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The treatment of copper-gold ores by ammonium thiosulfate leaching Molleman, Ellen 1998

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T H E T R E A T M E N T OF C O P P E R - G O L D ORES B Y A M M O N I U M THIOSULFATE LEACHING by  ELLEN MOLLEMAN  M . S c , Delft University o f Technology, The Netherlands, 1995  A THESIS S U B M I T T E D IN P A R T I A L F U L F I L L M E N T OF  THE REQUIREMENTS F O R T H E D E G R E E OF  M A S T E R OF A P P L I E D SCIENCE in T H E F A C U L T Y OF G R A D U A T E STUDIES  Department o f Metals and Materials Engineering  We accept this thesis as conforming to the lequired standard  T H E U N I V E R S I T Y OF BRITISH C O L U M B I A June 1998 © Ellen Molleman, 1998  In presenting  this  degree at the  thesis  in  partial fulfilment  of  University of  British Columbia,  I agree  freely available for reference copying  of  department  this or  and study.  thesis for scholarly by  publication of this  his  or  her  the  requirements that the  I further agree  purposes  representatives.  may be It  thesis for financial gain shall not  that  allowed  permission.  Department  of f W f l k  QrvJ  MinWv^K £ l g / A g p ^  The University of British Columbia Vancouver, Canada  Date  DE-6 (2/88)  advanced  Library shall make it  by the  understood be  an  permission for extensive  granted  is  for  that without  head  of  my  copying  or  my written  ABSTRACT The application o f ammonium thiosulfate for the treatment o f copper-gold ores has been investigated. Leaching studies were conducted with copper minerals, copper minerals with gold addition to solution and copper-gold samples o f different copper and gold grades. The behaviour of thiosulfate, tetrathionate and sulfate in solution was studied using ion chromatography.  The copper sulfide minerals chalcopyrite and enargite  seem to be unreactive toward an  ammonium thiosulfate leach. Covellite and chalcocite leach to a slight extent in this leaching system. The copper extractions o f the sulfide minerals seem to be independent o f the availability o f complexing agents. The copper oxide minerals cuprite and malachite showed high copper extractions in the presence o f sufficient lixiviant.  Experiments showed that both gold extraction and thiosulfate stability are influenced by a combination o f aeration and cupric ions in solution. H i g h initial gold extractions were achieved in an aerated solution in the presence o f cupric ions. However, these conditions simultaneously catalyzed thiosulfate degradation, resulting in gold precipitation. Therefore, it is important to establish a balance between providing sufficient air and cupric ions for fast gold dissolution, and to minimize the amount o f air in the presence o f cupric ions to prevent excessive thiosulfate degradation. A promising potential alternative to these conditions is a 24 hour leach without forced aeration.  ii  TABLE OF CONTENTS ABSTRACT  ii  T A B L E OF CONTENTS  iii  LIST OF TABLES  vi  LIST OF FIGURES  vii  LIST OF SYMBOLS  xi  ACKNOWLEDGMENTS  xii  1 INTRODUCTION  1  2 LITERATURE REVIEW  5  2.1 Aqueous Chemistry  5  2.1.1 Gold-water system  5  2.1.2 Copper-water system  8  2.1.3 Thiosulfate  10  2.2 Overview of Thiosulfate Technology for Gold Recovery  13  2.2.1 Thiosulfate leaching for gold extraction  13  2.2.2 G o l d recovery from thiosulfate solutions  19  2.3 Chemistry of the Ammonium Thiosulfate System  21  2.3.1 Thermodynamic considerations  21  2.3.2 Copper catalysis  24  2.3.3 Thiosulfate degradation reactions...  27  2.4 Ammonia technology for copper/gold leaching  33  2.4.1 A m m o n i a technology for gold recovery  33  2.4.2 A m m o n i a technology for copper recovery  35  2.5 Summary of literature  37  2.5.1 Thiosulfate technology  37  2.5.2 A m m o n i a Technology  40  2.5.3 Treatment o f copper-gold ores by ammonium thiosulfate  3 EXPERIMENTAL METHODS  ;  40  42  3.1 Materials  42  3.1.1 Copper Minerals  42 iii  3.1.2 Copper-Gold Samples  44  3.1.3 Reagents  46  3.2 Experimental Apparatus  47  3.3 Experimental Procedures  49  3.3.1 Preliminary testwork  49  3.3.2 Leaching of copper minerals  50  3.3.3 Leaching o f copper minerals with 5 ppm gold addition  51  3.3.4 Leaching o f copper-gold samples  51  3.4 Experimental Design  52  3.4.1 Preliminary experiments  52  3.4.2 Leaching o f copper minerals  52  3.4.3 Leaching o f copper minerals with 5 ppm gold addition  53  3.4.4 Leaching o f copper - gold samples  53  3.5 Sample Analysis  54  3.5.1 Copper and gold analysis  54  3.5.2 Thiosulfate, tetrathionate and sulfate analysis 4 RESULTS AND DISCUSSION  •  55 57  4.1 Eh-pH diagrams  57  4.2 Preliminary experiments  64  4.2.1 Summary  69  4.3 Leaching of copper minerals  70  4.3.1 Baseline experiments  70  4.3.2 Effect o f aeration  82  4.3.3 Summary  86  4.4 Leaching of copper minerals with gold addition  87  4.4.1 Baseline experiments  87  4.4.2 Effect o f aeration  96  4.4.3 Effect o f temperature  100  4.4.4 Effect o f reagent addition  103  4.4.5 G o l d precipitation  107  4.4.6 Summary  108 iv  4.5 Leaching of copper-gold samples  110  4.5.1 Overall Lobo Composite  ,111  4.5.2 Guanaco Composite  120  4.5.3 M 4 0 Pyrite Feed  127  4.5.4 M 1 0 Pyrite Concentrate  133  4.6 Sulfur balance  140  5 GENERAL SUMMARY AND CONCLUSIONS  146  6 RECOMMENDATIONS  151  7 REFERENCES  154  Appendix A: Screen analysis of copper-gold samples  160  Appendix B: Ion Chromatography  163  Appendix C: Thermodynamic data  173  Appendix D: Experimental data  174  Appendix E: Cyanidation procedures  182  Appendix F: Sulfate determination  186  v  LIST OF TABLES Table 2. A Standard reduction potentials for gold i n aqueous solution]  7  Table 2.B Stability constants for some gold complexes  8  Table 3 A Composition o f copper mineral samples for testwork  43  Table 3.B Composition copper-gold concentrate and ore samples  44  Table 3.C P  46  8 0  o f the copper-gold samples  Table 3.D Overview o f experiments with copper minerals  52  Table 3.E Overview o f experiments with copper minerals and gold addition  53  Table 3.F Overview o f experiments with copper-gold samples  54  Table 4. A Copper extractions (%) o f copper minerals at different times during baseline experiments  109  Table 4.B Results o f 24 hour cyanide leach o f the copper-gold samples  Ill  Table 4.C G o l d and copper extraction (%) o f Overall Lobo Composite at different times  119  Table 4.D G o l d and copper extraction (%) o f Guanaco Composite at different times  126  Table 4.E G o l d and copper extraction (%) o f M 4 0 Pyrite Feed at different times  132  Table 4.F G o l d and copper extraction (%) o f M 1 0 Pyrite Concentrate at different times  140  Table A . I Screen analysis o f copper-gold samples  160  vi  LIST OF FIGURES Figure 2.1 E h - p H diagram for the system A u - H 0 at 25°C  6  Figure 2.2 E h - p H diagram for the system C u - H 0 at 25°C  9  2  2  Figure 2.3 E h - p H diagram for the system S - H 0 at 25°C  10  2  Figure 2.4 E h - p H diagram for the metastable system S - H 0 at 25°C  11  Figure 2.5 Structure o f the thiosulfate ion  11  2  Figure 2.6 Effect o f copper concentration and temperature on the dissolution o f gold i n 0.25 M thiosulfate, 1.0 M N H , 196 k P a 0 , stirring velocity 200 rpm 3  2  14  Figure 2.7 Gold-ammonia-thiosulfate-water system at 25°C  22  Figure 2.8 Effect o f thiosulfate concentration on the rest potential o f gold  22  Figure 2.9 Copper-ammonia-thiosulfate-water system at 25°C  23  Figure 2.10 Effect o f copper sulfate concentration on gold leaching rate  25  Figure 2.11 Effect o f the concentration ratio o f ammonia to thiosulfate on gold leaching rate ....26 Figure 2.12 The model o f electrochemical-catalytical mechanism o f amrnoniacal thiosulfate leaching o f gold  27  Figure 2.13 Oxidation state diagrams for sulfur at several p H values  28  Figure 2.14 Effect o f pH/ammonia concentration on thiosulfate decomposition  32  Figure 2.15 E h - p H diagram for the A u - N H - H 0 system at 25°C  33  3  2  Figure 2.16 Leach extraction as a function o f time on three concentrates o f different mineralogy  36  Figure 3.1. Schematic diagram o f the experimental set-up  48  Figure 4.1. E h - p H diagram o f the gold-thiosulfate-ammonia-water system at 25°C  59  Figure 4.2 E h - p H diagram o f the gold-thiosulfate-ammonia-water system at 25°C  60  Figure 4.3 E h - p H diagram o f the copper-thiosulfate-ammonia-water system at 25°C  62  Figure 4.4 E h - p H diagram o f the copper-thiosulfate-ammonia-water system at 25°C  63  Figure 4.5 E h - p H diagram o f the copper-ammonia-water system at 25°C  64  Figure 4.6. Covellite, 5 g/L copper, 0.1 M S 0 , 0.47 M N H , p H 9.5, aeration, 35°C  67  Figure 4.7 Covellite, 5 g/L copper, 0.20 M ( N H ) S 0 , aeration, 35°C  71  Figure 4.8 Chalcocite, 5 g/L copper, 0.20 M ( N H ) S 0 , aeration, 35°C  72  Figure 4.9 Chalcopyrite, 5 g/L copper, 0.20 M ( N H ) S 0 , aeration, 35°C  73  2  3  3  4  2  4  2  2  2  4  vii  3  2  3  2  3  Figure 4.10 Enargite, 1 g/L copper, 0.20 M ( N H ) S 0 , aeration, 35°C  74  Figure 4.11 Cuprite, 5 g/L copper, 0.20 M ( N H ) S 0 , aeration, 35°C  75  Figure 4.12 Malachite, 5 g/L copper, 0.20 M ( N H ) S 0 , aeration, 35°C  76  Figure 4.13 Chalcocite, 5 g/L copper, 0.20 M ( N H ) S 0 , no aeration, 35°C  83  Figure 4.14 Cuprite, 5 g/L copper, 0.20 M ( N H ) S 0 , no aeration, 35°C  84  Figure 4.15 Covellite, 5 ppm gold, 1 g/L copper, 0.20 M ( N H ) S 0 , aeration, 35°C  88  Figure 4.16 Chalcocite, 5 ppm gold, 5 g/L copper, 0.20 M ( N H ) S 0 , aeration, 35°C  90  Figure 4.17 Chalcocite, 5 ppm gold, 1 g/L copper, 0.20 M ( N H ) S 0 , aeration, 35°C  92  Figure 4.18 Chalcopyrite, 5 ppm gold, 1 g/L copper, 0.20 M ( N H ) S 0 , aeration, 35°C  93  Figure 4.19 Enargite, 5 ppm gold, 1 g/L copper, 0.20 M ( N H ) S 0 , aeration, 35°C  94  Figure 4.20 Cuprite, 5 ppm gold, 5 g/L copper, 0.20 M ( N H ) S 0 , aeration, 35°C  95  Figure 4.21 Malachite, 5 ppm gold, 5 g/L copper, 0.20 M ( N H ) S 0 , aeration, 35°C  97  Figure 4.22 Chalcocite, 5 ppm gold, 5 g/L copper, 0.20 M ( N H ) S 0 , no aeration, 35°C  98  Figure 4.23 Cuprite, 5 ppm gold, 1 g/L copper, 0.20 M ( N H ) S 0 , no aeration, 35°C  99  Figure 4.24 Chalcocite, 5 ppm gold, 5 g/L copper, 0.20 M ( N H ) S 0 , aeration, 20°C  101  Figure 4.25 Cuprite, 5 ppm gold, 1 g/L copper, 0.20 M ( N H ) S 0 , aeration, 20°C  102  4  4  2  2  2  2  4  3  2  4  4  3  2  2  2  3  2  2  3  3  4  2  2  4  3  2  4  2  2  2  4  4  4  2  2  4  2  2  2  4  4  2  3  3  2  2  2  3  4  4  3  2  2  2  3  3  2  3  3  2  2  2  3  3  Figure 4.26 Chalcocite, 5 ppm gold, 5 g/L copper, 0.20 M ( N H ) S 0 , 0.05 M S 0 \ 2  4  2  2  3  3  aeration, 35°C  104  Figure 4.27 Chalcocite, 5 ppm gold, 5 g/L copper, 0.20 M ( N H ) S 0 , 0.05 M S 0 " , 2  4  2  2  3  4  aeration, 35°C  105  Figure 4.28 Chalcocite, 5 ppm gold, 5 g/L copper, 0.20 M ( N H ) S 0 , 0.4 M N H , 4  2  2  3  aeration, 35°C  3  106  Figure 4.29 Effect of dilution factor on detected thiosulfate concentration  108  Figure 4.30 Overall Lobo Composite, 0.20 M ( N H ) S 0 , 1 g/L cupric addition, 4  2  2  3  aeration, 35°C  113  Figure 4.31 Overall Lobo Composite, 0.20 M ( N H ) S 0 , 1 g/L cupric addition, 4  2  2  3  aeration, 35°C, reground  114  Figure 4.32 Overall Lobo Composite, 0.20 M ( N H ) S 0 , 1 g/L cupric addition, 4  2  2  3  no aeration, 35°C  115  Figure 4.33 Overall Lobo Composite, 0.20 M ( N H ) S 0 , no cupric addition, 4  aeration, 35°C  2  2  3  116 viii  Figure 4.34 Overall Lobo Composite, 0.20 M ( N H ) S 0 , no cupric addition, aeration, 4  2  2  3  35°C, duplicate  117  Figure 4.35 Guanaco Composite, 0.20 M ( N H ) S 0 , 1 g/L cupric addition, aeration, 35°C 4  2  2  3  121  Figure 4.36 Guanaco Composite, 0.20 M ( N H ) S 0 , 1 g/L cupric addition, 4  2  2  3  no aeration, 35°C  122  Figure 4.37 Guanaco Composite, 0.20 M ( N H ) S 0 , no cupric addition, aeration, 35°C 4  2  2  3  123  Figure 4.38 Guanaco Composite, 0.20 M ( N H ) S 0 (Thiogold™ grade), 1 g/L cupric addition, 4  2  2  3  aeration, 35°C  125  Figure 4.39 M 4 0 Pyrite Feed, 0.20 M ( N H ) S 0 , 1 g/L cupric addition, aeration, 35°C  128  Figure 4.40 M 4 0 Pyrite Feed, 0.20 M ( N H ) S 0 , 1 g/L cupric addition, no aeration, 35°C  129  Figure 4.41 M 4 0 Pyrite Feed, 0.20 M ( N H ) S 0 , no cupric addition, aeration, 35°C  130  Figure 4.42 M 4 0 Pyrite Feed, 0.20 M ( N H ) S 0 , no cupric addition, no aeration, 35°C  131  4  4  4  4  2  2  2  2  2  3  2  3  2  2  3  3  Figure 4.43 M 1 0 Pyrite Concentrate, 0.20 M ( N H ) S 0 , 1 g/L cupric addition, 4  2  2  3  aeration, 35°C  134  Figure 4.44 M 1 0 Pyrite Concentrate, 0.20 M ( N H ) S 0 , 1 g/L cupric addition, 4  2  2  3  no aeration, 35°C  135  Figure 4.45 M 1 0 Pyrite Concentrate, 0.20 M ( N H ) S 0 , 1 g/L cupric addition, 4  2  2  3  no aeration, 35°C, 24 hours  136  Figure 4.46 M 1 0 Pyrite Concentrate, 0.20 M ( N H ) S 0 , no cupric addition, 4  2  2  3  aeration, 35°C  137  Figure 4.47 M 1 0 Pyrite Concentrate, 0.72 M ( N H ) S 0 , 1 g/L cupric addition, 4  2  2  3  aeration, 35°C  139  Figure 4.48 Ion Chromatograms o f the A S 4 A column representing the sulfate concentration at different times during the leaching o f 1 g/L Chalcocite with 5 ppm gold addition  142  Figure 4.49 Ion Chromatograms o f the Omnipax column showing (from left to right) thiosulfate, trithionate and tetrationate at different times during the leaching o f l g / L chalcocite with 5 ppm gold addition  143  Figure A . l Particle size distribution o f Overall Lobo Composite  161  Figure A . 2 Particle size distribution o f Guanaco Composite  161  Figure A . 3 Particle size distribution o f M 4 0 Pyrite Feed  162  Figure A . 4 Particle size distribution o f M 1 0 Pyrite Concentrate  162  ix  Figure D . l (a) Diagram showing the separation of a mixture o f components A and B by column elution chromatography, (b) The output of the signal detector at the various stages of elution shown i n (a)  164  Figure D.2 Chromatogram of the A S 4 A column  165  Figure D.3 Chromatogram of the Omnipax column  166  LIST OF SYMBOLS E°  Standard potential versus a reference electrode, V  F  Faraday constant; 96500 C /mol  AG °  Standard free energy o f formation; kJ/mol  AG  Standard free energy change; kJ/mol  f  K  0  s t a b  Stability constant  M  Molarity, m o l / L  R  Ideal gas constant; 8.314 J/mol K  T  Temperature; K  xi  ACKNOWLEDGMENTS This research project would not have been possible without the help o f a great number o f people. First o f all, I would like to thank Dr. David Dreisinger for his guidance and support. I am very grateful to Berend Wassink, for his assistance i n dealing with the ion chromatograph and for sharing his knowledge o f sulfur chemistry. Special thanks to Anita L a m , for helping me with the X R D . Thanks to my fellow graduate students o f the hydrometallurgy group for their assistance in various areas, and with whom I enjoyed working with. Thanks to Onno Rutten and B e n Saito for their valuable comments. A s last, I would like to thank my parents and sister for their encouragement and support.  xii  1 INTRODUCTION Cyanide leaching is the conventional method used to extract gold from ores; it is economical, biodegradable and achieves excellent recoveries from a wide range o f ores [Michaelis, 1987]. However, the use o f lixiviants other than cyanide for gold recovery is a subject that has received considerable attention i n the past few years. There are two reasons for this increased interest: 1) Potential environmental restraints on the use o f cyanide i n some areas; 2) Poor leaching characteristics o f some ore-bodies to cyanidation, the so-called refractory ores.  Refractory ores are not amenable to cyanide treatment due to the presence o f impurities (such as copper, arsenic, antimony, tellurium, and manganese), sulfide or silicate encapsulation or pregrobbing characteristics, which cause unsatisfactory gold recovery and high cyanide consumption. Refractory ore-bodies are becoming an increasingly important source o f gold, because o f the decreasing supply o f ores which are amenable to cyanide treatment.  The application o f other lixiviants for gold recovery, such as thiourea, chloride, bromide, and ammonium thiosulfate has been extensively researched in the last few years. These lixiviants would be particularly attractive i f they improve the gold recovery or result i n a savings in reagent cost [Michaelis, 1987]. Ideally the lixiviant is cheap or recyclable, selective, non-toxic, and compatible with down-stream recovery processes [Avraamides, 1982]. In practice, meeting all these criteria is difficult.  Thiosulfate has the ability to complex gold and silver. Moreover, ammonium thiosulfate has been used for many decades as a fertilizer and consequently, from an environmental standpoint, 1  has a definite advantage over cyanide [Atluri, 1987]. Comparing reagent unit costs, ammonium thiosulfate is far cheaper than sodium cyanide (US $0.13/kg vs. U S $1.80/kg). Consequently, with similar or even slightly higher lixiviant consumption, the application o f thiosulfate for gold recovery can be economical and compete directly with cyanidation.  The ammonium thiosulfate system has three indispensable components: thiosulfate, ammonia and copper ions. Thiosulfate stabilizes gold i n solution, while copper and ammonia accelerate the leach reaction. Research into the ammonium thiosulfate system for gold recovery has shown that the chemistry is complex and not yet fully understood. Another complication o f the system is that thiosulfate is prone to degradation. Thiosulfate is a metastable sulfur species which w i l l eventually decompose in aqueous solutions.  A m m o n i u m thiosulfate has been proposed for the treatment o f refractory gold ores i n which carbon components  are intimately associated with the gold, the carbonaceous  ores. The  carbonaceous matter i n the ore adsorbs the solubilized gold complexes from the leach solutions back into the ore and thus reduces gold extraction. Several patents [Kerley, 1981, 1983; Perez and Galaviz, 1987; W a n et al, 1994; Marchbank et al, 1996] describe a process development that uses an ammonium thiosulfate leach solution containing copper as a catalyst for gold recovery. A soluble gold-thiosulfate complex is formed, which is not adsorbed by the carbon content o f the ore. G o l d can then be recovered from solution by several methods including cementation by zinc or copper and ion exchange methods.  2  Copper-gold ores form another type o f refractory gold ore. Large quantities o f this ore type are available for processing. Cyanide treatment o f ores with a high cyanide soluble copper to gold ratio results in unfavourable economics due to high cyanide consumption resulting from the formation o f copper-cyanide complexes.  The objective o f this work is to investigate the application o f ammonium thiosulfate for the leaching o f copper-gold ores. A m m o n i u m thiosulfate is particularly attractive as an alternative to cyanide for the treatment o f copper-gold ores; the presence o f copper is desired i n a thiosulfate leach, whereas it is detrimental in a cyanide leach. Moreover, copper leaching may occur at the same time as gold is leached, provided enough lixiviant (thiosulfate and ammonia) is available, and thus gold and copper may be recovered. A t the moment copper recovery from a cyanide leach solution has not found wide commercial application.  To study the application o f ammonium thiosulfate for the treatment o f copper-gold ores an experimental program consisting o f three different sets o f experiments was performed. First, the leaching behaviour o f sulfide (chalcocite, covellite, chalcopyrite and enargite) and oxide (cuprite and malachite) copper minerals in the ammonium thiosulfate system was studied. It has been reported that copper plays a catalytic role during a thiosulfate leach, however, no information is available regarding the leaching o f copper minerals in an ammonium thiosulfate solution.  To observe the behaviour o f gold in combination with the copper minerals, gold was added to the leach solutions in a second set o f experiments. The experimental program was completed by subjecting several copper-gold concentrates o f different  3  copper and gold grades to  the  ammonium thiosulfate system. Special attention is given to the behaviour o f thiosulfate during the different experimental sets, in order to gain an understanding o f the leach reactions taking place.  This thesis contains five chapters. The second chapter gives a review o f existing literature on the ammonium thiosulfate system for gold recovery, and a brief background on ammonia technology for the recovery o f copper and gold. The experimental methods are described in chapter three, followed by a discussion o f the experimental results in chapter four. The summary and conclusions are presented in chapter five, followed by the recommendations i n the final chapter.  4  2 LITERATURE REVIEW The chemistry o f the ammoniacal thiosulfate  system for gold recovery involves many  interrelated chemical equilibria which are not yet fully understood. This complexity can be attributed to the presence o f three essential components which define the ammoniacal thiosulfate leaching environment: ammonia, thiosulfate and copper. Many reactions may occur between these species, dissolved metal ions, atmospheric gases and ore minerals.  This chapter gives a brief review o f the aqueous chemistry o f copper and gold, followed by a summary o f the research conducted over the years in the field o f ammonium thiosulfate leaching for the recovery o f gold. Some background is also given regarding the recovery o f gold from thiosulfate solutions. The chemical reactions involved i n ammonium thiosulfate leaching and the degradation o f thiosulfate are discussed next.  Ammonia, itself, is an important lixiviant for copper recovery and research has been conducted on the recovery o f gold with ammonia. Both topics are discussed briefly. The chapter ends with a summary o f the presented literature.  2.1 Aqueous Chemistry 2.1.1 Gold-water system G o l d occurs in two oxidation states, as +1 (aurous) and +3 (auric) in aqueous solutions. A s can be seen in Figure 2.1, the E h - p H diagram o f the gold-water system shows no area o f stability o f Au  3 +  relative  to  water,  i.e.  the  standard reduction potential for A u 7 A u is greater than 3  5  the upper stability limit o f water (Equation (II. 1) and (II.3)) [Bard et al, 1985]. The value o f the standard reduction potential for A u 7 A u is even larger (Equation (II.2)) which makes this ionic species even less stable in water (the dashed lines delineate the stability limits o f water).  1.5  1  [  1  1  1  PLOT LRBELS  1  c -  '—  Temp = 2 9 8 . 1 5 K  _  IHul  1.1  = 1.0E-0M  B_ 5TRBLE  ^ —____  ^  ——  '  —  -—-  —  q  flu  B  H Ru 0 3 < 2 - > [RQ)  C  flu  H20  LU  RREH5  I D H 13  STABILITY  1  02/H20  2  H2/H20  LIMITS  R  ^  ^~  ^  —  — ~—•  1.5,  1  1  !  1  1  1  14 PH  Figure 2.1 Eh-pH diagram for the system Au-H 0 at 25°C. The activity of soluble gold species is 10' . 2  4  Au  3 +  Au  + 3e" = A u  E = +1.52 V  S H E  + e" = A u  E = +1.83 V  S H E  (II.2)  E = +1.23 V  S H E  (II.3)  +  0  0  0 +4H +4e" =2H 0 +  2  2  0  (HI)  A s predicted by the Nernst equation for A u (Equation II.4), a decrease in the activity o f free +  gold ions w i l l lead to a lowering o f the standard reduction potential. E = 1.83 + 0.0591og[Au ]  (II.4)  +  6  Complexing the dissolved gold ions with ligands such as cyanide or thiosulfate therefore results in a dramatic lowering o f the reduction potentials (see Table 2.A).  Table 2.A Standard reduction potentials for gold in aqueous solution, [Bard et al, 1985] Ligand  E°, Au /Au (Volts)  E°, Au 7Au (Volts)  H 0  +1.83  +1.52  cr  +1.154  +1.002  Bf  0.960  0.854  r  0.578  0.56  +0.563 ± 0.006  +0.325 ± 0.003  S 0 "  ca. +0.153  ca. +0.10*  CN"  -0.595 ± 0.002  -0.50*  +  2  NH  3  2  2  3  3  *Avraamides [1982]  From Table 2 . A it can be concluded that all the ligands listed form water-stable complexes with gold, i.e. the resulting standard reduction potentials are below that o f water.  Therefore, from a thermodynamic viewpoint there are two basic requirements for a successful leaching system for gold [Avraamides, 1982]: 1. The presence o f an oxidizing agent; 2. The presence o f a suitable complexing ligand.  Some stability constants associated with these gold complexes are shown in Table 2 . B .  7  Table 2.B Stability constants for some gold complexes [Avraamides, 1982] Au  Au  +  Complex Au(CN) " 2  2xl0  3 8  5 x 10  28  Auiy  4 x 10  19  AuBr '  10'  AuCl "  10  3  3  2  2  2  Complex  ^stab.  Au(CN) "  ca. 10  Aul "  5 x 10  AuBr "  10  32  AuCV  10  26  4  Au(S 0 ) " 2  3+  4  2  4  56  47  9  From this data, ligands for complexing gold may be chosen, which, when combined with a suitable oxidant, could form a successful leaching system. Then, parameters such as kinetics, selectivity, reagent stability and reagent consumption need to be determined to assess the applicability o f the system.  2.1.2 Copper-water system In the absence o f complexing substances, copper is a relatively noble metal. Copper can exist i n oxidation states o f +1 (cuprous) and +2 (cupric) i n aqueous solutions. Copper forms a large number o f complexes [Atluri, 1987]: •  for the +1 valency complexes with CI", CN", N H and S 0 " , which are colorless;  •  for the +2 valency the yellow C u C F complexes, the intense blue ammine complexes and the  2  3  brown complexes with SCN".  8  2  3  PLOT LRBELS Temp = 2 9 8 . 1 5 K ICu]  1.0  = ].0E-04  STABLE ARER5 fl  0. 5  Cu  B  Cu2 0  C  Cu 0  E  H Cu 0 2 < 2 - >  D Cu <2*> (RQI (RQI  H20 S T A B I L I T Y L I M I T S 1 D2/H2D 2 H2/H20  -0. 5 h  Figure 2.2 Eh-pH diagram for the system Cu-H 0 at 25°C. The activity of soluble copper species is 10". 2  4  The E h - p H diagram for the copper-water system is presented in Figure 2.2. The equilibrium constant for the disproportionation reaction in water 2 C u <-> C u +  2 +  + Cu°  (II.5)  is 10 (at 25°C) and therefore only small amounts o f C u can exist unless it is stabilized by 6  +  complexing agents [Baes and Mesmer, 1976]. The only significant hydrolysis product o f C u is +  C u 0 , which has a low solubility. C u S is very insoluble ( [ C u ] [ S ] = 10" ) and w i l l likely +  2  2  2  49  2  form in a reducing environment in the presence o f sulfide [Baes and Mesmer, 1976].  The C u  2 +  ion at practical concentrations begins to hydrolyze above p H 4 and precipitates as  oxides or hydroxides soon thereafter, i.e. solution hydrolysis occurs to a slight extent at moderate concentrations before precipitation occurs [Atluri, 1987]. 9  Copper complexes with N H are notably stable (Equation (II.6) and (II.7)) [Sillen, 1971]. 3  C u + 2 N H = Cu(NH )2 +  3  Cu  2 +  + 4NH  3  3  = Cu(NH )  2 +  3  K = 10  1 0 4  (II.6)  K = 10  1 2 3  (II.7)  Other copper(II)ammonia complexes form in addition to the ones given i n Equation (II.6) and (II.7). However, these w i l l not dominate in the p H range o f interest ( p H 9 to 10).  2.1.3 Thiosulfate The E h - p H diagram o f the sulfur-water system demonstrates that thiosulfate has no dominant region in aqueous solution (Figure 2.3). It is only when sulfate formation is omitted i n the construction o f the E h - p H diagram for the sulfur-water system, that the metastable sulfur species appear (Figure 2.4).  1. 5  1.0  0. 5  1  1  i  I  PLOT  I  —  -  "r:  0  IS)  -  ~.  £  LRBELS'  Temp = 2 9 8 . 1 5 K = 0.M  0  H S 04 <->  B  S 014 < 2 - >  C  S  D  H2 5 (RQ)  E  H S <->  F  S < 2 - > IRQ)  H20  -5-  -0. 5  (HQ)  STABILITY  1  02/H2D  2  H2/H20  (RQ) IfiQI  LIMITS  ~ ~ —  :;  E  -1.0  F  2  i  LI  i  6  1  1  1  8  10  12  m  PH  Figure 2.3 Eh-pH diagram for the system S-H 0 at 25°C. The activity of sulfur species is 0.4. 2  10  1.5  PLOT L A B E L S Temp = 2 9 8 . 15 K IS]  = B.M  STABLE AREAS R  S  B  S <2->  C  S 0 3 < 2 - > . IflQl  IRQ)  •  52 03 < 2 - >  E  S2 0 8 < 2 - >  (RQ) IRQ!  F  OB < 2 - >  IRQ)  C  H S < - > [RQI  H  H2 S [AQ1  1  H S 0 3 < - > IRQ)  H20 S T A B I L I T Y 1  02/H20  2  H2/H20  LIMITS  Figure 2.4 Eh-pH diagram for the metastable system S-H 0 at 25°C. The activity of sulfur species is 0.4 M (AG ° (S 0 ) = -518.8 kJ/mol). 2  2  f  2  3  Thiosulfates are compounds containing the group S 0 " which is a structural analog o f sulfate 2  2  3  with one oxygen atom replaced b y a sulfur atom. The two sulfur atoms are not equivalent (Figure 2.5). s~  O— S— 0" o Figure 2.5 Structure of the thiosulfate ion [Kirk Othmer, 1983]  The unique chemistry o f the thiosulfate ion, S 0 " or S S 0 " , is dominated b y the sulfide-like 2  2  3  2  3  sulfur atom which is responsible for the reducing properties and complexing abilities o f thiosulfates [Kirk-Othmer, 1983]. 11  Chemical properties o f thiosulfate include: I.  Tendency to be oxidized (by 0 , C u  II.  