recovery of gold from thiosulfate solutions and pulps with ......leach pulp to yield loadings of...

Recovery of Gold from Thiosulfate

Solutions and Pulps with Anion-

Exchange Resins

This Thesis is Presented for the Degree of

Doctor of Philosophy in Extractive Metallurgy

From Murdoch University,

Western Australia

Glen Peter O'Malley

2002

i

I declare that this thesis is an account of my own research and contains work that has not previously been submitted for a degree

at any other educational institution.

Glen O'Malley

March, 2002

ii

ACKNOWLEDGEMENTS

A special thanks goes to Professor Mike Nicol for his guidance and support throughout the

PhD. Thanks also go to Rob Dunne (Newcrest Mining) and Dr. Steven La Brooy (Gold

Resource and Development Macraes Ltd) for their support and feedback as the SPIRT

sponsors for this PhD.

I am grateful for the financial support of my two Australian Research Council SPIRT

sponsors, Newcrest Mining and GRD Macraes. Also for the financial support of the

sponsors of AMIRA Project P420A Module 3 which include Anglogold Australasia, Alcoa

World Alumina Australia, Carpentaria Gold, Goldfields, Newcrest Mining, Normandy

Mining, Placer Dome, Rio Tinto Research & Technology Development, WMC Resources

and Worsley Alumina.

The author is also grateful to Dr. Gamini Senanayake, Dr. Kathryn Hindmarsh, Dr. Peter

Lye and Karen Barbetti for their suggestions and help in preparing the initial stages of the

thesis. Special thanks go to Dr. Nimal Perera, Dr. Peter Lye and Dr. Hongguang Zhang for

their suggestions during the final stages of the thesis. Thanks also go to J. Avilla

(ResinTech Inc.), I. Needs and S. Stewert (Rohm and Haas) for providing the resin samples

and G. Wardell-Johnson (Boddington Gold Mine) and G. Kelly (Kandowna Belle Gold

Mine) for the ore samples.

Lastly I would like to thank the A.J. Parker Cooperative Research Centre for

Hydrometallurgy for the courses and seminars conducted as part of the PhD student

development program.

iii

ABSTRACT

With growing environmental and occupational safety concerns over the use of cyanide in

gold processing, more acceptable alternatives are receiving increased interest. The most

promising of the possible alternatives is thiosulfate. However, as activated carbon is not an

effective substrate for the adsorption of the gold thiosulfate complex, the thiosulfate

process lacks a proven in-pulp method for recovering dissolved gold. Anion exchange

resins offer a possible route for in-pulp recovery. This thesis describes work aimed at

evaluating the effectiveness of commercially available anion exchange resins for the

recovery of gold from thiosulfate leach liquors and pulps.

It was found that Strong-base resins are superior at accommodating the gold thiosulfate

complex compared to Weak-base resins, which means Strong-base resins have a greater

capacity to compete with other anions in leach solutions. Strong-base resins were therefore

the preferred choice of resin for recovery of gold from thiosulfate leach solutions and

pulps. Work with a selected commercial Strong-base resin showed that competing

polythionates (particularly tri- and tetrathionate) lower the maximum possible loading of

gold but that gold is selectively recovered over other base-metal anions in typical leach

solutions. From kinetic experiments, it was found that competing polythionates did not

affect the initial rate of loading of gold but displaced the loaded gold at long times. Thus it

would be important to minimise the contact time of the resin with the pulp.

Equilibrium loading isotherms of gold in the presence of competing anions could be

analysed by treating the ion exchange reaction as a simple chemical reaction. However, a

stoichiometry and equilibrium quotient which does not follow that normally used for anion

exchange, was required to describe the experimental data. A single value for the

equilibrium constant also cannot be used to describe the data over the range of

concentrations for a given competing anion. The order of selectivity of the anions for the

anion exchange resin could be explained by the difference in structure and the charge of

each anion.

iv

The rate of loading of gold is controlled by mass transport in the aqueous phase in the

presence of weakly competing anions such as sulfate and thiosulfate. An attempt was

made to describe the more complex loading curves obtained in the presence of stronger

competing anions such as sulfite, trithionate and tetrathionate in which it was found that

the loading of gold increased to a maximum before declining to a lower equilibrium value.

The difference in the rate of loading between the macroporous and gel anion exchange

resins was explained by the difference in the location of their functional groups.

Operation of a small-scale resin-in-pulp plant showed that gold could be recovered from a

leach pulp to yield loadings of gold of up to 6000 mg L-1 and loadings of copper below 100

mg L-1. Under ideal conditions, the gold concentration in the barren pulp could contain

less than 0.01 mg L-1. Throughout the trial it was shown that loaded copper would be

displaced by gold which would result in the loading of copper falling from 2000 mg L-1 in

the last stage to lower than 100 mg L-1 on the resin in the first stage. It was observed that

some of the dissolved gold precipitated or adsorbed on the solids during leaching. Some of

this adsorbed gold was found to be recovered by the anion exchange resins that would have

reported to the tails if a solid/liquid separation method was employed.

Gold was efficiently eluted with a nitrate solution and the two-step elution process using

aerated ammonia followed by nitrate effectively stripped all the copper and gold from the

resin. This process was found not to materially affect the equilibrium gold concentration

on the resin after eight cycles, thus allowing the resin to be recycled without the need for

regeneration. Electrochemical studies showed that the gold thiosulfate complex was

reduced on stainless steel from a nitrate solution. Conventional electrowinning could

therefore be used to recover the gold from the eluant.

A copy of the thesis and the appendices are provided on CD.

