Yongfeng & Demopoulos, 2005) and AuCl4-) is described in Equation 4:

/XOH + A3- + H+ OA2- + H2O (4)

Reports of gold chloride adsorption on iron oxide surfaces do exist. Thus according to Karasyova et al. (1998) gold(III) chloride adsorbs on hematite via inner sphere coordination. But in addition to complexes, colloidal gold has been reported to adsorb on iron oxide particles due to electrostatic attraction. This phenomenon is known as mutual coagulation or heterocoagulation (Enzweiler & Joekes, 1991). This type of electrostatic interaction may lead to re-encapsulation hence constituting a possible cause of gold losses during pressure oxidation processes. This solubilisation-encapsulation phenomenon apparently is observed in natural mineralization processes theorized to be responsible in the genesis of refractory gold ores (gold encapsulated within iron sulphides upon cooling) (Ranet al., 2001). In this case the adsorption mechanism of gold(III) chloride on the surface of iron sulphides (see pyrite) is of electrostatic (physisorption) type followed by reduction on the mineral surface as represented by Equation 5 (Mycroft et al., 1995).

FeS2 + 5AuCl4- + 8H2O Fe3+ + 2SO42- + 16H+ + 5Au(s) + 20Cl- (5)

By extrapolating this geochemical information to the case of pressure oxidation of refractory gold ores, it may be deduced that partially oxidized sulphides (e.g. pyrite or arsenopyrite) plus in-situ forming hematite or other oxidized products can cause equivalent to preg-robbing effect if chloride ions are present.


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Adsorption of Gold(III) Chloride onto Activated Carbon

Conventionally in gold mining, the gold(I) cyanide complex Au(CN)2- adsorbs onto activated carbon via Van der Walls forces, i.e., via physisorption (Adams, 1995; Lagerge et al., 1999). This adsorption is reversible and at high temperature, the gold complex can be stripped off the activated carbon. However, in the case of gold(III) chloride adsorption onto activated carbon (Sun & Yen, 1993; Huges & Linge, 1989), chemical reduction occurs on the surface of the activated carbon resulting in metallic gold formation and apparent production of carbon dioxide as described by Equation 6:

4AuCl4- + 3C + 6H2O 5Au(on carbon) + 12H+ + 16Cl- + 3CO2 E = 0.796 V (6)

It has been further established that the adsorption/chemical reduction of gold chloride onto activated carbon is diffusion-controlled (Sun & Yen, 1993; Huges & Linge, 1989).

RESULTS

Four types of experiments were conducted: (1) neutralization tests involving iron sulphate solutions containing a small amount of soluble gold chloride to monitor its co-precipiation behaviour; (2) autoclave precipitation experiments involving similar solutions; (3) carbon adsorption tests involving dilute gold chloride solutions over the temperature range 20 to 95C; and (4) pressure leaching/oxidation tests involving pyrite and cemented gold as model refractory gold material in the presence of activated carbon. A sample of results from these tests is presented here.

Gold (III) Chloride Co-precipitation During Neutralization

During the neutralization of autoclave discharge solutions, base is added in order to precipitate the residual iron left in solution. During this process any soluble gold could react with the forming solids. In order to study the behaviour of gold(III) chloride during precipitation of iron, various neutralization tests were performed. Initially, acidic solutions with various ferric (added as a sulphate salt) concentrations were treated. In all these solutions approximately 5 10-4 M (100 mg/L) Au (as HAuCl4) plus 0.05MHCl was added to provide for a Cl/Au molar ratio equal to 100. The tests involved a gradual increase of pH with the addition of MgO (3M) until the target precipitation pH (pH 4). The behaviour of gold during a typical test with an initial iron concentration of 0.3M is shown in Figure 4 (left). It can be seen that the gold concentration variation parallels that of iron. At low pH levels, before iron begins hydrolyzing (< 30 minutes) for all practical purposes gold can be said to remain in solution. That no gold precipitates when iron stays in solution was further verified by agitating a solution of pH 1.2 with 0.15M dissolved iron and the same Cl/Au content, for one hour. No gold concentration drop was observed.