Tendency to hydrolyse at p H < 5.5  2  etc., see section 2.3.3)  A.  to S° and H S 0 - at mildly acid p H [Skoog and West, 1976]  B.  to more complex mixtures i n strong acid [Smith and Hitchen, 1976]  2  3  III.  Relative hydrolytic stability in basic solution [Pryor, 1960]  IV.  Thiosulfate forms complex ions with a variety o f metals, e.g. gold, silver, copper ( C u  +  and C u ) , iron (Fe ) etc. [Burns et al, 1981]. The stability o f these complexes depend on 2+  3+  the solution conditions. V.  Formation o f metal sulfides, e.g. with copper, silver, mercury [Burns et al, 1981]  VI.  Reductive stability toward reduction to free sulfide or sulfur; the E° o f Equation (IV.8) is -0.643 V ,SHE (IV.8)  S 0 " + 2 0 F T + 2e" -> H S " + S 0 ~ + O F T 2  2  3  The best known reaction o f thiosulfate is its oxidation to tetrathionate (S 0 ") by iodine, which is 2  4  6  widely used i n analytical chemistry. The two most important salts o f thiosulfate are sodium thiosulfate, N a S 0 2  2  3  or N a S 0 . 5 H 0 , and ammonium thiosulfate, ( N H ) S 0 . The latter one 2  2  3  2  4  was used for this research.  12  2  2  3  2.2 Overview of Thiosulfate Technology for Gold Recovery  2.2.1 Thiosulfate leaching for gold extraction The use o f thiosulfate for the recovery o f precious metals was first proposed in the 19 century th  and is known as the "Patera Process". It was applied in 1858 by V o n Patera for silver recovery. The process utilizes the solubility o f silver chloride i n a solution o f sodium thiosulfate. The ore was treated by a chloridizing roast followed by a sodium thiosulfate leach. Sodium sulfide was used to precipitate silver sulfide [Liddell, 1945].  Berezowsky and Gormely [1978] revived the interest i n the thiosulfate leaching system for precious metals recovery by developing an atmospheric ammoniacal thiosulfate leach system to recover gold and silver from ammoniacal oxidative pressure leach residues o f sulfidic copper concentrates.  The residues were comprised predominantly o f hydrated ferric oxides and  unleached sulfides. Extractions o f 88-95% for gold and 83-98% for silver were achieved after 2 to 4 hours leaching with 0.4 to 0.8 M S 0 ' at 40 to 60°C. The parameters affecting the gold 2  2  3  solubilization were thiosulfate, ammonia and cupric ion concentrations, temperature,  and  residence time. To minimize oxidation o f the thiosulfate leach conditions initial testwork was performed under a nitrogen atmosphere.  In some experiments, a decrease in gold extraction with time was observed. Berezowsky and Gormely  [1978]  attributed  this  phenomenon  to  thiosulfate  degradation;  thiosulfate  decomposition in the presence o f copper causes the precipitation o f copper sulfide which coats the gold particles and thus inhibits gold leaching and recovery. Therefore, an optimal leach time exists; the leach time should not be needlessly prolonged.  13  A direct thiosulfate leaching test was performed on a chalcopyrite concentrate with 3 to 5 g/L Cu  2 +  addition and m i l d air sparging. A gold extraction o f 97% was obtained after 1 hour.  Analysis o f the leach liquors for thiosulfate indicated that 90-95% o f the thiosulfate  was  accounted for and could be recycled [Berezowsky and Sefton, 1979].  The dissolution o f metallic gold i n ammoniacal thiosulfate solution containing copper ions was investigated by Tozawa et al [1981]. The experiments were conducted i n an autoclave under various conditions. The following results were reported (Figure 2.6): 1. In the absence o f C u ( N H ) 3  2 + 4  the kinetics o f dissolution o f gold are very slow.  2. A t leaching temperatures between 65 to 100°C the dissolution o f gold is inhibited by the formation o f a copper sulfide coating. 3. Vigorous stirring and high temperature (>140°C) decrease the gold dissolution due to excessive oxidation o f thiosulfate ions.  '  •  20  30  < 40  •  '  '  '  '  i  i  '  50  60  70  80  90  100  110  120  T e m p e r a t u r e  '  '  130 140  ' 150  i  i  i_  160  170  180  (*C)  Figure 2.6 Effect of copper concentration and temperature on the dissolution of gold in 0.25 M thiosulfate, 1.0 M NH ,196 kPa0 , stirring velocity 200 rpm, [Tozawa et al, 1981] 3  2  14  The same authors also investigated the effect o f oxygen partial pressure on the dissolution o f gold in thiosulfate medium. M a x i m u m gold dissolution occurred at 98 k P a oxygen partial pressure. A t higher oxygen pressures the dissolution o f gold decreased because o f thiosulfate oxidation.  Kerley [1981,1983] patented a process for the recovery o f precious metals from refractory ores, particularly those containing manganese and/or copper, by lixiviation using an ammonium thiosulfate leach solution. The patent claims to improve upon the thiosulfate leaching o f the patent o f Berezowsky and Gormely [1978] by providing better control o f the stability o f the thiosulfate ion (see section 2.3.3).  The leach solution contains 1.2 to 1.35 M S 0 " , 1 to 4 g/L C u , sufficient ammonia to maintain 2  2  2 +  3  a p H o f 7.5 or higher, and a minimum o f 0.05% sulfite ions to control the stability o f the solutions during leaching. Kerley [1981] claims that sulfite ions inhibit the decomposition o f thiosulfate according to Equation (II.9) and thus prevent precipitation o f metal sulfides. 4 S 0 ' + 2 S " + 6 H <-> 3S 0 f 2  2  +  2  2  + 3H 0 2  (II.9)  This reaction is however unlikely to occur at p H 7.5, since the sulfide ion is not stable i n this p H region (see Figure 2.4 and section 2.3.3). The precious metals can be recovered by conventional methods, such as cementation or electrolysis.  The process patented by Kerley was carried out in a plant i n M e x i c o , but the scale-up from laboratory to plant-size failed. Perez and Galaviz [1987] describe the modifications required to  15  make plant operation feasible. The most important process adjustment was the p H value which should be maintained at a minimum level o f p H 9.5, instead o f 7.5 as suggested by Kerley [1981].  This higher p H inhibits the action o f substantial amounts o f metallic iron and ferric salts that are present i n the lixiviating solution as a result o f grinding the ore i n a ball m i l l prior to lixiviating. According to Perez and Galaviz [1987] the ferric ion accelerates the oxidation o f thiosulfate to tetrathionate (Equation 11.10) which has no lixiviating action on gold or silver. 2S 0l~ + 2 F e  2Fe  3+  2  2+  +S 0 "  (11.10)  2  4  The ferrous ion displaces silver and gold from solution according to: FeO + A g S 0 2  2  -> A g S + F e S 0  3  2  (11.11)  4  A t high p H (9.5 and higher) ammonia hydrolysis produces hydroxide ions which react with the ferrous ions according to Equation (11.12) and (11.13), thus preventing iron from displacing silver and gold from solution. Fe  2 +  + 2 N H + 2 H 0 = Fe(OH) + 2 N H ; 3  4Fe(OH) + 0 2  2  2  2  + 2 H 0 -> 4Fe(OH) (s) 2  3  (11.12) (11.13)  Zipperian and Raghavan [1988] identified the parameters o f importance in the dissolution o f gold and silver values from a rhyolite ore with a high manganese content using ammoniacal thiosulfate solutions containing copper. The effect o f thiosulfate, ammonia concentration, temperature, and copper sulfate addition was researched. Optimum conditions were established at 2 M S 0 \ 4.1 M N H , 6 g/L C u , 50°C, and 2 hours leaching i n the absence o f oxygen. 2  2  3  2+  3  16  In the absence o f cupric ions only 14% gold was solubilized. Initial rates o f gold extraction were enhanced by increasing cupric ion concentration, but the ultimate extraction was not influenced by the cupric ion concentration in the range investigated (up to 6 g/L Cu). Further, it was concluded that maintaining optimal p H and E h conditions (pH 10 and 200 m V ) are necessary to prevent precipitation o f copper as C u S . H a l f o f the thiosulfate in the lixiviant solution at p H 9.52  10 was reported to be consumed during the dissolution process.  The consumption o f the reagents and the recycling o f the lixiviant over a long period o f time was researched by Gong and H u [1990]. It was reported that the consumption o f thiosulfate depends on the amount o f air used, the temperature, agitation and configuration o f the reactor. I f these variables are optimized the lixiviant solution can be recycled and the losses are minimal. However, details regarding agitation and reactor configuration are not described. Over 9 5 % gold extraction was achieved under the conditions o f 0.8 to 1 M S 0 " , 1.8 to 2.2 M N H , 0.1 M 2  2  3  3  N a S 0 , 1 g/L C u , 40°C, 1.5 hours and an oxygen atmosphere. The p H was adjusted to 10. 2 +  2  3  Hemmati et al [1989] studied the application o f ammoniacal thiosulfate leaching for gold recovery on various types o f ores. It was found that the highest gold extractions were achieved at 0.7 M S 0 " , 3 M N H , p H o f 10.5, 35°C (over the range o f 25 to 85°C) and an oxygen partial 2  2  3  3  pressure o f 103 kPa. G o l d extraction o f 73% was accomplished under these conditions. N o impact was found on either gold extraction or thiosulfate consumption by variation o f the oxygen pressure i n the range o f 0 to 206 kPa. Hemmati et al [1989] reported that the efficiency o f thiosulfate leaching depends upon ore type. For treating carbonaceous ores, thiosulfate was found to be chemically superior and economically advantageous over cyanide.  17  Langhans et al [1992] focused on maximizing gold extraction while minimizing thiosulfate consumption at low reagent concentrations and ambient temperatures and a p H between 9 and 11. The range o f the thiosulfate concentration tested was 0.05 to 0.2 M . The research was conducted on low-grade oxidized gold ores with application to heap, dump, or in-situ leaching techniques and found to be competitive with conventional cyanidation. After leaching for 48 hours, 83% gold extraction was achieved with 0.2 M S 0 \ 0.09 M N F L O H , 0.00625 M N a S 0 , 2  2  63.5 ppm C u  2 +  3  2  3  with 0.4 kg S 0 " consumed per tonne o f ore. Langhans et al [1992] concluded 2  2  3  that these results compare favorably with 86% gold extraction after 24 hours with 0.21 k g C N " consumed per tonne o f ore using standard cyanidation methods.  The effectiveness o f low thiosulfate concentrations was confirmed by Cao et al [1992], who studied the effects o f the concentrations o f thiosulfate, copper and ammonia on the extraction o f gold and silver from a sulfide concentrate. Leaching was performed with a 0.2 M thiosulfate solution at 2 hours retention time which yielded 95% gold extraction. The sparge air was treated with ammonia in order to keep the ammonia concentration constant. Cao et al [1992] also researched the influence o f sulfate addition to the system, i n order to reduce the thiosulfate consumption, but found no effect (see section 2.3.3).  W a n et al [1994] patented  a process using a thiosulfate lixiviant for the treatment o f  carbonaceous gold ores on a heap leach without applied pre-treatment. The p H o f the solution should be i n the range o f 9 to 10 with a thiosulfate concentration o f 0.1-0.2 M .  18  Newmont G o l d Company evaluated the application o f ammonium thiosulfate on a demonstration heap leach o f 327,000 metric tonnes o f low-grade carbonaceous sulfidic ore that was pretreated by bio-oxidation. The average gold recovery for thiosulfate heap leach at a particle size o f minus 1.9 cm was found to be approximately 55 per cent. A m m o n i u m thiosulfate consumption was about 5 kg/tonne for low sulfide carbonaceous ores (without bio-oxidation) and 12-15 kg/tonne for bio-oxidised ores. Typical leach solutions used contained 0.1 M S 0 \ 0.1 M N H and 30 2  2  ppm C u  2 +  3  3  [Wan, 1997]. Construction started for a commercial size heap leach operation, but was  halted due to the current low gold price.  The application o f a thiosulfate salt lixiviant to recover gold from an oxidative pressure leach slurry is described i n a patent granted to Marchbank et al [1996]. A n ore slurry o f refractory sulfidic and refractory carbonaceous ore is subjected to pressure oxidation i n an autoclave under neutral or alkaline conditions followed by leaching with a thiosulfate salt in stirred tank reactors. Typical leaching conditions are 0.025 to 0.1 M ( N H ) S 0 , 50 to 100 ppm C u , 40 to 55°C and a 2 +  4  2  2  3  minimum sulfite concentration o f 0.001 M , while maintaining a p H between 7 and 8.7.  2.2.2 Gold recovery from thiosulfate solutions Process options for the recovery o f gold from thiosulfate solutions are: •  Cementation  •  Carbon adsorption  •  Direct electrowinning  •  Reductive precipitation  •  Ion exchange  •  Solvent extraction 19  Most researchers report the application o f cementation for the recovery o f gold from thiosulfate solutions [Berezowsky and Sefton, 1979; Kerley, 1983; Perez and Galaviz, 1987; W a n et al, 1994; Marchbank et al, 1996]. Suggested precipitants are zinc, copper, aluminum, iron or soluble sulfides. The use o f copper as precipitant is preferred since it enables the precipitation o f gold without also causing precipitation o f copper from the lixiviant solution.  Gallagher [1987] found that activated carbon has a very low affinity for the gold(I)thiosulfate complex. M c K e e and Lulham [1991] reported that gold was effectively recovered by the addition o f a near stoichiometric amount o f cyanide to the leach solution, enabling the recovery o f the gold cyanide complex by activated carbon or an anion exchange resin.  The direct electrowinning o f a gold thiosulfate leach solution onto a steel cathode at 40°C was investigated by Abbruzzese et al [1995]. The kinetics o f electrowinning were reported to be fast and the precious metal recovery was quantitative (99%). N o qualitative information was given. The current efficiency was 4% due to the concurrent parasitic reactions at the electrodes (oxygen evolution at the anode and reduction o f water and dissolved oxygen at the cathode).  Awadalla and Ritcey [1991] reported the use o f sodium borohydride for the reduction o f gold and silver in acidic solutions o f thiosulfate. Complications o f this process option are: l o w p H (around 6) and Cu ions decrease the efficiency o f borohydride. 2+  Atluri [1987] studied the use o f three anion exchange resins (Amberlite I R A - 4 0 0 , Amberlite I R A - 6 8 and Amberlite IRA-94) for the selective recovery o f gold and silver from simulated  20  thiosulfate leach liquors containing copper, gold and silver. A l l the three resins investigated were not selective to silver and gold over copper. Another process option mentioned is solvent extraction [Marchbank et al, 1996]. However, no literature was found which covered the selective extraction o f gold from a thiosulfate solution.  This brief review o f recovery methods indicates that the recovery o f gold from thiosulfate solutions is not easily accomplished, and offers opportunities for more research.  2.3 Chemistry of the Ammonium Thiosulfate System This section reviews some  fundamental  aspects o f the  ammonium thiosulfate  system:  thermodynamics, copper catalysis and the degradation pathways o f thiosulfate.  2.3.1 Thermodynamic considerations Thermodynamically, the ammonium thiosulfate system is not stable. Both ammonia and thiosulfate w i l l be lost through escape or decomposition. Therefore, true equilibrium diagrams cannot be constructed for this system. B y omitting the more stable species such as sulfate ions from the system, the features o f the metastable species can be examined (see Figure 2.4). The species considered are: N H , N H , S 0 " , S°, S 0 " (x>2), S 0 \ H S 0 \ The stability o f the +  4  2  3  2  2  3  x  6  2  3  2  3  thiosulfate ion w i l l be discussed in section 2.3.3.  The E h - p H diagram for the gold-ammonia-thiosulfate-water system at 25°C is presented in Figure 2.7.  21  -1.0 h  0  2  4  6  10  8  12 14  PH  Figure 2.7 Gold-ammonia-thiosulfate-water system at 25°C, Activity of species: 0.1 M thiosulfate, 0.1 M ammonia, 5x10^ M gold [Li et al, 1995] According to Figure 2.7, under alkaline conditions (pH>9), Au(NH ) is expected to be the most +  3  2  stable gold species. However, rest potential measurements performed by L i et al [1996] suggest that the predominant gold species is Au(S 0 ) " since the gold rest potentials varied with 3  2  3  2  thiosulfate concentrations and not with ammonia concentrations (in the range of 0 to 0.5 M N H ) 3  (Figure 2.8). -0.10  .0.30 -1.6 1  1  -1.4  1  -1.2  1  -1.0  ' -0.8  L  -0.6  1  -0.4  1  -0.2  0.0  log[S 0 ] (M) 2  2  3  Figure 2.8 Effect of thiosulfate concentration on the rest potential of gold [Li et al, 1996] 22  The half-cell reaction for gold thiosulfate is: Au(S 0 ) 2  3  3_  + e" -> A u + 2 S 0  E°= 0.153 V  2 _  2  S H E  (11.14)  It can be noted from Figure 2.7 that gold remains in the metallic state i f potentials are too low throughout the p H range of 6-12. Thus thermodynamics predicts that potentials exceeding 0.05 V  S H E  are required for gold leaching.  Figure 2.9 shows that copper(I)thiosulfate is the predominant species in solution at most common potentials and appears to be in equilibrium with copper(II)ammine (see section 2.3.2). Thus the excess copper(I)thiosulfate ions will be converted back to copper(II)ammine ions to maintain the equilibrium as dictated by potential [Li et al, 1996]. At low potentials and high pH, the decomposition of thiosulfate may lead to the precipitation of copper sulfides.  1.8 1.0  as | ao •0.5 -1.0 -1.5  O  2  4  6  10  8  12  14  PH  Figure 2.9 Copper-ammonia-thiosulfate-water system at 25°C, 0.1 M thiosulfate, 0.1 M ammonia, SxlO" M copper [Li, 1995] 4  23  2.3.2 Copper catalysis The rate o f gold leaching in an ammoniacal thiosulfate solution is enhanced greatly by a catalytic copper reaction [Berezowsky and Sefton, 1979, Zipperian et al, 1988, Tozawa et al, 1981]. However, conflicting opinions exist on the exact role o f the copper in the leach chemistry.  Agreement does exist concerning the dissolution reaction o f gold in thiosulfate solutions i n the presence o f oxygen (Equation 11.15). 2Au + 4S 03 + y 0 _  2  2  +H Oo> 2Au(S 0 )'- +20H" 2  2  (11.15)  3  The above reaction can be modified to incorporate the influence o f the catalytic action o f the cupric ion during gold dissolution [ L i et al, 1995; Marchbank et al, 1996]]:  A u + 5S Q j 2  2  + C u (_N H ) 3  2 +  -> A u ( S Q ) ~ + C u ( S Q ) " + 4. N H 3  2  3  2  3  3  3  (11.16)  Copper oxidation by oxygen or another appropriate oxidant  Copper minerals i n some ores can act as a source o f cupric ion or copper sulfate may be added to the leach solutions. The appropriate copper half cell reaction is: Cu(NH ) 3  2 +  + 3 S 0 " +e" -> Cu(S 0 )*~ + 4 N H 2  2  2  3  3  E°=0.225 V o l t  (11.17)  L i et al [1996] states that oxygen in the absence o f a catalyst may be an oxidant for gold leaching, but with much slower kinetics. It was found that i n the absence o f copper, the gold leaching rate i n ammonium thiosulfate becomes negligible (Figure 2.10).  24  0.00  0.01  0.02  0.03  0.04  0.05  [ C u s c g (M)  Figure 2.10 Effect of copper sulfate concentration on gold leaching rate [Li et al, 1996]  At low copper levels, an increase in the copper concentration results in a dramatic rise in the gold dissolution rate. However, too high copper concentrations significantly inhibit gold leaching. L i et al [1996] explained this by a deficiency of lixiviant for complexing gold, because the copper combines with most of the ammonia and thiosulfate in solution. Furthermore, thiosulfate degradation is accelerated by high copper concentrations (see section 2.3.3). Either way, loss of reactants results in slower leaching kinetics.  It is evident from Equation (11.17) that to enable the regeneration of the cupric ion, it is necessary to keep the concentration ratio of ammonia to thiosulfate in a certain range. Increasing the concentration of only one of the ligands will have a limited positive effect on the gold leaching process. However, this may have a negative effect, as an excess of either ammonia or thiosulfate may make the reaction depicted in Equation (11.17) less reversible [Li et al, 1996]. This effect is illustrated in Figure 2.11. •25  0.10  I  I  I I I 11 T |  1  1  1 I I I I I |  1  1 I I I l l |  1  1  1  1 I I I I I  0.08 h •a  "a  g; 0.06  2 S> 0.04  •Xi•-4 o cts oo  o  0.02 0.00* 0.01  LJ  i  i  i  i  i  1111  0.1  i  1  .  1  >  .  .  . . 4  •  10  100  [NH^H]/[(NHJ S 0 ] 2  2  3  Figure 2.11 Effect of the concentration ratio of ammonia to thiosulfate on gold leaching rate [Li, 1996]  Jiang et al [1993] proposed a model for gold dissolution in ammoniacal thiosulfate solutions on the basis of electrochemical investigations (Figure 2.12). According to this study, ammonia preferentially complexes gold, and the Au(NH ) , which forms at the anodic surface then reacts +  3  with S 0 " and is converted to the more stable Au(S 0 ) ". The Cu(NH ) 2  2  3  3  2  the cathodic surface and is reduced to Cu(NH ) 3  Cu(NH ) 3  2+ 4  + 2  3  2  3  2+ 4  gains an electron at  which is then oxidized by oxygen into  after entering the bulk solution.  In this model, a prominent role is ascribed to oxygen in the system, and the effect of the ratio of ammonia to thiosulfate does not seem to play a role on the regeneration of the cupric ion. This is in contradiction to the observations of L i et al [1996]. The two models display the complexity of the ammonia thiosulfate leaching system and the need for more fundamental research.  26  Gold Surface  Solution  Anodic Area  NH  3  Au = Au*Ve AuV2NH =Au(NH3>2 3  + Au(NH )| 3  Au  e • C u ( N H )2* | 3  Cathodic  Area  t 0 -*OH~  Cu(NH3)4*+e=Cu(NH )2  2  3  Cu(NH ) 3  2  Figure 2.12 The model of electrochemical-catalytical mechanism of ammoniacal thiosulfate leaching of gold [Jiang et al, 1993]  2.3.3 Thiosulfate degradation reactions The variables affecting the stability of thiosulfate solutions are [Skoog and West, 1976] •  pH;  •  The presence of microorganisms and redox catalysts;  •  The concentration of the solution;  •  The presence of oxygen;  •  Exposure to sunlight.  Experiments indicate that in the absence of strong oxidants or catalysts, the stability of thiosulfate solutions is at a maximum in the p H range between 9 and 10 [Skoog and West, 1976]. Bacterial activity appears to be at a minimum at this pH, which explains the maximum stability in this pH region since bacterial activity is reported to be the most prominent cause of instability.  27  Thiosulfate decomposition is catalyzed by copper(II), and iron(III) ions as well as by the reaction products of the decomposition. Skoog and West [1976] report that the decomposition rate is greater in more dilute solutions.  In the case of oxidative decomposition of thiosulfate, sulfur may exist in many different oxysulfur species in the +2 or higher oxidation states. These include S 0 " (+2), S 0 " (+2), S 0 " 2  5  2  6  2  2  3  4  6  (+5/2), S 0 - (+3), S 0 - (+10/3 ), S0 " (+4), S 0 " (+5), S0 " (+6), S 0 " (+7), S0 " (+8). 2  2  4  2  3  6  2  2  3  2  6  2  4  2  2  2  8  5  Thermodynamically, sulfate is the most stable sulfur species under the preferred leaching conditions (alkaline pH, oxidative potential). This is visible in Figure 2.13, the oxidation state diagram for sulfur at several pH values. By drawing a line between HS" and S0 " at pH 10, it is 2  4  visible that most sulfur species are unstable (all species lie above the line). The nearest species to the line are S° and S 0 " which indicates that relative stability for these species in the pH 10 2  2  3  range.  MEAN OXIDATION STATE Of SULPHUR  Figure 2.13 Oxidation state diagrams for sulfur at several pH values [Peters].  28  Smith and Hitchen [1976] described the oxidation o f thiosulfate to tetrathionate in the presence of cupric ion: 2S 0 f  +2Cu  2  2  2Cu + \ 0  2 +  -> S 0 " + 2 C u 2  + 2 H -> 2 C u  +  +  2  2S 0 -+2FT+{0 2  2  (fast)  +  4  +H 0  2 +  (measurable)  2  (II. 19)  ->S 0 ~+H 0  (11.20)  ^S 0  (11.21)  2  2  (11.18)  4  2  or, i n basic solution: 2S 0 "+H 0 + y0 2  2  2  2  +OH'  2 _  4  Tetrathionate may then decompose according to a variety o f pathways, leading to the following stoichiometry for the overall oxidation o f thiosulfate v i a the tetrathionate pathway: S 0 " + 2 0 F T + 2 0 -> 2 S 0 " + H 0 2  (11.22)  2  2  2  2  Byerley et al [1973] cited the following reactions to describe the disproptionation o f S 0 " to 2  4  6  higher and lower polythionates: S 0 - + S 0 " -> S 0 " + S O "  (11.23)  SO  11.24)  2  2  4  2  2 -  2  2  5  + S 0 - -> S 0 " + S 0 "" 2  4  2  2  3  2  S 0 "+30H" ^ | S 0 " + | H 0 2  (11.25)  2  5  2  2  2 S 0 " + 3 0 H " -> f S 0 " + S 0 " + f H 0 2  4  2  2  2  3  2  (11.26)  These reactions w i l l take place at p H > 7 [Smith and Hitchen, 1976]. The trithionate (S 0 ") 2  3  6  species only decomposes under harsh conditions (pH>13 and boiling) to form thiosulfate and  29  sulfite. Marsden and House [1992] used Equation (11.26) to explain why S 0 " is relatively stable 2  2  3  under basic conditions.  Pryor [1960] showed that S 0 " is relatively stable in basic solution, i.e. i n the absence o f 2  2  3  catalysts. The disproportionation o f thiosulfate in water or aqueous buffers at 250-280°C follows Equation (11.27) and is irreversible. S 0 " + OH" -» SO 2  2 -  2  (11.27)  + HS~  This base hydrolysis was found to be very slow in basic solution, even at 250°C, and sulfite was found not to be an intermediate i n the reaction.  According to several authors the addition o f sulfite ions (S0 ") to an ammoniacal thiosulfate 2  3  solution has a beneficial effect on the stability o f thiosulfate [Kerley, 1981; Perez and Galaviz, 1987; Zipperian et al, 1988; Gong and H u , 1990]. The sulfite ions react with sulfide as to regenerate thiosulfate and consequently prevent the precipitation o f precious metal values with the sulfide ions. Different reactions are proposed to explain this effect: •  Zipperian et al [1988]: 3SO3" + 2S ~ + 3 H 0 <r> 2S 0 ~ 2  2  2  •  2  + 60H  0  (11.28)  Gong and H u [1990]: 4S0 " + 2S " + 3 H 0 o 2  2  2  3S 0 " + 60H" 2  2  (11.29)  Kerley [1981, 1983], Perez and Galaviz [1987]: 4 S 0 " + 2 S " + 6 H <-> 3 S 0 " + 3 H 0 2  2  +  2  2  2  30  (11.30)  It is questionable that the reactions described in Equation (11.28) to (11.30) take place under typical thiosulfate leaching conditions (around p H 9-10, E h 150 to 250 raV ). Little free sulfide SHE  ions and hydrogen sulfide ions w i l l exist at these p H and E h values (Figure 2.4). Equation (11.30) describes the formation o f thiosulfate under acidic conditions. This is unlikely since thiosulfate tends to hydrolyze under acidic conditions [Skoog and West, 1976].  There are a few reactions in which sulfite can participate. A n advantage o f sulfite addition can be the oxidation o f sulfite to sulfate (Equation 11.31). When this reaction is kinetically favored over the oxidation o f thiosulfate to sulfate, sulfite can improve the stability o f thiosulfate i n solution. SO " + j 0 2  2  -> S O  (11.31)  2 -  A further beneficial effect o f sulfite addition is described by Wasserlauf and Dutrizac [1982]. Sulfite ions accelerate the decomposition o f S 0 " , x > 4, in alkaline to mildly acidic solutions. 2  x  6  The products are trithionate, thiosulfate and protons. The reaction with tetrathionate is given in Equation (11.24). A l s o , elemental sulfur may react with sulfite under alkaline conditions, to form thiosulfate (Equation 11.32) [Wasserlauf and Dutrizac, 1982]: S0 -+iS ->S 0 " 2  (11.32)  2  x  2  Several authors described the substitution o f sulfate for sulfite, in order to stabilize thiosulfate [Gong and H u , 1994; Cao et al, 1992]. Since sulfate is quite stable toward reduction [Burns et al, 1980], no effect o f sulfate with respect to thiosulfate degradation would be anticipated.  31  L i et al [1996] reported the effect of ammonia concentration on thiosulfate degradation. Increasing ammonia concentration increased the rate of thiosulfate decomposition significantly (Figure 2.14). Since this effect is more evident after several days, it is unlikely to play a significant role in stirred tank leaching, but has to be accounted for in heap leach operations. No explanation was given for this phenomena.  100  g, 8  3 < < < -  80 60  o c  Y  pH 7.3/0.1 M ATS/0.0 M NH OH  A • O • •  pH 8.1/0.1 M ATS/0.02 M NH OH pH 9.2/0.1 M ATS/0.1 M NH OH pH 9.7/0.1 M ATS/0.2 M NH Otf^ pH 10.1/0.1 M ATS/0.5 M>ift OH pH 11.0/0.1 M NajSjO^il M NH OH  4  4  4  4  4  4  .2 40 a,  e  9 '  a  a  20 6  9  12  15  Time (day)  Figure 2.14 Effect of pH/ammonia concentration on thiosulfate decomposition [Li et al, 1996].  It can be concluded from the above overview that the exact degradation pathways of thiosulfate are very complex. Many reactions are possible between existing and generated species in solution. The reactions quoted by the literature do not reflect this complexity, and even tend to simplify the reactions taking place. A good example of this simplification is the explanation repeatedly cited in the literature regarding the beneficial effect of sulfite addition.  