v

TABLE OF CONTENTS

Page

DECLARATION i

ACKNOWLEDGEMENTS ii

ABSTRACT iii

TABLE OF CONTENTS v

LIST OF TABLES xiii

LIST OF FIGURES xvi

LIST OF APPENDICES xxvi

ABREVIATIONS xxvii

CHAPTER 1 REVIEW OF THE LEACHING OF GOLD USING

THIOSULFATE 1

1.1 INTRODUCTION 1

1.2 THE LEACHING OF GOLD WITH THIOSULFATE 2

1.2.1 History 2

1.2.2 Limitation and Advantages of Leaching of Gold with Thiosulfate 3

1.2.3 Condition Employed in Thiosulfate Leaching with Thiosulfate 4

1.3 THE CHEMISTRY OF THIOSULFATE 8

1.3.1 Oxidation State of Sulfur 8

1.3.2 Stability of Thiosulfate 10

1.3.3 Metal Complexation with Thiosulfate 12

1.3.4 Generation and Stabilization of Thiosulfate 13

1.3.5 Determination of Thiosulfate 14

1.3.6 Corrosion Properties 15

vi

1.4 THERMODYNAMICS OF THE LEACHING OF GOLD AND SILVER

BY THIOSULFATE 16

1.4.1 Stability of Gold and Silver Complexes 16

1.4.2 Stability of Copper Complexes 18

1.4.3 Stability of Sulfur Species 19

1.4.4 Alternative Stability Diagram for Gold 20

1.5 FACTORS INFLUENCING THE RATE OF DISSOLUTION OF GOLD

AND SILVER 22

1.5.1 Effect of Thiosulfate Concentration 22

1.5.2 Effect of Ammonia Concentration 23

1.5.3 Effect of Copper Concentration 24

1.5.4 Effect of Temperature 25

1.5.5 Effect of pH 25

1.5.6 Effect of the Partial Pressure of Oxygen 26

1.5.7 Effect of Sulfite and Sulfate 26

1.5.8 Effect of Leach Time 27

1.5.9 Effect of Pulp Density 27

1.5.10 Effect of Pre-Oxidation 28

1.6 ELECTOCHEMICAL STUDIES ON THE MECHANISM OF GOLD

DISSOLUTION IN THIOSULFATE SOLUTION 29

1.7 CONCLUSION 33

vii

CHAPTER 2 REVIEW OF THE RECOVERY OF GOLD FROM

THIOSULFATE LEACH SOLUTIONS AND PULPS 34

2.1 INTRODUCTION 34

2.2 REDUCTION OF THE GOLD THIOSULFATE COMPLEX 35

2.2.1 Cementation of Gold on Copper, Iron, Zinc or Aluminium 35

2.2.2 Chemical Reduction 37

2.2.3 Electrowinning 39

2.3 CHEMICAL PRECIPITATION 41

2.4 SOLVENT EXTRACTION 42

2.5 ACTIVATED CARBON 44

2.6 ADSORPTION ON ANION EXCHANGE RESINS 46

2.6.1 History 46

2.6.2 Structure and Synthesis of Anion Exchange Resins 47

2.6.3 Strong-Base and Weak-Base Anion Exchange Resins 50

2.6.4 Physical Properties of Anion Exchange Resins 52

2.6.5 Adsorption Equilibria and Loading Capacity 52

2.6.6 Equilibrium Gold Loading Models 54

2.6.7 Kinetic Studies and Models for Gold Loading 55

2.6.8 Selectivity of Anion Exchange Resins 61

2.6.8.1 Solution Chemistry 61

2.6.8.2 Properties of Anions 62

2.6.8.3 Structure of the Resin 69

2.6.9 Application of Resins to Recover the Gold Thiosulfate Complex 76

2.6.9.1 Strong-Base Resins 76

2.6.9.2 Weak-Base Resins 76

viii

2.6.10 Summary of Published Work on Recovering Gold from Thiosulfate

Solutions and Pulps 78

2.6.11 Elution of Metal Complexes from Anion Exchange Resins 81

2.6.11.1 Strong-Base resins 81

2.6.11.2 Weak-Base resins 83

2.6.12 Characteristics of a Resin for the Recovery of Gold Thiosulfate from Pulps 84

2.7 FURTHER DEVELOPMENT 86

CHAPTER 3 MATERIALS AND METHODS 87

3.1 MATERIALS, REAGENTS, PROCEDURES AND APPARATUS 87

3.2 PROCEDURES FOR THE PREPARATION, CHARACTERISATION

AND ANALYSIS OF THE RESINS 90

3.2.1 Preparation of Resins 90

3.2.2 Particle Size Distribution 90

3.2.3 Resin Density 91

3.2.4 Theoretical Exchange Capacity 91

3.3 SYNTHESIS OF TRITHIONATE AND TETRATHIONATE 93

3.3.1 Synthesis of Tetrathionate 93

3.3.2 Synthesis of Trithionate 93

3.4 DETERMINATION OF METAL CONCENTRATION IN SOLUTION

BY ATOMIC ADSORPTION SPECTROPHOTOMETER (AAS) 95

3.4.1 Stability of Metal Complexes in a Thiosulfate Leach Solutions 95

3.4.2 Determining Very Low Concentrations of Gold in Solution 95

ix

3.5 DETERMINING OF SULFUR SPECIES 96

3.5.1 Standardisation of Thiosulfate and Iodine 97

3.5.2 Determination of the Sulfite or Thiosulfate Concentration in Solution 98

3.5.3 Determination of Sulfate Concentration in Solution 98

3.5.4 Determination of Tetrathionate Concentration in Solution 99

3.5.5 Determination of Trithionate Concentration in Solution 99

3.6 CHEMISTRY OF GOLD LEACHING WITH THIOSULFATE 100

3.6.1 The Rate of Conversion of Copper (II) to Copper (I) 100

3.6.2 Stability of Copper (II) Amine in a Thiosulfate Solution 100

3.7 ADSORPTION EXPERIMENTS 102

3.7.1 Gold Loading on Various Commercial Resins 102

3.7.2 Loading of Sulfur Anions 102

3.7.3 Effect of Competing Sulfur Anions on the Loading of Gold 103

3.7.4 Effect of Competing Metal Thiosulfate Complexes on the Loading of Gold 103

3.7.5 Consecutive Metal Loadings in a Synthetic Leach Solution 104

3.8 KINETICS OF ADSORPTION 105

3.8.1 The Rate of Loading of Gold and Copper 106

3.8.2 Effect of Copper on the Rate of the Loading of Gold 107

3.9 ELUTION EXPERIMENTS 107

3.9.1 Ammonium Nitrate Elution Process 108

3.9.2 The Effect of Nitrate Ions on the Loading of Gold 108

3.9.3 Gold Elution Rate in Nitrate Solution 108

3.9.4 Elution of Thiosulfate 109

x

3.10 COUNTER-CURRENT RESIN-IN-PULP ADSORPTION 110

3.10.1 Origin of the Ore 110

3.10.2 Sample Preparation 110

3.10.3 Leaching Experiments 111

3.10.4 Laboratory Scale Counter-Current Adsorption Apparatus 112

3.10.5 Experiments Undertaken in the Adsorption Apparatus 115

3.11 ELECTROWINNING INVESTIGATION 116

CHAPTER 4 RESULTS 117

4.1 CHEMISTRY OF GOLD LEACHING WITH THIOSUFLATE 117

4.1.1 Reduction of Copper (II) by Thiosulfate 117

4.1.2 The Initial Reduction Reaction of Copper (II) 119

4.1.3 Stability of Copper (II) Over 24 hours 120

4.1.4 Stability of Metal Complexes in a Thiosulfate Leach Solution 121

4.2 ADSORPTION OF GOLD ON COMMERCIAL RESINS 123

4.3 THE EFFECT OF COPPER (I) ON THE LOADING OF GOLD 125

4.3.1 Effect of Ammonia Concentration 125

4.3.2 Effect of Oxygen 127

4.4 EFFECT OF COMPETING ANIONS ON THE RATE OF LOADING

OF GOLD 128

4.4.1 The Effect of Sulfur species 128

4.4.2 The Effect of Other Metal Complexes 130

xi

4.5 EFFECT OF COMPETING ANIONS ON THE EQUILIBRIUM

LOADING OF GOLD 132

4.5.1 The Effect of Sulfur Species 132

4.5.2 The Effect of Other Metal complexes 135

4.6 CONSECUTIVE LOADING AND ELUTION CYCLES 136

4.7 LEACHING EXPERIMENTS 141

4.8 LABORATORY SCALE COUNTER-CURRENT ADSORPTION

EXPERIMENTS 144

4.8.1 First Run 144

4.8.2 Second Run 149

4.8.3 Third Run 153

4.8.4 Summary of the Adsorption Runs 156

4.9 ELUTION OF COPPER AND GOLD 157

4.9.1 Comparison Between Different Eluting Agents 157

4.9.2 Equilibrium Loading Isotherms of Gold in the Presence of

Ammonium Nitrate 159

4.9.3 Elution Kinetics 161

4.9.4 Selective Elution of Copper (I) and Gold (I) 163

4.9.5 Elution of Thiosulfate 165

4.10 ELECTROWINNING 167

xii

CHAPTER 5 DISCUSSION 168

5.1 CHEMISTRY OF GOLD LEACHING WITH THIOSULFATE 168

5.2 LEACHING EXPERIMENTS 174

5.3 ADSORPTION OF GOLD ON ION EXCHANGE RESINS 177

5.4 THE EFFECT OF COMPETING ANIONS ON THE EQUILIBRIUM

LOADING OF GOLD 181

5.5 EQUILIBRIUM MODELS FOR THE ADSORPTION OF GOLD 185

5.6 THE EFFECT OF COMPETING ANIONS ON THE LOADING OF GOLD 207

5.7 MINIMISING THE EFFECT OF COPPER ON THE LOADING OF GOLD 211

5.8 MODELING THE ADSORPTION KINETICS OF GOLD ON ANION

EXCHANGE RESINS 216

5.9 LABORATORY SCALE COUNTER-CURRENT ADSORPTION

EXPERIMENTS 225

5.10 ELUTION OF COPPER AND GOLD FROM STRONG-BASE RESINS 228

5.11 CONSECUTIVE LOADING AND ELUTION CYCLES 233

5.12 ELECTROWINNING 236

CHAPTER 6 CONCLUSION AND RECOMMENDATIONS 237

CHAPTER 7 REFERENCES 241

xiii

LIST OF TABLES

Page

Table 1.1 The Leaching Conditions and Recovery from Gold and Gold Ores. 5

Table 1.2 The Leaching Conditions and Recovery from Sulfide Concentrates. 6

Table 1.3 The Leaching Conditions and Recovery from Sulfidic and Carbonaceous

Ores. 7

Table 1.4 Oxidation States of Sulfur. 9

Table 1.5 Stability of Metal Complexes in an Ammoniacal Thiosulfate Solution. 12

Table 2.1 References to Various Metals Studied. 35

Table 2.2 Commercial Anion Exchange Resins (Riveros, 1993). 51

Table 2.3 Summary of the Metal Loadings on Various Anion Exchange Resin

from a Cyanide Solution (Riveros, 1993). 62

Table 2.4 Predicted Loading of Sulfur Species on Resins Presented in Declining

Order. 64

Table 2.5 Predicted Loading of Metal Thiosulfate Complexes on Resins

Presented in Declining Order. 65

Table 2.6 Elution Characteristics of Metal Ions on Dowex 1-X8 using 0.06 M

Sodium Thiosulfate Solution as Eluent. 67

Table 2.7 Elution Characteristics of Metal Ions on Dowex 1-X8 using 1.0 M

Sodium Thiosulfate Solution as Eluent. 68

xiv

Table 2.8 The Predicted Selectivity Order of Strong-Base Resins for Gold

Thiosulfate Presented in Declining Order. 74

Table 2.9 The Predicted Selectivity Order of Weak-Base Resins for Gold

Thiosulfate Presented in Declining Order. 75

Table 2.10 pKb Values for Alkylamines in Water at 25 oC

(Clifford and Weber, 1983). ` 78

Table 2.11 Extraction Results from the RIP Experiment with Head Grade of 6.28 g.t-1