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However, upon reaching pH 4 (at ~35 min), complete iron precipitation (apparently as ferrihydrite) occurred, accompanied by partial co-precipitation of gold(III) chloride. The gold uptake occurs apparently over a short period of time and then plateaus at a gold concentration of 0.17 mmol/L. The mechanism by which gold co-precipitates could be attributed to either reduction of gold(III) into metallic gold, or adsorption of gold(III) chloride onto the surface of ferrihydrite or the incorporation of gold in the structure of ferrihydrite. As there are no other ions present in solution to act as reducing agents for gold(III) chloride the more likely mechanism for the loss of gold is via adsorption. The adsorption of gold(III) chloride to goethite via inner-sphere bidentate surface complex formation has been reported in literature (Machesky et al., 1990). The present results, upon further analysis were found to involve indeed a similar mechanism. Thus for example we found a linear correlation between the relative surface area of the ferrihydrite precipitate and the amount of gold(III) chloride removed from solution.

The effect of temperature on gold co-precipitation at fixed iron concentration is reported in Figure 4 (right). The plotted data reveals that the loss of gold increases with increasing temperature. Not only is the amount of gold removed affected (as deduced from the final gold concentration) but also the kinetics of gold co-precipitation.

In order to determine the mechanism by which gold co-precipitates, diagnostic leaching on the produced precipitates was performed. By independently digesting solids obtained from iron and gold co-precipitation experiments in a non oxidizing acid (3M HCl) and in aqua regia (AR) a distinction between metallic gold and adsorbed gold was made. The diagnostic leaching results (in the form of dissolved gold concentration) are reported in Table 1. As it can be seen there is no difference in the gold concentration in HCl and AR. Therefore, as was hypothesized in the previous section, no reduction to metallic gold occurs during gold(III) chloride co-precipitation with ferrihydrite.

Table 1 Diagnostic leaching results obtained from iron and gold co-precipitation solids

HCl AR Temperature (mmol/L) (mmol/L)

22C 0.067 0.066 60C 0.117 0.121 90C 0.129 0.127

Gold(III) Chloride Behaviour During Autoclave Precipitation

The possible preg-robbing effect of precipitated iron phases on gold(III) chloride complexes under pressure oxidation conditions was investigated for the cases of hematite and jarosite precipitation. Thus in this study hematite precipitation from sulphate solutions was carried out at 200C in the presence of 100 mg/L Au (5 10-4 M HAuCl4) plus 0.05 M HCl (Cl/Au molar ratio = 100).

As it can be seen in Figure 5 (left) all gold was found to coprecipitate with hematite. After characterization of the gold/hematite precipitate, using the diagnostic leaching technique mentioned earlier, it was concluded that at least 50% of gold was present in the precipitate as adsorbed gold(III) chloride species and the rest in metallic state. The adsorption of gold(III) chloro-complexes onto the surface of hematite is suggested to be described by the following reaction (where =FeOH symbolizes the surface of iron (hydroxyl)oxide:

=FeOH + AuCl4- FeOHAuCl3 + Cl- (7)

It is interesting to note that similar results to hematite were also observed in the case of the three iron-arsenate phases: scorodite, FAsH and BFAS (Gomez et al., 2011) although in this case substitution rather than adsorption is suspected.


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Figure 5 Gold concentration change during hydrolytic precipitation of hematite (left) and jarosite (right) at 200C

In contrast to hematite, no gold was found to co-precipitate in the tests favouring the formation of natrojarosite, again at 200C. This is clearly seen with the data (triplicate tests) plotted in Figure 5(right). In other words in this case, gold(III) chloride was neither adsorbed nor apparently chemically reduced. This finding was utilized in evaluating the in-situ recovery of solubilised gold(In the presence of small amount of chloride ions) by activated carbon as described later.

