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OXIDATIVE PRESSURE LEACHING

OF CHALCOCITE IN SULPHURIC ACID

by

Ishwinder Singh Grewal

B.A.Sc, The University ofBritish Columbia, 1989

A THESIS SUBMITTED IN PARTIAL FULFILLMENT OF

THE REQUIREMENTS FORTHE DEGREE OF

MASTEROF APPLIED SCIENCE

in

THE FACULTY OF GRADUATE STUDIES

Department of Metals and Materials Engineering

We accept this thesis as conforming

to the required standard

THE UNIVERSITY OF BRITISH COLUMBIA

October 1991

Ishwinder Singh Grewal, 1991

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In presenting this thesis in partial fulfilment of the requirements for an advanced

degree at the University of British Columbia, I agree that the Library shall make it

freely available for reference and study. I further agree that permission for extensive

copying of this thesis for scholarly purposes may be granted by the head of my

department or by his or her representatives. It is understood that copying or

publication of this thesis for financial gain shall not be allowed without my written

permission.

Department of (V\gjbn\< \ - Mc\e* .qU Bvqc^e^nWj

The University of British Columbia

Vancouver, Canada

Date O C T - M

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ABSTRACT

At INCO's Copper CliffCopper Refinery, a copper sulphide residue containing precious

metals is subjected to a pressure leach at 115C Copper occurs predominantly as G12S and

Cu : %S. The leach produces a slurry of copper sulphate solution and basic copper sulphate and

precious metals solids. The basic copper sulphate is dissolved in spent acid from the

electrowinning tankhouse leaving a precious metals residue for further processing. The leach

periodically develops a problem referred to as a "slow cook" where leaching times are greatly

extended and/or incomplete leaching ofcopper is encountered.

Based on earlier work, chalcocite leaching was proposed to occur sequentially via the

following reactions:

1. Cw,S+ //2S04+102 - CuS+ CuSCuS+ Cu{OH\ CuSOA

3. Cw1S+^02+H20 -*^CuS+^(2Cu(OH) 2'CuSOt)

4. CuS + 202-+CuSOt

5. CuS+H^ + ^-^CuSO^+H^ +S

The exact reaction path is determined by the initial solution conditions (copper sulphate

and sulphuric acid concentrations). Under normal batch makeup conditions, all of the CU2S is

oxidized to cupric ions and sulphate via reactions 1,2 and 4. If the solution becomes depleted in

copper and acid, reaction 3 could occur. Elemental sulphur can be produced via reaction 5.

Experimental studies showed that the reactions were nearly sequential. Reactions 1 and 2

were found to be very fast relative to the rate of reaction 4. No slow cook conditions were

observed in the laboratory under normal leaching conditions. There is evidence suggesting that

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the slow cooks are caused by oxygen mass transfer limitations under conditions where the

slurry becomes highly viscous and pseudoplastic due to formation of finely divided basic

copper sulphate.

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Table of Contents

ABSTRACT ii

Table of Tables vi

Table of Figures viiAcknowledgements xi

CHAPTER 1 - Introduction 1

CHAPTER2 - Background and Literature Review 4

2.1 Overview of the INCO-CRED Process 4

2.1.1 Background of the Second Stage Leach 7

2.1.2 The Second Stage Leach 10

2.2 The UBC Screening Model 12

2.3 Scope of the Study 21

2.4 Literature Review 22

2.4.1 Copper Sulphides - Chalcocite to Covellite 22

2.4.2 Leaching of Chalcocite and Covellite 23

2.4.3 Electrochemical Studies 30

2.4.4 Eh -pH Relationships and Phase Systems 33

2.4.5 Gas-Liquid Mass Transfer in Oxidative Leaching 38

2.5 Summary 41

CHAPTER3 - Experimental Methods 43

3.1 Part A: Study of the sequential nature of the reactions 44

3.1.1 Batch make-up chemistry 44

3.1.2 Experimental procedure 45

3.1.3 Additional experiments in Part A 46

3.2 Part B: Kinetic Experiments 47

3.3 Part C: Leaching CuS in the presence of basic copper sulphate 48

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CHAPTER4 - Results and Discussion 50