(Thomas et al., 1998). 80

Table 3.1 Details of the Various Strong- and Weak-Base Resins Tested. 87

Table 3.2 Description of the Chemicals used. 88

Table 3.3 Composition of the Stability of Copper (II) Amine Solutions. 101

Table 3.4 Composition of the Copper Solutions. 106

Table 3.5 Reagents used in the Leaching Study. 111

Table 4.1 Initial Absorbance of the Two Copper Species for Each Solution. 119

Table 4.2 Average Percentage of Recovery of Each Metal in Elution. 138

Table 4.3 Metal Ion Concentration on the Resin. 139

Table 4.4 Fire Assay Analysis of Gold on the Resin Samples. 140

Table 4.5 Accountability of Gold and Silver Precipitated on the Resin. 140

Table 4.6 Gold Extraction under the Various Leaching Conditions. 142

xv

Table 4.7 Resin Loadings from the Leach Solutions. 143

Table 4.8 Extraction Results from the First Run. 147

Table 4.9 Extraction Results from the Second Run. 151

Table 4.10 Extraction Results from the Third Run. 153

Table 5.1 Freundlich and Langmuir Constants for the Loading of Anions

on the Strong-Base Resin Amberjet 4200. 188

Table 5.2 Freundlich and Langmuir Parameters for Modelling the Loading

of Gold in the Presence of a Competing Anion. 190

Table 5.3 Derived Equilibrium Constants for the Exchange of the Sulfate

Anion. 196

Table 5.4 Determined Equilibrium Constants for the Exchange of the

Competing Anion and Gold Thiosulfate. 206

Table 5.5 Structure of Species Expected in a Thiosulfate Leach Solution. 209

Table 5.6 Change in the Rate of Loading of Gold with Copper and Ammonia

Concentration under a Constant Thiosulfate Concentration (0.2 M). 214

Table 5.7 Kinetic Modelling Parameters. 221

xvi

LIST OF FIGURES

Page Figure 1.1 Oxidation State Diagram for Sulfur at pH 10 (Peters, 1976). 9

Figure 1.2 EH-pH Diagram of Au-NH3-S-H2O System at 5 x 10-4 M Au,

1 M S2O32- and 1 M NH3/NH4+ (Wan, 1997). 17

Figure 1.3 EH-pH Diagram of Ag-S-H2O System at 5 x 10-4 M Ag, 1 M S2O32-

and 1 M NH3/NH4+ (Present Study). 18

Figure 1.4 EH-pH diagram of Cu-NH3-S-H2O system at 5 x 10-4 M Cu,

1 M S2O32- and 1 M NH3/NH4+ (Wan, 1997). 19

Figure 1.5 EH-pH Diagram for S-H2O System at 1 M soluble species

and without consideration of sulfate (Michel and Frenay, 1999). 20

Figure 1.6 EH-pH Diagram for the Au-S-H2O System at 5 x 10-4 M Au,

1 M S2O32- and 1 M NH3/NH4+ (Jiayong et al., 1996). 21

Figure 1.7 Electrochemical Model for the Dissolution of Gold in the Thiosulfate

Leaching System (Toa et al., 1993). 29

Figure 1.8 The Proposed Model for the Dissolution of Gold in the Thiosulfate

Leaching System (present study). 30

Figure 1.9 Proposed Reaction Scheme for Copper Speciation (Present Study). 31

Figure 1.10 Electrochemical Reaction Scheme for Thiosulfate Leaching (Michel and

Frenay, 1998). 32

Figure 2.1 Current-Potential Curve for the Reduction of Gold Thiosulfate on

a Gold Electrode (Sullivan and Kohl, 1997). 40

xvii

Figure 2.2 Adsorption of Gold Complexes on Activated Carbon

(Gallagher et al., 1990). 45

Figure 2.3 Synthesis of Polystyrene Anion Exchange Resins (Hayes, 1995). 48

Figure 2.4 Synthesis of Acrylic Anion Exchange Resins (Hayes, 1995). 49

Figure 2.5 Gold Equilibrium Isotherm. 54

Figure 2.6 Adsorption Isotherm Models (K = 0.5, a = 18, b = 0.5 and Cmax = 95). 55

Figure 2.7 Representation of the Shell Progressive Mechanism in a Spherical Ion

Exchange Bead (Nativ et al., 1975). 58

Figure 2.8 Simultaneous Extraction of Gold at 0.5 mmol.L-1 and Zinc at

0.5 mmol.L-1 on Dowex MSA-1 with free Cyanide = 2 mmol.L-1

and Sodium Hydroxide = 2.5 mmol.L-1 (Riveros and Cooper, 1988). 60

Figure 2.9 Elution Curve of a Mixed Sample (Flow rate 0.3 mL min-1)

(Iguchi, 1958). 68

Figure 2.10 The Effect of Resin Matrix on the Selectivity of −24SO over −3NO

for Polyamine or Weak-Base Anion Exchangers

(Clifford and Weber, 1983). 69

Figure 2.11 The Effect of Resin Matrix on the Selectivity of −24SO over −3NO

for Quaternary Amine or Strong-Base Anion Exchangers


Figure 2.12 The Effect of Amine Functionality on the Selectivity of −24SO over

−3NO for Polystyrene-Divinylbenzene Anion Exchangers


xviii

Figure 2.13 The Effect of Amine Functionality on the Selectivity of −24SO over

−3NO for Acrylic Anion Exchangers (Clifford and Weber, 1983). 72

Figure 2.14 Titration Curves for Polystyrene Macroporous Resin with a Tertiary

Amine Functionality (Clifford and Weber, 1983). 77

Figure 3.1 Trithionate Crystals Magnified Five Times Their Original Size. 96

Figure 3.2 Set-up for Kinetic Experiments. 105

Figure 3.3 Diagram of the Elution Set-up. 107

Figure 3.4 Side Elevation View of Contactor. 113

Figure 3.5 Plan View of Contactor. 113

Figure 3.6 Diagram of the Laboratory Scale Adsorption Apparatus. 114

Figure 3.7 The Laboratory Scale Resin-In-Pulp Adsorption Apparatus. 114

Figure 3.8 Electrochemical Apparatus. 116

Figure 4.1 The Effect of Ammonia and Thiosulfate Concentration on the First

Order Rate Constant for the Disappearance of the Copper (II) Amine

From Solution. 118

Figure 4.2 The Percentage of Copper (II) Amine after 24 hours. Effect of Ammonia

and Thiosulfate concentration with a Fixed Copper (0.05 M), Sulfate (0.1 M)

and Sulfite (0.1 M) Concentrations, in the Presence of Air at Ambient

Temperature. 120

xix

Figure 4.3 Change in the Slope of the Calibration Curve for the Combined Metal

Standard Solutions Containing 0.2 M Ammonia and 0.05 M Thiosulfate

at pH 9.5 and Ambient Temperature. 122

Figure 4.4 Change in the Slope of the Calibration Curve for Individual Metal

Standard Solutions Containing 0.2 M Ammonia and 0.05 M Thiosulfate

at pH 9.5 at Ambient Temperature. 122

Figure 4.5 Equilibrium Loading of Gold on Various Commercial Strong-Base

Anion Exchange Resins from a Synthetic Solution Containing only

Gold Thiosulfate Complex, at Ambient Temperature. 124

Figure 4.6 Equilibrium Loading of Gold on Various Commercial Weak-Base

Anion Exchange Resins from a Synthetic Solution Containing only

Gold Thiosulfate Complex, at Ambient Temperature. 124

Figure 4.7 The Effect of Ammonia and Copper Concentration on the Rate of

Loading of Gold on Amberjet 4200 in the Presence of 0.2 M

Thiosulfate. 126

Figure 4.8 The Effect of Ammonia and Copper Concentration on the Rate of

Loading of Gold on Vitrokele 911 in the Presence of 0.2 M

Thiosulfate. 126

Figure 4.9 Effect of Oxygen on the Loading of Copper (I) on Amberjet 4200

at 0.05 M Thiosulfate, 0.2 M Ammonia, and 0.3 mM Copper 127

Figure 4.10 Effect of Various Sulfur Anions at 0.05 M on the Kinetics of the

Loading of Gold on Amberjet 4200 at an Initial Gold

Concentration of 20 mg L-1. 129

Figure 4.11 The Effect of Metal Thiosulfate Complexes on the Loading of

Gold Kinetics on Amberjet 4200 at 0.05 M Thiosulfate, 0.2 M

Ammonia, and 20 mg L-1 for Each Metal Ion. 131

xx

Figure 4.12 The Effect of Metal Thiosulfate Complexes on the Loading of Gold

Kinetics on Amberjet 4200 at 0.05 M Thiosulfate, 0.2 M Ammonia,

5 mM Trithionate, and 20 mg L-1 for Each Metal Ion. 131

Figure 4.13 The Equilibrium Loading of Sulfur Compounds on Amberjet 4200. 132

Figure 4.14 Effect of Various Sulfur Anions at 0.05 M on the Equilibrium

Loading of Gold on Amberjet 4200. 133

Figure 4.15 The Effect of Anion Concentration on the Maximum Loading of

Gold on Amberjet 4200. 134

Figure 4.16 Equilibrium Loading of Various Metal Thiosulfate Complexes from

a Synthetic Thiosulfate Solution Containing 0.05 M Thiosulfate and

0.2 M Ammonia on Amberjet 4200. 135

Figure 4.17 Consecutive Equilibrium Loadings with a Simulated Leach Solution

Containing 0.05 M Thiosulfate, 0.2 M Ammonia, 10 mM Trithionate

and 10 mg L-1 Each for Gold, Silver and Copper. 136

Figure 4.18 Consecutive Equilibrium Loading with a Simulated Leach Solution


and 10 mg L-1 Each for Gold and Copper. 137

Figure 4.19 Consecutive Equilibrium Loading with a Simulated Leach Solution


and 10 mg L-1 Each for Silver and Copper. 137

Figure 4.20 Leaching Curves for Gold under Various Leaching Conditions. 142

Figure 4.21 Change in Gold Concentration in Each Stage during the First Run. 144

Figure 4.22 Change in Copper Concentration in Each Stage during the First Run. 145

xxi

Figure 4.23 Distribution of Resin in a Laboratory Adsorption Tank. 145

Figure 4.24 The Loading of Gold Determined from Solution Mass Balance,

Elution and Resin Assays for the First Run. 147

Figure 4.25 Copper and Gold Concentration on the Resin Leaving Contactor 1

during the First Run. 148

Figure 4.26 Change in Gold Concentration in Each Stage during the Second Run. 149