32  2.4 Ammonia technology for copper/gold leaching Leaching of copper using ammonia/ammonium salts has been well established. Up to now ammonia has found no application for the extraction of gold and silver from ores due to the very slow dissolution rate of gold and silver in the absence of proper oxidants, and high temperatures and pressures.  A brief overview of ammonia technology for gold and copper is given, followed by a discussion of relevant chemistry.  2.4.1 Ammonia technology for gold recovery Gold dissolves in ammonia as the gold(I)ammonia complex, [Au(NH ) ] , within the range of +  3  2  water stability over a range of pH values (Figure 2.15). The gold(III)ammonia complex, [Au(NH ) ] is thermodynamically unstable in the region of water stability. 3+  3  4  4.0  Au02  3.0  2.0 >  1.0  0  -2  0  2  4  6  8 pH  10  12  14  16  Figure 2.15 Eh-pH diagram for the Au-NH -H O system at 25°C. Activity of ions is 10 ; NH =1.0 M [Meng and Han, 1993] 4  3  3  33  z  The dissolution o f gold is thermodynamically feasible at room temperature. However, kinetic experiments showed that gold was essentially not leached. A practically acceptable dissolution rate was not observed unless the leaching temperature was greater than 120°C i n the presence o f an oxidant [Meng and Han, 1993].  Various oxidants such as oxygen, cupric, cobaltic ( C o ) and manganic ( M n ) , alone or i n 3+  4+  combination, can be used in a solution containing ammonia and ammonium salts (e.g. ammonium chloride, ammonium carbonate or ammonium sulfate) for the leaching o f gold [Han and M e n g , 1992]. Oxygen i n itself insufficient as an oxidant. Han and M e n g [1992] reported the most desirable combination o f oxidants to be: oxygen, hypochlorite and cupric ion at an appropriate ratio (the ratio was not further specified). M e n g and Han [1993] found oxygen (100 kPa) and cupric ion (10 g/L C u ) as the preferred combination. 2+  Other important variables apart from time and particle size, are temperature and ammonia concentration. M e n g and Han [1993] reported that there appears to be a temperature-controlled solubility limit for the gold(I)ammonia complex: at 180°C this value is about 80 m g / L A u .  Extractions o f 80-96% A u from three gold ores and one concentrate in 3-4 hours were obtained by using an autoclave. Conditions were 190°C, C u  2 +  5 to 10 g/1, free N H 5.5 M , ( N H ) S 0 3  4  2  4  0.5  M , oxygen pressure 600-1000 kPa, particle size minus 0.147 mm, and a pulp density o f 10 to 25% solids.  34  2.4.2 Ammonia technology for copper recovery A m m o n i a leaching o f copper and nickel has been well established. Sherritt Gordon developed an ammonia pressure leach process for extracting copper, nickel, cobalt and sulfur from high grade nickel concentrates [Forward and M a c k i w , 1955].  Anaconda designed the Arbiter process for the treatment o f copper concentrates with a lowpressure ammonia-ammonium sulfate leach [Kuhn et al, 1974]. A suitable mixing technology was applied to overcome the necessity o f high oxygen partial pressure rather than using high pressure reactors. The process requires the use o f oxygen rather than air. The leach takes place at temperatures ranging from 60 to 90°C.  The simplified leaching chemistries o f chalcocite, covellite and chalcopyrite are shown i n Equations (II. 33) to (11.35) [Kuhn et al, 1974]: jCu S + f 0 2  2  + 4 N H + { H 0 <-> C u ( N H ) 3  2  2 +  3  + JS0 ." + O H "  (11.33)  2  CuS + 2 0 + 4 N H + H 0 <-> C u ( N H ) f + S O " + H 0  (11.34)  2  2  3  2  3  2  C u F e S + j 0 + 4 N H + H 0 <-> C u ( N H ) 2  2  3  2  2 +  3  + iFe 0 2  3  +2S0  2 -  + 2FT  (11.35)  Chalcocite is probably the most amenable copper sulfide mineral to ammonia leaching. It has been demonstrated that half o f the copper in chalcocite is essentially free and dissolves readily, leaving a covellite mineral matrix. A s long as there is sufficient ammonia available to complex the copper and sufficient oxygen provided, it is expected that the leaching o f chalcocite and covellite reaches completion [Kuhn et al, 1974].  35  When leaching chalcopyrite, the iron present in the mineral forms a hematite layer on the copper particle, thus becoming the rate-controlling factor in the leaching of chalcopyrite [Williams and Light, 1978].  A general formulation of enargite leaching is given in Equation (11.36) [Kuhn et al, 1974]: C u A s S + 13NH + 8 l 0 + f H 0 <-> 3 C u ( N H ) 3  4  3  2  2  3  2+  +NH H As0 +4S0 4  2  4  2 -  + 2H  +  (11.36)  Figure 2.16 shows the copper extraction as a function of leaching time for three different minerals. Kuhn et al [1974] did not explain the slow leaching characteristics of the enargite mineral, other than that the rate of enargite leaching appears to be dependent on the relative specific surface area of the mineral. Percent 10 0 , _  130*1 I hour 15' 49'  2  3 4 L E A C H TIME (Hour*)  5  6  Figure 2.16 Leach extraction as a function of time on three concentrates of different mineralogy. OWeed chalcocite; A Twin Buttes chalcopyrite; • Butte enargite [Kuhn et al, 1974]  In summary, according to Kuhn et al [1974] the leaching of copper sulfide minerals in the arnmoma-aminonium sulfate environment depends mainly on oxygen mass transfer through the bulk solution to the surface and diffusion through an iron oxide layer, i f present. Several studies 36  support this theory [Halpern, 1953, Fisher and Halpern, 1956]. According to K u h n et al [1974], the solution to this mass-transfer problem lies in the mixing system. B y increasing the stirrer speed, the chemical reaction rate is made to be the rate-controlling mechanism, instead o f oxygen mass transfer, and ammonia leaching can be accomplished in a low-pressure system.  However, Beckstead and M i l l e r [1978] stated that a change i n mixing system changes the morphology o f the hematite deposit which causes a change in surface reaction kinetics, rather than improving the oxygen mass transfer. This was noted before by Forward and M a c k i w [1955] who found that intense agitation tends to reduce the iron oxide layer thickness.  For copper oxide minerals, the following relations can be written [Fisher and Halpern, 1956]: C u O + 2 N H + 2 N H ; -> C u ( N H ) 3  C u 0 + 2 N H + 2NH 2  2 +  3  3  In the presence o f N H  + 4  + 4  +H 0 2  ->• 2 C u ( N H ) £ + H 0 3  2  (11.37) (11.38)  the dissolution o f the copper oxide is favored.  2.5 Summary of literature  2.5.1 Thiosulfate technology From the literature review it was realized that the thiosulfate concentrations applied for the recovery o f gold have decreased over the years o f study. The patents granted in the early 1980's applied thiosulfate concentrations ranging from 0.4 to 2 M thiosulfate, while recent studies mention the addition o f 0.1 to 0.2 M thiosulfate. The application o f lower  37  thiosulfate  concentrations while still achieving high gold extractions render the ammonium thiosulfate process more economical and should therefore be pursued in current and future investigations.  The p H o f the recently developed leaching processes is generally maintained between 9 to 10. This p H range is dictated by the ammonia/ammonium buffer point (9.25 at 25°C), since the presence o f ammonia has to be ensured in order to solubilize copper as the copper(II)ammonia complex. Testing at higher temperature and lower p H values are possible because the buffer point shifts to lower values with increasing temperatures. This is probably the reason why Marchbank et al [1996] were successful in leaching at p H ' s o f 7 to 8.7 at 50°C. The thiosulfate stability w i l l however decrease at higher temperatures and lower p H values. Therefore, a p H o f 9 to 10 is generally preferred at ambient temperature because thiosulfate appears to be less prone to degradation in this region and the copper(II)ammonia complex is stable.  When evaluating the optimum ammonia concentration two factors should be considered. Firstly, there has to be sufficient ammonia present in the leaching system to solubilize the copper partially as an ammonia complex. Secondly, the ammonia to thiosulfate ratio should be kept i n a certain range, preferably around 1 to 2. This effect is illustrated i n Figure 2.9. A too high ammonia concentration might stabilize copper as the copper(II)ammonia complex, or a too high thiosulfate concentration might result in the stabilization o f copper as the copper(I)thiosulfate complex according to Equation (11.39), thus limiting the catalytic action on gold extraction:  Cu(NH ) 3  2 +  + 3 S 0 " + e" -> C u ( S 0 ) " + 4 N H 2  2  5  2  3  3  3  (11.39)  Another significant difference between early studies and recent ones is the difference i n cupric 38  ion addition. In the patent o f Kerley [1981, 1983] a cupric addition o f 1 to 4 g/L is described and Berezowsky and Gormely [1978] mention an addition o f 1-10 g/L. Cupric ion concentrations o f 50-100 ppm are mentioned in the patents granted more recently [Wan et al, 1994; Marchbank et al, 1996]. L o w cupric ion concentrations  seem to be favourable, since high cupric ion  concentrations accelerate thiosulfate degradation.  The addition o f sulfite ions to an ammonium thiosulfate solution i n order to stabilize the thiosulfate ion is widely practiced. A s discussed i n section 2.3.3., the reactions referred to i n the literature seem unlikely to take place. Sulfite might, however, participate in different reactions, such as the reaction with tetrathionate to form trithionate and thiosulfate, and thus influence the leaching reactions (see section 2.3.3).  The main factors effecting the thiosulfate stability i n an ammoniacal thiosulfate leach solution are: cupric ions, oxygen pressure and temperature. Some authors reported a negative effect on thiosulfate consumption with high oxygen pressures, others however, did not observe any negative effect when applying higher oxygen pressures. A t high temperatures (>60°C) excessive thiosulfate consumption and the formation o f a copper sulfide coating on gold particles have been observed. A t slightly higher than ambient temperatures, the kinetics o f leaching are favoured, without causing excessive thiosulfate consumption.  In conclusion, ammonium thiosulfate leaching for gold recovery preferably should be performed at: •  L o w thiosulfate concentrations, ranging from 0.1 to 0.2 M ;  •  A n ammonia to thiosulfate ratio o f 1 to 2;  39  •  Copper concentrations ranging from 50 to 100 ppm;  •  A n alkaline p H , preferably 9 to 10;  •  Ambient temperatures to 50°C.  2.5.2 Ammonia Technology The leaching o f copper sulfide minerals in ammonia solution has been extensively studied throughout the years. Less information is available concerning the behaviour o f copper oxide minerals in an ammonia solution.  Most studies investigated the leaching o f copper minerals at higher temperatures and pressures, only in the Arbiter process ammonia leaching is accomplished in a low-pressure system with sufficient oxygen availability and high stirrer speed [Kuhn et al, 1974].  The optimum conditions for the leaching o f gold in ammonia thiosulfate solution are at l o w temperatures and atmospheric pressure (section 2.5.1). It can be concluded that there is no data available regarding the behaviour o f copper minerals under these conditions.  2.5.3 Treatment of copper-gold ores by ammonium thiosulfate Copper-gold ores are available i n large quantities. However, treatment o f these ores by cyanide is unfavourable for several reasons. Firstly, the use o f cyanide is restricted i n many areas. Secondly, cyanidation o f gold ores containing cyanide soluble copper, results in high cyanide consumption. Since cyanide is an expensive reagent (US$ 1.80/kg), and the recovery o f cyanide from copper cyanide complexes is not widely applied, this reagent consumption can be detrimental for the economics o f a project. 40  A m m o n i u m thiosulfate might be a potential alternative for the treatment o f these ores. Copper is required as a catalyst in the leaching reaction, and thiosulfate is cheap compared to cyanide (US$0.13/kg). Furthermore, ammonium thiosulfate is used as a fertilizer, thus from an environmental standpoint it has an advantage over cyanide.  Little research is available regarding the treatment o f copper-gold ores by ammonium thiosulfate. The research reported by Berezowsky and Sefton [1979] concerned the treatment o f ammonia pressure oxidation residues, which contained 0.54 to 2.58 % copper. Several studies were performed by Gong and H u [1990] and Cao et al [1992] on a gold sulfide concentrate containing 3.16 % C u .  A l l o f these studies were performed at high thiosulfate concentrations (0.4 to 1 M ) , and high ammonia concentrations (1 to 2 M ) . The catalytic effect o f copper on gold extraction was recognized by these researchers,  but only Berezowsky and Sefton [1979] mentioned the  detrimental effect o f copper on thiosulfate stability. None o f the studies attempted to investigate the behaviour o f copper and thiosulfate during the leach.  This study is directed towards developing an economical leach process for copper-gold ores using ammonium thiosulfate. Therefore, the reagent concentrations should be as low as possible. Furthermore, an understanding should be obtained o f the behaviour o f copper, gold and thiosulfate in an ammonium thiosulfate leach solution.  41  3 EXPERIMENTAL METHODS This chapter first discusses the materials and reagents used for testing, followed by a description o f the experimental apparatus, the experimental procedures and an overview o f the performed test work. The chapter concludes with a description o f the analytical methods used in this work.  3.1 Materials  3.1.1 Copper Minerals To study the behaviour o f copper minerals in ammonium thiosulfate solutions, samples o f natural occurring copper minerals were obtained from Ward's Natural Science Establishment, Inc. These include covellite, chalcocite, chalcopyrite, cuprite and malachite. A mineral sample o f enargite was received from the Montana Bureau o f Mines and Geology.  The mineral samples were reduced in size by subjecting the samples to a series o f cone crushers, followed by a disk grinder. A narrow particle size fraction between 230 (63 urn ) and 270 mesh (53 urn) was obtained by screening. This fraction was analysed for copper and iron content. X-ray diffraction ( X R D ) was used to determine the main mineral phases present. The results o f these analyses are summarized in Table 3.A.  42  Table 3.A Composition of copper mineral samples for testwork Mineral Sample  %Cu*  %Fe*  Mineral phases  Covellite (CuS)  63.1 (66.4)  2.5(0)  Covellite, Chalcocite  Chalcocite (Cu S)  53.9 (79.8)  8.43 (0)  Chalcocite, Bornite ( C u F e S )  Chalcopyrite (CuFeS )  28.5 (34.6)  27.7 (30.4)  Enargite ( C u A s S )  23.2 (48.4)  7.0 (0)  2  2  3  4  5  4  Chalcopyrite Enargite, Chalcopyrite, Tennantite ((Cu,Fe)] As S, ) 2  Cuprite ( C u 0 ) 2  Malachite ( C u ( C 0 ) ( O H ) ) 2  3  2  4  3  38.9 (88.8)  13.9 (0)  Cuprite, Goethite ( F e O O H )  47.7 (57.5)  0.6 (0)  Malachite  in brackets represents the percentage of C u or Fe i n a 100% pure mineral  From this table, it can be concluded that all mineral samples contain several mineral phases other than the desired copper mineral phase. For example, X R D showed the presence o f goethite (FeOOH) in the cuprite mineral sample which explains the high iron content. Upgrading these copper minerals by flotation was deemed very difficult; therefore, it was decided to conduct testing with the samples in the as-received condition.  Oxidation o f the copper sulfide minerals after particle size reduction was prevented by purging the sample bottles containing the copper minerals with nitrogen and sealing the lids.  43  3.1.2 Copper-Gold Samples The following copper-gold concentrates and ores were obtained for the leaching studies: •  Overall Lobo Composite from Teck Corporation  •  Guanaco Composite from A m a x G o l d Inc.  •  M 4 0 Pyrite Feed and M 1 0 Pyrite Concentrate from Newcrest M i n i n g L t d .  The Overall L o b o Composite and Guanaco Composite are whole ores, whereas the M 4 0 Pyrite Feed and M 1 0 Pyrite Concentrate are flotation products. G o l d and copper content and the copper mineral phases present in the samples are listed in Table 3.B.  Table 3.B Composition copper-gold concentrate and ore samples Au (ppm)*  Cu (ppm)*  Overall L o b o Composite  2.11  1,120  Chalcopyrite, Covellite, Chalcocite  Guanaco Composite  2.06  3,320  Enargite, Cuprite, Malachite  M 4 0 Pyrite Feed  7.10  5,920  Chalcopyrite  M 1 0 Pyrite Concentrate  45.5  16,600  Chalcopyrite  Copper-gold samples  Cu mineral phases  Assayed by Chem M e t Consultants, Vancouver, B . C .  Mineralogical examination  o f the  Overall Lobo Composite revealed  that the  ore  was  characterized by highly altered rock fragments and lesser granular vein quartz with minor finely disseminated sulfides and iron oxides/oxyhydroxides. Copper mineralization was dominated by chalcopyrite (CuFeS , 76%), with minor covellite (CuS, 24%), trace chalcocite ( C u S , <1%) and 2  2  rare bornite (<0.2%). Trace gold / electrum was identified as small grains (<15um), and was found associated with silicates and iron oxides / oxyhydroxides [Lakefield Research Limited, 1997]. 44  The Guanaco Composite was comprised o f rather strongly oxidized ore with minor remnant pyrite (FeS ), enargite ( C u A s S ) , and sphalerite (ZnS). The oxidation suite includes goethite 2  3  4  (FeOOH), with lesser amounts o f psilomelane ( B a M n M n 2 +  4 + 8  0 ( 0 H ) ) , cuprite ( C u 0 ) , and 1 6  4  2  copper oxysalts minerals (malachite ( C u ( C 0 ) ( 0 H ) ) , azurite ( C u ( C 0 ) ( O H ) ) , and olivenite 2  3  2  3  3  2  2  ( C u ( A s 0 ) ( O H ) ) [Honea, 1998]. N o information is available about the gold occurrence or 2  4  association i n this composite.  The M 4 0 Pyrite Feed was characterized by 78% pyrite (FeS ) and 10% quartz ( S i 0 ) . Some 2  2  minor chalcopyrite (CuFeS , 1.6%) and calcite ( C a C 0 , 1%) was present and the remainder was 2  3  comprised o f mainly clay minerals. The M 1 0 Pyrite Concentrate was found to be similar with 76% pyrite and 7% quartz as the main minerals present. M i n o r chalcopyrite (1.6%) and calcite (2.5%) and some dolomite ( C a M g ( C 0 ) , 5.0%) 3  2  were identified [Oretest, 1997]. Details  regarding gold association are not available for these concentrates.  The L o b o composite and Guanaco composite samples were reduced i n particle size by a series o f cone crushers, followed by pulverization using a disk pulverizer. The product was split into smaller quantities (about 400 grams) suitable for testwork. Size reduction for the M 4 0 Pyrite Feed and M 1 0 Pyrite Concentrate was not necessary. The concentrates were also split into smaller quantities for testwork. The screen analyses o f the copper-gold samples are given i n Appendix A ; the 80% passing size (P ) was determined for every material (Table 3.C). 80  45  Table 3.C P of the copper-gold samples 80  Copper-gold samples  P o (M-m) 8  Overall Lobo Composite  141  Guanaco Composite  160  M 4 0 Pyrite Feed  75  M 1 0 Pyrite Concentrate  46  The P  8 0  o f the Guanaco and Overall Lobo Composite is high for leaching experiments, however,  both size fractions are already the result o f a regrind. Further reduction o f the P  8 0  was deemed  difficult with the equipment available, therefore, testing was continued with this particle size distribution.  3.1.3 Reagents •  Soluble gold was added to some leaching solutions. A 1,000 ppm A u reference standard solution was used. In this solution (10% hydrochloric acid) gold is present as a chloride complex (AuCl "). 4  •  Unless otherwise stated, analytical grade ammonium thiosulfate i n powder form was used. Tessenderlo-Kerley provided Thiogold™ grade ammonium thiosulfate solution. The solution contains between 5 0 - 60 % ammonium thiosulfate ( ( N H ) S 0 ) , 40 - 50 % water and 4  2  2  3  0 - 4 % ammonium sulfate ( ( N H ) S 0 ) . 4  •  2  4  Other reagents used during testing were cupric sulfate pentahydrate, C u S 0 . 5 H 0 , anhydrous 4  2  sodium sulfite, N a S 0 , and anhydrous sodium sulfate, N a S 0 . Concentrated sodium 2  3  2  4  hydroxide (50% solution i n water) was used for p H adjustments. A m m o n i a solutions (1 and 3  46  M ) were prepared from concentrated reagent grade ammonium hydroxide ( N H O H ) solution. 4  Preliminary experiments used anhydrous ammonia gas (99.99% purity) to adjust the p H o f the leach solution and 1 M sulfuric acid ( H S 0 ) prepared from reagent grade sulfuric acid. 2  4  A i r (19.5-23.5% 0 ) was supplied v i a a gas cylinder. 2  3.2 Experimental Apparatus  A schematic diagram o f the experimental set-up used for the batch leaching tests is given in Figure 3.1. The experiments were carried out in a 1500 m L cylindrical glass vessel with a removable high density polyethylene lid. The vessel was vertically baffled with a baffle width o f 11 m m which is 1/10 o f the tank diameter. The baffles were made o f acrylic. The solutions were stirred by a stainless steel Rushton impeller (35 m m diameter, 1/3 o f the vessel diameter, 6 blades) on a stainless steel shaft connected to a variable speed D C motor. A l l experiments were performed at an impeller speed o f 1695 R P M . It was experimentally determined that this was the minimum impeller speed required for complete particle suspension.  47  Thermocouple Sparger  Heater  Redox probe  p H probe  ' Baffles  Rushton turbine  Figure 3.1.  Magnetic stirrer  Schematic diagram of the experimental set-up.  Unless otherwise stated, experiments were performed at 35°C. The glass vessel was placed i n a water bath heated by an immersion heater connected to a P I D controller. A thermocouple was placed i n the reactor to provide temperature control. A i r was added to the solutions v i a a stainless steel sparger placed directly under the Rushton impeller. A gas flowmeter controlled the flow o f air into the reactor.  Controlled p H experiments were performed by ammonia gas and sulfuric acid addition. T w o solenoid valves, connected to the p H controller, regulated the flow o f the ammonia gas from a gas cylinder and the flow o f sulfuric acid from a burette. The p H was controlled in a range o f ± 0.10 p H . The ammonia gas was added to the solution v i a the air sparger.  48  A gel-filled combination p H probe with a silver/silver chloride reference electrode was used to monitor the p H o f the solutions. The redox potential was measured using a platinum combination redox electrode with a silver/silver chloride reference electrode. The E° o f this reference electrode at the experimental temperature o f 35°C was 189 m V . In order to report the experimentally observed E h values versus the potential o f the standard hydrogen electrode (SHE), all values were corrected by adding 189 m V .  3.3 Experimental Procedures  Four sets o f experiments were performed: 1. Preliminary testwork 2. Leaching o f copper minerals 3. Leaching o f copper minerals with 5 ppm gold addition  4. Leaching o f copper-gold samples  The procedure for each set varied and w i l l be discussed separately.  3.3.1 Preliminary testwork The preliminary testwork involved the leaching o f the copper minerals i n one litre o f deionized water. Depending on the copper content o f the different minerals (Table 3. A ) , sufficient mineral feed was added to the solution such that upon 100% copper dissolution, the copper concentration would be 5 g/L. Due to the limited amount o f the 53-63 jam material available, the preliminary testwork was conducted with the <53 jam fraction. 49  The initial slurry o f copper mineral feed was heated to the desired temperature. Once the desired temperature was attained, aeration o f the slurry was initiated (0.23 L/min), and the ammonium thiosulfate was added. The p H was adjusted to the desired value with either ammonia gas or 3 M ammonia solution. This time was noted as time zero.  In p H uncontrolled experiments, the p H was allowed to drift. During p H controlled experiments, the p H was adjusted with either ammonia or sulfuric acid i n a range o f ± 0.1 p H . Slurry samples of 7 m L were withdrawn at selected intervals and centrifuged for liquid/solid separation. The resulting solutions were prepared for copper, gold, thiosulfate, tetrathionate and sulfate analyses.  Experiments continued until the redox potential o f the solution stabilized (typically 6 - 8 hours). A t the end o f the experiment the solution was filtered, the residue was washed with deionized water, dried, and collected.  3.3.2 Leaching of copper minerals After the preliminary testwork, some adjustments were made in the experimental procedure. A s above, experiments were performed in one litre o f deionized water. Sufficient mineral feed was added to the solution such that upon 100% copper dissolution, the copper concentration would be either 5 g/L or 1 g/L copper. The fraction 53-63 urn o f the copper minerals was used for this set of experiments.  Heating, aeration and reagent addition followed the same procedure described above. The p H o f the slurry was adjusted to 10 for all experiments with concentrated sodium hydroxide and the time was set to zero. The p H was allowed to drift in all experiments. 50  Slurry samples o f 7 m L were withdrawn at 30, 60, 120, 180, 240 and 360 minutes, centrifuged and analysed for copper, thiosulfate, tetrathionate and sulfate. In several experiments, a solution sample was taken after 5 minutes. However, this sample time was discontinued at a later stage i n order to reduce the time necessary to perform the analytical procedure with the ion chromatograph after the experiment (section 3.5). The experiments lasted 6 hours, after which the solution was filtered. The residues and the wash water were collected. The amount o f residue remaining after the experiments was not deemed sufficient for copper analysis (see section 3.5).  3.3.3 Leaching of copper minerals with 5 ppm gold addition The experimental procedure for this series o f experiments is similar to that described i n section 3.3.2 with the exception that 5 ppm o f gold from the standard reference was added to the deionized water. Accordingly, solution samples were also analysed for gold.  3.3.4 Leaching of copper-gold samples The experiments involving the concentrates were performed at a pulp density o f 30% by weight (360 grams o f concentrate and 840 grams o f deionized water). The pulp was heated to desired temperature, after which the reagents were added. Experiments were continued as described i n section 3.3.2. One adjustment was the sample volume. Because o f the higher pulp density 8 m L samples were taken. Solution samples for copper, gold, thiosulfate, tetrathionate and sulfate analysis were prepared. Wash water samples were analysed for gold content. The residues were analysed for copper and gold.  51  3.4 Experimental Design The experimental design is discussed according to the four different series o f experiments mentioned i n section 3.3.  3.4.1 Preliminary experiments Preliminary experiments (at controlled p H and uncontrolled pH) were conducted with the copper minerals at 0.05 and 0.1 M ammonium thiosulfate at p H ' s ranging from 8 - 9 . 5 . A l l experiments were performed at a temperature o f 35°C.  3.4.2 Leaching of copper minerals So-called baseline experiments were first conducted, after which the effects o f several parameters could be studied (Table 3.D). Baseline tests included: aeration o f the system, 0.2 M ammonium thiosulfate, p H 10 (uncontrolled), a copper mineral addition which would achieve 5 g/L copper or 1 g/L copper upon 100% dissolution, and 35°C. The experiments lasted 6 hours. T w o experiments were performed without aeration.  Table 3.D Overview of experiments with copper minerals Copper minerals  Baseline  Covellite  X  Chalcocite  X  Chalcopyrite  X  Enargite  X  Cuprite  X  Malachite  X  No aeration  X  X  52  3.4.3 Leaching of copper minerals with 5 ppm gold addition The experiments were performed under the same conditions as described i n section 3.4.2. However, several other parameters were evaluated in this test series; aeration, leach temperature, ammonia concentration, sulfite concentration and sulfate concentration (Table 3.E).  Table 3.E Overview of experiments with copper minerals and gold addition Copper  Baseline  mineral Covellite  X  Chalcocite  X  Chalcopyrite  X  Enargite  X  Cuprite  X  Malachite  X  No  Temperature  0.4 M  aeration  20°C  NH  X  X  X  X  x  3  0.05 M so  X  2 3  0.05 M so 2  4  X  3.4.4 Leaching of copper - gold samples Baseline experiments o f this test series included: aeration o f the system, 0.2 M ammonium thiosulfate, p H 10 (uncontrolled), 30 % pulp density, and 1 g/L copper addition (as cupric ( C u S 0 . 5 H 0 ) ) and 6 hours. The parameters studied in this series are (see Table 3.F): aeration o f 4  2  the system, leaching time, ammonia thiosulfate concentration, copper addition and ammonium thiosulfate grade.  