4.1 Part A: Study of the sequential nature of the reactions 50

4.1.1 The behavior of copper dissolution 50

4.1.2 Iron and Arsenic 58

4.1.3 Nickel and Cobalt 61

4.1.4 Additional Observations ofPart A Experiments 63

4.2 Part B: Kinetic Experiments 63

4.3 Part C: Leaching CuS in the presence of basic copper sulphate 66

4.3.1 Discussion ofORP Measurements 71

4.4 Comparison ofLeaching Rates - Part A and Part C 72

4.5 Additional Observations 73

CHAPTER5 - Conclusions and Recommendations 76

5.1 Conclusions 76

5.2 Recommendations for further work 77

REFERENCES 79

APPENDIX A - Detailed Flowsheet of the CRED Plant 82

APPENDIX B - Planned Experiments for Part B-1 83

APPENDIX C - Assay results ofPart A experiments 84

APPENDIX D - Part B experimental results 90

APPENDIX E - Part C experimental results 101

APPENDIX F - Part C experimental results - Effect ofiron plots 107

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Table of Tables

Table 2.1 Typical Assay ofIPC Residue 4

Table 3.1 Assay ofFirst Stage Residue used in Part A Experiments 44

Table 3.2 List ofpart C experiments performed 49

Table 4.1 X-Ray Diffraction Results 54

Table 4.2 Degree ofsulphur oxidation 65

Table 4.3 Time taken to leach to 10,15 and 20% oxygen consumption 66

Table 4.4 Potential and pH measurements of the leach slurry after leaching and the

approximate temperatures at which they were measured 69

Table 4.5 Comaparison of oxygen flow rates 73

Table Bl. List of the planned experiments for part B 83

Table Cl. Amount of the indicated species in the leach solution at various oxygen

consumption levels '. 84

Table C2. Amount of the indicated species in the releach solution at various oxygen

consumption levels : 85

Table C3. Amount of the indicated species in the releach cake at various oxygen

consumption levels 86

Table C4. Total amount ofspecies (calculated as a sum oftables C1-C3) in the process

87

Table C5. Distribution ofspecies of experiments performed with the "copper

depleted cake" 88

Table C6. The pH and ORP values of the slurry after leaching to a given level of

oxygen consumption 89

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Table of Figures

Figure 1.1 Classification of leaching methods 2

Figure 2.1 General processing route of the IPC residue at the INCO-CRED plant in

Copper Cliff 6

Figure 2.2 Second Stage Leaching Circuit at INCO's - Copper CliffCopper Refinery... 10

Figure 2.3 Approximate shape and dimensions of the Second Stage Autoclaves 11

Figure 2.4 Distribution ofspecies as predicted by model for case (i) conditions 16

Figure 2.5 Oxygen flow rates as predicted by model for case (i) conditions 16

Figure 2.6 Distribution ofspecies as predicted by model for case (ii) conditions 18

Figure 2.7 Oxygen flow rate as predicted by model for case (ii) conditions 18

Figure 2.8 Distribution ofspecies as predicted by model for case (iii) conditions 19

Figure 2.9 Oxygen flow rate as predicted by model for case (iii) conditions 19

Figure 2.10 Distribution ofspecies as predicted by model for case (iv) conditions 20

Figure 2.11 Oxygen flow rate as predicted by model forcase (iv) conditions 20

Figure 2.12 Crystal structures of copper sulphide minerals relevant to this study 22

Figure 2.13 Leaching morphology for a chalcocite particle (a)0-20% copper extraction

03)20-50% copper extraction (c) 50-100% copper extraction 26Figure 2.14 Evans diagram of applicable polarization curves during oxygen pressure

leaching of chalcocite 27

Figure 2.15a-b Potential-pH diagram for the Cu-S-H20 system 34

Figure 2.16a-b Thermal precipitation diagrams for the CUSO4-H2SO4-H2O system ....... 35

Figure 2.17 Thermal precipitationdiagram for the CUSO4-H2SO4-H2O system at 100C

for 3cu2+= a

S04 2- 36

Figure 2.18a-b Phase diagrams of the Cu-S system. The blaubleinder covelUte is

abbreviated as "be" 37

Figure 2.19 Models for oxygen adsorption during oxidative leaching 40

Figure 3.1 Schematic of the experimental setup 43

Figure 4.1 The distributionof copper during leaching of chalcocite 52

Figure 4.2 Similar to Figure 4.1 - no copper in solids shown to magnify scale 52

Figure 4.3 Comparison of the model results to the actual behavior 53

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Figure 4.4"Reaction" temperature versus pH^ change during precipitation from 1 M