Figure 4.27 Change in Copper Concentration in Each Stage during the Second Run. 149


Elution and Resin Assays for the Second Run. 151


during the Second Run. 151

Figure 4.30 Change in Gold Concentration in Each Stage during the Third Run. 153

Figure 4.31 Change in Copper Concentration in Each Stage during the Third Run. 153


Elution and Resin Assays for the Third Run. 154


during the Third Run. 155

Figure 4.34 The Effect of the Loading of Gold on the Amount of Copper Extracted

by Amberjet 4200 for all Stages during Each Run. 156

Figure 4.35 The Effect of the Loading of Gold on the Extraction of Copper in

Stage 1 After Each Transfer for Each Run. 156

xxii

Figure 4.36 Profile for the Elution of Gold with Different Eluants at 2 M from

Amberjet 4200 with an Initial Loading of Gold at 1284 mg L-1. 157

Figure 4.37 Profile for the Elution of Gold with Different Nitrate Salts at 2 M

from Amberjet 4200 with an Initial Loading of Gold at 1167 mg L-1. 158

Figure 4.38 Effect of Ammonia Nitrate Concentration on the Equilibrium Loading

of Gold Thiosulfate on Amberjet 4200. 159

Figure 4.39 Effect of Ammonia Nitrate Concentration on the Equilibrium Loading

of Gold Thiosulfate on Vitrokele 911. 159

Figure 4.40 Effect of Ammonium Nitrate Concentration on the Elution of Gold

from a Column of Amberjet 4200 Loaded with 1167 mg L-1 and

1250 mg L-1 of Gold. 160

Figure 4.41 Effect of Ammonium Nitrate Concentrations on Rate of Gold Elution

from Amberjet 4200 Loaded with 85 g L-1 Gold. 161

Figure 4.42 Effect of Amminium Nitrate Concentration on Rate of Gold Elution

from Vitrokele 911 Loaded with 72 g L-1 Gold. 161

Figure 4.43 Selective Elution at a Flow Rate of 5 BV hr-1-of Copper and Gold

from Vitrokele 911 and Amberjet 4200 Resins Loaded with 1000 mg L-1

of copper and 1600 mg L-1 of Gold. 163

Figure 4.44 Selective Elution at a Flow Rate of 5 BV hr-1-of Copper and Gold

from Amberjet 4200 Resins Loaded with 550 mg L-1 of Copper and

1700 mg L-1 of Gold. 164

Figure 4.45 Elution of Thiosulfate with 1 M Ammonium Sulfate and Nitrate from

Amberjet 4200 Strong-Base Resin Loaded with 23 g L-1 Thiosulfate. 165

xxiii

Figure 4.46 Limiting Current Curves for the Reduction of Gold Thiosulfate on

Stainless Steel at pH 7.5. The Solution Contained 0.6 mM Gold

Thiosulfate and 2 M Ammonium Nitrate at a Temperature of 25 oC. 167

Figure 4.47 The Effect of pH on the Limiting Current Compared to that

Calculated from the Levich Equation. 167

Figure 5.1 The Reaction Pathway Investigation of the Copper (II) Amine to

Copper (II)Amine Thiosulfate Species. 170

Figure 5.2 Titration Curves for Polystyrene Macroporous Resin with a Tertiary

Amine Functionality (Clifford and Weber, 1983). 179

Figure 5.3 The Equilibrium Quotient for the Exchange of Sulfate Ions by Various

Anions. The Sulfate Line is Generated with a Nitrate Resin. 182

Figure 5.4 Comparison of Observed and Calculated Isotherms for the Loading of

Gold. 186

Figure 5.5 Comparison of Observed and Calculated Isotherms for the Loading of

Sulfite. 186

Figure 5.6 Equilibrium Loading of Gold Isotherms in the Presence of a Competing

Anion Predicted by the Freundlich and Langmuir Models. 189

Figure 5.7 Analysed Published Data Obtained From Clifford and Weber (1983)

for the Exchange of Nitrate ions by Sulfate or Chloride on

Strong-Base Anion Exchange Resins. 192

Figure 5.8 The Equilibrium Quotient Model for the Exchange of Sulfate Ions by

the Gold Complex. The Experimental Data are the Points and the Line

is Drawn with a Slope of Unity. 194

xxiv


the Gold Complex for the Various Strong-Base Anion Exchange

Resins Tested. The Experimental Data are the Points and the Line

is Drawn to Summarise the Slope for the Two Types of Resin. 195


Various Anions. The Experimental Data are the Points and the Line is

Drawn with a Slope of Unity. 196

Figure 5.11 The Equilibrium Loading Ratios of Gold Thiosulfate and Sulfate on the

Resin and in Solution. 198

Figure 5.12 The Equilibrium Loading Ratios of Gold Thiosulfate and Thiosulfate on the


Figure 5.13 The Equilibrium Loading Ratios of Gold Thiosulfate and Sulfite on the


Figure 5.14 The Equilibrium Loading Ratios of Gold Thiosulfate and Nitrate on the


Figure 5.15 The Equilibrium Loading Ratios of Gold Thiosulfate and Trithionate

on the Resin and in Solution. 200

Figure 5.16 The Equilibrium Loading Ratios of Gold Thiosulfate and Tetrathionate

on the Resin and in Solution. 200

Figure 5.17 Variation of the Equilibrium Constant for the Exchange of Sulfate by

Gold in the Presence of Added Sulfate. 201


Gold in the Presence of Added Thiosulfate. 201

xxv


Gold in the Presence of Added Sulfite. 202


Gold in the Presence of Added Nitrate. 202


Gold in the Presence of Added Trithionate. 203


Gold in the Presence of Added Tetrathionate. 203

Figure 5.23 The Change in the Rate of Gold Loading on Two Resins In Different

Solutions. 215

Figure 5.24 The Rate of Loading of Gold in the Presence of Varying Concentrations of

Sulfate. Experimental Data (Points) and Model (Line). 220


Thiosulfate. Experimental Data (Points) and Model (Line). 222


Sulfite. Experimental Data (Points) and Model (Line). 223


1 M S2O32- and 1 M NH3/NH4+ (Wan, 1997). 231

Figure 5.28 Comparison of the Elution of Two Resin Columns Filled With

Amberjet 4200. 232


1 M S2O32- and 1 M NH3/NH4+ (Jiayong et al., 1996). 235

xxvi

LIST OF APPENDICES

Page Appendix I Sample Calculations 1

Appendix II All Raw Data 10

Provided on CD.

xxvii

ABREVIATIONS

AAS Atomic absorbance spectrophotometer

BV Bed volume

CIC Carbon in column

CRC Corporate research centre

DIBK Diethyl isobutyl ketone

Gr Grashot number

HPLC High performance liquid chromatography

LSSS Lime sulfur synthetic solution

LHS Left hand side

NA Not applicable

NK Not known

NR Not reported or not recorded

RHS Right hand side

RIP Resin-in-pulp

RIL Resin-in-leach

rpm Revolutions per minute

Sc Schmidt number

SCE Standard calomel electrode

Sh Sherwood number

SHE Standard hydrogen electrode

TBP Tributyl phosphate

TRAO Trialkyl amine oxide

UV Ultraviolet

CHAPTER 1: Review of the Leaching of Gold Using Thiosulfate 1

1

CHAPTER 1 REVIEW OF THE LEACHING OF GOLD USING

THIOSULFATE

1.1 INTRODUCTION

MacArthur and the Forrest brothers pioneered research on the use of cyanide for the

leaching of gold ores in the 1880’s (La Brooy et al., 1994). Since then cyanidation has

superseded all other processes for the extraction of gold and silver from a wide range of

ores. While the majority of ores can be effectively treated using cyanide there is a growing

quantity of carbonaceous, telluriferous, pyritic, arsenical, manganiferous and cupriferous

gold and silver ores that are refractory to cyanidation (Hiskey and Alturi, 1988).

These ores will cause an increase in cyanide consumption, a decrease in the recovery of

gold, an increase in the overall cost of the process, and an increase in the discharge of toxic

cyano-complexes into tailings dams. Over the past twenty years, there has been an

increase in research into alternate lixiviants such as the halogens ( −−− I,Br,Cl ),

ammonia ( 3NH ), thiocyanate (−SCN ), thiourea ( 22 )NH(CS ), thiosulfate (

−232OS ),

polysulfides ( −−2

62S ), sulfite (−2

3SO ), diethylamine ( NH)HC( 252 ) and nitriles

[malononitrile ( −2)CN(CH ) and lactonitrile (−2n )CN(CH )] for the treatment of problem

ores (Sparrow and Woodcock, 1995).

Increased environmental pressure to ban or limit the use of cyanide in plants throughout

the world is also a prime motivator for research into alternatives to cyanide. Some of these

alternatives not only offer a safer and environmentally sound method of extraction but, for

some ores, the use of these lixiviants can also increase the recovery of gold (Block-Bolten

and Torma, 1986; Yen et al., 1998). With the exception of chlorine, little commercial use

has been made of alternative lixiviants but several have been tested to pilot scale plants.

The most favoured current alternative to cyanide for the treatment of problem ores is

thiosulfate (Abbruzzese et al., 1995; Michel and Frenay, 1999).


2

1.2 THE LEACHING OF GOLD WITH THIOSULFATE

1.2.1 History

It has been known for over a hundred years that gold can be leached with thiosulfate.