53  Table 3.F Overview of experiments with copper-gold samples Copper-gold  Baseline  samples  No  No  No  No  Extra  Thiogold™  aeration  aeration,  cupric  aeration,  ATS*  Grade  24 hours Overall Lobo  no cupric  X  X  X  X  X  X  X  X  X  X  X  Composite Guanaco  X  Composite M 4 0 Pyrite  X  Feed M 1 0 Pyrite  X  X  X  Concentrate * A T S = A m m o n i u m Thiosulfate  3.5 Sample Analysis  3.5.1 Copper and gold analysis A l l the copper and gold assays on solution and solid samples were performed by C h e m M e t Consultants i n Vancouver. Selected solution samples were analysed for gold and copper content. The solution samples (exactly 20 m L ) for copper were prepared spectrophotometry  for atomic  absorption  ( A A S ) by treatment with hydrogen peroxide to destroy any thiosulfate  present since thiosulfate interferes with A A S measurements. After this, nitric acid was added to dissolve any precipitates resulting from the hydrogen peroxide addition. The solution volume was adjusted to known volume and the copper content measured by A A S .  54  Between 3-5 m L o f sodium cyanide was added to the samples for gold analysis, to keep gold in solution. The gold content was determined by evaporating 25 m L o f solution to dryness i n a lead foil boat, followed by fire assay and cupellation. The resulting bead was digested and the gold content o f the resulting solution was determined by A A S .  The residues o f the copper-gold samples were analysed for copper and gold for mass balance purposes. A representative sample o f the residues was obtained by coning and splitting. Copper content o f the residues was determined by acid digestion and A A S analysis o f the resulting solution. G o l d content o f the residues was determined by fire assay and cupellation. The resulting bead was digested and the gold content o f the resulting solution was determined by AAS.  3.5.2 Thiosulfate, tetrathionate and sulfate analysis A l l solution samples were analysed for thiosulfate, tetrathionate and sulfate by high performance liquid chromatography ( H P L C ) methods directly following the experiment. A Dionex Series 4500i Chromatograph was used for the analysis. The ion chromatography methods and procedures are discussed in more detail i n Appendix B .  Solution samples produced from the experiments were diluted to a concentration suitable for detection by the ion chromatograph (< 20 ppm). The typical dilution factor was 1,000 times. The diluted samples were kept refrigerated and in the dark to reduce the rate o f thiosulfate degradation. The samples were allowed to warm to room temperature just before analysis.  55  It should be noted that only the free ion concentrations are detected by ion chromatography. Sulfur species complexed with metal ions w i l l either not be retained by the column (when the complex has no charge) or have such a high affinity for the ion exchange sites on the column that the complex does not elute; for example the copper(I)tri-thiosulfate complex which has a -5 charge.  Further, for most experiments the presence o f sulfite and trithionate was detected by ion chromatography. However, these species could not be quantified.  56  4 RESULTS AND DISCUSSION The experimental results are presented and discussed in this chapter. A s mentioned in Chapter 3, four different sets o f experiments were conducted: preliminary testwork, leaching o f copper minerals, leaching o f copper minerals with gold addition to solution and leaching o f copper-gold samples. The experimental results are discussed accordingly. First, the construction o f E h - p H diagrams which represent the experimental conditions are discussed.  4.1 E h - p H diagrams To  understand  and  characterize  the  chemistry  o f copper-thiosulfate-ammonia  solutions  containing gold, E h - p H diagrams were constructed which reflect the experimental conditions used in this research. E h - p H diagrams represent the thermodynamic equilibria in a system and do not give information regarding the kinetics o f the reactions represented i n the diagram. This has to be considered when interpreting E h - p H diagrams.  The thermodynamic data used for the various copper and gold species expected to be present in these solutions are presented i n Appendix C . The diagrams were constructed with the use o f the computer program C S I R O Thermochemistry [Turnbull, 1986]. A n E h range o f -1.5 to 1.5 V  S H E  and a p H range o f 6 to 14 were considered. The p H range was chosen between 6 and 14; below p H 6 complex reactions involving the hydrolyses o f thiosulfate start to take place which are not pertinent to this research (see section 2.3.3). The p H values o f interest are around 9-10 (see section 2.5). According to the literature, the E h values during thiosulfate leaching are generally in the range o f 150 to 250 m V  S H E  [Atluri, 1987].  57  The E h - p H diagrams were constructed for 25°C, while the experiments were conducted at 35°C. Entropy data are required to calculate the free energy o f a species at 35°C. This data is not available for several sulfur species (e.g. S 0 " , S 0 " , x>2). The diagrams were 2  2  3  2  x  6  therefore  constructed at 25°C. The change o f the ammonia/ammonium buffer point with temperature can be calculated and changes from p H 9.25 at 25°C to p H 8.96 at 35°C. This change i n p H value is not incorporated on the E h - p H diagrams, but has to be considered when interpreting the experimental results.  Moreover, E h - p H diagrams are theoretical thermodynamic diagrams and should therefore be constructed using ionic activities. In dilute solutions these activities can be replaced by molalities (mol/kg o f solvent). Due to the complex nature o f the solutions i n this system, representative activity coefficients are not readily available. The experimental solutions cannot be considered to be dilute either, since at least 0.4 M N H and 0.2 M S 0 " are present. The diagrams were 2  3  2  3  constructed using molarities, which in this instance, is the most practical alternative.  The E h - p H diagram for the gold-thiosulfate-ammonia-water system appears i n Figure 4.1. The gold activity is 2.5 10" M (5 ppm), the thiosulfate activity is 0.2 M and the ammonia activity is 5  0.4 M . It can be seen that the gold(I)thiosulfate complex is stable i n the whole p H range shown.  58  Figure 4.1. Eh-pH diagram of the gold-thiosulfate-ammonia-water system at 25°C. The activities of the species are 2.5 x 10 Au (5 ppm), 0.2 S 0 and 0.4 N H . AG (S O ) = 518.8'kJ/moL 5  2  2  3  0  3  f  2  2  3  Comparing Figure 4.1 with the Eh-pH diagrams given in the literature (see Figure 2.7), it can be observed that the gold(I)ammonia complex has no region of stability in Figure 4.1, while Figure 2.7 shows a region of stability for this complex. A closer examination of the literature revealed that the free energy of formation value of the thiosulfate species was different. In Figure 4.1 a value of-518.8 kJ/mol was used for the thiosulfate species [Bard et al, 1985]. The Handbook of Chemistry and Physics [Weast, 1975] lists a value of -532.2 kJ/mol for thiosulfate, resulting in the Eh-pH diagram presented in Figure 4.2, which resembles Figure 2.7.  59  PLOT LABELS Temp  = 298.15 K  I Ru1  = 2.5E-05  IS2031  =0.2  INM31  = 0.U  STABLE BRERS fl flu B  H Ru 0 3 < 2 >  (BQ)  C  H2 flu 0 3 < - >  (BQ)  D  Ru [ 0 H  E  Ru [N H3 12 « - >  F  Ru ( S 2 0 3 12 < 3 - >  13  H20 S T A B I L I T Y  (BO) (BQ)  LIMITS  1 02/H2D 2  6  8  10  12  H2/H2D  14  PH  Figure 4.2 Eh-pH diagram of the gold-thiosulfate-ammonia-water system at 25°C. The activities of the species are 2.5 x 10 Au (5 ppm), 0.2 S 0 and 0.4 N H . AG (S O ) = 532.2 kJ/mol. 5  2  2  0  3  3  f  2  2  3  The gold(I)ammonia complex now appears next to the gold(I)thiosulfate complex. This large effect o f this relatively small change i n free energy is explained when the following equation is considered: Au(S 0 ) ~ +2NH = Au(NH ) +2S 0 ~ 2  3  2  3  3  2  2  (IV. 1)  3  The free energy change for this reaction, calculated using the AG ° value o f -518.8 kJ/mol for f  thiosulfate, is + 22.5 k J which indicates that this reaction is not favoured. W h e n the free energy change is recalculated using the AG ° value o f -532.2 kJ/mol for thiosulfate, the value -4.3 k J is f  obtained which indicates that the formation o f the gold(I)ammonia complex is now favoured. It can be concluded that a small change i n the free energy o f the thiosulfate species has a great 60  impact on the thermodynamic equilibria presented in the E h - p H diagram o f the gold-thiosulfateammonia-water system.  According to Figure 4.2, the gold(I)ammonia complex is the most stable species at p H 10, which does not agree with experimental observations. A s L i et al [1996] remarked, it is generally accepted that the gold(I)thiosulfate species is the more stable species at p H 10 and L i et al [1996] confirmed this by rest potential measurements (see  section 2.3.1). These rest potential  measurements are easily explained by Figure 4.1. This example illustrates that E h - p H diagrams have to be interpreted with the necessary precautions.  In Figure 4.1 a decrease in thiosulfate concentration results in a slight decrease i n area o f stability for the gold(I)thiosulfate complex over the entire p H range considered. In Figure 4.2 a decrease in thiosulfate concentration results in a larger area o f stability for the gold(I)ammonia complex. When thiosulfate disappears, thermodynamically the gold(I)ammonia complex becomes the stable species over the whole p H range. In this study the concentrations o f gold i n ammoniacal thiosulfate solutions are usually very low (in range o f 5 to 30 ppm). Hence, a variation i n gold concentration should not significantly affect the thermodynamics o f the system.  The copper-thiosulfate-ammonia E h - p H diagram, calculated using the value o f -518.8 kJ/mol for thiosulfate, is presented i n Figure 4.3. The copper(I)tri-thiosulfate complex occupies the whole p H range and is stable i n the area o f the leaching conditions; p H 10 and an E h range o f 150-250 mV  S H E  . The copper(I)di-thiosulfate complex is stable at low p H values. A slight lowering o f the  E h w i l l result in the precipitation o f copper as C u S .  61  Figure 4.3 Eh-pH diagram of the copper-thiosulfate-ammonia-water system at 25°C. The activities of the species are 0.015 Cu (0.95 g/L)„ 0.2 S 0 and 0.4 N H . AG ° (S 0 ) = -518.8 kJ/mol. 2  2  3  2  3  f  2  3  The copper(II)ammonia complex has a small stability region around the ammonia/ammonium buffer point. A s mentioned before, this buffer point w i l l shift towards p H 8.96 at 35°C. This area lies slightly above the 0 / H 0 line, and w i l l therefore not exist at significant concentrations. This 2  2  complex w i l l have a region o f stability within the water lines, when the thiosulfate concentration decreases. A decrease i n copper concentration increases the width o f the region o f the copper(II)ammonia complex. The dominant copper species at the E h - p H conditions o f the experiments is not affected by a decrease in thiosulfate concentration.  Calculating the same diagram with the thiosulfate value of-532.2 kJ/mol, results in Figure 4.4.  62  PLOT  LABELS  Temp  = 298.15 K  ICul  = 0.015  IS2D3!  = 0.2  INH31  = 0.U  STABLE P B  flRERS  Cu S Cu  C  Cu2 S  D  Cu2 0  F  Cu 0  F  Cu [ S 2 0 3 1 3 < 5 - >  C  Cu [ N H3|L| <2+> IRQ)  H20 S T A B I L I T Y  6  8  10  12  1  02/H20  2  H2/H20  (RQ)  LIMITS  m  PH  Figure 4.4 Eh-pH diagram of the copper-thiosulfate-ammonia-water system at 25°C. The activities of the species are 0.015 Cu (0.95 g/L)„ 0.2 S 0 and 0.4 N H . AG (S O ) = -532.2 kJ/mol. 2  2  0  3  3  f  2  2  3  When comparing Figure 4.4 to Figure 4.3, two observations can be made. The region o f interest is p H 10. In Figure 4.4, the area for the copper(I)thiosulfate complex is greatly reduced compared to Figure 4.3. A t E h values lower than 0 V  S H E  , the C u S species has an area o f stability 2  directly below the copper(I)thiosulfate species instead o f CuS i n Figure 4.3. Furthermore, the region o f stability o f the copper(II)ammonia complex increases to lower E h values when compared to Figure 4.3.  Figure 4.5 shows the copper-ammonia-water system. This E h - p H diagram represents the leaching conditions when thiosulfate is no longer present in solution.  63  Figure 4.5 Eh-pH diagram of the copper-ammonia-water system at 25°C. The activities of the species are 0.015 Cu (0.95 g/L) and 0.4 NH . 3  The copper(I)ammonia complex now appears as a stable species in the area o f interest. The areas o f the copper(I)ammonia and copper(II)ammonia decrease  i n width with higher copper  concentrations and lower ammonia concentrations. A t very l o w ammonia concentrations the stability region o f the copper(II)ammonia complex completely disappears.  4.2 Preliminary experiments  The preliminary experiments were performed to obtain an understanding o f the behaviour o f copper minerals i n an ammonium thiosulfate solution under conditions that are typical for the leaching o f gold (see section 2.5.1). The results o f these experiments were used to determine the experimental conditions for further testwork. 64  Conditions A s described in section 3.3.1 the preliminary experiments were performed with the copper minerals (section 3.1.1), particle size o f minus 270 mesh (53 jam). A copper mineral addition was employed which would achieve 5 g/L copper i n solution upon 100% dissolution. In the remainder o f this chapter this w i l l be referred to as a 5g/L copper addition. 5 g/L copper was chosen as a representative amount for an industrial operation. This would be equivalent to 0.5% copper i n an ore, with leaching at 50% pulp density.  A s outlined i n the literature review (section 2.5.1) it is necessary to apply l o w ammonium thiosulfate concentrations, ranging from 0.05 - 0.20 M , to achieve an economical process for the recovery  o f gold  by  ammonium  thiosulfate  leaching. Therefore,  testing  started  with  concentrations o f 0.05 and 0.1 M ammonium thiosulfate. Since the thiosulfate to copper ratio i n the copper(I)thiosulfate complex is 3 to 1 (the C u ( S 0 ) " complex is not stable under these 3  2  3  2  conditions), there was not enough thiosulfate present in solution to complex all the copper (maximum o f 5 g/L copper (=0.079 M copper)).  The same applies to the ammonia concentration; the ammonia to copper ratio i n the copper(II)ammonia complex is 4 to 1 and thus not enough ammonia is present to complex all the copper i f 100% dissolution would occur. However, ammonia was used for p H adjustment, resulting i n sufficient ammonia concentrations at high p H values (>9.0).  The p H range tested was from 8 to 9.5. Testing was started at lower p H values, considering the shift o f the ammonia/ammonium buffer point to p H 8.96 at 35°C. First, the experiments were p H  65  controlled in a narrow range (+0.1 pH). After several experiments, however, it was realized that by allowing the p H to drift during the experiment, a better understanding o f the chemical reactions occurring in the leach solution can be obtained. Accordingly, the p H was allowed to drift in all further experiments.  Results A l l preliminary experiments showed that the thiosulfate disappeared quickly. Thiosulfate was not detected in the leach solutions after 2 to 3 hours. A n example is presented i n Figure 4.6., the leaching o f covellite with 0.10 M ammonium thiosulfate at p H 9.5.  The first graph reflects the p H (primary y-axis) and E h (secondary y-axis) changes during the experiment. The negative time on this plot reflects the heating o f the solution to 35°C, i n the absence o f reagents. The addition o f the reagents to the solution was noted as time zero. The large increase o f p H and sharp decrease o f the E h result from the reagent addition. The bottom graph shows the results o f thiosulfate and tetrathionate analyses by ion chromatograph on the primary y-axis, and the copper extraction on the secondary y-axis. Sulfate concentrations are not available for this experiment.  66  12  600  *>—• •  M 400  a  ^  -50  50  150  250  ^^^^  200 W  450  350  Time (minutes) pH  -0—  Eh  10  60  40 c  DC  o «  w o c •Si  u  20  |  a o  U ^0 100  200  300  400  500  Time (minutes) • Thiosulfate  Tetrathionate  Copper  Figure 4.6. Covellite, 5 g/L copper, 0.1 M S 0 , 0.47 M NH , pH 9.5, aeration, 35°C 2  3  67  3  A more detailed discussion o f the leaching o f covellite is given in a later section (4.3.1). This graph is presented here for two reasons: •  The thiosulfate concentration in solution decreases quickly. This was also observed during the leaching o f the other copper minerals.  •  Copper extraction is low, just above 20% (of 5 g/L), which is about 1 g/L o f copper.  Based on these results it was decided to try to maintain thiosulfate i n solution by increasing the thiosulfate concentration to 0.2 M . The p H was set to 10 for further experiments. A t this p H thiosulfate is quite stable (see section 2.3.3) and the presence o f ammonia instead o f ammonium in solution is ensured. A t lower p H values the leaching o f the copper sulfide minerals might be prevented because o f an ammonia deficiency.  Moreover, the experimental procedure was changed. Exposing the copper minerals to deionized water resulted i n a unique solution p H for each mineral. B y adjusting the p H with ammonia, a different ammonia concentration was obtained for each experiment. To ensure the same ammonia concentration for all experiments, concentrated sodium hydroxide was used for p H adjustment. B y adding the same amount o f ammonium thiosulfate to every experiment and raising the p H to 10 by sodium hydroxide, the concentrations o f thiosulfate and ammonia w i l l be similar for every experiment (small differences in concentration w i l l exist due to the fact that the amount o f sodium hydroxide required to raise the p H to 10 is mineral dependent).  Furthermore, some difficulties were experienced with the dilution o f solution samples to concentrations within the range o f the ion chromatograph (2-20 ppm). Several samples turned  68  turbid upon dilution. A m m o n i a was added to the solutions to redissolve solid precipitates, enabling the measurement o f the samples by ion chromatograph. It was realized that this addition changed the solution chemistry. For the next series.of experiments it was recorded to which samples ammonia was added (Appendix D).  The ammonia addition may affect the solution chemistry i n the following way: ammonia could displace any thiosulfate complexed with copper, resulting in a higher measured free thiosulfate concentration than actually present in the leach solution. In most experiments, ammonia was only added to the last samples (240, 360 minutes) resulting i n no effect since i n most experiments thiosulfate was no longer present after 180 minutes. For experiments with high initial copper extraction, ammonia was added to all samples. Again, this resulted i n a m i n i m u m effect, since thiosulfate degraded quickly in the leach solutions because o f the high copper concentration.  4.2.1 Summary The main conclusion that can be drawn from the preliminary experiments is that thiosulfate was not detected after 2 to 3 hours under the experimental conditions applied. To extend the presence o f thiosulfate i n solution the thiosulfate concentration was increased to 0.2 M for further testwork.  The copper extractions for the sulfide minerals were not very high and this was thought to be due to l o w p H values, i.e. at low p H values ammonium w i l l be the dominant species instead o f ammonia. To ensure the presence o f ammonia in the leach solution, the p H was increased to 10  69  in later testwork. This p H value might also positively affect the thiosulfate stability, since it is reported to be at a maximum i n solutions o f p H 9 to 10 (see section 2.3.3).  4.3 Leaching of copper minerals  In these experiments copper minerals were leached i n an ammonium thiosulfate solution (see Table 3.D). The baseline experiments included a mineral addition which upon 100% copper dissolution would result in 5 g/L copper in solution, 0.20 M ammonium thiosulfate (~ 22 g/L thiosulfate), p H 10, 35°C and aeration. The p H was allowed to drift. The behaviour o f thiosulfate, tetrathionate, sulfate and copper was recorded. The effect o f aeration was studied. Experimental data and observations made during the experiments, such as solution colours are presented i n Appendix D . A l l % extraction calculations are based on the head assays o f the copper minerals.  4.3.1 Baseline experiments The results o f the baseline experiments are presented in Figures 4.7 to 4.12. The top graphs represent the p H and E h values during the leach, whereas the bottom graphs show the sulfur species and copper extractions. Instrumental difficulties prevented the determination o f sulfate concentrations for some experiments.  70  12  800  10 I 600 8 X c  6+'  400 >  4 200 2 0  0  + -50  0  50  100  150  200  250  300  350  400  Time (minutes) -*-pH  -*-Eh  50  25  40 Co *2 ©  30 «  Vi V  "3 a  20 Z  Vi  U  a  a 10 t j  100  200  300  400  Time (minutes) Thiosulfate  • Tetrathionate  -•— Copper  Figure 4.7 Covellite, 5 g/L copper, 0.20 M (NH ) S 0 , aeration, 35°C 4  71  2  2  3  0  100  200  300  400  Time (minutes) —•— Thiosulfate  —•— Tetrathionate  —•>— Copper  Figure 4.8 Chalcocite, 5 g/L copper, 0.20 M (NH ) S 0 , aeration, 35°C 4  72  2  2  3  12  600  10 8  a  400  s  6  >  4  200 W  izi  2 0 -50  1  1  1  1  1  1  1  1  0  50  100  150  200  250  300  350  0 400  Time (minutes) •pH  -•— E h  25  50 40  3  30  VI  a | «  20 « O  a a  s  10 °  tZ2  100  200  300  400  Time (minutes) -•— Thiosulfate  Tetrathionate  - * - Sulfate  ~#— Copper  Figure 4.9 Chalcopyrite, 5 g/L copper, 0.20 M (NH ) S Q , aeration, 35°C 4  73  2  2  3  12  600  10  -*>«-a—•—•—•—•  8  400  6  Q  s  £  4  200 W ^ ^  ^ -  HMt  0  0  0 ^  2 40  H -50  0  h  H  50  100  150  0  (200  250  300  350 400  Time (minutes) -0— E h  •pH  25  50  20  40 Is  S "3  a u  15  30  |  10  20 2  a a  "a  10 cj  100  200  300  400  Time (minutes) -•— Thiosulfate  • Tetrathionate  - A - Sulfate  Copper  Figure 4.10 Enargite, 1 g/L copper, 0.20 M (NH ) S 0 , aeration, 35°C 4  74  2  2  3  12  600  10 4- 400  8 +  c.  6 200 W  4 2  0  0 -50  0  50  100  150  200  250  300  350  400  Time (minutes) -0—Eh  pH  100 80 60 40 20  100  200  300  a | cs S-  a a o  400  Time (minutes) Thiosulfate  Tetrathionate  Sulfate  Copper  Figure 4.11 Cuprite, 5 g/L copper, 0.20 M (NH ) S 0 , aeration, 35°C 4  75  2  2  3  0  100  200  300  400  Time (minutes) -•— Thiosulfate  -m- Tetrathionate  Sulfate  -#— Copper  Figure 4.12 Malachite, 5 g/L copper, 0.20 M (NH ) S 0 , aeration, 35°C 4  76  2  2  3  A n examination o f the top graphs o f the figures reveals that all figures, except Figure 4.10, show an initial p H increase followed by a p H decrease after a certain time period. The solution p H is influenced by several reactions which occur simultaneously:  •  The oxidation o f thiosulfate to tetrathionate (Equation IV.2), coupled with the reduction o f oxygen to hydroxide (Equation IV.3), resulting in the production o f hydroxide ions: 2S 03 ->S 0 ,"+2e"  (IV.2)  \0  (IV.3)  _  2  2  •  4  + H 0 + 2e" -> 2 0 F T  2  2  Concurrently, hydroxide w i l l be consumed by the hydrolysis o f tetrathionate i n a basic solution: 2S 0l 4  •  + 3 0 P T -> f S Ol~ + S 0 ~ + | H 0 2  2  3  6  2  (IV.4)  Hydroxide is also consumed because ammonia is removed from the system due to evaporative losses and complexation reactions. The equilibrium depicted i n Equation (IV.5) changes accordingly: NH  3  + H 0 <r> N H ; + O H "  (IV.5)  2  The reactions depicted in Equations (IV.2) and (IV.3) are dominant while thiosulfate is present. When the thiosulfate is gone, the hydroxide consuming reactions become more prominent and consequently, the p H decreases. This sequence o f reactions is visible i n for example, Figure 4.7, the leaching o f covellite. The p H increases as long as thiosulfate is present, followed by a p H decrease around the same time (about 240 minutes) that thiosulfate was no longer detected i n solution (bottom graph). 77  A closer look at Figure 4.7 reveals that the E h remains constant for a long time and then suddenly increases around 240 minutes. This time coincides with the instant that the p H starts to drop and thiosulfate is no longer present. Further, the colour o f the leach solution changes: after about 180 minutes the solution colour changes from colourless to light blue indicating that the copper(II)ammonia complex has formed (see Appendix D ) . Moreover, the copper extraction curve shows a steady increase after thiosulfate is gone, which can only be explained by copper complexing with ammonia, since there is no thiosulfate present in solution. It can be remarked that the copper extraction is low.  Figure 4.5, the E h - p H diagram o f the copper-ammonia system at 25°C, can be used to interpret these results. It can be seen from the bottom graph that the copper concentration is only about 0.006 M . This means that the stability regions o f the copper(I)ammonia and copper(II)ammonia complex are slightly larger than represented in Figure 4.5. The p H and E h value measured indicate that it is likely that the copper(I)ammonia complex forms under these conditions. The copper(I)ammonia complex oxidizes rapidly, in the presence o f sufficient oxygen, to form the copper(II)ammonia complex, explaining the colour change o f the solution and the rise i n E h .  Figures 4.7 to 4.12 show that the p H and E h trends described above are applicable to all figures, except for the leaching o f enargite (Figure 4.10). However, this figure too confirms the reactions postulated above. N o decrease in p H value is observed, which indicates that thiosulfate was present throughout the entire experiment. Consequently, no change i n E h value should occur, which is confirmed by the E h curve in Figure 4.10.  78  Chalcocite (Figure 4.8) shows an increase i n potential after about 40 minutes, thereafter stabilizing until at 180 minutes the E h increases again to a higher potential. A s discussed in the literature review, section 2.4.2, half o f the copper i n chalcocite is essentially free and leaches rapidly leaving a covellite matrix behind ( A G o f Equation (IV.6) is -313.1 kJ/mol, using -518.8 0  kJ/mol for S 0 ) : 2  2  3  2Cu S + 8S 0 ~ + 4 N H ; + 0 2  2  2  -> 2CuS + 2 C u ( S 0 ) ' + S 0 5  2  2  3  3  2 _  4  + 4NH + 2H 0 3  2  (IV.6)  Figure 4.8 shows that copper extraction decreases slightly and then rises again. It was observed that the solution colour changed from colourless to light blue after 30 minutes. This indicates that the copper(I)thiosulfate  complex, the copper(I)ammonia and the  copper(II)ammonia  complex could be present under these conditions. According to the E h - p H diagram i n Figure 4.3 the copper(I)thiosulfate species is dominant. However, the E h - p H diagram presented i n Figure 4.4. shows that under these conditions the copper(II)ammonia complex can also be present, which is indicated by the solution colour.  When most o f the copper is present as the copper(I)thiosulfate complex, free thiosulfate has to be present in solution to keep the equilibrium depicted i n Equation (IV.6) to the right. When the thiosulfate  degrades,  the  copper(I)thiosulfate  complex also  decomposes  following  the  equilibrium. When thiosulfate is no longer present, copper dissolves again forming complexes with ammonia, which results in an E h rise (copper(I)ammonia to copper(II)ammonia) and an increase i n copper extraction. Chalcopyrite shows similar behaviour with an increase i n copper extraction after the thiosulfate is no longer present (Figure 4.9).  79  The patterns are different for the leaching o f the copper oxide minerals, cuprite and malachite (Figure 4.11 and 4.12). H i g h copper extractions are achieved at the start o f the experiment, followed by a decrease in extraction and then a steady increase once the E h changes. Cuprite leaches according to Equation (IV.7) ( A G o f Equation (IV.7) is -117.5 kJ/mol, using -518.8 0  kJ/mol for S 0 "): 2  2  3  C u 0 + 6S 0 ~ + 2NH 2  2  2  + 4  -> 2 C u ( S 0 ) 2  3  5  _ 3  + 2NH + H 0 3  (IV.7)  2  The high concentration o f sulfate after 5 minutes and its steady decline does not seem to correlate with the patterns previously observed. B y summing up the concentrations o f thiosulfate, tetrathionate and sulfate, a sulfur balance can be calculated at every moment o f the experiment (Appendix D ) . This balance w i l l be discussed in more detail i n section 4.6. The resulting sulfur balance for this experiment is quite close to the 0.4 M sulfur (0.2 M S 0 ") which was originally 2  2  3  added to the system. A closer examination o f the graph revealed that the results can be explained by sample degradation.  