CuS0 4 solution. The "equihbrium" boundary for 1 M solution isalso shown 56

Figure 4.5 Comparison ofsulphur levels in the leach residue between the model and

actual results 57

Figure 4.6 Distribution ofiron during leaching. 59

Figure 4.7 SEM photomicrographps - Effect ofiron on agglomerate size of the

precipitate 60

Figure 4.8 Distribution of arsenic during leaching 61

Figure 4.9 Distribution ofnickel during leaching 62

Figure 4.10 Distribution of cobalt during leaching 62

Figure 4.11 Oxygen consumption rate at various initial copper concentrations with

[Fe]=0.25g/L. Each curve represents a different level ofoxygen consumption

.- 67

Figure 4.12 Oxygen consumption rate at various copper concentrations with [Fe]=0

g/L. Each curve represents a different level ofoxygen consumption 68

Figure 4.13 The effect ofinitial copper concentration on the measured ORP at [Fe]=0

and [Fe]=0.25 g/L 70

Figure 4.14Evans diagram of applicable polarization curves during pressure

leaching ofCuS (schematic) 70

Figure 4.15 A schematic highlighting areas believed to be well mixed in the second

stageautoclave. The poorly mixed zones are thought to be the result of theobserved pseudoplastic behavior ofthe slurry 75

Figure Al . Detailed flowsheet of the CRED Plant 82

Figure DI. Rate ofoxygen consumption for the condition where (x2,x3,x4)=(l,l,0) and

xt=0 91

Figure D2. Rate ofoxygen consumption for the condition where (x2,x3,x4)=(l,0,l) and

x,=0 91

Figure D3. Rate ofoxygen consumption for the condition where (x2,X3/x4)=(l/-l/0) and

x,=0 92

Figure D4. Rate ofoxygen consumption for the condition where (x2,x3/x4)=(l/0,-l)and

x,=0 92

Figure D5. Rate ofoxygen consumption for the condition where (x2/X3/X4)=(0,l,-l) and

x1=0 93

Figure D6. Rate ofoxygen consumption for the condition where (x2,x3/x4)=(0,0,0) and

x1=0 93

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Figure D7. Rate of oxygen consumption for the condition where (x2/x3/X4)=(0,l/l) and

x,=0 94

Figure D8. Rate of oxygen consumption for the condition where (x2/x3,X4)=(0/-l/l) and

x 1= 0 94

Figure D9. Rate of oxygen consumption for the condition where (x2,x3,x4)=(0,-l,-l)

andx^O 95

Figure D10. Rate of oxygen consumption for the condition where (x2/x3/X4)=(0,0,0)

andx,=0 95

Figure DTI. Rate of oxygen consumption for the condition where

( x ^ x ^ M - l A L O ) 96

Figure D12. Rate of oxygen consumption for the condition where

(x1,x

2,x

3/x

4)=(-l,0,-l,0) 96

Figure D13. Rate of oxygen consumption for the condition where(x1/x2/x3^4)=(-l,0J3/-l) 97

Figure D14. Rate of oxygen consumption for the condition where

( x ^ x ^ M - l A O , ! ) 97

Figure D15. Rate of oxygen consumption for the condition where (x2/x3/x4)=(-l,0,l)

and x:=0 98

Figure D16. Rate of oxygen consumption for the condition where (x2,x

3,x

4)=(-l

/0,-l)

andx,=0 98

Figure D17. Rate of oxygen consumption for the condition where (x2,x3/X4)=(-l,l,0)and x1=0 99

Figure D18. Rate of oxygen consumption for the condition where (x2,x

3/x

4)=(-l,-l,0)

andx 1=0 99

Figure D19. Rate of oxygen consumption for the condition where

(x:/x2,x3/X4)=(-l,-l,0,0) 100

Figure El. Rate of oxygen consumption where initial [Cu]=0 g/L and [Fe]=0.25 g/L

102

Figure E2. Rate of oxygen consumption where initial [Cu]=80 g/L and [Fe]=0.25 g/L102

Figure.E3. Rate of oxygen consumption where initial [Cu]=10 g/L and [Fe]=0.25 g/L

> 103

Figure E4. Rate of oxygen consumption where initial [Cu]=40 g/L and [Fe]=0.25 g/L

103

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Figure E5. Rate ofoxygen consumption where initial [Cu]=l g/L and [Fe]=0.25 g/L

104

Figure E6. Rate ofoxygen consumption where initial [Cu]=0 g/L and [Fe]=0 g/L 104

Figure E7. Rate ofoxygen consumption where initial [Cu]=80 g/L and [Fe]=0 g/L 105

Figure E8. Rate ofoxygen consumption where initial [Cu]=10 g/L and [Fe]=0 g/L 105