Thiosulfate was the main competitor to cyanide in the 1880’s when there was an increase

in research to improve gold leaching and recovery from the existing gravity and mercury

amalgamation processes. Both Russell and Von Patera developed processes that used

thiosulfate to leach gold and silver (Von Michaelis, 1987; La Brooy et al., 1994).

The Von Patera process involved subjecting the gold and silver ore to a chloridising roast

before being leached with thiosulfate. This process was used successfully in South

America for many years prior to World War II for treating largely silver sulfide ores but

very little process information was published in the literature (Flett et al., 1983). Due to

being a more complex leach system, thiosulfate was largely forgotten until it was

reinvestigated for the treatment of carbonaceous and refractory ores.

The Newmont Gold Company investigated the Von Petra process for the possible

treatment of their stockpiles of carbonaceous gold ores in the 1970’s (Jiayong et al., 1996).

Cyanide was not a viable option due to the carbonaceous content of the ore. However the

technique was found to be economic only for high-grade ores. Bacterial oxidation

followed by leaching with thiosulfate in a pilot scale heap leach operation was also studied

for treatment of a low-grade refractory stockpiles. The overall recovery from a 300,000

tonne low-grade ore heap crushed to less than 1.9 cm was approximately 55 % with a

thiosulfate consumption of 5 kg t-1. The Newmont Gold Company has since patented their

research on thiosulfate heap leaching and recovery with copper cementation (Brierley and

Hill, 1993; Wan et al., 1994).

A similar treatment process was also carried out at the La Colorada Mine at Sonora,

Mexico in the 1980’s (Von Michaelis, 1987). This large pilot scale thiosulfate plant

operated for four years treating tailings from an old cyanidation plant and also incorporated

the ideas of Kerley (Li et al., 1995; Perez and Galaviz, 1987; Qian and Jiexue, 1989; Von

Michaelis, 1987). The pulp was leached at 40 % solids with a retention time of two hours

with an ammonium thiosulfate concentration of 100 g L-1 and a copper concentration of

3 g L-1.


3

The overall recovery of silver and gold was 85 and 75 % respectively with gold and silver

recovered by cementation on copper. Unfortunately the plant was plagued with problems

which included pollution and mechanical corrosion with the result that it did not run

smoothly during operation.

More recently, Barrick Gold Corporation has re-examined leaching of gold with thiosulfate

over the past ten years and patented agitated leaching and recovery processes for the

treatment of a range of ores (Marchbank et al., 1996; Thomas et al., 1998). Another

Canadian Mining Company, Placer Dome, has also started research into leaching gold with

thiosulfate. Both companies have only tested leaching and recovery with thiosulfate in

laboratory and pilot scale plants but aim to trial a small commercial scale plant in

2002/2003.

1.2.2 Limitation and Advantages of the Leaching of Gold with Thiosulfate

Gold is dissolved by thiosulfate in the presence of ammonia and copper (II) (Equation 1):

)1(NH4)OS(Cu)OS(Au)NH(CuOS4Au 33232

3232

243

232

o ++→++ −−+−

Copper (II) is the oxidant for gold dissolution, while oxygen is needed to replenish the

oxidant in solution (Equation 2):

)2(OH4OS8)NH(Cu4NH16OHO)OS(Cu4 232243322

3232

−−+− ++→+++

Despite the fact that both thiosulfate and cyanide react with copper (II), thiosulfate has a

number of advantages over cyanide. These include being non-toxic at concentrations

tested for leaching, lower in unit cost in certain locations, and the potential to leach gold

more rapidly under normal process conditions. Thiosulfate also has the potential to

recover more gold than cyanide from ores that are high in copper, are cyanide consuming,

or contain carbonaceous materials (Flett et al., 1983; Marchbank et al., 1996; Wan et al.,

1994). The main limitations of thiosulfate as a leaching agent are the need for high reagent

concentrations compared to cyanide, its instability which leads to high reagent

consumption, the complexity of its chemistry, and the lack of a viable in-pulp method for

recovering the dissolved gold.

Compared to the cyanidation process, the thiosulfate process has a more complex

chemistry which could be expected to pose considerable control problems. Variables that

must be considered are the thiosulfate, ammonia, copper and oxygen concentrations, pH


4

and temperature. There are also a number of additives such as sulfate and sulfite that have

been suggested to assist in increasing the leaching rate of gold with thiosulfate (Jiexue and

Qian, 1991; Kerley, 1981; Kerley and Barnard, 1983). However there have been

contradictory reports in the literature on the mechanism of leaching and still disagreement

on the most appropriate combination and concentration of each reagent required to

optimise leaching.

1.2.3 Condition Employed in Leaching Gold with Thiosulfate

A wide range of conditions have been reported by researchers investigating the leaching of

copper-silver-gold bearing materials with thiosulfate are presented in Tables 1.1 to 1.3.

Thiosulfate concentrations range from 0.02-2 M, copper (II) between 0.5-120 mM and

ammonia from 0.03-4 M. The pH ranges from 8.5 to 10.5, the temperature between 20 oC

and 60 oC and the oxygen flow rate from zero to 2 L min-1. Leaching time varies from

several hours for agitation leaching to 116 days for heap leaching and the material studied

varies from solid gold, to gold ore containing sulfidic and/or carbonaceous material and

sulfide concentrates.

In general it is apparent that a higher concentration ratio of thiosulfate to ammonia results

in higher recovery of copper and silver. However, gold recoveries appear to be increased

by a higher concentration ratio of ammonia to thiosulfate at the expense of silver and

copper dissolution. Increasing temperatures up to 60 oC clearly increase the leaching

kinetics and decreases the time needed for leaching to less than 4 hours. The results also

show there is no apparent difference between leaching in the presence of air or oxygen.

Overall, because there is a wide range of conditions and reagent concentrations tested on a

number of different ores and concentrates, no obvious relationships between recovery and

the leaching conditions can be drawn. It appears that the optimum conditions reported for

leaching gold or silver could be specific to the particular ore studied. Some of the

conditions reported are extreme in terms of reagent concentrations and more economical

levels need to be considered if the process is to be an attractive alternative to cyanide.

Furthermore, research needs to be focused on ambient temperatures, as leaching at elevated

temperatures is generally uneconomical for low-grade ores. In addition, this would allow

for a more direct comparison to cyanidation.


5

Cu

(II)

(M)

Au

Ext

(%)

(Toz

awa

et a

l., 1

981)

A

nnea

led

gold

pl

ates

99.9

9 %

Au

0.5

1.0

NR

0.04

0.04

100a

NR

653

NR

NR

NA

(Des

chen

es, 1

998)

Gol

d Fo

il99

.99

% A

u0.

10.

5N

RN

R0.

040.

2110

2524

NR

55N

A

(Mur

thy

and

Pras

ad,

1996

)D

ross

752

gt-1

Au

7019

gt-1

Ag

0.32

% C

u

0.5

NR

NR

NR

NR

12N

R60

620

38ca

lcin

e17

97ca

lcin

e16

(Abb

ruzz

ese

et a

l., 1

995)

Gol

d or

e51

.6 g

t-1A

u2.

04.

0N

R0.

10.

10.

218.

5-10

.525

440

79N

A

(Ker

ley

and

Bar

nard

, 19

83)

Gol

d or

e0.

43 g

t-1A

u37

5.5

gt-1

Ag

2.1

%M

n

0.73

0.26

0.01

0.06

0.06

0.21

950

630

8793

(Lan

ghan

s et a

l., 1

992)

Gol

d or

e (O

xidi

zed)

1.65

gt-1

Au

0.02

% C

u

0.2

0.09

0.00

60.

001

0.00

10.

2111

2548

5083

NA

(Per

ez a

ndG

alav

iz, 1

987)

Add

ed a

s cop

per

amm

oniu

m th

iosu

lfate

Gol

d or

e1.

33 g

t-1A

u18

6 g

t-1A

g3.

8 %

Mn

0.4

NR

0.86

NR

0.05

NR

1050

340

9890

(Zip

peria

net

al.,

198

8)G

old

ore

(Ryo

lite)

3 g

t-1A

u11

3 g

t-1A

g7

gkg

-1M

nO2

24

0.1

NR

0.1

0.21

1050

240

9060

-75

Tab

le 1

.1 T

he L

each

ing

Con

ditio

ns a

nd R

ecov

ery

From

Gol

d an

d G

old

Ore

s.

a=

kP

aN

R=

N

ot R

epor

ted

NA

=

Not

App

licab

le

Ref

eren

ceM

ater

ial

S 2O

32-

(M)

NH

3(M

)SO

32-

(M)

SO42

-

(M)

O2

flow

rate

(Lm

in-1

)pH

Tem

p (o

C)

Tim

e (h

ours

)So

lids

(%)

Ag

Ext

(%)

Cu

(II)

(M)

Cu

(II)

(M)

Au

Ext

(%)

Au

Ext

(%)

(Toz

awa

et a

l., 1

981)

A

nnea

led

gold

pl

ates

99.9

9 %

Au

0.5

1.0

NR

0.04

0.04

100a

NR

653

NR

NR

NA

(Des

chen

es, 1

998)

Gol

d Fo

il99

.99

% A

u0.

10.

5N

RN

R0.

040.

2110

2524

NR

55N

A

(Mur

thy

and

Pras

ad,

1996

)D

ross

752

gt-1

Au

7019

gt-1

Ag

0.32

% C

u

0.5

NR

NR

NR

NR

12N

R60

620

38ca

lcin

e17

97ca

lcin

e16

(Abb

ruzz

ese

et a

l., 1

995)

Gol

d or

e51

.6 g

t-1A

u2.