The samples are analyzed as soon as possible by ion chromatograph. Practically, however, the time between taking the sample and analysis by ion chromatograph can be up to 7 to 10 hours for the first few samples (taken at 5, 30, 60 minutes) and less than 7 hours for the last samples. When a lot o f copper is present in solution, as is the case when leaching the copper oxide minerals, and the cupric ion is generated in the presence o f air, the degradation o f thiosulfate is accelerated (see section 2.3.3).  Re-examining Figure 4.11, the leaching o f cuprite results in a high copper extraction. The graph  80  also shows that there is a relatively high tetrathionate concentration and very high sulfate concentration. Both are the products o f thiosulfate degradation. The previous graphs show similar trends, but to a lesser extent because the copper concentrations are lower.  Figure 4.11 shows that after the initial high copper extraction the copper concentration declines and stabilizes after about 240 minutes. The same explanation given for the leaching o f chalcocite applies to cuprite. A t 30 minutes copper is complexed with both thiosulfate and ammonia, because there is not enough thiosulfate in solution to complex copper and give a free thiosulfate concentration o f 5 g/L thiosulfate (see bottom graph, Figure 4.11). This is confirmed by the blue colour o f the solution, indicating the presence o f the copper(II)ammonia complex. From this, it can be concluded that Figure 4.4. seems to represent the actual solution more accurately, instead of Figure 4.3.  If the copper was mainly complexed with ammonia during the entire experiment, the decline in copper extraction would not have been observed, since the ammonia concentration remains fairly constant during the experiment. There w i l l only be a slight ammonia loss to the atmosphere. The same explanation can be applied to malachite (Figure 4.12).  For the enargite mineral sample the conditions were different. After several experiments it was concluded that the 5 g/L addition was too high in order to observe the leaching o f copper with both thiosulfate and ammonia, since there was a deficiency o f thiosulfate. It was decided to lower the mineral addition such that upon 100% copper dissolution, the copper concentration would be 1 g/L. In the remainder o f this chapter this w i l l be referred to as a 1 g/L copper addition.  81  From Figure 4.10 it can be observed that enargite barely dissolves in ammonium thiosulfate under the applied experimental conditions. The effect o f a l o w copper concentration i n solution on thiosulfate degradation can be clearly observed from this experiment. The copper that dissolves is probably converted to cupric i n the presence o f air, but the concentration is too l o w to significantly affect thiosulfate stability. The p H o f the solution changes only slightly and only small amounts o f tetrathionate and sulfate are detected.  4.3.2 Effect of aeration In these experiments no forced aeration o f the solution was applied, however, the vessel was not sealed airtight. It can be assumed that air was able to enter the reactor and the leach slurry, especially since a high stirring speed was applied (see section 3.2).  Comparing Figure 4.8 and Figure 4.13 clearly illustrates the effect o f aeration on thiosulfate degradation. Figure 4.13 shows that the thiosulfate concentration remains high; degradation does not occur appreciably. This is confirmed by the constant p H value. Because the system is not aerated the potential o f the solution is lower. The erratic nature o f the potential indicates that the system has not yet stabilized. The copper extraction shows a steady increase but remains fairly low. Compared to the leaching o f chalcocite in an aerated system the final copper extractions do not differ significantly. It appears that the thiosulfate concentration i n solution does not affect the final copper concentration i n solution. The erratic behaviour o f the sulfur species is explained by sample degradation.  The leaching o f cuprite i n a non-aerated system is shown in Figure 4.14. Thiosulfate remains i n  82  12  600  10 8 X  a  400  6  =  4  200 W  2 0 -50  0  +  +  50  0  100  150  200  250  300  350  400  Time (minutes) •pH  -©-Eh  50  25  40 30  a |  « 20 » at  a a  10 ^  100  200  300  400  Time (minutes) Thiosulfate  - A - Sulfate  - » - Tetrathionate  -©— Copper  Figure 4.13 Chalcocite, 5 g/L copper, 0.20 M (NH ) S 0 , no aeration, 35°C 4  83  2  2  3  12  600 tee e e  10  i-e—e-e—e—ee  8 a  400  6  x s  4  ^ ^ <?> 'fo$^-0-<^^....fo  ^  200 W  2 4 0 -50  0  -I  h  H  50  100  150  0  h  200  250  300  350  400  Time (minutes) -e-pH  V5  -•-Eh  25  100  20  80  15  60  10  40  3  20  CZ3  100  200  300  | s-  a a o  U  400  Time (minutes) -•— Thiosulfate  —•— Tetrathionate  —A— Sulfate  —e— Copper  Figure 4.14 Cuprite, 5 g/L copper, 0.20 M (NH ) S 0 , no aeration, 35°C 4  84  2  2  3  solution. The l o w thiosulfate concentrations in the first half o f the experiment are explained by sample degradation. The later samples are more representative o f the true leaching conditions, because the length o f time between sampling and analyses was less for these samples.  A s for copper extraction, it can be observed that the initial extraction is just as high as with aeration. The major difference is that, without aeration, the copper extraction remains high throughout the course o f the experiment. This indicates that the decline in copper concentration presented i n Figure 4.11 is due to the degradation o f thiosulfate. The colour o f the solution is light blue during the experiment, indicating copper is complexed by both thiosulfate and ammonia. The stable E h and p H o f the solution reflect the minimal thiosulfate degradation. The E h shows a small increase directly after the start o f the experiment, indicating the quick dissolution o f cuprite.  From this study, it can be concluded that both air and cupric ion concentration play an important role in thiosulfate degradation. With copper in solution as cupric, thiosulfate degradation occurs readily and the cupric ion is converted to cuprous (Equation IV.8). The cuprous ions i n solution w i l l be easily re-oxidized by air to cupric ions resulting i n a fast rate o f thiosulfate degradation (Equation IV.9). The overall reaction is represented by Equation (IV. 10). When no aeration is applied, the cuprous ion is not as readily oxidized to the cupric ion, resulting in greater thiosulfate stability. 4Cu(NH ) 3  2 +  + 1 6 S 0 " -> 4 C u ( S 0 ) ' + 1 6 N H + 2 S 0 "  4Cu(S 0 )^ + 0 2  3  5  2  2  3  + 4NH  2  + 4  3  (IV.9)  2  3  3  + 1 2 N H -> 4 C u ( N H ) 3  3  85  4  2 +  6  + 2 H 0 + 12S 0 " 2  2  2  (IV.8)  Overall reaction:  4S 0^ +4NH; +0 2  2  -> 2 S 0 g " + 4 N H + 2 H 0 4  3  2  (IV. 10)  4.3.3 Summary In general it can be observed that the leaching o f copper sulfide minerals i n an ammonium thiosulfate solution with and without aeration, results i n low copper extractions, around 0.5 to 1 g/L copper. Chalcopyrite and enargite seem to be unreactive toward ammonium thiosulfate leaching. The copper oxide minerals show high initial copper extractions, followed by a decline in extraction in the case o f aeration. The copper extractions remain high without aeration. The E h - p H diagram o f the copper-thiosulfate-ammonia-water system presented i n Figure 4.4 seems to give a more accurate representation o f the actual solution conditions, then the E h - p H diagram given i n Figure 4.3.  The following effects were observed regarding thiosulfate degradation: •  A combination o f aeration and copper, whereof a significant amount w i l l be present as the cupric ion i n the presence o f air, results in fast thiosulfate degradation.  •  In the absence o f aeration, even a substantial amount o f copper in solution (~ 1 to 4 g/L) does not constitute a detrimental effect on thiosulfate stability.  •  In the presence o f a small amount o f copper (-0.1 g/L) and aeration, thiosulfate remains longer in solution (leaching o f enargite (Figure 4.10)).  •  The continuing occurrence o f thiosulfate degradation in samples has to be taken into consideration when interpreting the ion chromatograph analyses.  86  4.4 Leaching of copper minerals with gold addition In this set o f experiments, the copper minerals were leached with a 5 ppm gold addition to solution (see Table 3.E). The effect o f the different variables on the leach reaction are only investigated with regard to chalcocite and cuprite.  For the next two sections the following comment regarding the gold extraction curves is highlighted. Three solution samples were analyzed for gold content per experiment (30, 120 and 360 minutes). In the figures lines are drawn through those three sample points. This is to indicate a trend; it is not intended to imply that the gold extraction actually follows this curve. For a few experiments more samples were analyzed, resulting i n a more accurate understanding o f the behaviour o f gold. A l l % copper extraction calculations are based on the head assays o f the copper minerals. G o l d extractions (%) are based on the 5 ppm gold addition.  4.4.1 Baseline experiments Figure 4.15 shows the leaching o f covellite at 1 g/L copper addition. Comparing this figure to Figure 4.7, the leaching o f covellite at 5 g/L without gold, a few differences can be observed. Firstly, the p H during heating (at "negative"time) is lower. This is due to the addition o f the gold to the solution (as chloride complex in 10% hydrochloric acid, see section 3.1.3). The p H o f the solution was immediately raised to p H 10 with concentrated  sodium hydroxide after  the  ammonium thiosulfate addition, to avoid any hydrolysis o f the ammonium thiosulfate at the low p H (see section 2.1.3).  87  12  800  10 8 + X cu  A  a 600  400  f X!  4 + 200  0 -50  0  50  100  150  200  250  300  350  400  Time (minutes) pH  25 100 -o  o 4- 80  fl fl o o 7— 40  « s  « .O ft "  20  100  200  300  a. o U  400  Time (minutes) Thiosulfate  - a - Tetrathionate  Sulfate  Copper  Gold  Figure 4.15 Covellite, 5 ppm gold, 1 g/L copper, 0.20 M (NH ) S 0 , aeration, 35°C 4  88  2  2  3  The p H follows the pattern described previously. Since the thiosulfate is still present at the end o f the experiment, the reduction o f oxygen is the dominant reaction which produces hydroxide ions resulting in a steady increase in p H . The potential o f the solution remains fairly stable, and at the same value as the potential i n Figure 4.7 before the thiosulfate disappeared.  The total amount o f copper extracted is lower in the experiment with 1 g/L addition compared to the 5 g/L addition (0.15 g/L copper vs. 0.85 g/L copper). It seems that the copper extraction is independent o f thiosulfate or ammonia availability. Further, it can be noted that the gold remains in solution during the leaching o f covellite.  The bottom graph illustrates the effect o f lower copper concentration on thiosulfate degradation. Comparing the detected thiosulfate concentration with the concentration plotted in Figure 4.7 it can be concluded that the thiosulfate degrades at a slower rate when less copper is present in solution. Thiosulfate is still detected after 6 hours with lower copper concentration. This effect has already been discussed in the leaching o f enargite.  The leaching o f chalcocite in the presence o f gold shows the same patterns for p H , E h , thiosulfate, tetrathionate and copper compared to the leaching o f chalcocite without gold addition (Figure 4.16). The high sulfate concentration and its subsequent decline can be attributed to sample degradation. The behaviour o f the gold in solution is quite different from the previous discussed experiment (Figure 4.15). The line is dashed to indicate that it is unknown how the gold extraction behaves between the points. More samples were not available for analysis.  89  600  400 an  200  a  2 + 0 -50  0  50  100  150  200  250  300  350  400  Time (minutes) -0—Eh  •pH  25 100 2 &n 80  e e o o  .2-  a 40  * £ « s  20  a * a o  1X1  £3  "3  >- "S <y ©  U 100  200  400  300  Time (minutes) - • — Thiosulfate - * — Copper  - « — Tetrathionate  - A — Sulfate  Gold  Figure 4.16 Chalcocite, 5 ppm gold, 5 g/L copper, 0.20 M (NH ) S 0 , aeration, 35°C 4  90  2  2  3  This experiment was also performed with an addition o f 1 g/L copper (Figure 4.17). It has to be remembered that gold was added as a solution. The fact that the gold extraction at time 120 minutes is higher than at time 30 minutes is therefore difficult to explain. It might be that gold precipitated during heating and that it was resolubilized by the thiosulfate. When thiosulfate is no longer present in solution, gold precipitates, although more gold remains in solution compared to the experiment with 5 g/L copper addition.  Before this phenomena is discussed the results o f the leaching o f the other copper minerals with gold addition are examined. Chalcopyrite shows a stable E h and p H which only rises at the end of the experiment indicating that almost no thiosulfate degradation takes place until the end o f the experiment (Figure 4.18). This is confirmed by the measured thiosulfate concentrations. Copper extraction is low, even lower compared to the experiment with a 5 g/L copper addition.  The gold extraction shows the same pattern as with chalcocite; slightly lower extraction at 30 minutes than at 120 minutes. The gold remains in solution throughout the entire experiment, similar to covellite. Enargite (Figure 4.19) shows almost exactly the same behaviour as the experiment without gold addition (Figure 4.10). The gold also remains i n solution throughout the entire experiment.  The leaching o f cuprite (Figure 4.20) follows the same patterns as those presented i n Figure 4.11. The initial tetrathionate concentration is very high. Coupled with the very l o w initial thiosulfate concentration and its subsequent increase and a high copper concentration, this pattern indicates sample degradation. It can be seen that gold precipitates slightly.  91  600  12 - • — •  10  4- 400 Q a  8  CO  6  a  5  • 0 -  200 W  4 2  0  0 -50  0  50  100  150  200  250  300  350 400  Time (minutes) -©-Eh  •pH  25 100 -o ©  20 S  80  s e o o 60 --g  15  u  V O. S-  li  10  40  si "3  20  100  200  300  a <*> a o u  400  Time (minutes) Thiosulfate  -•>- Tetrathionate  —A- Sulfate  Copper  Gold  Figure 4.17 Chalcocite, 5 ppm gold, 1 g/L copper, 0.20 M (NH ) S 0 , aeration, 35°C 4  92  2  2  3  12  800  10 600 8 s  6  a  f  400  4 200  0 -50  0  50  100  150  200  250  300  350  400  Time (minutes) -*-pH  Eh  25 100 -o  o Ml  80  fl s  o o  60  a  ~ « N  Vi  + 40 SB  a *  ~5 t/5  20  100  200  300  a  U  400  Time (minutes) Thiosulfate  • Tetrathionate  —A— Sulfate  -#— Copper  Gold  Figure 4.18 Chalcopyrite, 5 ppm gold, 1 g/L copper, 0.20 M (NH ) S 0 , aeration, 35°C 4  93  2  2  3  12  800  10 600 8 s  6  f  400  a  4 4 200 2 0  + -50  0  0  + 50  100  150  200  250  300  350  400  Time (minutes) -•-Eh  •pH  25 4- 100 -o o Ml 80 s e a a  20  .8  1 5  <u U  s CZ2  10 4  ©  i  40  3 S  a  .a  ©  60  ~  » s a> o  5  20  a  «  a  o U  0 0  100  200  300  400  Time (minutes) • Thiosulfate  Tetrathionate  —A— Sulfate  Copper  Gold  Figure 4.19 Enargite, 5 ppm gold, 1 g/L copper, 0.20 M (NH ) S 0 , aeration, 35°C 4  94  2  2  3  12  600 —•  10  •—c-o  8  400  x  SO  6  X  c  >-0-0—0——0—  4  200 W  2 0  + -50  0  50  100  0  + 150  200  250  300  350  400  Time (minutes) Eh  •pH  25 100 -o ^  o  20 80  J  a GB S-  1  II 3  5  3  ©O  10  60 4- 40  .3 3 !/3  20 -0100  200  300  400  ~  ~  *  ^  «  3  (U  o  a « a o  U  Time (minutes) -•— Thiosulfate  Tetrathionate  Copper  Gold  Figure 4.20 Cuprite, 5 ppm gold, 5 g/L copper, 0.20 M (NH ) S 0 , aeration, 35°C 4  95  2  2  3  Sample degradation is also visible in the leaching o f malachite (Figure 4.21). The very l o w initial thiosulfate  concentration,  its  increase  later,  the  high tetrathionate,  sulfate  and  copper  concentration indicate sample degradation. G o l d is not affected and remains in solution. It has to be commented that the overall copper extraction for cuprite and malachite is lower with the 1 g/L copper addition than the 5 g/L copper addition.  It can be seen for chalcocite, cuprite and malachite that there is no thiosulfate left i n solution after  180 to 240 minutes, this opposed to the covellite, chalcopyrite and enargite  where  thiosulfate remains present for 6 hours. The experiments with the first three minerals showed gold precipitation, whereas gold remains i n solution during the experiments o f the latter three minerals.  4.4.2 Effect of aeration Experiments without aeration were performed for chalcocite (Figure 4.22) and cuprite (Figure 4.23). The same conditions apply as described i n section 4.3.2. Chalcocite shows a stable p H and potential, which is confirmed by the thiosulfate concentration which barely changes.  The  resulting copper extraction is lower than previous experiments. This is probably due to the l o w potential, which is lower than previous experiments with chalcocite (~50 m V  S H E  vs. 2 5 0 m V  SHE  ).  A s opposed to the experiment with aeration, gold remains in solution during the whole experiment.  96  12  1000 s—Ad (Jim)  10  800  8  1  600  >  6 400  4  200  2 0  0 -50  0  50  100  150  200  250  300  350  400  Time (minutes) •pH  -•-Eh  25 100 -a  o  20 80  w o  a  15  60  c  .2 .2  a  40  « "5  "3  20  a « a. o U  100  200  300  400  Time (minutes) -•— Thiosulfate  •Tetrathionate  —A—Sulfate  —©—Copper  Gold  Figure 4.21 Malachite, 5 ppm gold, 5 g/L copper, 0.20 M (NH ) S 0 , aeration, 35°C 4  97  2  2  3  12  800  10 600 8 6  a  CO  400  >  4 200 2  -50  0  —0  #0#—#0 ^0-0#$0-0~0>^0-O-  1 50  1 100  1 150  1 200  1— 250 300  "0^^  0  350  400  Time (minutes) pH  -•-Eh  25 100 -a  o  20 80 CO  '3  60  e c o o i *  40  si i  15  cu  a  10  CO  i  U  3 o 20 —I 100  a  08  a  U  1 300  200  400  Time (minutes) • Thiosulfate  Tetrathionate  • Copper  Gold  Figure 4.22 Chalcocite, 5 ppm gold, 5 g/L copper, 0.20 M (NH ) S 0 , no aeration, 35°C 4  98  2  2  3  12  600  10 400  8 +  Q  s  6  o.  200 W  4 -0  2 40 -50  0 0  1  1  1  1  1  1  1  1  0  50  100  150  200  250  300  350  0 400  Time (minutes) •pH  25 100 2 80  "3 a u  s  15  60  10  40  a "3  g  Copper exitractic )lubi  50  and  o OX)  20  w  so  5  20 -Hh  0  » — • — • -  100  200  300  400  Time (minutes) Thiosulfate  -a—Tetrathionate  -A—Sulfate  -#— Copper  Gold  Figure 4.23 Cuprite, 5 ppm gold, 1 g/L copper, 0.20 M (NH ) S 0 , no aeration, 35°C 4  99  2  2  3  The sulfur analysis o f Figure 4.23 shows high tetrathionate and sulfate concentrations and a thiosulfate concentration which increases after  180 minutes. This trend indicates  sample  degradation o f the first 5 samples. Copper extraction is high (1 g/L copper addition), however, the copper extraction o f cuprite at 5 g/L with aeration is higher. It was observed that copper and gold extraction declined with time when aeration was applied. Figure 4.23 shows that copper and gold remain in solution without aeration. The above results indicate that the retention o f gold in solution is ensured by the presence o f thiosulfate.  4.4.3 Effect of temperature Two experiments were performed at 20°C to investigate the effect o f temperature on thiosulfate stability. Comparing the figures representing the leaching o f chalcocite and cuprite at 35°C (Figures 4.17 and 4.20 ) with those obtained at 20°C (Figures 4.24. and 4.25), it can be observed that the thiosulfate degradation is slightly delayed in the case o f chalcocite. The effect is barely noticeable for the leaching o f cuprite. The ion chromatograph analysis for cuprite suggest the occurrence o f sample degradation for the initial samples indicated by the high tetrathionate and sulfate concentrations.  The recorded p H and potential show that the patterns for both minerals are similar to those obtained during leaching at 35°C, only slightly delayed. The behaviour o f gold and the copper extraction follow the same pattern at 20 and 35° for chalcocite. For cuprite, the results cannot be compared since the mineral feed was not the same (5 g/L copper vs. 1 g/L copper).  100  12  600  10 8 x  400  =  6 4  200 W  2 0  0  + -50  0  50  100  150  200  250  300  350  400  Time (minutes) -^Eh  •pH  100 -o o 6D  80  100  300  200  60  O O ~ ~ |  |  40  i  |  S  ©  20  §"  a  »  U  400  Time (minutes) • Thiosulfate  —•— Tetrathionate  Sulfate  Copper  Gold  Figure 4.24 Chalcocite, 5 ppm gold, 5 g/L copper, 0.20 M (NH ) S 0 , aeration, 20°C 4  101  2  2  3  12  600  10 400  8+  s  6  s  4  200 W  2 0  0  + -50  0  50  100  150  200  250  300  350  400  Time (minutes) pH  -©-Eh  25 100 73 o + 80 CO CU  w <U C  4- 60  i-  40  CO  !§ C73  20  100  200  300  1 * a e .2 .2 w cs  "S  a « a o  U  400  Time (minutes) Thiosulfate  -HB— Tetrathionate  Sulfate  Copper  Gold  Figure 4.25 Cuprite, 5 ppm gold, 1 g/L copper, 0.20 M (NH ) S 0 , aeration, 20°C 4  102  2  2  3  4.4.4 Effect of reagent addition A s discussed i n chapter 2, section 2.3.3, several papers attribute a positive effect to the addition o f sulfite and sulfate to an ammonium thiosulfate leach solution. The additions are supposed to stabilize the thiosulfate i n solution. The effects o f the additions were investigated with the copper mineral chalcocite and the results are presented i n Figure 4.26 (addition o f sulfite) and 4.27 (addition o f sulfate).  Comparing these results to Figure 4.17, it can be observed that the additions did not significantly improve the stability o f thiosulfate. The E h increases are slightly delayed i n time, indicating that thiosulfate remains slightly longer i n solution. Copper and gold behaviour are similar i n all three experiments. The only visible difference is the increase in sulfate concentration when sulfite and sulfate are added. This endorses the statement postulated in section 2.3.3 that sulfite acts as a sacrificial reagent, oxidizing to sulfate. The effect o f sulfate cannot be explained with the data available.  L o w copper extractions were observed during the baseline experiments. It was thought that this might be due to a low ammonia concentration. One experiment was performed with an additional 0.4 M ammonia resulting i n a total ammonia concentration o f 0.8 M (Figure 4.28). It can be seen that the extra ammonia had no beneficial effect on the copper extraction, and decreased the thiosulfate stability.  103  800  600 s  X c  400  >  4- 200  -50  0  50  100  150  200  250  0  +  •+-  300  350  400  Time (minutes) -0—Eh  •pH  25 100 -o o  20  .a *o «u c in  U  15 10  80  li  60  a a ~© ~ o  2 -a fl s  40  cu ©  "3 C/2  5 4  a « a  20  o  u 100  200  300  400  Time (minutes) -•— Thiosulfate  Tetrathionate  -A—Sulfate  -#-Copper  Gold  Figure 4.26 Chalcocite, 5 ppm gold, 5 g/L copper, 0.20 M (NH ) S 0 , 0.05 M S 0 , aeration, 35°C 2  4  104  2  2  3  3  12  800  10 600 8  fad £  6  400  |  4 200 2 0  0 -50  0  50  100  150  200  250  300  350  400  Time (minutes) -0—Eh  pH  25 100 -a ©  -J w  20 480 15  60  o .3  ss a o o «  2 -a  10  CZ3  200  300  3  4>  O  a « a  20  100  *  » "s  40  o U  400  Time (minutes) • Thiosulfate  -m- Tetrathionate  Sulfate  Copper  Gold  Figure 4.27 Chalcocite, 5 ppm gold, 5 g/L copper, 0.20 M (NH ) S 0 , 0.05 M SO 4 1 aeration, 35°C 4  105  2  2  3  12  600  10 8 a  400  w X  >  6  E 200 W  4  0 -50  0  50  100  150  200  250  300  350  400  Time (minutes) pH  Eh  25 100 -o ©  20 80  .a  1 5  a  c c  '  60  O CS  10  4- 40  "5  02  o o ~ ~  5  20  » -s >- ~x o ©  a * a  o u  0 100  200  300  400  Time (minutes) • Thiosulfate  Tetrathionate  Sulfate  Copper  Gold  Figure 4.28 Chalcocite, 5 ppm gold, 5 g/L copper, 0.20 M (NH ) S 0 , 0.4 M NH , aeration, 35°C 4  106  2  2  3  3  4.4.5 Gold precipitation A t the instant that thiosulfate is no longer present in solution, the following reaction is driven to the left and gold precipitates, unless it is complexed by something else i n solution: 2 A u + 4S 03~ + y 0 + H 0 <-> A u ( S 0 ) " + 2 0 H " 2  2  2  2  3  2  (IV.ll)  Thermodynamically, ammonia should be able to complex the gold, but that does not seem to occur. The discussion i n section 4.1 already indicated that it seems that Figure 4.1 represents more accurately the actual solution conditions instead o f Figure 4.2. However, it can be noticed that in several cases the gold extraction does not completely go to zero, although there is no thiosulfate present i n solution. Several explanations can be given for this phenomena: •  A m m o n i a complexes gold to a certain extent. The literature reports  10 to 20% gold  extraction i n the presence o f ammonia and cupric ions [Zipperian et al, 1988; Cao et al, 1992]. •  G o l d precipitates as colloidal gold, and is therefore detected in solution.  •  There  is still a small amount  o f thiosulfate  left in solution not detected  by ion  chromatographic analysis, providing enough background for gold to remain i n solution to a certain extent  The presence o f a background concentration o f thiosulfate is confirmed by ion chromatograph measurements (Figure 4.29). Typical dilution factor for all experiments was 1000 times. This dilution factor ensured that the samples could be measured within the calibration range (2-20 ppm, see Appendix B ) . A disadvantage was that concentrations below 1-2 ppm were not detected. This implies that, with a 1000 times dilution factor, concentrations below 1-2 g/L  107  thiosulfate could still be present in solution. Figure 4.29 shows the measurement of a sample at 1000, 500 and 50 times dilution. At 1000 and 500 times no thiosulfate is detected, whereas a small thiosulfate concentration is detected at 50 times dilution. This indictates that about 0.1 g/L of thiosulfate is in solution, providing enough background for gold to remain in solution. More research has to be performed to determine the nature of the gold precipitates.  lOOOx dilution  0  0.5  1  1.5  2 Minutes  2.5  3  500x dilution  AU  0.5  1  1.5  2 Minutes  2.5  3  3.5  4  3  3.5  4  50x dilution  Mi  0  0J  1  1.5  2 Minutes  2.5  Figure 4.29 Effect of dilution factor on detected thiosulfate concentration (retention time thiosulfate is 1.9 minutes)  4.4.6 Summary It can be concluded that in the presence of chalcocite, cuprite and malachite gold precipitates out of solution, while gold remains in solution in the presence of covellite, chalcopyrite and enargite. This seems to correlate with the fast degradation of thiosulfate observed during the leaching of  108  the first three minerals. H i g h copper concentrations were achieved during the leaching o f those three minerals, thus increasing the presence o f cupric ion i n solution, which enhances the degradation o f thiosulfate. The copper extractions achieved during the baseline experiments o f this series and the series discussed i n section 4.3 are summarized i n Table 4 . A . It is evident from Table 4 . A that the copper extractions did not improve by lowering the mineral feed addition and thus increasing lixiviant availability.  Table 4.A Copper extractions (%) of copper minerals at different times during baseline experiments. Mineral  % Copper leached at time (in minutes) Feed  Leaching of copper  Feed  Leaching of copper  addition  minerals  addition  minerals + gold addition  30  120  360  30  120  360  Covellite  5 g/L C u  5.8  6.7  16.5  1 g/L C u  8.8  14.3  15.1  Chalcocite  5 g/L C u  26.0  17.8  24.0  5 g/L C u  29.5  26.3  18.3  Chalcocite  -  -  -  -  1 g/L C u  19.7  35.7  46.4  Chalcopyrite  5 g/L C u  2.7  5.1  8.9  1 g/L C u  3.78  9.9  14.5  Enargite  1 g/L C u  3.8  6.3  9.2  1 g/L C u  3.19  5.7  9.2  Cuprite  5 g/L C u ,  79.1  36.2  22.8  5 g/L C u ,  80.9  32.2  24.4  Malachite  5 g/L C u  64.8  62.5  44.5  1 g/L C u  87.0  86.1  83.5  N o beneficial effect o f lower temperature, sulfite and sulfate addition on thiosulfate stability was observed. Lower copper extractions with 1 g/L copper addition were observed compared to the copper extractions achieved with 5 g/L copper addition. This cannot be attributed to an ammonia deficiency since an extra addition o f ammonia did not improve copper extractions. It may be 109  that the copper minerals are not provided with sufficient oxidizing agents and/or the leach temperature is too low, resulting in poor copper extractions.  Regarding the use o f the E h - p H diagrams given i n Figure 4.1 to Figure 4.4 to explain the experimental results presented in section 4.3 and 4.4. the following can be concluded: •  For the gold-thiosulfate-ammonia-water system, Figure 4.1 (using the free energy value o f 518.8 kJ/mol for thiosulfate) seems to give a more accurate representation o f the actual leaching conditions  •  For the copper-thiosulfate-ammonia-water system, Figure 4.4 (using the free energy value o f -532.2 kJ/mol for thiosulfate) seems to explain the experimental results better.  More fundamental research is necessary to investigate this phenomenon.  