Figure E9. Rate ofoxygen consumption where initial [Cu]=40 g/L and [Fe]=0 g/L 106

Figure E10. Rate ofoxygen consumption where initial [Cu]=l g/L and [Fe]=0 g/L 106

Figure Fl . Rate ofoxygen consumption for various initial copper concentrations at

the 10% oxygen consumption point 108

Figure F2. Rate ofoxygen consumption for various initial copper concentrations at

the 20% oxygenconsumption point 108

Figure F3. Rate ofoxygenconsumption for various initial copper concentrations atthe 30% oxygenconsumption point 109

Figure F4. Rate ofoxygen consumption for various initial copper concentrations at

the 40% oxygen consumption point 109

Figure F5. Rate ofoxygenconsumption for various initial copper concentrations at

the 50% oxygen consumption point 110

Figure F6. Rate ofoxygen consumption for various initial copper concentrations at

the 60% oxygen consumption point 110

Figure F7. Rate ofoxygenconsumption for various initial copper concentrations atthe 70% oxygen consumption point I l l

Figure F8. Rate ofoxygen consumption for various initial copper concentrations at

the 80% oxygen consumption point I l l

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ACKNOWLEDGEMENTS

I would like to thankDr. David Dreisinger for his constant encouragement and support

and Dr. Ernest Peters for his thought provoking discussions and ideas throughout the course of

this project.

I wish to thankINCO; without their financial support and faculties, this project would not

have been possible. Thanks are also extended to all of the people associated with INCO who

helped in various ways in the completion of the experimental work.

I wish to thankmy mother who has always encouraged me to press on and has shown me

the value of perseverance. I also wish to thank my wife for believing in me more than I believed

in myself throughout this endeavor.

And a final thanks is extended to the Cy and Emerald Keyes Foundation for their financial

support.

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CHAPTER1 -Introduction

Copper is one of the less abundant base metals found in the earth's crust occurring at

levels of approximately 7 ppm (compared to aluminum and iron at approximately 80 000 and 60

000 ppm respectively) [1]. Approximately 90% of the world's supply of copper occurs as

sulphidic ores. Pyrometallurgical techniques have historically been the dominant processing

method. However, pollution problems associated with sulphur dioxide emissions from

pyrometallurgical operations has resulted in a considerable research effort in

hydrometallurgicalprocessing of copper and other sulphide ores.

Some of the main advantages that hydrometallurgical processes offer are that:

1. Hydrometallurgy allows the processing of complex ores with multiple recoverable metals.

By controlling solution conditions, it is possible to recover various metals in separate unit

operations. These metals can be sold for additional revenues.

2. Hydrometallurgicaloperations are performed at lower temperatures and generally use less

energy compared to the high temperatures often employed in pyrometallurgical

operations. This is especially true for low grade ores.

3. Hydrometallurgy has often been found to be more economically viable in the treatment of

low grade ores especially if the crushing and grinding steps can be minimized as in

percolation leaching methods.

4. Hydrometallurgical operations produce little or no air pollution. The liquid waste

generated at hydrometallurgical plants is often easier to contain and treat than effluent

gases.

5. Solutions and slurries in hydrometallurgical plants are easily transported by pipeline

systems as opposed to moving of molten slags and mattes between furnaces using heavy

refractory ladles in pyrometallurgical processes.

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Successful hydrometallurgical processes for copper extraction from sulphide concentrates

have been proposed and designed but they have not been able to compete commercially with

pyrometallurgical processes. Although hydrometallurgical processes have demonstrated many

advantages as listed above, they are not a panacea for extractive metallurgy. Buthydrometallurgy does have an important role to play in the treatment ofspecific ores such as

low-grade ores, complex mineral ores and secondary materials produced from other

pyrometallurgical and hydrometallurgical operations.

The most common hydrometallurgical step is the leaching process which serves to free the

desired constituents from the gangue material via dissolution. In general, leaching methods can

be classified into percolation leaching and agitated leaching (see Figure 1.1). The method used

depends upon the nature of the ore and the mineral deposit.

f

In-situ

Leaching

Percolation Leaching

t

Heap or

Dump

Leaching

Leaching Method

t

Vat

Leaching

Agitated Leaching

Thin Layer

LeachingSlime (Pulp)

Leaching

Pressure

Leaching

t

Baking

Process

Figure 1.1 Classification of leaching methods [2].

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The leaching reagents used must dissolve the ore minerals as rapidly as possible and be

substantially inert towards the gangue minerals. Extensive reaction with the gangue minerals

causes excessive reagent consumption and causes the solution to become fouled with

impurities. The reagent must also be readily available and as inexpensive as possible.