04.

0N

R0.

10.

10.

218.

5-10

.525

440

79N

A

(Ker

ley

and

Bar

nard

, 19

83)

Gol

d or

e0.

43 g

t-1A

u37

5.5

gt-1

Ag

2.1

%M

n

0.73

0.26

0.01

0.06

0.06

0.21

950

630

8793

(Lan

ghan

s et a

l., 1

992)

Gol

d or

e (O

xidi

zed)

1.65

gt-1

Au

0.02

% C

u

0.2

0.09

0.00

60.

001

0.00

10.

2111

2548

5083

NA

(Per

ez a

ndG

alav

iz, 1

987)

Add

ed a

s cop

per

amm

oniu

m th

iosu

lfate

Gol

d or

e1.

33 g

t-1A

u18

6 g

t-1A

g3.

8 %

Mn

0.4

NR

0.86

NR

0.05

NR

1050

340

9890

(Zip

peria

net

al.,

198

8)G

old

ore

(Ryo

lite)

3 g

t-1A

u11

3 g

t-1A

g7

gkg

-1M

nO2

24

0.1

NR

0.1

0.21

1050

240

9060

-75

Tab

le 1

.1 T

he L

each

ing

Con

ditio

ns a

nd R

ecov

ery

From

Gol

d an

d G

old

Ore

s.

a=

kP

aN

R=

N

ot R

epor

ted

NA

=

Not

App

licab

le

Ref

eren

ceM

ater

ial

S 2O

32-

(M)

NH

3(M

)SO

32-

(M)

SO42

-

(M)

O2

flow

rate

(Lm

in-1

)pH

Tem

p (o

C)

Tim

e (h

ours

)So

lids

(%)

Ag

Ext

(%)

(Toz

awa

et a

l., 1

981)

A

nnea

led

gold

pl

ates

99.9

9 %

Au

0.5

1.0

NR

0.04

0.04

100a

NR

653

NR

NR

NA

(Des

chen

es, 1

998)

Gol

d Fo

il99

.99

% A

u0.

10.

5N

RN

R0.

040.

2110

2524

NR

55N

A

(Mur

thy

and

Pras

ad,

1996

)D

ross

752

gt-1

Au

7019

gt-1

Ag

0.32

% C

u

0.5

NR

NR

NR

NR

12N

R60

620

38ca

lcin

e17

97ca

lcin

e16

(Abb

ruzz

ese

et a

l., 1

995)

Gol

d or

e51

.6 g

t-1A

u2.

04.

0N

R0.

10.

10.

218.

5-10

.525

440

79N

A

(Ker

ley

and

Bar

nard

, 19

83)

Gol

d or

e0.

43 g

t-1A

u37

5.5

gt-1

Ag

2.1

%M

n

0.73

0.26

0.01

0.06

0.06

0.21

950

630

8793

(Lan

ghan

s et a

l., 1

992)

Gol

d or

e (O

xidi

zed)

1.65

gt-1

Au

0.02

% C

u

0.2

0.09

0.00

60.

001

0.00

10.

2111

2548

5083

NA

(Per

ez a

ndG

alav

iz, 1

987)

Add

ed a

s cop

per

amm

oniu

m th

iosu

lfate

Gol

d or

e1.

33 g

t-1A

u18

6 g

t-1A

g3.

8 %

Mn

0.4

NR

0.86

NR

0.05

NR

1050

340

9890

(Zip

peria

net

al.,

198

8)G

old

ore

(Ryo

lite)

3 g

t-1A

u11

3 g

t-1A

g7

gkg

-1M

nO2

24

0.1

NR

0.1

0.21

1050

240

9060

-75

Tab

le 1

.1 T

he L

each

ing

Con

ditio

ns a

nd R

ecov

ery

From

Gol

d an

d G

old

Ore

s.

a=

kP

aN

R=

N

ot R

epor

ted

NA

=

Not

App

licab

le

(Toz

awa

et a

l., 1

981)

(T

ozaw

aet

al.,

198

1)

Ann

eale

d go

ld

plat

es99

.99

% A

u

Ann

eale

d go

ld

plat

es99

.99

% A

u

0.5

0.5

1.0

1.0

NR

NR

0.04

0.04

0.04

0.04

100a

100a

NR

NR

656533

NR

NR

NR

NR

NA

NA

(Des

chen

es, 1

998)

(Des

chen

es, 1

998)

Gol

d Fo

il99

.99

% A

uG

old

Foil

99.9

9 %

Au

0.1

0.1

0.5

0.5

NR

NR

NR

NR

0.04

0.04

0.21

0.21

10102525

2424N

RN

R5555

NA

NA

(Mur

thy

and

Pras

ad,

1996

)(M

urth

yan

dPr

asad

, 19

96)

Dro

ss75

2 g

t-1A

u70

19 g

t-1A

g0.

32 %

Cu

Dro

ss75

2 g

t-1A

u70

19 g

t-1A

g0.

32 %

Cu

0.5

0.5

NR

NR

NR

NR

NR

NR

NR

NR

1212N

RN

R6060

662020

38ca

lcin

e1738

calc

ine

17

97ca

lcin

e1697

calc

ine

16(A

bbru

zzes

eet

al.,

199

5)(A

bbru

zzes

eet

al.,

199

5)G

old

ore

51.6

gt-1

Au

Gol

d or

e51

.6 g

t-1A

u2.

02.

04.

04.

0N

RN

R0.

10.

10.

10.

10.

210.

218.

5-10

.58.

5-10

.52525

444040

7979N

AN

A

(Ker

ley

and

Bar

nard

, 19

83)

(Ker

ley

and

Bar

nard

, 19

83)

Gol

d or

e0.

43 g

t-1A

u37

5.5

gt-1

Ag

2.1

%M

n

Gol

d or

e0.

43 g

t-1A

u37

5.5

gt-1

Ag

2.1

%M

n

0.73

0.73

0.26

0.26

0.01

0.01

0.06

0.06

0.06

0.06

0.21

0.21

995050

663030

87879393

(Lan

ghan

s et a

l., 1

992)

(Lan

ghan

s et a

l., 1

992)

Gol

d or

e (O

xidi

zed)

1.65

gt-1

Au

0.02

% C

u

Gol

d or

e (O

xidi

zed)

1.65

gt-1

Au

0.02

% C

u

0.2

0.2

0.09

0.09

0.00

60.

006

0.00

10.

001

0.00

10.

001

0.21

0.21

11112525

48485050

8383N

AN

A

(Per

ez a

ndG

alav

iz, 1

987)

Add

ed a

s cop

per

amm

oniu

m th

iosu

lfate

(Per

ez a

ndG

alav

iz, 1

987)

Add

ed a

s cop

per

amm

oniu

m th

iosu

lfate

Gol

d or

e1.

33 g

t-1A

u18

6 g

t-1A

g3.

8 %

Mn

Gol

d or

e1.

33 g

t-1A

u18

6 g

t-1A

g3.

8 %

Mn

0.4

0.4

NR

NR

0.86

0.86

NR

NR

0.05

0.05

NR

NR

10105050

334040

98989090

(Zip

peria

net

al.,

198

8)(Z

ippe

rian

et a

l., 1

988)

Gol

d or

e (R

yolit

e)3

gt-1

Au

113

gt-1

Ag

7 g

kg-1

MnO

2

Gol

d or

e (R

yolit

e)3

gt-1

Au

113

gt-1

Ag

7 g

kg-1

MnO

2

2244

0.1

0.1

NR

NR

0.1

0.1

0.21

0.21

10105050

224040

909060

-75

60-7

5

Tab

le 1

.1 T

he L

each

ing

Con

ditio

ns a

nd R

ecov

ery

From

Gol

d an

d G

old

Ore

s.

a=

kP

aN

R=

N

ot R

epor

ted

NA

=

Not

App

licab

le

Ref

eren

ceM

ater

ial

Mat

eria

lS 2

O32

-

(M)

S 2O

32-

(M)

NH

3(M

)N

H3

(M)

SO32

-

(M)

SO32

-

(M)

SO42

-

(M)

SO42

-

(M)

O2

flow

rate

(Lm

in-1

)O

2flo

wra

te(L

min

-1)

pHpHTe

mp

(oC

)Te

mp

(oC

)Ti

me

(hou

rs)

Tim

e (h

ours

)So

lids

(%)

Solid

s (%

)A

g Ex

t (%

)A

g Ex

t (%

)


6

Ref

eren

ceM

ater

ial

S 2O

32-

(M)

NH

3(M

)SO

32-

(M)

SO42

-

(M)

Cu

(II)

(M)

O2

flow

rate

(Lm

in-1

)pH

Tem

p (o

C)

Tim

e (h

ours

)So

lids

(%)

Au

Ext

(%)

Ag

Ext

(%)

(Ber

zow

sky

and

Sefto

n, 1

978)

Sulfi

de C

once

ntra

te3.

46-7

.27

gt-1

Au

115-

454

gt-1

Ag

23.2

-25.

3 %

Cu

32.8

-36.

4 %

S

0.3-

0.7

NR

NR

0.05

-0.0

80.

05-0

.08

N2

1035

-50

3-5

40-6

088

-95

83-9

8

(Blo

ck-B

olte

nan

dTo

rma,

198

6)Su

lfide

Con

cent

rate

1.75

gt-1

Au

22.5

gt-1

Ag

0.4

% C

u9.