However, the mechanism responsible for the reduction of part of the gold chloride to metallic gold during autoclave precipitation of hematite is still elusive. As per OLI theoretical calculations, the stability (resistance to reduction) of gold tetrachloride ions decreases with increasing temperature on one hand and increases with increasing excess chloride ion concentration on the other. Thus as per theoretical calculations made, a 100 mg/L Au (as AuCl3) concentration is supposed to remain stable in solution at 200C, when the excess chloride ion concentration is 100 times higher as was the case with the tests of Figure 5. The fact that gold was nevertheless reduced in the actual experiment may reflect error with the thermodynamic data on which the theoretical OLI calculations are based on. This is not unlikely given the inevitable uncertainty with high temperature data extrapolation. Another possible cause for this behaviour may have been the presence of a reducing agent, such as ferrous ions. Given the small concentration of gold (5 10-4 M) it would take three times as much iron, see Equation 8, to be present as ferrous, which corresponds to approximately 0.4-0.5% of the 0.15M ferric sulphate used, not an unrealistic scenario.

AuCl4- + 3Fe2+ Au + 3Fe3+ + 4Cl- (8)

Carbon Adsorption of Gold(III) Chloride

Atmospheric adsorption experiments were performed with a solution containing 1.5 10-4 M Au(III) and 0.05 M chloride concentration at pH of 1.5. Tests were done at three different carbon loadings (10, 20 and 30 g/L) and at three different temperatures (20, 60 and 95C).

The variation of soluble gold(III) concentration with time for the room temperature test is shown in Figure 6 (left). As it can be seen complete gold removal was achieved within a few minutes of contact. The kinetics was found to increases with increasing carbon loading as well as temperature but somewhat to slow down with increasing chloride background concentration. The beneficial effect of temperature on gold chloride adsorption (in the form of first-order plots) is shown in Figure 6(right). From these plots, an activation energy of 8.69 kJ/mol was determined consistent with previous assertions of a diffusion-controlled process (Huges & Linge, 1989).


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Figure 6 (left) Adsorption kinetics of gold(III) chloride on activated carbon at 20C ; (right) First-order adsorption kinetic plots (10 g/L carbon) at 20, 60 and 95C

Upon detailed XPS characterization of gold-loaded carbon surfaces it was determined the mechanism of adsorption to be complex involving a redox reaction. Thus it was found for example at room temperature 45-50% of gold loaded on carbon was in the form of AuCl and the remaining as metallic gold. The amount of metallic gold increased with increasing temperature (from 55% at RT to 70% at 200C). At the same time the amount of Au(I)Cl increased with increasing chloride concentration reflecting the apparent increase in complex stability.

Carbon-in-autoclave

Gold Recovery

In this part of the investigation, tests were performed involving carbon addition in autoclave in order to evaluate the extent of in-situ gold recovery in a simple pressure oxidation-leaching system. In a typical carbon-in-autoclave test a simulated refractory feed material was used consisting of ground pyrite as model sulphide mineral along with metallic gold deposited in the starting slurry by cementation. The starting solution contained as well different amounts of free chloride and sodium ions. The presence of adequate chloride and sodium ion concentrations was required to ensure the solubilisation/adsorption of gold and suppression of hematite formation in view of the results reported in Figure 6. The presence of sodium favours the formation of jarosite instead of hematite thus avoiding the losses of gold due to adsorption on the latter.