Leaching in the presence ofsulphuric acid is one of the most common methods in leaching

copper sulphides. Sulphides are insoluble in dilute sulphuric acids but can be solubilized if

oxidizing species are present in solution. These oxidizing agents include oxygen, sulphate and

chloride salts ofiron and copper and aqueous chlorine (ie. hypochlorous acid and hypochlorite).

Acids, present as concentrated solutions, are sometimes also powerful oxidizing agents.

Increasing temperature and oxygen pressures have been found to contribute significantly to the

rate of copper extraction.

One process for the leaching of copper from a sulphide residue is the CRED1 Second Stage

Leach at INCO's Copper Cliff Copper Refinery. The second stage leach processes secondary

material, mostly CujS, generated from a preceding metathetic leach. The objective of this thesis

is to investigate the poor leaching behavior that has occasionally been encountered in this

process. The problem ofslow copper leaching kinetics has existed for approximately 15 years

occasionally becoming severe enough to warrant investigation. The research workcontained in

this thesis was designed following some initial mathematical modelling work done by

Dreisinger and Peters [3] at U.B.C. which pointed to a possible metallurgical explanation for the

poor leaching behavior.

This thesis is organized in the following way. Chapter 2 contains a brief literature survey

on the leaching behavior of copper sulphides. Chapter 2 also contains the investigations done

in the past by INCO and the model developed at U.B.C. Chapter 3 covers the experimentalmethods and chapter 4 contains the results and discussions. In chapter 5, some conclusions and

recommendations are offered.

1 CRED - Copper Refinery Electrowinning Department

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CHAPTER2 - Background and Literature Review

2.1 Overview of the INCO-CRED Process

INCCs Copper Cliff Nickel Refinery extracts nickel via the INCO Pressure Carbonyl

(IPC) process. This process produces a residue that contains mostly copper along with the other

constituents shown in Table 2.1. Approximately 50 tons of this residue are processed daily in a

hydrometallurgical plant (CRED-Copper Refinery Electrowinning Department) at the Copper

CliffCopper Refinery. The purpose of the CRED plant is to separate the various constituents of

the IPC residue through a number ofhydrometallurgical operations (see Figure 2.1) [4]. The

process steps include pressure leaching, cementation, precipitation, thickening, filtration andelectrowinning. A detailed process flowchart is provided in Appendix A.

Table 2.1 Typical Assay ofIPC Residue [4].

Weight Percent Oz/Ton

Cu Ni Co Fe S Se Te PGM + Au Ag

55-60 6-10 4-8 4-9 13-19 0.06-0.10 0.06-0.10 20-30 25-45

The IPC residue is first treated via a metathetic pressure leach in sulphuric acid (100-200

g/L) and copper sulphate (40-90 g/L) solution at 150C. This batch process, referred to as First

Stage Leaching, is used to dissolve nickel, cobalt and iron and separate these metals from

copper, selenium, tellurium and precious metals which remain in the solid phase. The overallreactions taking place in the leaching process are:

MeO(s) + H2SOA = MeS04+H20

Me(s) + CuSOt= MeSOA + Cu(s)

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MeS(s) + CuSOA =MeSOA + CuS(s)

where Me =Ni,Co,Fe

Approximately 95-98% of the base metals, other than copper, are leached out of the solids

in this step. The copper entering the first stage leach is present predominantly in the form of

CujS (chalcocite) and passes through to the second stage leach unmodified. Thefirststage leach

slurry is filtered. The filtrate is put through a copper clean-up circuit to remove some copper

which still remains in solution after the leach. Thefirststage leach residue is treated in a total

oxidative pressure leach at 115C. This step is referred to as Second Stage Leaching.

Thefirststage cake obtained from thefiltrationstep is combined with water and spent

electrolyte from the plant to produce a slurry of approximately 30% solids. This slurry is

charged into the second stage autoclaves. The chalcocite is batch-leached to form a slurry of

neutral copper sulphate (CuS04) solution (pH of -2.5-3.0) and basic copper sulphate

(CuS04-2Cu(OH)2) solids. The slurry is then mixed with spent electrolyte to dissolve the basic

copper sulphate to leave a residue containing precious metals and lead sulphate. Selenium,

tellurium and most of the base metals are also solubilized during the leaching process. The

unleached solids arefilteredout for processing at another INCO plant to recover the precious

metals. The solution is treated for selenium-tellurium removal and then pumped to the

tankhouse for copper recovery via electrowinning.