8 %

S

0.1-

0.5

1N

RN

RN

R2

10.5

2nd

step

9 1

stst

ep50

125

99 2

ndst

ep89

1st

step

27 2

nd

step

(Blo

ck-B

olte

net

al.,

19

85 a

, b)

Sulfi

de C

once

ntra

te1.

75 g

t-1A

u22

.5 g

t-1A

g0.

4 %

Cu

9.8

% S

0.5

1N

RN

RN

R2

1048

330

95N

A

(Cha

nglin

et a

l.,

1992

)

Sulfi

de C

once

ntra

te62

gt-1

Au

3.19

% C

u

0.2-

0.3

2-4

0.05

NR

0.05

0.21

1060

1-2

2095

NA

(Cha

nglin

et a

l.,

1992

)Su

lfide

Con

cent

rate

62 g

t-1A

u60

gt-1

Ag

3.2

% C

u20

.6 %

S

0.2-

0.3

2-4

0.5-

0.8

0.05

0.05

19.

5-10

.560

1-2

2595

NR

(Gro

udev

et a

l.,

1996

)aSu

lfide

Con

cent

rate

70 g

t-1A

u37

.8 %

S

0.2

NR

NR

NR

NR

0.1

9.5

20-5

012

1697 12

un

treat

ed

NA

(Jie

xue

and

Qia

n,

1991

)Su

lfide

Con

cent

rate

50.4

gt-1

Au

0.04

8 %

MnO

23.

19 %

Cu

20.6

% S

1.0

2.0

NR

0.32

0.02

1N

R40

117

96N

A

(Qia

nan

dJi

exue

, 19

88)

Sulfi

de C

once

ntra

te50

.4 g

t-1A

u3.

19 %

Cu

0.42

% C

20.6

% S

0.8-

1.0

1.8-

2.2

0.1

0.02

0.02

110

.540

-50

1-2

33-7

588

-96

NA

Tab

le 1

.2

The

Lea

chin

g C

ondi

tions

and

Rec

over

y Fr

om S

ulfid

e C

once

ntra

tes.

a=

3-

5 g.

L-1

prot

ein

hydr

olys

ate

and

60 %

bac

teria

oxi

dize

dN

R=

N

ot R

epor

ted

NA

=

Not

App

licab

le

Ref

eren

ceM

ater

ial

Mat

eria

lS 2

O32

-

(M)

S 2O

32-

(M)

NH

3(M

)N

H3

(M)

SO32

-

(M)

SO32

-

(M)

SO42

-

(M)

SO42

-

(M)

Cu

(II)

(M)

Cu

(II)

(M)

O2

flow

rate

(Lm

in-1

)O

2flo

wra

te(L

min

-1)

pHpHTe

mp

(oC

)Te

mp

(oC

)Ti

me

(hou

rs)

Tim

e (h

ours

)So

lids

(%)

Solid

s (%

)A

u Ex

t (%

)A

u Ex

t (%

)A

g Ex

t (%

)A

g Ex

t (%

)(B

erzo

wsk

yan

d Se

fton,

197

8)Su

lfide

Con

cent

rate

3.46

-7.2

7 g

t-1A

u11

5-45

4 g

t-1A

g23

.2-2

5.3

% C

u32

.8-3

6.4

% S

Sulfi

de C

once

ntra

te3.

46-7

.27

gt-1

Au

115-

454

gt-1

Ag

23.2

-25.

3 %

Cu

32.8

-36.

4 %

S

0.3-

0.7

0.3-

0.7

NR

NR

NR

NR

0.05

-0.0

80.

05-0

.08

0.05

-0.0

80.

05-0

.08

N2

N2

101035

-50

35-5

03-

53-

540

-60

40-6

088

-95

88-9

583

-98

83-9

8

(Blo

ck-B

olte

nan

dTo

rma,

198

6)(B

lock

-Bol

ten

and

Torm

a, 1

986)

Sulfi

de C

once

ntra

te1.

75 g

t-1A

u22

.5 g

t-1A

g0.

4 %

Cu

9.8

% S

Sulfi

de C

once

ntra

te1.

75 g

t-1A

u22

.5 g

t-1A

g0.

4 %

Cu

9.8

% S

0.1-

0.5

0.1-

0.5

11N

RN

RN

RN

RN

RN

R22

10.5

2nd

step

9 1

stst

ep10

.5 2

ndst

ep9

1st

step

505011

252599

2nd

step

89 1

stst

ep99

2nd

step

89 1

stst

ep27

2nd

step

27 2

nd

step

(Blo

ck-B

olte

net

al.,

19

85 a

, b)

(Blo

ck-B

olte

net

al.,

19

85 a

, b)

Sulfi

de C

once

ntra

te1.

75 g

t-1A

u22

.5 g

t-1A

g0.

4 %

Cu

9.8

% S

Sulfi

de C

once

ntra

te1.

75 g

t-1A

u22

.5 g

t-1A

g0.

4 %

Cu

9.8

% S

0.5

0.5

11N

RN

RN

RN

RN

RN

R22

10104848

333030

9595N

AN

A

(Cha

nglin

et a

l.,

1992

)(C

hang

linet

al.,

19

92)

Sulfi

de C

once

ntra

te62

gt-1

Au

3.19

% C

u

Sulfi

de C

once

ntra

te62

gt-1

Au

3.19

% C

u

0.2-

0.3

0.2-

0.3

2-4

2-4

0.05

0.05

NR

NR

0.05

0.05

0.21

0.21

10106060

1-2

1-2

20209595

NA

NA

(Cha

nglin

et a

l.,

1992

)(C

hang

linet

al.,

19

92)

Sulfi

de C

once

ntra

te62

gt-1

Au

60 g

t-1A

g3.

2 %

Cu

20.6

% S

Sulfi

de C

once

ntra

te62

gt-1

Au

60 g

t-1A

g3.

2 %

Cu

20.6

% S

0.2-

0.3

0.2-

0.3

2-4

2-4

0.5-

0.8

0.5-

0.8

0.05

0.05

0.05

0.05

119.

5-10

.59.

5-10

.56060

1-2

1-2

25259595

NR

NR

(Gro

udev

et a

l.,

1996

)a(G

roud

evet

al.,

19

96)a

Sulfi

de C

once

ntra

te70

gt-1

Au

37.8

% S

Sulfi

de C

once

ntra

te70

gt-1

Au

37.8

% S

0.2

0.2

NR

NR

NR

NR

NR

NR

NR

NR

0.1

0.1

9.5

9.5

20-5

020

-50

12121616

97 12

untre

ated

97 12

untre

ated

NA

NA

(Jie

xue

and

Qia

n,

1991

)(J

iexu

ean

dQ

ian,

19

91)

Sulfi

de C

once

ntra

te50

.4 g

t-1A

u0.

048

% M

nO2

3.19

% C

u20

.6 %

S

Sulfi

de C

once

ntra

te50

.4 g

t-1A

u0.

048

% M

nO2

3.19

% C

u20

.6 %

S

1.0

1.0

2.0

2.0

NR

NR

0.32

0.32

0.02

0.02

11N

RN

R4040

111717

9696N

AN

A

(Qia

nan

dJi

exue

, 19

88)

(Qia

nan

dJi

exue

, 19

88)

Sulfi

de C

once

ntra

te50

.4 g

t-1A

u3.

19 %

Cu

0.42

% C

20.6

% S

Sulfi

de C

once

ntra

te50

.4 g

t-1A

u3.

19 %

Cu

0.42

% C

20.6

% S

0.8-

1.0

0.8-

1.0

1.8-

2.2

1.8-

2.2

0.1

0.1

0.02

0.02

0.02

0.02

1110

.510

.540

-50

40-5

01-

21-

233

-75

33-7

588

-96

88-9

6N

AN

A

Tab

le 1

.2

The

Lea

chin

g C

ondi

tions

and

Rec

over

y Fr

om S

ulfid

e C

once

ntra

tes.

a=

3-

5 g.

L-1

prot

ein

hydr

olys

ate

and

60 %

bac

teria

oxi

dize

dN

R=

N

ot R

epor

ted

NA

=

Not

App

licab

le


7

pH

(Cha

i, 19

97)

Sulfi

de o

re1.

06 g

t-1A

u0.

73 %

S

0.3

2N

R0.

90.

121

10.5

602.

54

90N

A

(Gro

udev

et a

l., 1

994,

Gro

udev

et a

l., 1

995)

48 %

bac

teria

l oxi

dize

d1

g.L-

1pr

otei

n hy

drol

ysat

e1

g.L-

1io

n ca

taly

st

Sulfi

de o

re3.

2 g

t-1A

u15

.2 g

t-1A

g0.

8 %

S

0.07

NR

NR

NR

0.01

80.

219.

5N

R20

bH

eap

leac

h70

.720

.3

untre

ated

51.2

14.3

un

treat

ed

(Bha

duri,

198

7)Su

lfide

/Car

bona

ceou

s ore

5

gt-1

Au

0.62

% C

0.67

% S

0.2

30.

090.

040.

0410

3a10

.535

220

80N

A

(Mar

chba

nket

al.,

199

6)Su

lfide

/Car

bona

ceou

s or

e3-

7 g

t-1A

u1.

2 %

C1.

22-2

.54

% S

0.02

-0.1

0.03

0.01

-0.0

5N

R0.

010.

217-

955

440

70-8

5N

A

(Wan

and

Brie

rley,

199

7)Su

lfide

/Car

bona

ceou

s or

e1-

3 g

t-1A

u0.