Some representative results are summarized in Table 2. In this series of tests, a 300 mL capacity titanium Parr autoclave, with a Pyrex liner, was used. The Pyrex liner was required to avoid the loss of gold by plating/cementation on the metallic titanium surfaces. To minimize further gold losses (due to cementationmore on this issue later) all internal metallic parts, including the stirrer, were removed. Tests were conducted at two temperatures corresponding to the medium (CESL) and high temperature (POX) processing options. Typical conditions were: 53.3 g/L pyrite (80% -38 m), 30 mg/L metallic gold, 40 g/L carbon, initial pH 1.5, and P = 10 atm. In each test 75 mL of solution was used in which variable amounts of Na2SO4 and NaCl were added. Once the target temperature was reached, oxygen was supplied to the autoclave and oxidation was allowed to take place for 2 hours, after which the solution was quickly cooled and immediately screened with a sieve of 80 m size. The solution was analyzed for gold; the final residue (oxidation product) was digested in aqua regia, and then analyzed to determine its gold content (loss). The activated carbon was analyzed via neutron activation, which yields results in the ppb scale.

Upon review of the results summarized on Table 2, it can be seen that there was partial reporting of gold on activated carbon varying from 37.8% (Test 1) to 75.6% (Test 4). Test 1, performed at 155 oC, was associated with about 20% of the gold remaining with the residue. Of the remaining gold, 3% was in solution and only 38% was recovered (captured) by the activated carbon. Some 40% was unaccounted for. This unaccounted type of gold was thought to have been partially lost during sample handling and


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analysis but the majority to have plated on internal autoclave surfaces-a hypothesis verified later in the next series of tests. Interestingly the solution (approximately 40% of the total) that collected outside the liner due to evaporation and subsequent condensation upon cooling carried no soluble gold at low chloride concentration. It can be seen conditions that favour jarosite precipitation (confirmed by XRD) like sodium concentration and seeding resulted in higher gold recoveries. Comparison of Test 3 to Test 1 reveals that gold recovery on carbon increased with temperature. This indicates an endothermic mechanism behind gold adsorption on carbon, i.e. a similar behaviour with that observed over the temperature range 22-95C mentioned earlier. This behaviour is the opposite of that exhibited in conventional gold adsorption practice that involves gold-cyanide complexes, where elevated temperatures are known to promote de-sorption. This observation proves that the mechanism is not based on physisorption but rather on a endothermic chemical reaction that as shown by XPS characterization leads to reduction of gold(III) into gold(I) and metallic gold. Interestingly enough gold loss in the residue was again limited to 3% but 22% of gold remained unaccounted meaning that it was deposited on internal autoclave surfaces. The results serve as proof-of-concept of the effectiveness of in-situ carbon adsorption during pressure leaching/oxidation or atmospheric leaching for that matter!

Table 2 Summary of results from selected pressure oxidation experiments (300 mL Ti autoclave)

TEST 1 TEST 2 TEST 3 TEST 4 TEST 5

CONDITIONS Sodium1 (g/L) 8.5 6.8 8.5 10.2 4.3 Chloride (g/L) 2.5 2.5 2.5 5 5 Temperature (C) 155 155 200 200 200 Jarosite seed (g/L) 0 6.7 0 0 0 RESULTS

Gold left in solution (%) 2.9 3.2 4.4 2.6 3.6 Gold lost in residue (%) 18.9 1.9 3.2 0 21.8 Gold recovered on carbon (%) 37.8% 64.6% 70.2% 75.6% 46.6 Unaccounted gold2 (%) 40.4 30.3 22.2 21.8 28

1This is total sodium including sodium sulphate and sodium chloride additions. 2Assumed to have plated on metallic surfaces inside the autoclave; this assumption is verified in the next series of tests.

Gold Plating on Metal Surfaces

As noted above a significant amount of gold (20-30%) was unaccounted, i.e., it was neither found to be on the carbon, in the solution or in the precipitated residue. This unaccounted gold was determined to have plated out of solution on internal autoclave surfaces via electrochemical coupling similar to cementation reactions. It would be interesting to know if such plating takes place in brick lined industrial autoclaves (highly unlikely). Nevertheless, in similar systems to the present one, like the CESL and PLATSOL processes (where chlorides are added in the aut

Hydrometallurgia a Presion

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