Selenium and tellurium removal is essential because these impurities tend to co-deposit

with the copper during electrowinning and contaminate the cathodes. The removal of Se and

Te is achieved by heating the solution to 95 C and passing it through a column filled with

copper shot. This promotes the formation ofselenide and telluride precipitates that form as fine

blackparticulates. The solution and solids are passed through four aging towers in series in

which the solids settle out. The Se and Te concentrations are reduced to less than 1 mg/1 in

solution. The overflow from the aging towers is passed through polishingfiltersand sent to the

electrowinning circuit.

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PM ResidueSpent Electrolyte

Sulphuric Acid

Oxygen . . Second StagePressureLeaching

First

Stage

Cake

IPC Residuefrom INCOPressure CarbonylProcess ^

First StagePressureLeaching

Spent

Electrolyte

First

Stage

Filtrate

NaSH '

Filter Aid

CopperClean-up

CuS to SecondStage PressureLeaching

Se, Te Residue

Filtrate Selenium,TelluriumRemoval

CuShot

Filter Aid

Steam

Sulphuric Acid

SteamOxygen

Lime

Filter Aid

Filtrate

Iron/ArsenicRemoval

Iron hydroxide-Gypsum Solidsto Effluent

Filtrate CopperElectrowinning

Titanium Blanks

Lead Anodes

Reagents

Water/Steam

Soda Ash

Steam

Filtrate

Nickel-CobaltRecovery

CopperCathodes

Spent

Electrolyte

Nickel/Cobalt* Carbonate

Vacuum

Bosh Pond

Figure 2.1 General processing route ofthe IPC residue at the INCO-CRED plant in Copper

Cliff

The filtrate from the first stage leach is processed through a copper clean-up circuit to

remove any copper which exists in solution in the form ofcopper sulphate. The removal is

necessary to prevent copper losses to the effluent during the iron/arsenic removal step and to

prevent copper contamination of the nickel/cobalt carbonate. The copper is removed by the

addition of a 30% NaHS solution at 70 C. Most ofthe copper is precipitated as CuS. The process

is controlled so as to prevent the evolution ofH2S gas. A thickener is used to thicken the CuS

precipitate. The solids are returned to the first stage filters and the overflow solution is sent to

the iron/arsenic removal circuit.

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The objective of the iron/arsenic removal circuit is to produce environmentally stable

compounds of iron, arsenic and sulphur and discard them to the effluent stream without

excessive loss of nickel and cobalt. The feed solution to this circuit contains 0.02 g/L copper,

25-35 g/L nickel, 15-25 g/L cobalt, 20-35 g/L iron, 1-2 g/L arsenic, and 40-70 g/L sulphuric

acid. It is processed continuously at a rate of 150 L/min. Lime slurry is added to the solution as

it is passed through two autoclaves operating under oxygen pressure at 90-95 C. The iron and

arsenic precipitate out and the solids are separated by filtration to produce an iron-gypsum

cake. The filtrate is sent to the nickel/cobalt recovery step. The iron-gypsum cake is partially

redissolved to remove co-precipitated nickel and cobalt and then filtered. The filter cake is sent

to the tailings and the filtrate is returned ahead of the iron precipitation circuit.

The filtrate from the iron removal circuit is mixed with a 200 g/L solution of sodium

carbonate (Na2C03) in two reaction vessels in series. The pH is controlled in the ranges 7.6-7.8

and 8.1-8.3 in the two vessels respectively. The reaction product is a precipitate of basic nickel

and cobalt carbonates which is thickened to 20-25% solids. The nickel/cobalt carbonate is

shipped to the cobalt refinery and the barren solution is pumped to the waste pond.

2.1.1 Background of the Second Stage Leach

When the CRED plant first became operational in the early 70's, the second stage leach

was designed to leach CU2S completely in the presence ofexcess acid (H2SO4) and under

oxygen pressure at 110C. The products of the leaching process were CuS04(aq) and

elemental sulphur. The equation governing the process was reported as [5]:

Cu + 2HOt + 02 -> 2CuSO, + 2H20 +S

The leach products typically analyzed 90% elemental sulphur and 10% precious metals and

unreacted sulphides. The factors affecting the reaction rate were determined to be the iron

content of the solution, oxygen pressure, temperature and feed particle size. The acid and

copper concentrations were not found to be critical factors in the leaching rate. The iron

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content of the solution was identified as beingan important variable. With iron in solution,

the reaction proceeds rapidly according to the reaction above to form elemental sulphur.