4.5 Leaching of copper-gold samples The experimental results o f the copper-gold samples are discussed per copper-gold sample, to directly observe the influence o f the different variables (Table 3.F). In this series baseline experiments were performed and the effects o f aeration and cupric addition were studied. The gold extraction calculations are based on calculated head, the copper extraction calculations on solids head assay.  The cyanide consumption and copper and gold extraction o f the copper-gold materials were determined by a 24 hour cyanide leach at a constant 1 g/L cyanide level. The experimental conditions are described in Appendix E . The results are summarized in Table 4 . B .  110  Table 4.B Results of 24 hour cyanide leach of the copper-gold samples Copper-gold samples  Gold  Copper  Cyanide consumption  extraction (%)  extraction (%)  (kg/t ore)  Overall Lobo Composite  81.4  41.4  4.38  Guanaco Composite  84.3  70.4  6.77  M 4 0 Pyrite Feed  94.7  34.9  6.40  M 1 0 Pyrite Concentrate  47.5  28.3  9.91  M 1 0 Pyrite Concentrate*  81.2  50.3  n.a.  *Data obtained from Newcrest M i n i n g Ltd. [Oretest, 1997]  Cyanide leaching o f all copper-gold samples resulted i n high cyanide consumption, and consequently a high reagent cost must be anticipated when treating these samples with cyanide. It can be seen that a substantial amount o f the copper i n all samples is cyanide soluble. Two test numbers are listed for the M 1 0 Pyrite Concentrate. The first was obtained during the leach described i n Appendix E , the second was provided by Newcrest M i n i n g L t d ; a 24 hour cyanidation at an initial cyanide concentration o f 1.25 g/L cyanide, maintained at 0.75 g/L cyanide. The first numberis believed to be low because o f the following reaction:  C u ( C N ) " -> C u ( C N ) " + CIST 2  (IV. 12)  which indicated the presence o f free cyanide, while in reality it was complexed with copper.  4.5.1 Overall Lobo Composite Figure 4.30 gives the results o f the leaching o f the Overall Lobo Composite with a 1 g/L cupric ion addition in an aerated system. Similar patterns as described i n the previous sections are 111  observed for the p H and E h . The p H initially increases, and starts decreasing around 180 minutes which coincides with an increase in E h . The initial E h , during heating o f the solution without any reagents present, shows a sharp decrease. Apparently, a component is present i n the ore which strongly consumes oxygen.  The line representing the thiosulfate concentration is dashed, the 180 minutes sample was not analyzed for sulfur species. The line implies that the thiosulfate concentration is about 5 g/L after 180 minutes and zero around 240 minutes. However, the p H and E h patterns indicate that most thiosulfate was gone after 180 minutes. The erratic pattern o f tetrathionate is attributed to sample degradation.  G o l d extractions are stable around 40%. Initially copper is high which can be explained by the l g / L cupric addition. The decline in copper extraction with time can be attributed to the disappearance o f thiosulfate, a pattern that has been observed previously. The experiment was conducted with coarse material ( P  80  = 253 um) and the aeration o f the system did not work  properly (the sparger plugged). The experiment was duplicated with reground material ( P  80  =141  um) and the results are presented i n Figure 4.31.  Most patterns are comparable to those o f Figure 4.30. The thiosulfate concentration goes to zero after 180 minutes and the tetrathionate, sulfate and copper values are resembling those o f Figure 4.30. T w o differences can be observed. The E h value shows a steady increase during the second half o f the experiment, and the gold extraction is lower during the second experiment.  112  12  600  10  + 400  8  a  200  a  6  0 4  ffi  U  c«  -200  2 0  -400 -50  0  50  100  150  200  250  300  350  400  Time (minutes) -«— p H  Eh  25  80  20 15 cu  60  c • -.  40  CU  a  £  10  g CU  CO  "8  t/5  20  100  200  300  2  400  Time (minutes) • - - Thiosulfate  - a — Tetrathionate  -®— Copper  —  - A — Sulfate  Gold  Figure 4.30 Overall Lobo Composite, 0.20 M (NH ) S 0 ,1 g/L cupric addition, aeration, 35°C 4  113  2  2  3  12  600  10 8 X  a  400  6  X s  4  200 W  2 0 -50  0  50  100  150  200  250  300  350  400  Time (minutes) •pH  Eh  80  25 20  60  £ Is ©  15 4 40  C u si  2 es  20  t/5  -H-  100  200  300  400  Time (minutes) -*>— Thiosulfate  •Tetrathionate  -A—Sulfate  -e-Copper  Gold  Figure 4.31 Overall Lobo Composite, 0.20 M (NH ) S 0 ,1 g/L cupric addition, aeration, 35°C, reground. 4  114  2  2  3  12  600  10  -•  0  8  400  Q  s  6  >  4  200 W  a,  »# •—  2 +  0-  0  0 -50  0  50  100  150  200  250  300  350  400  Time (minutes)  25  80  20  Vi  a u  £ "a  03  60  £ o  15 40  2  10 + OS 20  5  w=^k  ^ A = -  2  a ^ = ^  0 0  100  200  300  400  Time (minutes) • Thiosulfate  -m- Tetrathionate  - A — Sulfate  Copper  Gold  Figure 4.32 Overall Lobo Composite, 0.20 M (NH ) S 0 ,1 g/L cupric addition, no aeration, 35°C 4  115  2  2  3  12  600  10  e •  400  8  Ld X  a c.  6  200  ^  4 2 0  H -50  0  50  100  h  H  h  -200  150 200 250 300 350 400  Time (minutes) •pH  -0—Eh  25  80  20 s 1X1  °3 a  a "3  60  15 40 es 10 + 20  5 0  ®  fjh  ill"*  100  200  300  400  Time (minutes) -•— Thiosulfate  —•— Tetrathionate  —A— Sulfate  —•— Copper  Gold  Figure 4.33 Overall Lobo Composite, 0.20 M (NH ) S 0 , no cupric addition, aeration, 35°C 4  116  2  2  3  12  600  10  > •  —  a —  m  m  i  a  8  400 K  S  >  6  =  0-—000 0 0Q0 0 0-0-0-0 0, ...^  4  200 W  2 0 -50  1  1  1  1  1  1  1  1  0  50  100  150  200  250  300  350  0 400  Time (minutes) -©-Eh  •pH  25  ^ -J  J  "3  2  80  0  60  .2  15 40  CU  a CO  x  2  10  .3  s  20  <Z3  100  200  300  400  Time (minutes) -•— Thiosulfate  Tetrathionate  —A— Sulfate  —®— Copper  Gold  Figure 4.34 Overall Lobo Composite, 0.20 M (NH ) S 0 , no cupric addition, aeration, 35°C, duplicate 4  117  2  2  3  The experiment with 1 g/L cupric addition without aeration is shown i n Figure 4.32. A s shown with the previous series o f experiments without aeration, thiosulfate remains i n solution during the whole experiment, copper is present at a constant value. G o l d slowly dissolves and the final extraction is similar to the one achieved i n the first experiment. The irregular pattern o f the E h cannot be explained; after the increase it stabilizes again at its previous value.  Figure 4.33 shows the effect o f aeration without cupric addition. A higher gold extraction was achieved and the thiosulfate concentration remains very high. This is confirmed by the stable p H and E h values. A duplicate o f this experiment was performed to ensure that sufficient aeration had been applied (Figure 4.34). The results are very similar, except for gold extraction, which indicates that the reproducibility o f the experiment is good. When no cupric is added to the solution, the resulting copper extraction is very low; the composite contains mainly chalcopyrite (see section 3.1.2) which shows comparable low copper extraction (Figure 4.18).  Summary The leaching o f the Overall Lobo Composite results i n moderate gold extractions. The absence o f both aeration and cupric ion addition have a positive effect on the thiosulfate stability and result in the case o f no cupric ion addition (with aeration) i n a gold extraction o f 66%. Copper i n this composite seems to be unreactive towards ammonium thiosulfate leaching. The beneficial effect o f cupric ions addition on the gold extraction is not demonstrated i n these experiments. The copper and gold extractions o f the experiments are summarized in Table 4.C.  118  From Table 4.C, it can be concluded that most o f the copper(II)sulfate initially added to the leach solution, precipitates. Compared to the cyanidation results (81.4 % A u , 41.4 % C u after 24 hours) it can be concluded that comparable gold extractions are achieved i n the experiment with aeration and no cupric ion addition. Copper extractions with the ammonium thiosulfate leach are very l o w compared to the cyanidation.  Table 4.C Gold and copper extraction (%) of Overall Lobo Composite at different times. Experimental  %Gold extracted at time  %Copper extracted at  (minutes):  time (minutes):  Conditions 30  120  360  Total*  30  120  360  Aeration  42.0  41.6  41.2  62.1  36.9  20.4  11.6  Aeration, duplicate  16.2  15.9  15.5  44.8  35.0  13.8  11.3  N o aeration  b.d.  b.d.  35.8  n.a.  48.0  48.6  48.7  Aeration, no cupric  34.6  67.9  66.0  73.5  9.3  11.7  12.6  Aeration, no cupric,  b.d.  34.2  66.5  74.3  7.9  10.4  12.6  duplicate * Total represents total gold extraction, including wash water b.d.= below detection limit n.a.= not available  The thiosulfate consumption for the 24 hours experiment was around 15.7 kg/tonne ore, which results i n a reagent cost o f US$2.04/tonne ore (US$0.13/kg). The concentrate contains 2.11 g/tonne gold which yields US$20.3/tonne (at US$300/oz gold). Thus, the reagent cost constitutes about 10% o f the total metal value. Cyanidation would result i n a reagent cost o f US$7.88 (cyanide consumption o f 4.38 kg/tonne ore). Thiosulfate leaching compares favourable to cyanidation for this experiment, however both leaches are not very economical. 119  4.5.2 Guanaco Composite The leach behaviour o f the Guanaco Composite presented i n Figures 4.35 to 4.37, resembles closely the patterns observed during the leaching o f the Overall Lobo Composite. Aeration o f the solution with cupric ion addition resulted in the disappearance o f thiosulfate which is confirmed by the p H and E h patterns (Figure 4.35). The gold extraction declines to zero when the thiosulfate is no longer present. The initial copper extraction is slightly higher than the final extraction.  When cupric is added and no aeration is applied, the E h o f the system is very erratic and shows an upward trend (Figure 4.36). The lowering o f the thiosulfate concentration is not reflected i n the p H curve, and thus indicates sample degradation. Initially, there is not a lot o f copper i n solution, and thus sample degradation should not significantly occur. The samples taken after 120 minutes contain more copper and therefore sample degradation is catalyzed.  Both gold and copper follow an upward trend. Especially copper shows a large increase. It seems that due to the l o w E h value at the start o f the experiment, the added cupric ions precipitated. The solution colour changed around 240 minute from colourless to very light blue. This indicates that initially the copper could be complexed as the thiosulfate and copper(I)ammonia complex, which are both colourless. The oxidation o f the copper(I)ammonia to the copper(II)ammonia complex progresses slow in the absence o f an abundant amount o f air, but explains the steady rise i n E h and copper extraction.  120  12  600  10  —  8  +#0  •  —  •  6  0  —  e  e  400  #0  + 200  6  a  >E  4 -200  2 0  -400 -50  0  50  100  150  200  250  300  350  400  Time (minutes) •pH  -•-Eh  25  80  20 4  60  6X1  o  °3  40 s. 5 8  « 2s  £  10  20  (Z3  2  5 0  100  200  400  300  Time (minutes) Thiosulfate  • Tetrathionate  —A— Sulfate  Copper  Gold  Figure 4.35 Guanaco Composite, 0.20 M (NH ) S 0 ,1 g/L cupric addition, aeration, 35°C 4  2  121  2  3  Figure 4.36 Guanaco Composite, 0.20 M (NH ) S 0 ,1 g/L cupric addition, no aeration, 35°C 4  122  2  2  3  600  400 UJ  0-0  X  0—00  200  0  |  -200 -50  0  50  100  150  200  250  300  350  400  Time (minutes) •pH  -0—Eh  25  80  20 4-  cu  60  ©  15  w  40 2  CU  a : 3 «  £ la  CU  io 20 £  5  100  200  300  400  Time (minutes) • Thiosulfate  Tetrathionate  - A — Sulfate  Copper  Gold  Figure 4.37 Guanaco Composite, 0.20 M (NH ) S 0 , no cupric addition, aeration, 35°C 4  2  123  2  3  Aeration and no cupric ion addition resulted in thiosulfate degradation, as opposed to the result o f the leaching o f the Overall Lobo Composite (Figure 4.37). The copper concentration is higher in this solution explaining the occurrence o f thiosulfate degradation. However, compared to Figure 4.35, leaching o f the composite i n the presence o f air and with cupric ion addition, the thiosulfate degradation is slightly delayed. This implies that the addition o f cupric ions results in faster thiosulfate degradation. Copper extraction remains fairly constant and gold extraction decreases at the end o f the experiment.  When comparing the three figures it can be concluded that the same gold extractions are achieved every experiment, however, at different times. When aeration is applied, initial gold extractions o f around 40% are achieved, which is the same as the final gold extraction o f the experiment without aeration. However, the experiment without aeration implies that higher gold extractions can be achieved when the experiment is allowed to last longer, since thiosulfate is still present i n solution.  One experiment was performed with Thiogold™ grade ammonium thiosulfate (see section 3.1.3). The results are presented i n Figure 4.38. The high initial gold extraction is due to an analytical error. The conclusion that can be drawn from this experiment is that the purity o f the thiosulfate solution has no effect on the gold extraction.  124  12  600 •*—*v  10  400 8  u. B  6i  a  200  f  4 2  N  0  -200 -50  0  50  100  150  200  250  300  350  400  Time (minutes) •pH  •-Eh  25 ^  20  !& u  15  t  10  120 100 80  Is  o  60  a  40  -4 -"^  tZ5  20 5 0 0  100  200  300  400  Time (minutes) — • — Thiosulfate — • — Copper  — • — Tetrathionate  —A—  Sulfate  Gold  Figure 4.38 Guanaco Composite, 0.20 M (NH ) S 0 (Thiogold™ grade), 1 g/L cupric addition, aeration, 35°C 4  125  2  2  3  Summary Similar gold extractions are achieved under all conditions, however, at different times. Initial gold extractions are higher i n an aerated system, while final extractions are higher i n a nonaerated system. Copper extractions are l o w when thiosulfate is not present during the entire experiment. In the presence o f thiosulfate higher copper extractions are achieved. L i k e the Overall Lobo Composite there is no obvious beneficial effect observed on gold extraction with cupric ion addition. Table 4.D lists the gold and copper extractions o f the experiments.  Table 4.D Gold and copper extraction (%) of Guanaco Composite at different times. Experimental Conditions  %Gold extracted at time  %Copper extracted at  (minutes):  time (minutes):  30  120  360  Total*  30  120  360  Aeration  33.6  33.0  b.d.  0.0  47.0  34.9  32.5  N o aeration  b.d.  b.d.  39.6  44.0  9.0  28.7  59.2  Aeration, no cupric  51.6  50.6  24.6  30.7  30.4  31.5  24.7  Aeration, Thiogold™  112.3  36.8  35.7  36.7  46.5  37.3  36.9  *Total represents total gold extraction, including wash water b.d. = below detection limit  Compared to the cyanidation results (84.3% A u , 70.4 % C u at 24 hours cyanidation) it can be seen that the extractions achieved with ammonium thiosulfate are low. For this composite, the experiment with no forced aeration should be extended to a 24 hours leach, enabling a more accurate comparison with the cyanidation test results.  126  4.5.3 M40 Pyrite Feed A s can be seen in Figures 4.39 to 4.42, the experimental results o f the M 4 0 Pyrite Feed generally follow the same patterns as observed previously. One additional experiment was performed i n which no cupric was added to the solution and no aeration was applied (Figure 4.42). Both the experiments shown i n Figure 4.41 and 4.42 were conducted at 0.24 M ammonium thiosulfate.  Comparison o f the experiments with aeration and cupric addition (Figure 4.39) and aeration without cupric addition (Figure 4.41) shows that the cupric addition accelerates the thiosulfate degradation. This is visible from the p H , E h and thiosulfate curves. The experiment without cupric addition shows a steady increase in copper extraction.  Comparison o f the experiments presented in Figure 4.40 and 4.42, both without aeration and one with cupric addition and one without, show that similar low gold extractions are achieved. The cupric addition does not seem to influence the leaching o f gold. H i g h tetrathionate concentrations were detected (Figure 4.40) and it is difficult to predict whether this occurred i n the leach solution or was due to sample degradation.  The low copper extraction i n Figure 4.42  demonstrates that the copper in the M 4 0 Pyrite Feed, which is mainly chalcopyrite, is unreactive toward an ammonium thiosulfate leach. This corresponds to the results o f the leaching o f chalcopyrite (Figure 4.18).  127  Time (minutes) ••— Thiosulfate  - • — Tetrathionate  — A - Sulfate  —@— Copper  Gold  Figure 4.39 M40 Pyrite Feed, 0.20 M (NH ) S 0 ,1 g/L cupric addition, aeration, 35°C 4  2  2  3  128  12  600  10 8  400  Q  x  CO  6  a  4  200 W  0 -50  0  50  100  150  200  250  300  350  400  Time (minutes) •pH  -•-Eh  25  60  20  VI  '3  40 "a  15  cs u  <u  a VI  U  10 20 *  ~3  in  5 0 100  200  300  400  Time (minutes) -•— Thiosulfate  -m- Tetrathionate  —A— Sulfate  Copper  —  Gold  Figure 4.40 M40 Pyrite Feed, 0.20 M (NH ) S 0 ,1 g/L cupric addition, no aeration, 35°C 4  2  2  3  129  12  600  10 8 w  a  400  6 0—o—  4  Q  s ;> s,  200 W  2 0  0 -50  0  50  100  150 200  250  300  350 400  Time (minutes) pH  -©-Eh  60  40  ^ _© *-+-» u 05 U  20 "oS <u  100  200  300  400  Time (minutes) Thiosulfate  Tetrathionate  - A — Sulfate  Copper  Gold  Figure 4.41 M40 Pyrite Feed, 0.20 M (NH ) S Q , no cupric addition, aeration, 35°C 4  2  2  130  3  12  600  10  y—m—9 •  —  8  a  4- 400  6  s  CO  ' r  4  200 W  2 0  —I  -50  0  1  1  50  100  1  1  150 200  1  1  250  300  1—  0  350 400  Time (minutes) -•— p H  -0—  Eh  30  60  25 20  40 a' .©  CO  cu cu  a CO  £"U3 t/3  cu  15  S3 ii 20 « J-  10  CU  5 0 100  200  300  400  Time (minutes) Thiosulfate  Tetrathionate  —A— Sulfate  Copper  Gold  Figure 4.42 M40 Pyrite Feed, 0.20 M (NH ) S 0 , no cupric addition, no aeration, 35°C 4  2  2  131  3  Summary Highest gold extractions were achieved i n an aerated system around 120 to 180 minutes. The gold precipitates when the thiosulfate starts degrading significantly. L o w gold extractions were achieved i n the non-aerated system, however, thiosulfate is still present. This implies that higher gold extraction could be achieved when the residence time is prolonged. Cupric addition did not show a pronounced effect on gold extractions, i n both an aerated and non-aerated system. The copper in the concentrate seems unreactive towards an ammonium thiosulfate leach. Copper and gold extractions for all experiments are summarized i n Table 4.E.  Table 4.E Gold and copper extraction (%) of M40 Pyrite Feed at different times. Experimental  %Gold extracted at time  %Copper extracted at  (minutes):  time (minutes):  Conditions 30  120  360  Total*  30  120  360  Aeration  11.1  32.8  10.6  14.1  30.4  29.3  32.0  N o aeration  10.5  10.3  19.9  21.3  30.2  29.1  28.6  Aeration, no cupric  13.3  26.1  12.7  14.5  11.5  13.9  21.4  N o aeration, no cupric  10.1  9.9  19.3  22.7  7.2  7.1  7.4  T o t a l represents total gold extraction, inc uding wash water  The results o f the ammonium thiosulfate leach are very l o w compared to the cyanidation test results (94.7% A u , 34.9% C u with 24 hours leach). It is uncertain why this concentrate shows such l o w overall extractions.  132  4.5.4 M10 Pyrite Concentrate The experiments involving the M 1 0 Pyrite Concentrate are presented i n Figures 4.43 to 4.47. From Figure 4.43, it can be observed that gold extractions o f 80% were achieved after 120 minutes. The same experiment without cupric addition resulted i n 50% gold extraction, indicating a positive effect o f the cupric ion addition (Figure 4.45). The final  copper  concentrations were l o w i n both cases and very similar. This indicates that a portion o f the cupric ions added to the solution precipitated during the experiment.  The experiment performed without aeration resulted i n a gold extraction o f 50% (Figure 4.43). It was decided to duplicate this experiment and extend the leaching time to 24 hours, to observe the effect on thiosulfate concentration and gold extraction. The results are presented i n Figure 4.45. A gold extraction o f 80% was achieved and thiosulfate is still present i n solution. A plausible explanation for the increase i n thiosulfate concentration from 5 to 500 minutes is sample degradation; the samples were analyzed after about 30 hours. The copper extraction remained stable during the entire 24 hours.  Two observations can be made from the series o f experiments presented i n section 4.15 to 4.46: •  G o l d precipitates when no thiosulfate is present i n solution.  •  The copper extractions remain low, perhaps because o f the presence o f insufficient complexing agents ( M 1 0 Pyrite Concentrate contains 16,600 ppm copper).  133  Time (minutes) pH  — E h  Time (minutes) —•—Thiosulfate  -H—Tetrathionate  —A—Sulfate  —•— Copper  ——Gold  Figure 4.43 M10 Pyrite Concentrate, 0.20 M (NH ) S 0 ,1 g/L cupric addition, aeration, 35°C 4  134  2  2  3  12  600  10 8  400 0  x  6 4  200 W  0 -50  0  50  100  150  200  250  300  350  400  Time (minutes) •pH  w <U  Eh  25  100  20  80  15  60  10  40  |  u  o. S-  "3  20  100  200  300  400  Time (minutes) Thiosulfate  Tetrathionate  Sulfate  - • - Copper  Gold  Figure 4.44 M10 Pyrite Concentrate, 0.20 M (NH ) S 0 ,1 g/L cupric addition, no aeration, 35°C 4  135  2  2  3  Vi  25  100  20  80  ^  15  60  -2  CU « s-  'cu CU  O. Vi  U  10 40  SB  s  t/2  20  500  1000  1 cu  1500  Time (minutes) -•— Thiosulfate  Tetrathionate  —±— Sulfate  —®— Copper  Gold  Figure 4.45 M10 Pyrite Concentrate, 0.20 M (NH ) S 0 ,1 g/L cupric addition, no aeration, 35°C, 24 hours 4  136  2  2  3  Time (minutes) pH  —•— E h  Time (minutes) —•—Thiosulfate  —•— Tetrathionate  —&— Sulfate  -#—Copper  —Gold  Figure 4.46 M10 Pyrite Concentrate, 0.20 M (NH ) S 0 , no cupric addition, aeration, 35°C 4  137  2  2  3  To investigate this, an experiment was performed with an initial ammonium thiosulfate concentration o f 0.36 M (Figure 4.47). Every hour, at 60, 120, 180, 240 and 300 minutes, an additional 0.072 M o f ammonium thiosulfate was added to the solution, resulting in a total ammonium thiosulfate concentration o f 0.72 M . Every ammonium thiosulfate addition gave rise to a slight decrease i n p H and E h . A s can be seen from the graph, 80% gold extraction was also achieved i n this experiment, and the final thiosulfate concentration is almost zero. The copper extraction is not affected by the additional thiosulfate and ammonia.  The M 1 0 Pyrite Concentrate contains a considerable amount o f copper (16,600 ppm), which means that 0.72 M ammonium thiosulfate does not provide sufficient thiosulfate to complex all the copper, but it does supply sufficient ammonia to complex all the copper. The copper is mainly present as chalcopyrite, which does not leach very well i n ammonium thiosulfate solutions, as observed in section 4.3 and 4.4.  Summary The combination o f aeration and cupric addition results i n high gold extraction (80%) early in the experiment (around 120 minutes), but is followed by gold precipitation due to the degradation o f thiosulfate. Similar gold extractions were obtained i n a 24 hour leach without aeration and a leach with a total addition o f 0.72 M thiosulfate. Cupric addition resulted in higher initial gold extractions and total copper extraction is not significantly influenced by any o f the variables tested: cupric addition, aeration, leaching time, reagent concentration. Table 4.F summarizes the copper and gold extractions achieved.  138  0  100  200  300  400  Time (minutes) —•— Thiosulfate  -m- Tetrathionate  —A— Sulfate  —•— Copper  —— G o l d  Figure 4.47 M10 Pyrite Concentrate, 0.72 M (NH ) S 0 , 1 g/L cupric addition, aeration, 35°C 4  139  2  2  3  Table 4.F Gold and copper extraction (%) of M10 Pyrite Concentrate at different times. Experimental  %Gold extracted at time  %Copper extracted at  (minutes):  time (minutes):  Conditions 30  120  360  Total*  30  120  360  Aeration  39.2  87.1  5.9  7.2  21.2  20.5  27.4  N o aeration  14.1  29.7  50.0  55.6  20.1  20.8  21.0  N o aeration, 24 hours  13.2  29.8  76.1**  83.3  20.3  21.6  21.7**  Aeration, no cupric  39.8  49.3  3.99  5.8  11.9  12.9  23.6  0.72 M ( N H ) S 0  49.1  74.3  76.1  84.4  23.6  25.5  29.2  4  2  3  * Total represents total gold extraction, including wash water ** at time 1440 minutes (24 hours)  The gold extractions achieved during the 24 hour leach compare very well to the cyanidation test results (81.4% A u , 50.3% C u at 24 hours). It should be noted that the resulting copper extractions are very low. The thiosulfate consumption for the 24 hours experiment was around 40 kg/tonne ore, which results in a reagent cost o f US$5.2/tonne ore (US$0.13/kg). The concentrate contains 45.5 g/tonne gold which yields US$436/tonne (at US$300/oz gold). Thus, the reagent cost constitutes about 1.2% o f the total metal value. The cyanide consumption was not available.  4.6 Sulfur balance A closer examination o f the curves representing the sulfur species in the figures discussed above, reveals that not all the sulfur that was added to the solution (0.2 M ammonium thiosulfate) is  140  accounted for by the sulfur species detected by the ion chromatograph. A good example is given in Figure 4.47, the leaching o f the M 1 0 Pyrite Concentrate with a total addition o f 0.72 M ammonium thiosulfate. After 360 minutes a negligible amount o f thiosulfate and tetrathionate is detected and about 5 g/L sulfate. Since the experiment indicates that thiosulfate degrades, it is expected that a large amount o f tetrathionate and sulfate are present. I f all the thiosulfate converted to sulfate, 138 g/L o f sulfate should be present. However, the ion chromatography analyses did not detect high concentrations o f either sulfate or tetrathionate.  In Appendix D the sulfur balances for all experiments are presented, which clearly indicates that the sulfur balance o f most experiments is not correct. To investigate where the sulfur reported to, several leach solutions were analyzed for sulfate using barium chloride (Appendix F). The experiments indicated that the sulfate levels were not significantly high compared to the ion chromatograph analyses. Next, several copper mineral residues were analysed for sulfur species (S  total  , S°, and S ") to determine whether the sulfur reported to the residues. The analyses showed 2  that the sulfur levels did not increase significantly i n the residues.  Several solutions were also analyzed for total sulfur and these results indicated that the sulfur was still present in solution. The ion chromatograph detected along with the thiosulfate, tetrathionate and sulfate species, two other sulfur species: sulfite (S0 ") and trithionate (S 0 "). 2  3  2  3  6  The sulfite peak appears just before the sulfate peak on the ion chromatogram (Figure 4.48) and the trithionate appears between the thiosulfate and tetrathionate peak (Figure 4.49). It should be noted that the identification o f these peaks was based on previous experience and not on an actual identification.  141  t = 30 minutes  I  I  |l/ I  X I I I I  I  I I  I I ! I I I I  1  I  I  I I I I 1 I I I I  4  t = 180 minutes  1  I I I I  I  I I I  5  0  11/  I  I II  I  I  1  I  I  I I I  2  I  I I I  3  Minutes  I  I I I  4  !I  AJV. I M  I  5  1  I  I I I I  7  8  Minutes  t = 60 minutes  t = 240 minutes  uS  I I I  llf I  3  I I II  1  i  I I I I  2  I  II  I I  3  I  M  I  I I  4  I I I I  5  I  I  I I II  6  I I I I  7  I  111 i/i 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 pi 1 1 1 1  0  8  1  2  3  5  t = 120 minutes  5r  6  7  8  Minutes  Minutes  t = 360 minutes  uS  uS  M  0  4  11/  I'  1  I  I I  II  2  I  I I  I  I I I I  3  II  4  I I I  I  I I I I  5  I  I  6  I M  I  J i i i 1/ h i i i I i i i i I i i i i I i i i i ! i i i i I i i i i I i i i i  I I I I  7  0 1  Minutes  2  3  4  5  6  7  Minutes  Figure 4.48 Ion Chromatograms of the AS4A column representing the sulfate concentration at different times during the leaching of lg/L Chalcocite with 5 ppm gold addition 142  8  t = 30 minutes  0.35f  t = 180 minutes  0.35f  AUI  AU  I II I  0  II 'iI i'  0.5  I  I I  I  1  I ! I I  i  I I  iiiIi  1.5 2  I !  1  I  I I I  2.5 3  I  I I ! I  0  Ii[I) I  3.5 4  4.5  I I I I  0  II  I I i  0.5  I II  I i  1  II  I I I  Minutes  II  I I i ! I I I ! i i !'  I I I  3.5 4  4.5  t = 240 minutes  ATJ  I  I I  II i  I I  0.5  II  I I !  