67-2

.42

% C

0.78

-1.3

5 %

S

0.1

0.1

0.00

1N

R0.

001

0.21

925

116b

Hea

p le

ach

65N

A

(Wan

et a

l., 1

994)

Sulfi

de/C

arbo

nace

ous

ore

2.4

gt-1

Au

1.4

% C

1 %

S

0.1-

0.2

0.1

0.00

1N

R0.

002

0.21

9-10

2512

-25b

Hea

p le

ach

40-6

0N

A

(Bha

duri,

198

7)C

arbo

nace

ous

ore

8.67

gt-1

Au

1 %

C

1.4

30.

230.

060.

0610

3a8

601.

520

75N

A

(Hem

mat

i, 19

87)

Car

bona

ceou

s or

e14

.74

gt-1

Au

2.5

% o

rgan

ic C

arbo

n

0.7

30.

150.

20.

1510

3a10

.535

420

71N

A

Tab

le 1

.3

The

Lea

chin

g C

ondi

tions

and

Rec

over

y Fr

omSu

lfidi

c an

d C

arbo

nace

ous O

res.

a=

kP

ab

=

days

NR

=

Not

Rep

orte

d

NA

=

Not

App

licab

le

Ref

eren

ceM

ater

ial

S 2O

32-

(M)

NH

3(M

)SO

32-

(M)

SO42

-

(M)

Cu

(II)

(M)

O2

flow

rate

(Lm

in-1

)Te

mp

(oC

)Ti

me

(hou

rs)

Solid

s (%

)A

u Ex

t (%

)A

g Ex

t (%

)pHpH

(Cha

i, 19

97)

Sulfi

de o

re1.

06 g

t-1A

u0.

73 %

S

0.3

2N

R0.

90.

121

10.5

602.

54

90N

A

(Gro

udev

et a

l., 1

994,

Gro

udev

et a

l., 1

995)

48 %

bac

teria

l oxi

dize

d1

g.L-

1pr

otei

n hy

drol

ysat

e1

g.L-

1io

n ca

taly

st

Sulfi

de o

re3.

2 g

t-1A

u15

.2 g

t-1A

g0.

8 %

S

0.07

NR

NR

NR

0.01

80.

219.

5N

R20

bH

eap

leac

h70

.720

.3

untre

ated

51.2

14.3

un

treat

ed

(Bha

duri,

198

7)Su

lfide

/Car

bona

ceou

s ore

5

gt-1

Au

0.62

% C

0.67

% S

0.2

30.

090.

040.

0410

3a10

.535

220

80N

A

(Mar

chba

nket

al.,

199

6)Su

lfide

/Car

bona

ceou

s or

e3-

7 g

t-1A

u1.

2 %

C1.

22-2

.54

% S

0.02

-0.1

0.03

0.01

-0.0

5N

R0.

010.

217-

955

440

70-8

5N

A

(Wan

and

Brie

rley,

199

7)Su

lfide

/Car

bona

ceou

s or

e1-

3 g

t-1A

u0.

67-2

.42

% C

0.78

-1.3

5 %

S

0.1

0.1

0.00

1N

R0.

001

0.21

925

116b

Hea

p le

ach

65N

A

(Wan

et a

l., 1

994)

Sulfi

de/C

arbo

nace

ous

ore

2.4

gt-1

Au

1.4

% C

1 %

S

0.1-

0.2

0.1

0.00

1N

R0.

002

0.21

9-10

2512

-25b

Hea

p le

ach

40-6

0N

A

(Bha

duri,

198

7)C

arbo

nace

ous

ore

8.67

gt-1

Au

1 %

C

1.4

30.

230.

060.

0610

3a8

601.

520

75N

A

(Hem

mat

i, 19

87)

Car

bona

ceou

s or

e14

.74

gt-1

Au

2.5

% o

rgan

ic C

arbo

n

0.7

30.

150.

20.

1510

3a10

.535

420

71N

A

Tab

le 1

.3

The

Lea

chin

g C

ondi

tions

and

Rec

over

y Fr

omSu

lfidi

c an

d C

arbo

nace

ous O

res.

a=

kP

ab

=

days

NR

=

Not

Rep

orte

d

NA

=

Not

App

licab

le

Ref

eren

ceM

ater

ial

S 2O

32-

(M)

NH

3(M

)SO

32-

(M)

SO42

-

(M)

Cu

(II)

(M)

O2

flow

rate

(Lm

in-1

)Te

mp

(oC

)Ti

me

(hou

rs)

Solid

s (%

)A

u Ex

t (%

)A

g Ex

t (%

)

(Cha

i, 19

97)

Sulfi

de o

re1.

06 g

t-1A

u0.

73 %

S

0.3

2N

R0.

90.

121

10.5

602.

54

90N

A

(Gro

udev

et a

l., 1

994,

Gro

udev

et a

l., 1

995)

48 %

bac

teria

l oxi

dize

d1

g.L-

1pr

otei

n hy

drol

ysat

e1

g.L-

1io

n ca

taly

st

Sulfi

de o

re3.

2 g

t-1A

u15

.2 g

t-1A

g0.

8 %

S

0.07

NR

NR

NR

0.01

80.

219.

5N

R20

bH

eap

leac

h70

.720

.3

untre

ated

51.2

14.3

un

treat

ed

(Bha

duri,

198

7)Su

lfide

/Car

bona

ceou

s ore

5

gt-1

Au

0.62

% C

0.67

% S

0.2

30.

090.

040.

0410

3a10

.535

220

80N

A

(Mar

chba

nket

al.,

199

6)Su

lfide

/Car

bona

ceou

s or

e3-

7 g

t-1A

u1.

2 %

C1.

22-2

.54

% S

0.02

-0.1

0.03

0.01

-0.0

5N

R0.

010.

217-

955

440

70-8

5N

A

(Wan

and

Brie

rley,

199

7)Su

lfide

/Car

bona

ceou

s or

e1-

3 g

t-1A

u0.

67-2

.42

% C

0.78

-1.3

5 %

S

0.1

0.1

0.00

1N

R0.

001

0.21

925

116b

Hea

p le

ach

65N

A

(Wan

et a

l., 1

994)

Sulfi

de/C

arbo

nace

ous

ore

2.4

gt-1

Au

1.4

% C

1 %

S

0.1-

0.2

0.1

0.00

1N

R0.

002

0.21

9-10

2512

-25b

Hea

p le

ach

40-6

0N

A

(Bha

duri,

198

7)C

arbo

nace

ous

ore

8.67

gt-1

Au

1 %

C

1.4

30.

230.

060.

0610

3a8

601.

520

75N

A

(Hem

mat

i, 19

87)

Car

bona

ceou

s or

e14

.74

gt-1

Au

2.5

% o

rgan

ic C

arbo

n

0.7

30.

150.

20.

1510

3a10

.535

420

71N

A

Tab

le 1

.3

The

Lea

chin

g C

ondi

tions

and

Rec

over

y Fr

omSu

lfidi

c an

d C

arbo

nace

ous O

res.

a=

kP

ab

=

days

NR

=

Not

Rep

orte

d

NA

=

Not

App

licab

le

(Cha

i, 19

97)

(Cha

i, 19

97)

Sulfi

de o

re1.

06 g

t-1A

u0.

73 %

S

Sulfi

de o

re1.

06 g

t-1A

u0.

73 %

S

0.3

0.3

22N

RN

R0.

90.

90.

120.

1211

10.5

10.5

60602.

52.

544

9090N

AN

A

(Gro

udev

et a

l., 1

994,

Gro

udev

et a

l., 1

995)

48 %

bac

teria

l oxi

dize

d1

g.L-

1pr

otei

n hy

drol

ysat

e1

g.L-

1io

n ca

taly

st

(Gro

udev

et a

l., 1

994,

Gro

udev

et a

l., 1

995)

48 %

bac

teria

l oxi

dize

d1

g.L-

1pr

otei

n hy

drol

ysat

e1

g.L-

1io

n ca

taly

st

Sulfi

de o

re3.

2 g

t-1A

u15

.2 g

t-1A

g0.

8 %

S

Sulfi

de o

re3.

2 g

t-1A

u15

.2 g

t-1A

g0.

8 %

S

0.07

0.07

NR

NR

NR

NR

NR

NR

0.01

80.

018

0.21

0.21

9.5

9.5

NR

NR

20b

20b

Hea

p le

ach

Hea

p le

ach

70.7

20.3

un

treat

ed

70.7

20.3

un

treat

ed

51.2

14.3

un

treat

ed

51.2

14.3

un

treat

ed

(Bha

duri,

198

7)(B

hadu

ri, 1

987)

Sulfi

de/C

arbo

nace

ous o

re

5 g

t-1A

u0.

62 %

C0.

67 %

S

Sulfi

de/C

arbo

nace

ous o

re

5 g

t-1A

u0.

62 %

C0.

67 %

S

0.2

0.2

330.

090.

090.

040.

040.

040.

0410

3a10

3a10

.510

.53535

222020

8080N

AN

A

(Mar

chba

nket

al.,

199

6)(M

arch

bank

et a

l., 1

996)

Sulfi

de/C

arbo

nace

ous

ore

3-7

gt-1

Au

1.2

% C

1.22

-2.5

4 %

S

Sulfi

de/C

arbo

nace

ous

ore

3-7

gt-1

Au

1.2

% C

recovery of gold from thiosulfate solutions and pulps with ......leach pulp to yield loadings of...

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