However in the absence ofiron, the reaction proceeds more slowly and consumes up to 4

times as much oxygen and the sulphide sulphur is oxidized to sulphate according to the

reaction:

CifjS + H2SOA+\o2^> 2CuS04+H20

The leaching process was operated by this method for a few years. However, in the

mid-seventies, historical reports [6] show thata proposal was made to change the leaching

process. It was suggested that all the sulphide sulphur should be oxidized to form sulphate

as opposed to elemental sulphur by leaching in a low-acid solution. Based on laboratory

work, the product from this leaching method would produce a precipitate of basic copper

sulphate (CuS04-2Cu(OH)2) in a solution of copper sulphate. The slurry would have to be

mixed with spent electrolyte containing sulphuric acid to dissolve the precipitateand leave

a residue containing precious metals and a small amount of gangue materials. The

operating conditions of the process were set at 150 psi oxygen pressure at 105'C. The batch

feed to the autoclave would be a mixture of spent electrolyte containing sulphuricacid,

copper and iron in solution with first stage residue to form a slurry containing -40% solids

by weight.

The process was eventually changed to a total oxidative leach in 1975 with some

modifications being continually made to improve the process. Howver, the secondstage

leach residue occasionally showed high levels of copper still remaining in the solids even

after extended leaching times [7]. Extensive examination ofIPC residue, first stage residue

and poorly leached second state leach residue was carried out to determine the nature of the

poor leaching behavior. The results showed that the presence ofCu 20 in the feed to second

stage, associated with high levels of oxygen in the IPC residue, was linked to the poorly

leached batches. The presence ofCu 20 was proposed to cause an adhering film of basic

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copper sulphate at the point of basic copper sulphate formation. This would block oxidizing

species from reaching the copper sulphide particles and hinder further leaching.

Photomicrographs were presented to support this theory.

During 1986, slow leaching and high copper in the residue became a severe problem

and warranted further investigation of the process [8]. A complete leach should nominally

take 5 hours based on plant reports and experimental tests. However, the length of leach

times almost always exceeded 5 hours with the occasional leach time taking longer than 20

hours. For example, from plant data for the month ofJanuary 1988 [9], out of93 leaches, 33%

ofthe leaches exceeded 8 hours. A calculation in one report [10] gives the cost of downtime

as approximately $25 000 /hr for deferred revenues from precious metals. It is apparent

from this calculation how even small improvements in leaching times can make a significant

difference to the revenues over one year.

One laboratory study [11] of the second stage leach shows some interesting results on

the behavior of the process with respect to pulp density, particle size of precipitate, and

agitation speed. It was found that higher solids density feeds lowered the leaching rate of

copper considerably due to higher viscosity of the solution. In lab tests with normal solids

densities, higher viscosities were observed on the material which leached poorly in the plant

as opposed to material which leached quickly. The higher viscosities observed in this case

were associated with afinerparticle size of the precipitate. The agitation speed of the slurry

also had a significant effect on the leaching rate. All these factors appear to indicate that

oxygen dispersion in the autoclave is severely affected by changing viscosities and agitation

speed.

The autoclaves have no level meter, so the impeller depth varies considerablybetween

leaches. Impeller depth was suggested to be an important parameter in the leaching rate in

one study and so the slurry level was lowered (amount unknown) and was found to

improve the leaching process for a short while. However, this did not cure the problem

permanently.

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2.1.2 The Second Stage Leach

The equipment used in the second stage leaching process consistsofa batch make-up

tank/2 titanium autoclaves, a dissolving tank, 2 pressure filters and a rotary vacuum filter

(Figure2.2) [4]. There are also product holding tanks between various stages of the process.

PM residueslurry storage

Figure 2.2 Second Stage Leaching Circuit at INCO's - Copper CliffCopper Refinery [4].

The autoclaves were originally designed for leaching in excess acid as described

earlier but were modified in the mid-seventies when the process was changed. The major

change to the autoclaves was the installation of vertical cooling coils around the inside

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perimeter of the vessels (see figure 2.3) [12]. The agitator system consists of two 45'

pitched-down, 4 bladed turbine impellers attached to a single 18 cm diameter shaft rotating

at 68 rpm. The internal parts are made out of 316L stainless steel.

Oxygen

Sparger

Figure 2.3 Approximateshape and dimensions of the Second Stage Autoclaves [12].