1  I  I I  I I  II  1.5 2  I I I  I  I I I I  II  2.5 3  I I I  I  I I I I  II  I  I I I  3.5 4  11111 I I 1 1 1 1 1 I I 1 1 11 11 1 1 1 1 1  4.5  0  0.5  1  1.5 2  Minutes  1111 11 11 11 11 i  2.5 3  11  11 i  3.5 4  4.5  Minutes  t = 120 minutes  0.35f  t = 360 minutes  0.35f  AUl  AU  fl 1 1 1 1 1 1 1 i  0  I I I I  2.5 3  0.3ft  AU|  0  i  II  Minutes  t = 60 minutes  0.351  II i  1.5 2  0.5  11II  1  i 1 1 1 1 1 1 1 i i 1 1 i II  1.5 2  fl  i 1 1 1 1 1 1 1 1 1 1 111 i 1 1  2.5 3  3.5 4  I  4.5  •0  Minutes  I I I  IM  0.5  I  I  II  1  I I i  II  I I I  II  1.5 2  I  I I  II  I I I  II  2.5 3  I I I  I  I I I I  I  I  I I I  3.5 4  Minutes  Figure 4.49 Ion Chromatograms of the Omnipax column showing (from left to right) thiosulfate, trithionate and tetrationate at different times during the leaching of lg/L chalcocite with 5 ppm gold addition 143  I  4.5  A t first hand it seems unlikely that these ions are present at high concentrations, since their peak area is very small (Figure 4.49). However, a study by Steudel [1986] indicated that the response factor o f trithionate is about 18 times lower than tetrathionate. This means that when trithionate and tetrathionate have the same area, the trithionate concentration is 18 times larger. U s i n g this knowledge, the sulfur balance was recalculated for several experiments (Appendix D ) . This resulted i n improved sulfur balances.  This indicates that the decomposition o f tetrathionate to higher and lower polythionates, as presented in Equation (IV. 13), takes place (section 2.3.3): 2S OJ:- + 3 0 H - ^ 4  f S 0 " +S 0^ +f H 0 2  2  3  2  (IV. 13)  A s mentioned i n chapter 2 (section 2.3.3), trithionate is a relatively kinetically stable sulfur species under the leaching conditions applied. It decomposes quickly under harsh conditions: boiling and pH>13. This indicates that i f trithionate formed, it is very likely that it remains for a while i n solution.  Two comments have to be made: 1. The determination o f the response factor in the study quoted, was performed using different equipment. This study was only used to show that the concentration o f trithionate might be higher than would be expected from the size o f the peak area. 2. The trithionate peak areas were determined with the tetrathionate calibration curve. The trithionate peak areas were all below 1 m g / L , which is below the m i n i m u m detection limit, resulting in inaccurate area measurements.  144  Summary The sulfur balances for the leach experiments did not account for all the sulfur added to the solutions. Further research revealed that it is very likely that most o f the sulfur ends up as the trithionate species i n solution.  145  5 GENERAL SUMMARY AND CONCLUSIONS This study investigated the application o f ammonium thiosulfate for the treatment o f copper-gold ores. Leaching studies were conducted with copper minerals, copper minerals with gold addition and copper-gold samples o f different copper and gold grade. O n the basis o f the experimental results, the following conclusions were drawn:  Leaching of copper minerals in ammonium thiosulfate solution  General observations •  The copper sulfide minerals covellite and chalcocite do not leach appreciably in an ammonium thiosulfate solution. Chalcopyrite and enargite seem to be unreactive toward an ammonium thiosulfate leach. Copper extraction seems to be independent o f the availability of complexing agents; leaching o f less material i n the presence o f the same amount o f lixiviant did not result in higher copper extractions.  •  The copper oxide minerals cuprite and malachite, showed fast dissolution resulting i n high initial copper extractions (around 80%). The final copper extractions were dependent on lixiviant availability; copper extractions decreased with declining thiosulfate concentrations, high copper extractions were achieved when thiosulfate remained i n solution. The exact role o f ammonia in copper solubilization is not clear. It can be further remarked that 80% copper extraction was achieved independent o f the amount o f material subjected to the leach; both mineral additions o f 5 g/L copper and 1 g/L copper resulted i n 80% copper extraction.  146  •  The amount o f copper solubilized was i n all cases, except enargite and chalcopyrite (both at a mineral feed o f 1 g/L copper), enough to catalyze the degradation o f thiosulfate in solution. After about 180 to 240 minutes thiosulfate was no longer present in solution. Because enargite and chalcopyrite barely leaches i n an ammonia thiosulfate solution, little copper was solubilized which resulted in increased thiosulfate stability; the thiosulfate was still present after 360 minutes.  It can therefore be concluded that the leaching o f the copper minerals i n the presence o f air resulted in the solubilization o f copper as cupric, which catalyzed thiosulfate degradation. Thiosulfate degradation reduces overall copper extraction.  The following observations were made when soluble gold was added to the leach solution in the presence o f copper minerals: •  Copper concentrations o f 0.3 g/L or higher result i n enhanced rates o f thiosulfate degradation. The consequent reduction o f thiosulfate concentration causes gold to precipitate.  •  In the presence o f l o w copper concentrations (around 0.1 g/L Cu) thiosulfate degradation is reduced and gold remains in solution.  Further study should be undertaken to determine exactly at which copper concentration a negative effect on thiosulfate stability is observed.  The above shows that gold solubilization is dependent on the presence o f thiosulfate. A m m o n i a is known as a complexing agent for gold, but does not appear to play a role under the experimental conditions investigated.  147  Effect of aeration A s stated above, an aerated solution in combination with as little as 0.3 g/L copper i n solution causes fast thiosulfate degradation, resulting i n gold precipitation. Thiosulfate degradation is less in the absence o f aeration, even in the presence o f high copper concentrations. Cuprous ions are easily regenerated into cupric in the presence o f air, resulting i n fast thiosulfate degradation. The oxidation o f cuprous to cupric is a lot slower i n the absence o f air, resulting i n greater thiosulfate stability, and thus gold remains in solution.  Effect of temperature A leach temperature o f 20°C instead o f 35°C did not result i n a measurable effect on thiosulfate degradation, copper extraction and gold solubilization.  Effect of reagent addition In this study, the addition o f sulfite and sulfate was not observed to affect thiosulfate stability, as suggested in the literature. A n increase i n ammonia concentration did not result i n higher copper extractions, and the thiosulfate stability was negatively affected by the increase i n ammonia concentration.  Leaching of copper-gold samples in ammonium thiosulfate solution  General observations These experiments confirmed the above conclusion: when the thiosulfate concentration declines, gold extraction decreases.  Thiosulfate degradation depends  148  on aeration and cupric ion  concentration. G o l d extractions ranged from low to high depending on the experimental conditions. The copper-gold samples tested did not show appreciable co-extraction o f copper during an ammonium thiosulfate leach. Compared to the cyanidation test results, the ammonium thiosulfate treatment o f the Overall Lobo Composite and M 1 0 Pyrite Concentrate showed that similar gold extractions were achieved at lower reagent cost.  Effect of aeration H i g h initial gold extractions are achieved in a well aerated system. Without aeration, but not i n an airtight system, gold dissolves a lot slower and low gold extractions are achieved after 6 hours o f leaching. A n extended residence time o f 24 hours for the M l 0 Pyrite Concentrate resulted in gold extractions o f 83%.  It can be concluded that a balance has to be found between providing enough oxidant for fast gold dissolution, and minimizing the amount o f oxidant i n the presence o f cupric ions to prevent excessive thiosulfate degradation. The alternative is slow gold dissolution and low thiosulfate consumption in a non-aerated system with extended residence time.  Effect of cupric ion addition A  high cupric ion concentration in combination with aeration results in fast  thiosulfate  degradation. The leaching o f the M 1 0 Pyrite Concentrate and M 4 0 Pyrite Feed showed higher initial gold extraction with the addition o f cupric ions in combination with aeration. This indicates a catalytic effect o f cupric ions on the dissolution o f gold i n an ammonium thiosulfate solution, which is described in the literature (see section 2.3.3). In the absence o f aeration, no  149  beneficial effect was observed from the addition o f cupric ions. For the two composites (Overall Lobo and Guanaco), higher gold extractions were observed without the addition o f cupric ions. This effect was not further investigated.  Analytical •  Part o f the sulfur could not be accounted for in the leach solutions. Initial research indicates that a large portion o f the sulfur initially present does not end up as sulfate,  the  thermodynamically predicted final degradation product o f thiosulfate, but as trithionate  (s o -). 2  3  •  6  It is very difficult to prevent early thiosulfate degradation in the samples, resulting i n an inaccurate portrayal o f the actual leaching conditions.  Eh-pH diagrams •  Different A G values are used for the thiosulfate species (-518.8 kJ/mol and -532.2 kJ/mol), 0  which result in a totally different representation o f the thermodynamic equilibria i n the ammonium thiosulfate system.  150  6 RECOMMENDATIONS This work investigated the application o f ammonium thiosulfate for the treatment o f copper-gold ores. From the previous chapter it can be concluded that the application o f ammonium thiosulfate for the treatment o f copper-gold ores offers definite opportunities; however, more research should be conducted in the following areas:  Process development •  H i g h gold extractions were achieved with a 24 hour leach i n a non-aerated system. Additional experiments should be conducted to optimize the gold extraction and at the same time minimize thiosulfate consumption. Factors to be studied to optimize the leaching are thiosulfate concentration and the effect o f the solution p H .  •  H i g h initial gold extractions were achieved i n an aerated system. It might be valuable to investigate the possibility o f separating solution and solids at that instant without causing thiosulfate degradation and thus gold precipitation. Further, the amount o f aeration should be studied, to optimize gold extraction and minimize the thiosulfate degradation.  •  L o w copper extractions were achieved with the experimental conditions specified. More research should be performed to determine why low copper extractions were achieved. The copper species i n solution should be more closely examined using U V spectrophotometry. This may reveal more information regarding the mechanism o f copper complexation i n solution; whether mainly copper thiosulfate or copper ammonia complexes form and which copper ammonia complexes form: the copper(I)ammonia or the copper(II)ammonia complex.  •  I f trithionate is the predominant degradation product o f thiosulfate i n this system, it should be researched how trithionate can be prevented from building up i n the system. Furthermore, 151  trithionate may have a significant effect on adsorption processes for gold and copper recovery and these effects must be quantified.  Fundamental Research •  Fundamental research should be performed to provide more insight i n the leaching chemistry. Because the chemistry o f the ammonia thiosulfate system is not yet fully understood and has proven to be very complex, it is difficult to optimize and control the leaching conditions, especially on an industrial scale.  •  Efforts should be undertaken to obtain correct values for the free energies o f the species present i n ammonium thiosulfate solutions. This w i l l enable a more accurate representation o f the thermodynamic equilibria.  •  More research is required to determine the mechanism o f gold precipitation when thiosulfate is no longer present in solution. This is coupled with the question o f whether or not ammonia plays a role in gold solubilization in this system.  Analytical •  Difficulties were encountered i n preventing thiosulfate degradation from occurring i n the solution samples for ion chromatography analyses. Sample treatment, with for example ion exchange resins, might offer a solution to this problem.  •  Moreover, it was only possible to detect the free species and not the complexed species by ion chromatography during this research. B y developing methods for sample treatment, it might be possible to quantify both the free and complexed species.  152  N o t all sulfur species could be quantified by ion chromatography, e.g. sulfite and trithionate. Since it appears that trithionate plays an important role in the solution, methods should be developed for quantifying these sulfur species by ion chromatography. Furthermore, the response factor for trithionate has to be determined for the ion chromatograph set up used in this research. U V spectrophotometry should be applied to qualitatively determine which copper ammonia complexes are present i n solution.  153  7 REFERENCES A B B R U Z Z E S E , C , Fornari, P., Massidda, R., Veglio, Ubaldini, S., Thiosulfate leaching for gold hydrometallurgy, Hydrometallurgy, 39 (1995), pp. 265-276  A T L U R I , V . P . , Recovery o f gold and silver from ammoniacal thiosulfate solutions containing copper by resin ion exchange method, Masters o f Science thesis, University o f Arizona, 1987  A V R A A M I D E S , J., Prospects for alternative leaching systems for gold " a review", i n Carbonin-pulp technology for the extraction o f gold, Australas. Inst. M i n . Metall. Melbourne, Murdoch, Australia, July 1982, pp. 369-391  A W A D A L L A , F.T., Ritcey, G . 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(1973): 889-893  C A O , C , H u , J., Gong, Q., Leaching o f gold by low concentration thiosulfate solution, Transactions o f NonFerrous Society o f China, v.2 (4), 21-25 N o v . 1992  D U B Y , P., The thermodynamic properties o f aqueous inorganic copper systems, International Copper Research Association, Inc., U S A , 1977  F I S H E R , J.I., Halpern, J., Corrosion o f copper-gold alloys by oxygen-containing solutions o f ammonia and ammonium salts, Trans. Electrochem. Soc. 103, 282-286(1956)  F O R W A R D , F . A . , M a c k i w , V . N . , Chemistry o f the A m m o n i a Pressure Process for Leaching N i , C u , and C o from Sherritt Gordon Sulfide Concentrates, Journal o f Metals, 7, (1955), pp. 457-463  G A L L A G H E R , N . P . , Interaction o f gold cyanide, thiocyanate, thiosulfate, and thiourea complexes with carbon matrices, Masters o f Science Thesis, University o f Nevada, Reno, 1987  G O N G , Q., H u , J., Treatment o f copper-containing gold sulfide concentrate using thiosulfate solution, Journal o f engineering chemistry and metallurgy, 1990 (11)2, 145-151  H E M M A T I , M . , Hendrix, J . L . , Nelson, J . H . , Milosavljevic, E . B . , Study o f the thiosulfate leaching o f gold from carbonaceous ore and the quantitative determination o f thiosulfate i n the leached solution, Extraction Metallurgy '89, (Institution o f M i n i n g and Metallurgy: London, 1989), pp. 665-678  H A L P E R N , J., Kinetics o f the dissolution o f copper in aqueous ammonia, J. Electrochem. Soc. 100, 421 (1953)  H A N , K . , Meng, X . , A m m o n i a extraction o f gold and silver from ores and other materials, U S Patent 5,114,687, M a y 19, 1992  H O N E A , R. M . , Polished section examination o f P2509 Guanaco Composite, February 6, 1998  155  J E F F R E Y , G . H . , Bassett, J., Mendham, J., Derrney, R . C . , V o g e F s textbook o f quantitative chemical analysis, 5 Edition, Longman Scientific and Technical, Great Britain, 1989, p.490 th  J I A N G , T., Chen, J., X u , S., Electrochemistry and mechanism o f leaching gold with ammoniacal thiosulfate, X V I I I Int. Mineral processing congres, 1993, Australia, V . 5., The australasian institute o f mining & metallurgy, pp. 1141-1146  K E R L E Y , B . J . , Recovery o f precious metals from difficult ores, U S Patent 4,269,622, M a y 26, 1981  K E R L E Y , B . J . , Recovery o f precious metals from difficult ores, U S Patent 4,369,061, Jan. 18, 1983  K I R K O T H M E R , Encyclopedia o f Chemical Technology, 3 N e w York, V o l . 22, 1983, pp. 974-988  rd  Edition, John W i l e y and Sons,  K U H N , M . C . , Arbiter, N . , K l i n g , H . , Anaconda's Arbiter process for copper, C I M Bulletin, (1) pp. 62-71, February, 1974  LAKEFIELD RESEARCH chromatography, 1990  LIMITED,  method  no. 9.1.9,  Sulfate  determination by ion  L A K E F I E L D R E S E A R C H LIMITED, determination by ion chromatography  method  no. 9.1.14, Thiosulfate and tetrathionate  L A K E F I E L D R E S E A R C H L I M I T E D , Mineralogical Examination o f Overall Lobo Composite, M a y 5, 1997  L A N G H A N S JR., J.W., L e i , K . P . V . L e i and Carnahan, T . G . , Copper-catalyzed thiosulfate leaching o f low-grade gold ores. Hydrometallurgy 29 (1992) 191-203  L I , J., M i l l e r , J.D., Wan, R . Y . , L e Vier, M . , The ammoniacal thiosulfate system for precious metal recovery, in: Proceedings o f the 19 International Mineral Processing Congress, San Francisco, 1995, Chapter 7 th  156  L I , J., M i l l e r , J.D., Wan, R . Y . , Important solution chemistry factors that influence the coppercatalyzed ammonium thiosulfate leaching o f gold, presented at the 125 S M E Annual meeting Phoenix, Arizona, M a r c h 11-14, 1996. th  L I D D E L L , D . M . , Handbook o f non-ferrous metallurgy, vol.2; Recovery o f metals, N e w Y o r k , M c G r a w - H i l l , 1945, p. 1052  M A R C H B A N K , A . R . , Thomas, K . G . , Dreisinger, D . , Fleming, C . , G o l d recovery from refractory carbonaceous ores by pressure oxidation and thiosulfate leaching, U S Patent # 5,536,297, July 16, 1996  M A R S D E N , J. House, I., The chemistry o f gold extraction, Ellis Horwood, N e w York, 1992  M C K E E , D . , Lulham, J.P., Separation process, Interational Publication # W O 91/11539, August 8, 1991  M E N G , X . , Han, K . N . , The dissolution behaviour o f gold i n ammoniacal solutions, i n : Hydrometallurgy, Fundamentals, Technology and innovation, S M E , 1993, pp. 205-221  M I C H A E L I S , H . V O N . , G o l d processing update, The prospects for alternative leach reagents, can precious metals producers get along without cyanide? Engineering & M i n i n g Journal , June 1987, pp. 42-47  O R E T E S T , Metallurgical testwork on Telfer M 4 0 and M l 0 Flotation samples, August 18, 1997  P E R E Z , A . E . , Galaviz, H . D . , Method for recovery o f precious metals from difficult ores with copper-ammonium thiosulfate., U S Patent # 4,654,078, March 31, 1987.  P E T E R S , E . , Oxidation state diagrams, Short course notes  P R Y O R , W . A . , The kinetics o f the disproportionation o & sodium thiosulfate to sodium sulfide and sulfate, Analytical Chemistry, The American Chemical Society, 82, 1960, pp.4794-4797  S I L L E N , L . G . , Stability constants o f metal-ion complexes, Supplement publication No.25, The Chemical Society, Burlington House, London 157  no.l.,  Special  S K O O G , D . A . , West, D . M . , Fundamentals o f analytical chemistry, 3 and Winston, N Y , 1976, pp.363-366  edition, Holt, Reinhart  S K O O G , D A . , Leary, J. J., Principals o f instrumental analysis, 4 edition, 1992, pp.579-667 th  S M I T H , C . W . , Hitchen, A . , Aqueous solution chemistry o f polythionates and thiosulfate: a review o f formation and degradation pathways, C A N M E T , 1976  S T E U D E L , R., Holdt, G . , Ion-pair chromatographic separation o f polythionates S 0 " with up to thirteen sulfur atoms, Journal o f Chromatography, 361, 1986, pp. 379 - 384 2  n  6  T O Z A W A , K . , Inui, Y . , Umetsu, Y . , Dissolution o f gold in ammoniacal thiosulfate solution, T M S paper selection A81-25, 1981  T U R N B U L L , A . G . , Wadsley, M . W . , The C S I R O Thermochemistry system (version V ) , Institute of Energy and Earth Resources, Division o f Mineral Chemistry, Australia, 1986  W A N , R . . Y . , L a V i e r , K . M a r c , Clayton, Richard, B . , ( Newmont G o l d Co.) Hydrometallurgical process for the recovery o f precious metal values from precious metal ores with thiosulfate lixiviant. United States Patent # 5,354,359, O c t . l l , 1994  W A N , R . Y . , Importance o f solution chemistry for thiosulfate leaching o f gold, unpublished, 1997.  W A N G , X . , Thermodynamic Equilibrium calculations on A u / A g lixiviant systems relevant to gold extraction from complex ores, in: Electrochemistry i n Mineral and Metal Processing III, St.Louis, Missouri, 17- 22 M a y 1992, The electrochemical society, Inc. ( U S A ) pp. 452-477, 1992  W A S S E R L A U F , M . , Dutrizac, J.E., The chemistry, generation and treatment o f thiosalts in milling effluents - a non-critical summary o f Canmet investigations 1976-1982, Canmet report 82-4E, March 1982  W E A S T , R . C . , C R C Handbook o f Chemistry and Physics, C R C Press, Ohio, 1975  158  W I L L I A M S , R . D . , Light, S.D., Copper concentrate dissolution chemistry and kinetics i n an ammonia-oxygen environment in: Chapman, T . W . , Tavlarides, L . L . , Hubred, G . L . , Wellek, R . M . (Eds.), Fundamental aspects o f hydrometallurgical processes, A I C h E Symposium Series, 1978, No.173, V o l . 7 4 , pp.21-27  W O O L F , A . A . , Anhydrous sodium thiosulfate as a primary iodometric standard, Analytical Chemistry, The American Chemical Society, 1982, 54, pp. 2134-2136  Z I P P E R I A N , D . , Raghavan, S., G o l d and silver extraction by ammoniacal thiosulfate leaching from.a rhyolite ore. Hydrometallurgy, 19 (1988) 361-375  159  Appendix A: Screen analysis of copper-gold samples  The particle size distributions o f the Overall Lobo Composite and Guanaco Composite were determined by taking a 500 gram representative sample o f the pulverized ore, and subjecting it to screening. Results are presented i n Table A . I .  Table A.I Screen analysis of copper-gold samples Size  Lobo  Guanaco  M40 Pyrite  M10 Pyrite  (um)  Composite  Composite  Feed  Concentrate  Mass(g)  Mass %  Mass(g)  Mass %  Mass(g)  Mass %  Mass(g)  Mass %  300  0.38  0.08  0.83  0.17  27.87  1.41  -  -  212  3.16  0.64  19.04  3.85  54.09  2.73  8.30  0.42  150  73.38  14.85  99.42  20.10  74.86  3.78  17.20  0.87  106  142.02  28.74  81.27  16.43  94.37  4.76  35.90  1.81  75  61.62  12.47  75.83  15.33  150.94  7.61  87.50  4.41  45  113.6  22.99  86.2  17.43  330.52  16.67  256.00  12.90  -45  100.02  20.24  132.02  26.69  1250.13  63.05  1578.85  79.59  Totals  494.18  100.0  494.61  100.0  1982.78  100.0  1983.75  100.0  P80(um)  141  160  75  46  The particle size distributions o f the M 4 0 Pyrite Feed and M 1 0 Pyrite Concentrate were obtained from Newcrest M i n i n g L t d (Table A.I). The P  8 0  o f the four materials was determined by plotting  the cumulative weight percentage passing against the screen sizes (see Figure A . 1 to A . 4 )  160  100 .  ^3  _  90 80 3?  70  § 60  '!>  i? 50 Q-  S E  40  O 30 .  20 10 . 0 ^  Size, microns  ^10  P  8 0  Figure A.2 Particle size distribution of Guanaco Composite  161  1000  mn  _  90 80 70 'assir  ? fin  Um  ^ 40 E  30  O  20 10 0  P  10  8 0  1 ()0 _.  Size, m i c r o n s  1000  Figure A.3 Particle size distribution of M40 Pyrite Feed  100 *  —  90 80 70  if 60 in  'to  S.  50  ^ 40  i O  30 20 10 0  10  80  100  Size, microns  Figure A.4 Particle size distribution of M10 Pyrite Concentrate  162  1000  Appendix B: Ion Chromatography  Chromatography is an important method that permits the separation and analysis o f ionic species. In this research ion exchange chromatography was used. A small amount o f liquid sample is injected into a moving stream o f liquid (termed the mobile phase), that passes through an immiscible stationary phase (resins with ion exchange sites), which is fixed i n a column or on a solid surface. Separation is based on ions partitioning into the ion exchange phase to varying degrees.  This process is illustrated in Figure D . l . A t time t , the sample, contained in the mobile phase, is 0  introduced at the head o f the column. The components o f the sample distribute themselves between the two phases. Introduction o f additional mobile phase (the eluent) forces the mobile phase containing a part o f the sample down the column, where further partition between the mobile phase and fresh portions o f the stationary phase occurs (time t,). Continuous flow o f the mobile phase carries analyte molecules down the column i n a continuous series o f transfers between the mobile and the stationary phases.  Because movement o f sample components can only occur i n the mobile phase, the average rate at which a species migrates depends upon the fraction o f time it spends i n that phase. For components that are strongly retained by the stationary phase (compound B i n Figure D . l ) this fraction is small, whereas the fraction is large for components that are weakly retained by the stationary phase (component A ) . Ideally, the components are separated into bands (time t ), 2  which can be detected at the end o f the column (time t and /,) [Skoog, D . A . , 1992]. }  163  Sample  Ca)  Mobile phase  •  7T J  m Packed column  41  HI  4  JfllS -Detector  21 '2 Time  Figure D . l (a) Diagram showing the separation of a mixture of components A and B by column elution chromatography, (b) The output of the signal detector at the various stages of elution shown in (a) [Skoog, D.A., 1992]  If the signal o f the detector is plotted as a function o f time, a series o f peaks is obtained, as shown i n the lower part o f Figure D . l . This plot is called a chromatogram, and is a tool for both qualitative and quantitative analysis. The positions o f peaks on the time axis may serve to identify the components o f the sample; the areas under the peaks provide a quantitative measure o f the amount o f each component. This is achieved by calibrating using standards that contain known components o f known concentrations.  164  Ion chromatography in this work  Two methods were used i n this work to measure the different sulfur species i n solution. The IonPac A S 4 A - S C analytical column and IonPac A G 4 A - S C guard column combined with a conductivity detector and anion micro membrane suppressor were used for sulfate analyses. A n example o f a chromatogram is given i n Figure D.2.  Figure D.2 Chromatogram of the A S 4 A column: sulfate (retention time of 4.7 minutes)  A n OmniPac P A X - 1 0 0 analytical and Omni Pac- P A X - 1 0 0 guard column with a U V detector was used for thiosulfate and tetrathionate determination. A n example chromatogram is shown i n Figure D.3. Detailed procedures for both columns are described on the next pages.  165  0.14 0.12  0.1  Figure D.3 Chromatogram of the Omnipax column: thiosulfate (retention time of 2 minutes) and tetrathionate (retention time of 5.1 minutes)  166  Determination of thiosulfate and tetrathionate (S 0 and S 0 ) 2  3  4  6  by Ion Chromatography [Lakefield Research Ltd.]  Objective  To determine thiosulfate and tetrathionate i n solutions by ion chromatography  Range: 1.0 m g / L to 20 m g / L (maximum detection limit), i f there are no interferences and no dilution factor. Concentrations higher than 20 m g / L should be diluted to below maximum detection limit. Interferences:  H i g h thiocyanate interferes with the thiosulfate peak  Sample Requirements:  The sample is not preserved, and should be analysed the same day as the experiment took place.  Apparatus: 1. L i q u i d chromatograph with U V detector 2.  OmniPac Pax-100 analytical column  3. OmniPac Pax-100 guard column 4.  1 m L syringe for sample injection  5. 0.45 u m syringe filter 167  Reagents: 1.  Sodium Perchlorate eluent (0.038 M ) with 10 % methanol: dissolve 10.675 g N a C 1 0 . H 0 4  2  ( H P L C grade) i n D I water, add 200 m L methanol, and dilute to 2 L with D I water using a 2 L beaker. Filter the solution through a 0.45um filter. 2.  Calibration standards (prepare fresh every day)  Thiosulfate Prepare 1000 mg/L S 0 standard by dissolving 1.4117 g o f Na S 0 (anhydrous sodium 2  3  2  2  3  thiosulfate) i n D I water and dilute to 1000 m L in a volumetric flask Tetrathionate Prepare  1000 mg/L stock solution by dissolving 0.1368 g N a S 0 . 2 H 0 ( s o d i u m 2  4  6  2  tetrathionate dihydrate) in water and dilute to 100 m L i n a volumetric flask  Prepare from these 1000 mg/L stock solutions calibration standards containing 2, 10, 15 and 20 ppm thiosulfate and tetrathionate  Procedure: 1. Set up the ion chromatograph according to the instructions in the ion chromatograph manual 2.  Guard column -OmniPac P A X - 100, analytical column - OmniPac P A X - 1 0 0  3.  Eluant: 0.038 M N a C 1 0 in 10% methanol 4  4. F l o w rate: 1.1 m L / m i n 5. sample loop volume - 50 u L 6. Detector: U V , measure at 215 nm 7. Inject one o f the calibration standards and run the chromatogram o f thiosulfate  168  and  tetrathionate to determine the retention times. Standardize the instrument on these conditions 8. Calibrate the machine with the 2, 10, 15 and 20 ppm calibration standards and check the calibration plot 9. If the calibration is successful, determine the chromatograms o f the samples (use syringe filters for injection)  Ion Retention Time* S 0 2  3  S 0  6  4  1.8 - 2.2 minutes 4 - 6 minutes  *Note: difference in retention time depends on condition o f the column  Quality Control: The following samples must be analyzed after an analytical run i.  Replicate  calibration standard:  choose  the  calibration standard  close  to  prevalent  concentration ii.  Replicate sample: select random a sample and perform a second analytical run  Note:  the anhydrous sodium thiosulfate can be prepared from hydrated sodium thiosulfate ( N a S 0 . 5 H 0 ) v i a a procedure described by W o o l f [1982]. It can also be purchased i n a 2  2  3  2  granular form.  169  Determination of sulfate by Ion Chromatography [Lakefield Research Ltd., 1990]  Objective:  To determine sulfate i n solutions by ion chromatography  Range: 0. 5 m g / L to 20 m g / L (maximum detection limit), i f there are no interferences and no dilution factor. Concentrations higher than 20 m g / L should be diluted to below maximum detection limit. Interferences:  There can be overestimation o f sulfate values with l o w p H samples  Sample Requirements:  The sample is not preserved, and should be analysed the same day as the experiment took place.  Apparatus: 1. L i q u i d chromatograph with conductivity detector and anion micro membrane suppressor (AMMS) 2.  IonPac A S 4 A - S C analytical column  3.  IonPac A G 4 A - S C guard column  4.  1 m L syringe for sample injection  5. 0.45 u m syringe filter  170  Reagents: 1. Sodium carbonate - sodium bicarbonate eluent with 10 % methanol: prepare 1.8 m M N a C 0 2  + 1.7 m M N a H C 0  3  by weighing 0.382 g o f N a C 0 2  3  anhydrous and 0.286 g o f N a H C 0  3  3  anhydrous into D I water. A d d 200 m L methanol, and dilute to 2 L with DI water using a 2 L beaker. Filter the solution through a 0.45um filter. 2.  Sulfuric acid (25 m N H S 0 ) : 100 m l I N sulfuric acid into 2 L beaker, dilute to 2 L . 2  4  3. Calibration standards Sulfate Prepare 1000 m g / L S 0 standard by dissolving 1.4796 g o f Na S0 (anhydrous sodium 4  2  4  sulfate) i n D I water and dilute to 1000 m L in a volumetric flask. Prepare from this 1000 mg/L stock solution calibration standards containing 2, 10, 15 and 20 ppm sulfate  Procedure: 1.  Set up the ion chromatograph according to the instructions i n the ion chromatograph manual  2. Guard column IonPac A G 4 A - S C , analytical column - IonPac A S 4 A - S C 3. Eluant: 1.8 m M N a C 0 + 1.7 m M N a H C 0 2  3  3  4. F l o w rate: 2.0 m L / m i n 5. Sample loop volume - 50 u L 6.  Suppressor: A n i o n M i c r o Membrane Suppressor ( A M M S ) Regenerant: 25 m N H S 0 at flow rate 3-5 m L / m i n 2  4  7. Detector: conductivity 8. Inject one o f the calibration standards and run the chromatogram o f sulfate to determine the retention time. Standardize the instrument on these conditions  171  9. Calibrate the machine with the 2, 10, 15 and 20 ppm calibration standards and check the calibration plot 10. I f the calibration is successful, determine the chromatograms o f the samples (use syringe filters for injection). If the sample contains thiosulfate, the peak w i l l appear around 15-20 minutes on the chromatogram, depending on the condition o f the column  Ion Retention Time* S0  S 0 2  4-6 minutes  4  3  15-20 minutes  *Note: difference i n retention time depends on condition o f the column  Quality Control: The following samples must be analyzed after an analytical run i.  Replicate  calibration standard:  choose  the  calibration standard  close  concentration i i . Replicate sample: select random a sample and perform a second analytical run  172  to  prevalent  Appendix C: Thermodynamic data Formula  AG , kJ/mol 0  Formula  H 0  -237.18  Au  +  OH"  -157.29  Au  3 +  NH  3  -26.6  NH  4  2  AG°,kJ/mol  176. 440. -51.9  Au0 " 3  3  -79.37  HAu0 "  -142.  0  H Au0 -  -218.  86.31  A u ( O H ) (aq)  -283.5  79.5  A u ( O H ) (s)  s "  73.6  Au(NH )  s "  69  s "  65.7  Cu  so -  -486.6  Cu  so -  -744.63  Cu  so -  -518.8  Cu 0  -148.10  so -  -532.2*, ***  CuO  -134.  S0 "  -600.4  SA -  +  S  s2  2  3  2  4  2  5  2  3  2  4  2  2  3  2  2  3  2  3  2  3  3  -317  3  -41 1**  +  3  2  -1048.**  Au(S 0 ) " 3  2  3  2  0 50.30  +  65.70  2 +  2  C u ( O H ) (s)  -359.50  -791.  HCu0 " 2  -258.90  S0 "  -966.  Cu0 "  -183.90  S0 "  -1110.4  Cu(NH )  so -  -958.  Cu(NH )  so -  -1022.2  Cu(NH )  2  so -  -956.  Cu(NH )  3  HS'  12.05  Cu(NH )  4  H S (aq)  -27.87  Cu(S 0 ) -  -1083.66***  HS0 '  -527.81  Cu(S 0 ) "  -1623.39***  HSCV  -756.01  Cu S  -87.60  0  CuS  -53.20  2  2  4  2  2  2  6  2  2  8  2  3  6  2  4  6  2  5  6  2  3  Au A l l data from Bard  2  2  2  3  -63.  + 2  15.60  2+  3  3  2  5  2  3  -111.50  2+  3  3  -73.20  2+  3  2  -30.50  2+  3  3  2  [1986], except * from C R C Handbook o f Chemistry and Physics [1975] and  ** from Atluri [1987] and Wang [1992] and *** from Duby [1977] 173  Appendix D: Experimental data Mineral  CuS  Cu S  CuFeSj  Cu AsS  Grams added Residue  Baseline 7.92 6.06  Baseline 9.28 7.78  Baseline 17.56 15.87  Baseline 4.31 3.97  Baseline 12.85 11.29  24.84 16.40  24.89  24.89  16.43  16.43  24.96 16.48  22.02 14.53  NaOH added (g) NaOH added (mL) Colour samples Time (min.) = 5  2  3  4  CujO  colourless  colourless  colourless  colourless  l.blue  30  colourless  l.blue  colourless  colourless  blue  60 120  colourless colourless  blue blue  colourless colourless  blue blue  very l.blue l.blue  colourless colourless colourless colourless  blue blue  180  l.blue  blue  240  blue  360  blue  blue blue  blue  blue  Addition of NH3 to samples: only t240, 360  all except t5,30  only O40, 360  to none  to all  Colour filtrate  d.blue  blue  v.l.blue  d.blue  colourless  colourless  colourless  colourless  g/L S-species 15.48  g/L S-species 21.55  g/L S-species  g/L S-species 6.72  blue  Colour washwater colourless Ion chromatograph analyses g/L S-species Thiosulfate 5 19.83 30 60  15.59 12.41 8.70  10.04  19.95  22.69  9.41 3.36  16.81  22.56  13.44  180 240 360  2.09 0.00  0.09 0.19  3.18 0.75  21.77 20.74 19.81  0.05  0.02  0.00  0.05  17.31  0.02  5  0.53  1.20  30 60 120  1.73  1.47  0.42 1.14  0.19  3.21 2.94  1.76 0.82  0.49 0.04  2.06  0.27  2.10  1.17  0.39  0.68  180  0.40  0.00  0.64  0.49  0.63  240  0.03  0.00  0.52  0.00  360  0.00  0.00 0.00  0.06  0.26  120  Tetrathionate  Sulfate  .4.80 2.67 3.30 1.31  0.00  5 30  0.24 0.56  60  1.54  120  0.78  180  3.65  0.00 0.00  240  5.23  0.00  3.62  0.77  4.07  g/LS  g/LS  0.00 0.00  360  5.46 Sulfur balance at every sample time (thiosulfate, tetrathionate and sulfate) g/LS 5 30  12.65  23.98 20.59 18.00 8.80 3.82  13.69 13.09  11.30  60  12.25 11.31  13.06  8.74  120  8.62  12.67  5.22  180  3.41  12.14  2.38  240  2.17  11.63  1.24  360  1.89  10.31  1.37  Trithionate 360  5.51  -  7.28  7.40  10.31  8.66  Total Sulfur balance at t=360, incl.trithionate: - = no trithionate present, or too small to measure Sample calculation:  CuFeS , baseline, t=5: 21.55*0.572 (% S in S 0 ) + 0.42*0.572 (% S in S O )-H).24*0.334 (% S in S0 )=12.65 2  2  3  4  174  6  4  Cu S  Cu 0  CuS+Au  Cu S+Au  no aeration 12.85 6.73  Baseline 1.58 1.24  Baseline  5.33  no aeration 9.28 7.42  NaOH added (g)  23.3  25.52  23.03  24.95  25.58  NaOH added (mL)  15.38  16.84  15.20  16.47  16.88  CuC0  Mineral  3  Baseline 10.48  Grams added Residue  2  2  Colour samples Time (min.) =  2  9.28 7.82  5 30  l.blue  colourless  l.blue  colourless  blue  colourless  l.blue  colourless  colourless l.blue  60  blue  colourless  l.blue l.blue l.blue l.blue  colourless  blue  colourless colourless colourless  blue blue  l.blue  v.l.blue  to none  to all  none l.blue colourless  120 180 240 360  blue blue blue blue  colourless colourless colourless colourless  blue blue  Addition of to all d.blue/purple  NH3 to samples:  to all, except t5 v.dark blue  colourless-l.blue  v.l.blue colourless  l.blue  Colour washwater Ion chromatograph analyses  g/L S-species  g/L S-species  g/L S-species  4.88  21.53  2.18  g/L S-species 21.88  g/L S-species  Thiosulfate  3.49 4.41  20.35  3.25  21.19  3.27  19.10  1.60  20.93  3.58  2.23 0.12 0.07  15.00 15.12 11.22  3.11 3.42 6.93  18.92 17.09 14.08  3.75 2.36 0.00  19.03 0.24  12.85  5 30  0.16 4.37 3.07  2.93 0.00  60 120 180  2.05 0.48 0.00  240  Colour filtrate  5 30 60 120 180 240 360  Tetrathionate  360 Sulfate  l.bl-green  2.99 2.60  0.55 1.08  13.31  0.06  0.22  2.20 2.41 1.55  2.09 1.99  3.11 2.94  0.51 0.72 0.47  0.76 0.08  0.00  2.88  2.56  0.13  0.00  0.00  1.26  1.84  0.00  0.00  1.63  23.56 22.73  0.15 0.15  11.44 21.40 9.03 5.01  5 30  18.88 15.73  60 120  10.03  180  2.91  2.58  23.61  4.41  4.48 12.77  22.26  0.15 0.15  2.93  13.97  22.61  0.15  19.27 17.59 3.15 240 17.85 6.99 3.79 360 Sulfur balance at every sample time (thiosulfate, tetrathionate and sulfate) g/LS g/LS g/LS  Trithionate  l.blue  '  16.26  0.15  5.40  1.72  6.07  g/LS  g/LS  10.82  12.57 12.30 12.31  12.69 10.40  13.04  10.94 10.47  8.37  14.04  10.99  11.28  5.59  1.05  14.45  11.19  10.10  3.07  240  1.09  13.94  360  1.36 4.93  13.94  11.86 14.37  8.18 2.25  -  -  3.73  2.06 8.34  13.94  14.37  5.98  10.40  5  11.60  30 60  9.01 7.04  120  3.03  180  360  6.29 Tot. S.balance at t=360: - = no trithionate present, or too small to measure  13.00 12.81  175  1.80  Mineral  Cu S+Au  CuFeS +Au  Cu AsS +Au  Cu 0+Au  CuCQ +Au  Grams added Residue  Baseline 1.86 1.16  Baseline 3.51 3.02  Baseline 4.31 3.89  Baseline 12.85 10.95  Baseline 2.1 0.36  NaOH added (g)  25.6  24.2  30.19  22  24.22  NaOH added (mL) Colour samples  16.90  15.97  19.93  14.52  15.99  Time (min.) =  2  2  3  4  2  3  5  colourless  colourless  colourless  blue  v.l.blue  30  colourless  colourless  colourless  blue  v.l.blue  60  v. l.blue  colourless  colourless  blue  l.blue  120 180 240  l.blue blue blue  colourless colourless colourless  colourless colourless colourless  blue blue blue  blue blue blue  blue  blue  colourless  360 Addition of NH3 to samples: Colour filtrate Colour washwater  to 240, 360  none  to none  all  to 120, 180.240  blue  v.l.blue  v.l.blue  blue  colourless  colourless  colourless  l.blue  d.blue/purple v.l.blue  g/L S-species  g/L S-species  g/L S-species  Ion chromatograph analyses Thiosulfate  Tetrathionate  Sulfate  5  22.47  g/L S-species 0.00  30  23.68  22.19  22.17  1.95  0.17  60 120 180  16.36 8.04 1.98  22.96 21.10  21.71  2.33  2.01  2.69 0.24  2.77  19.18  20.76 19.76  240 360  0.00 0.00  18.09 12.68  18.49 15.75  0.07 0.04  0.00 0.00  5 30 60 .  2.32 1.16  0.00 0.28 0.59  0.12  3.11 3.02  0.22  2.87 2.69 1.83  2.07  120. 180  0.73  0.95  0.00  0.59 ' 0.00  0.66 0.00  240  0.00  1.10 1.01  0.35 0.40 . 0.38  0.00  0.00  360 5  0.00  0.11  0.00  0.00 11.93  30 60  0.17 0.38  120  1.35  180 240 360  2.39 5.07 5.70  0.54 0.00  0.66 0.65  0.27  0.69 0.03  0.73 0.76 0.91  0.07 1.04  Sulfur balance at every sample time (thiosulfate, tetrathionate and sulfate) g/LS g/LS g/LS 5  Trithionate  g/L S-species 0.54  0.53  4.71 3.23 3.59 4.07 4.65 g/LS  30  14.93  12.85 12.85  60 120  10.15  13.56  12.77  3.91  5.47 1.93  12.61  180  11.61  12.31 11.78  3.04 1.50  240  1.69  10.95  11.04  1.36  360  1.90  7.91  360  5.94  2.99  9.38 -  6.38  9.38  7.94  5.76  Tot. S.balance at t=360: 7.84 10.90 - = no trithionate present, or too small to measure  176  12.97  1.83  1.55  Mineral Grams added Residue  CujS+Au  Cu 0+Au  CujS+Au  Cu 0+Au  Cu S+Au  Cu S+Au  no aeration 9.28  no aeration 2.57  20°C 1.86  20°C 2.57  0.4MNH3 1.86 1.14  Sulfite addn. 1.86 1.2  2  2  2  2  7.86  1.26  1.25  1.31  NaOH added (g)  25.54  23.54  21.27  19.31  33.58  18.33  NaOH added (mL)  16.86  15.54  14.04  12.75  22.17  12.10  5  colourless  colourless  colourless  colourless  l.blue  colourless  30  colourless  colourless  colourless  l.blue  l.blue  colourless  60  colourless  colourless  colourless  l.blue  blue  v.l.blue  120 180 240  colourless colourless  v.l.blue v.l.blue  blue blue  blue blue  l.blue blue  colourless  colourless colourless colourless  l.blue  blue  blue ,  blue  360  colourless  colourless  blue  blue  Addition of NH3 to samples:  none  none  to 360  to 120, 180.240 to none  to 240, 360  Colour filtrate  very l.blue  l.blue  d.blue  d.blue/purple  d.blue  blue  colourless  colourless  colourless  colourless  l.blue  g/L S-species  g/L S-species  g/L S-species  g/L S-species g/L S-species  0.00 0.00  22.08  0.00 0.00 0.00  15.93 11.92  20.20 17.17  Colour samples Time (min.) =  Colour washwater colourless Ion chromatograph analyses Thiosulfate  Tetrathionate  Sulfate  5  g/L S-species 21.08  30  21.21  60 120 180  20.99 19.06 17.59  0.00 0.00  19.86 12.60 8.62  0.00  5.16  2.05 0.94  1.93 0.00  2.11  240  17.33  360 5 30  19.83 0.22 0.39  0.84 4.32 2.51 2.65  1.59 0.00  0.00 0.00  0.00 0.00  0.30 0.00  0.32 1.07  3.39 4.32  0.58  2.72  2.75  0.99  1.79 0.90  0.03  60 120  0.96  2.71  2.71  2.04  0.49  180  2.81  1.04  0.00  240  1.19 1.20  0.15 0.00  2.91  0.00  0.00  0.00  0.00  360  0.77  3.13 11.38  0.00  0.00 14.95  0.00  0.00  5 30  0.56  7.73  0.24 0.21  0.65  12.46  0.86  1.92  60 120  12.55  1.24 1.69  6.35 2.79  0.65 1.41  2.91 4.39.  2.38  2.87  5.67  3.01 3.13 Sulfur balance at every sample time (thiosulfate, tetrathionate and sulfate)  3.22 3.68  8.72 9.48  180  5.73  1.69  240 360  5.21 0.92  1.67 3.28  g/LS 5.24 1.51  g/LS  g/LS  g/LS  g/LS  5 30  13.00 12.19  6.93 6.63  10.42  12.21  60  5.75  9.19  2.69  7.55  10.93  120  7.05 4.11  3.27  1.66 0.96  6.17  180  1.55 3.52  240  3.88  1.47  1.07  3.08  360  4.57 -  1.09  Trithionate 360 Tot. S.balance at t=360:  4.57 - = no trithionate present, or too small to measure  1.33 1.01  3.22  1.23  3.17  7.38  1.05 -  6.47  5.29  8.48  1.05  7.70  8.45  177  Mineral  Cu S+Au  Lobo  Lobo  Lobo  Lobo  Sulfate addn. 1.86 1.1  Baseline 360 342.66  Baseline, dupl. 360 336.53  no aeration 360 339.42  no Cu addn. 360 335.09  25.08 16.55  23 15.18  22.63 14.94  22.45 14.82  21.63 14.28  colourless  v.l.blue  v. l.blue  colourless  colourless  colourless  colourless  colourless  colourless  2  Grams added Residue NaOH added (g) NaOH added (mL)  2+  Colour samples Time (min.)  5 30  colourless  v.l.blue  60  v.l.blue  v.l.blue  v.l.blue v.l.blue  120 180  l.blue blue  l.blue l.blue  l.blue l.blue  colourless colourless  colourless colourless  240  blue  blue  colourless  colourless  360 Addition of NH3 to samples: Colour filtrate  to 240,  to none  to none  to none  to none  blue  blue  v.l.blue  Colour washwater  colourless  colourless  blue l.blue  colourless  colourless colourless  Ion chromatograph analyses Thiosulfate  Tetrathionate  Sulfate  g/L S-species  g/L S-species  g/L S-species  g/L S-species  g/L S-species  5 30  21.42  13.91  13.01  4.58  21.87  60  17.65  13.30  10.62  9.22  21.44  120 180 240  12.19 4.99  8.78 0.00 0.00  3.30 0.00  7.10 11.58  20.52 19.52  0.00  10.29  18.70  360  0.00  0.00  15.17  16.86  2.03 0.97 5.40  2.26 0.94 0.17  3.36  0.51  3.04 3.16  0.66 0.78  0.00  0.00  2.61  0.82  0.00  0.00  2.68  0.81  0.00  1.97  0.65  0.79  5 30  0.54  60 120 180  0.92 0.38 0.22  240  0.00  360  0.00  5 30  4.80  2.40  2.83  3.88  0.94  60  4.69  2.10  120  4.38 6.25  3.86 3.45  0.95 0.96  180  4.93 6.18  2.91 3.71 6.30  3.22  0.97  240  7.15  6.55  6.63 7.37  3.17  0.98  2.29  0.99  360 9.46 Sulfur balance at every sample time (thiosulfate, tetrathionate and sulfate) g/LS  g/LS  g/LS  g/LS  g/LS  30 60  14.17  9.92 8.87  9.68  5.84  13.12  12.19  7.58  8.30  12.96  120  8.84  9.58  3.23  7.02  12.51  C J  180  5.04  9.19  11.96  2.84  2.09 2.19  2.10  240  2.21  8.48  11.49  360  3.16  0.00  2.46  360  4.07  4.00  5.45  10.57 -  3.09  Tot. S.balance at t=360: 7.23 4.00 - = no trithionate present, or too small to measure  7.91  10.57  13.44  Trithionate  178  10.35  Mineral  Lobo  Grams added  no Cu addn.,dupl. 360  Guanaco  Guanaco  Guanaco  Guanaco  Baseline 360  no Cu addn. 360 338.65  Thiogold 360 341.97  Residue  338.78  343.13  no aeration 360 339.72  NaOH added (g)  21.74  22.76  22.76  20.69  24.11  14.35  15.02  15.02  13.66  15.91  colourless  l.blue  colourless  colourless  l.blue  2+  NaOH added (mL)  2+  Colour samples Time (min.) =  5 30  colourless  blue  colourless  v.l.blue  blue  60  colourless  blue  colourless  l.blue  blue  120 180  colourless colourless colourless  blue blue blue  blue colourless colourless/v.l.b. blue blue v.l.b.  240  blue blue blue  360 Addition of NH3 to samples:  to none  to all  to none  to 180, 240, 360  Colour filtrate  colourless  blue  l.blue  blue  d.blue/purple  blue  v.l.b.  l.blue  blue  g/L S-species  g/L S-species  g/L S-species  g/L S-species  colourless Colour washwater Ion chromatograph analyses g/L S-species Thiosulfate  Tetrathionate  Sulfate  to all  5 11.32  30 60 120 180 240  22.73 22.03 21.09 20.09  12.54 9.37 3.07 0.00  22.24 19.86  15.94 14.64  14.55 12.96  6.84 1.50  7.82 0.66 0.00  19.60  0.00  13.52  0.00  0.00  360  17.52  0.00  14.36  0.00  0.00  5 30  0.45  1.78  0.34  1.88  0.97  60 120  0.64 0.67  0.78 0.12  0.98  0.45  2.03  1.12 0.43  180  0.72  0.00  2.36  0.00  0.00  240  0.76  0.00  2.40  0.00  0.00  360  0.57  0.00  2.41  0.00  0.00  30 60  0.88  3.47  2.54  1.46  0.87  3.62  2.86  3.93 4.04  120  0.94  4.02  3.43  1.36 2.17  180  0.92  6.27  3.61  3.61  7.43  240 360  1.00 1.06  6.70  3.52  5.21  7.10  3.13  5.83  7.75 8.41  g/LS  g/LS  0.00  5  6.35  Sulfur balance at every sample time (thiosulfate, tetrathionate and sulfate) g/LS  g/LS  g/LS  9.35 7.01  13.77  10.69  3.17  12.88 10.63  9.47 4.88  8.35 6.08 2.50  c  J  30  13.55  60 120  13.26 12.75  180  12.21  2.09  9.97  2.07  2.48  240  11.98  2.24  10.28  2.59  360 360  10.70  2.37 8.34  10.64 -  1.74 1.95 5.17  2.81 6.15  10.71  10.64  7.12  8.96  Trithionate Tot. S.balance at t=360:  2.89  13.59 - = no trithionate present, or too small to measure  179  M10  Mineral  M40  M40  M40  M40  Grams added  Baseline 360  no aeration 360  Residue  334.39  336.6  no Cu addn. 360 336.35  no C u , no air 360 338.04  Baseline 360 333.4  NaOH added (g) NaOH added (mL)  21.5 14.19  19.85  25.63  24.72  20.94  13.10  16.92  16.32  13.82  v.l.blue  colourless  colourless  colourless  l.blue  l.blue  colourless  colourless  colourless  l.blue blue  2+  2+  Colour samples Time (min.)  5 30 60  blue  colourless  colourless  colourless  120 180  blue blue  colourless colourless  colourless l.blue  blue  240  blue  colourless  blue  colourless colourless colourless  360  blue  colourless  blue  colourless  blue  to all, except 5 d.blue/purple  to none  to 360  to all, exept 5 to none colourless, yellowis d.blue/purple  blue blue  Addition of NH3 to samples: Colour filtrate  l.blue Colour washwater Ion chromatograph analyses g/L S-species Thiosulfate  Sulfate  colourless  blue  g/L S-species  g/L S-species  g/L S-species  g/L S-species  1.60  9.04  23.87  19.67  8.67  7.32  12.11  17.67  8.06  9.45 11.95  15.25  180  4.51 0.66 0.00  25.13 25.81  11.71  10.34 10.28  26.18 25.04  240  0.00  12.02  1.36  25.05  0.00  360 5 30  0.00  12.95 3.89 3.41  2.20  24.78  1.09 2.37  5.45 0.63  60  3.67  2.32  0.52  120  0.62 0.00  0.00 0.00 4.27 3.07  3.07  180  0.00  3.09  1.57 1.38  0.66 0.52  1.49 0.30  240  0.00 0.00  2.82  0.12 0.24  0.53  0.00 0.00 4.46  5 30 60 120  Tetrathionate  colourless, yello blue colourless v.l.blue  3.68 3.08  7.02 2.55 0.00  7.93  2.68 4.54  1.94  0.72 0.27  30  4.66  4.12  2.33  2.09  4.75  60  5.04  5.09  2.63  0.46  3.93  120  5.00  4.24  0.15  180  6.62  4.41  3.31 3.88  0.55  3.97 5.66  240 360  7.39 8.76  3.98 3.81  6.86 7.95  0.00 0.00  6.55 8.22  g/LS 14.46  g/LS  360 5  Sulfur balance at every sample time (thiosulfate, tetrathionate and sulfate)  Trithionate  g/LS  g/LS  g/LS  5  5.67  8.91  14.92  30 60  7.50 4.62  10.26 9.21  12.24 10.93  15.44  8.64  15.21  7.08  120  2.05  10.00  7.92  15.40  3.64  180  2.21  9.94  7.97  14.81  2.06  240  2.47  9.82  3.14  14.63  2.19  360  2.93  10.21  4.05  -  5.68  14.59 -  2.74  5.17 8.09  10.21  9.73  14.59  10.85  360 Tot. S.balance at t=360:  - = no trithionate present, or too small to measure  180  6.45  8.10  M10  Mineral  M10  M10  M10  Grams added Residue  no aeration 360 335.52  no aeration 360 341.73  no Cu addn. 360 333.81  NaOH added (g)  21.58  21.48  21  38.49  NaOH added (mL) Colour samples 5 Time (min.) =  14.24  14.18  13.86  25.41  v.l.blue 30 v.l.blue 120  v.l.blue  colourless  l.blue  v.l.blue  v.l.blue  blue  v.l.blue 300  v.l.blue  l.blue  v.l.blue 480 v.l.blue 1440 v.l.blue  v.l.blue l.blue  blue blue blue  blue d.blue  30 60 120 180 240 360  2+  0.72 M A T S 360 337.71  Time  d.blue d.blue  blue  v.l.blue  Addition of NH3 to samples: Colour filtrate  to none 1. green  to 1440 l.blue  to 120,180,240 d.blue/purple  to none d.blue/purple  Colour washwater Ion chromatograph analyses  v.l. green  v.l.blue  blue  blue  g/L S-species  g/L S-species  g/L S-species  g/L S-species  0.51 2.09  12.87 17.56  Thiosulfate  Tetrathionate  5 30  7.37 4.10  4.90  60 120  10.71 9.06  5.76 7.11  2.67 0.14  12.55  180  10.13  9.57  8.75  240 360  9.42 12.54  7.97  0.00 0.00  5 30  4.10 3.92  60 120 180 240 Sulfate  360 5  0.00  6.04 1.09  3.81  3.18 2.67  5.25  3.42  3.29  0.93  3.14  3.52 3.12 3.01  2.92 2.46  0.00 0.00 0.00  3.74 3.24 2.14  2.48  0.00  0.24  5.95  6.92  1.14  30  7.61  6.81  3.43  6.15  60 120  4.65  6.40  4.28  5.12  5.72  5.72  5.33 5.63  180  4.55  5.03  7.07  5.24  5.35  8.02  5.46  240  8.31  9.72 3.05 360 Sulfur balance at every samp]e time (thiosulfate, tetrathionate and sulfate) g/LS g/LS g/LS  Trithionate  5 30 60  8.55 7.13  120  8.90  180 240 360  9.10 9.89 9.61  360  Tot. S.balance at t=360:  9.64  4.42 3.87 3.49  7.26 7.32 7.65  5.76 g/LS 12.42 13.62  1.99  11.20  8.56 7.00  2.36 2.68 3.25  8.61 6.50 2.69  -  2.95  5.86  21.08  9.61  2.95  9.11  23.76  - = no trithionate present, or too small to measure  181  Appendix E: Cyanidation procedures Cyanidation Test Report Sample: Overall Lobo Composite  Date: M a y 9/1998  Test Conditions: Slurry Weight [g]:  100.12  C a ( O H ) added [g]:  Target pulp density: H 0 to add [g]: Mass o f solids [g]:  30 233.61 100.12  N a C N t o a d d [g]: Leach time [firs]:  Mass o f liquids [g]:  233.613  Temperature [°C]:  19.7  Stirrer speed [rpm]:  512  2  Pre-lime p H :  Initial N a C N concentration [g/L]:  5.47 9.12  Post-lime p H :  0.11  2  1 0.234 24  Leach Data Time [hr]  N a C N [g/L]  0 2  1.00 0.38  4  0.75 0.76  8 12  NaCN added [g] 0.234  NaCN cumul. [g] 0.234  0.147  0.381  0.062  0.443 0.502  0.059 0.055  0.78 0.45  24  Solids Balance mass [g] Solids 100.12 Residue  99.39  0.557 0.557  0  pH  Ca(OH) added [g] 0.18 2  9.86 10.36  -  10.61 10.61 10.81  -  10.75  pH 10.50 10.54 10.66 10.67 10.85  -  Solution Balance A u [ppm] 2.11  C u [ppm]  0.4  640  1120  Filtrate Titr. waste  mass [g] 234.38  A u [ppm] 0.64  100  0.24  Au extraction [%]  81.4  Cu extraction [%]  41.1  assayed head [ppm]  2.11 2.14  assayed head [ppm]  1120  calculated head [ppm]  calculated head [ppm]  C u [ppm] 1.75 34  1080.27  Reagent Consumption N a C N [kg/t] C a ( O H ) [kg/t] 2  4.38 2.90  Notes: 1. Sodium cyanide concentration was determined by titration o f 5 n i L solution samples with silver nitrate and rhodanine indicator at selected time intervals. The solution samples were collected and analyzed for copper and gold content (Titr. Waste) 2. The extractions are based on calculated head  182  Cyanidation Test Report Date: M a y 9/1998  Sample: Guanaco Composite Test Conditions: Slurry Weight [g]:  100.04  0.03 0.233 24  2  30  Target pulp density:  C a ( O H ) added [g]: Initial N a C N concentration [g/L]:  H 0 to add [g]: Mass o f solids [g]:  233.43 100.04  N a C N t o a d d [g]:  Mass o f liquids [g]: Pre-lime p H : Post-lime p H :  233.427 7.93 9.72  Temperature [°C]: Stirrer speed [rpm]:  2  Leach time [hrs]:  1  20.7 514  Leach Data Time [hr]  N a C N [g/L]  NaCN added [g]  NaCN cumul. [g]  pH  0 2  1.00 0.09  4  0.233 0.446 0.577 0.684  10.49 10.27 10.58  8 12  0.45 0.55 0.68  0.233 0.213 0.131  24  0.33  0.107 0.079  0.763  0  0.763  Solids Balance mass [g] 100.04 99.78  Solids Residue  Ca(OH) added [g] 0.02 2  -  10.46 10.70 10.34  pH 10.59 10.74  -  10.78 10.64  -  -  10.76  Solution Balance A u [ppm]  C u [ppm]  2.06 0.28  3320 1500  Filtrate Titr. waste  mass [g]  A u [ppm]  C u [ppm]  236.697 100  0.6 0.08  1460 109  Au extraction [%]  84.3  Cu extraction [%]  70.4  assayed head [ppm]  1.6  assayed head [ppm]  3320  calculated head [ppm]  1.78  calculated head [ppm]  5061.47  Reagent Consumption N a C N [kg/t] C a ( O H ) [kg/t] 2  6.77 0.50  Notes: 1. Sodium cyanide concentration was determined by titration o f 5 m L solution samples with silver nitrate and rhodanine indicator at selected time intervals. The solution samples were collected and analyzed for copper and gold content (Titr. Waste) 2. The extractions are based on calculated head  183  Cyanidation Test Report Sample: M 4 0 Pyrite Feed  Date: M a y 10/1998  Test Conditions: Slurry Weight [g]:  100.06  Target pulp density:  C a ( O H ) added [g]: Initial N a C N concentration [g/L]: 2  30  0.07 1  H 0 to add [g]:  233.47  N a C N to add [g]:  0.233  Mass o f solids [g]:  100.06  Leach time [hrs]:  24  Mass o f liquids [g]: Pre-lime p H :  233.473 6.03  2  Post-lime p H :  Temperature [°C]: Stirrer speed [rpm]:  19.7 505  9.89  Leach Data N a C N [g/L]  NaCN added [g]  NaCN cumul. [g]  pH  Ca(OH) added [g]  pH  0  1.00  0.233  0.233  10.40  0.02  10.62  2 4  0.09 0.46 0.59 0.66 0.46  0.213 0.129 0.099 0.082  0.446  9.91 10.84  -  10.63 10.96 10.86 10.81  Time [hr]  8 12 24  Solids Balance mass [g] Solids 100.06 Residue 99.85  0.575 0.674  -  10.78 10.70  0.756 0.756  0  2  10.16  Solution Balance A u [ppm]  C u [ppm]  7.1 0.5  5920 3680  Filtrate Titr. waste  mass [g]  A u [ppm]  239.283 100  3.36 0.96  Au extraction [%]  94.7  Cu extraction [%]  assayed head [ppm]  7.1  assayed head [ppm]  calculated head [ppm]  9.5  calculated head [ppm]  Reagent Consumption N a C N [kg/t] C a ( O H ) [kg/t]  6.40 0.90  2  -  C u [ppm] 775 117  34.9 5920 5645.93  Notes: 1. Sodium cyanide concentration was determined by titration o f 5 m L solution samples with silver nitrate and rhodanine indicator at selected time intervals. The solution samples were collected and analyzed for copper and gold content (Titr. Waste) 2. The extractions are based on calculated head  184  Cyanidation Test Report Sample: M 1 0 Pyrite Concentrate  Date: M a y 10/1998  Test Conditions: Slurry Weight [g]: Target pulp density: H O t o a d d [g]:  100.06  C a ( O H ) added [g]: Initial N a C N concentration [g/L]: N a C N to add [g]:  Mass o f solids [g]:  233.47 100.06  Mass o f liquids [g]: Pre-lime p H :  233.473 5.76  2  Post-lime p H :  0.13  2  30  Leach time [firs]: Temperature [°C]: Stirrer speed [rpm]:  1 0.233 24 20 494  9.57  Leach Data Time [hr]  N a C N [g/L]  NaCN added [g]  NaCN cumul. [g]  0 2 4  1.00 0 0.07 0.16  0.233 0.233 0.217  0.233  10.34  0.466 0.683 0.879  10.20 10.86  8 12 24  0.196 0.182 0  0.23 0.29  Ca(OH) added [g] 0.04 2  -  11.01 11.08 10.25  1.061 1.061  Solids Balance  pH  pH 10.73 11.08 11.21  -  11.20 11.23  -  -  Solution Balance  mass [g]  A u [ppm]  C u [ppm]  Solids  100.06  45.5  16600  Filtrate  Residue  98.39  21.5  11400  Titr. waste  mass [g]  A u [ppm]  C u [ppm]  242.103  7.6 1.32  1750 192  100  Au extraction [%]  47.5  Cu extraction [%]  28.3  assayed head [ppm]  45.5  assayed head [ppm]  16600  calculated head [ppm]  41.54  calculated head [ppm]  15645.27  Reagent Consumption N a C N [kg/t] C a ( O H ) [kg/t] 2  9.91 1.70  Notes: 1. Sodium cyanide concentration was determined by titration o f 5 m L solution samples with silver nitrate and rhodanine indicator at selected time intervals. The solution samples were collected and analyzed for copper and gold content (Titr. Waste) 2. The extractions are based on calculated head  185  Appendix F: Sulfate determination  Determination of sulfate as barium sulfate (derived from Jeffery, [1989])  The method consists in slowly adding a solution o f barium chloride to a solution o f the sulfate: B a C l + S O " -> B a S 0 (s) + 2C1"  (F. 1)  2  2  4  Normally this procedure is carried out at boiling temperature, to obtain large barium sulfate crystals. For this study, the measurement was performed at room temperature.  To determine i f it was possible to obtain representative sulfate numbers at room temperature the following procedure was followed: •  Prepare standard solutions o f 0.1 M ammonium thiosulfate and 0.1 M ammonium sulfate and 1 M barium chloride  •  Measure out 100 m L o f the first two solutions and transfer them into an 300 m L erlenmeyer  •  While stirring add slowly 1 M o f barium chloride  •  Filter, wash and dry precipitate  The sulfate concentration was calculated from the resulting weight o f the barium sulfate precipitate. The amount o f sulfate added was 0.96 gram, the method gave 1.08 gram. The filtrates resulting from the leach experiments were subjected to the same procedure, only 50 m L o f filtrate was used.  186  

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