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The feed is prepared as a 30% solids slurry by mixing spent electrolyte from the

electrowinning plant, water and first stage residue in the batch make-up tank. The slurry is

then pumped into one of the second stage autoclaves. The autoclave is then pressurized

with oxygen to 150 psi while the slurry is being agitated. The temperature is maintained at

115C during the leach. The leaching process is assumed to be complete when there is no

more oxygen consumption, no more heat being generated and there is no temperature

increase when cooling is off. The slurry is then pumped to the redissolving tank where it is

mixed with spent electrolyte and filter aid. The acid in the spent electrolyte dissolves the

basic copper sulphate. The unleached solids are then recovered by pressure filtration and

the filtrate is sent to Se/Te removal. The solids are repulped in water and refiltered to

recover the precious metals residue. The final residue is shipped to Port Colborne for

precious metals recovery.

2.2 The UBC Screening Model

The studies in the past have been inconclusive in determining the cause of the long

leaching times and incomplete leaching in the second stage autoclaves. A mathematical model

of the CRED second stage leach was developed at UBC by Dreisinger and Peters [3] in an

attempt to evaluate possible metallurgical causes of the "slow cook" conditions occurring in the

process. This screening model was a first approximation to possibly highlight some of the

conditions that may lead to slow cook conditions. The model was developed on a number of

important assumptions based on work previously done by Peters and Mao [13] on the leaching

ofCu2S underslightly acidic conditions. Their workproduced the following results:

i. Cu^ leaches very quickly to CuS.

ii. The CuS produced by leaching Cu2S tends to fracture and become finely disseminated.

iii. Cu ^ conversion to CuS proceeds very quickly relative to the leaching ofCuS.

iv. A small amount ofiron in solution promotes elemental sulphur formation during leaching.

The leaches with no iron showed little or no elemental sulphur in the leach residue.

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Since CuS leaching proceeds slowly relative to Cu 2S conversion to CuS, it is possible to

separate the leaching steps of Cu 2S conversion to CuS and CuS leaching to CuSGv In

developing the model, a number of discrete leaching reactions were proposed which would

proceed one at a time depending on the prevailing solution conditions and the character of the

unreacted solids. These reactions are:

1. CuxS+{x- l]H2SO CuS+^(x - 1)2CM(0//)J CuS04

3. CuxS+^(x - \)02 + (x -\)H20 -> ^(3-x)CuS+^(x- l)(2Ca(0//), CuSOj

4. CuS + 202-+CuSOt

5. CuS+H2SOi+^02^,CuSOi+H20 + S

where Cu,S refers to the average Cu/S mole ratio in the feed, not to any particular mineral

form and the value of is approximately 2.

There are five reaction pathways possible under various initial batch recipes.

i. Begin with copper sulphate and acid in solution.

Reaction 1 proceeds until acid depletion.

Reaction 2 proceeds until only CuS is left in the residue

Reaction 4 proceeds to total oxidative endpoint.

ii. Begin with copper sulphate in solution and no acid.

Reaction 2 proceeds until only CuS is left in the residue.

Reaction 4 proceeds to total oxidative endpoint.

iii. Begin with copper sulphate andexcess acid in solution. ,

Reaction 1 proceeds until only CuS is left in the residue.

Reaction 5 proceeds to acid depletion.

Reaction 4 proceeds to total oxidative endpoint.

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iv. Begin with acid and low copper in solution.

Reaction 1 proceeds until acid depletion.

Reaction 2 proceeds until copper sulphate is depleted.

Reaction 3 proceeds until only CuS is left in the residue.

Reaction 4 proceeds to total oxidative endpoint.

v. Begin with low copper and no acid in solution.

Reaction 2 proceeds until copper sulphate is depleted.

Reaction 3 proceeds until only CuS is left in the residue.

Reaction 4 proceeds to total oxidative endpoint.

The kinetic routine in the model was divided into a two-step process. The first step is

gas-liquid mass transfer defined by the following equation,

d[02]- ^ - = *,([OJ -[Oz])

where kg is the gas-liquid mass transfer coefficient for the system

[OJ" is the saturated oxygen cone, in solution at the oxygen partial pressure in the

autoclave

[02] is the oxygen concentration in the bulksolution

The next step is controlled by chemical reaction defined by the following empirical equation:

=k^Cu2*][O J + ^[CM 2 +] [OJ [CuxS]

where kt and k2 are empirical rate constants.

The equation is defined on the basis of copper catalysis which is first order in cupric and

dissolved oxygen concentrations and a solids leaching term which is first order in cupric,

dissolved oxygen and unreacted CuJ$ concentrations. The first term recognizes the significant

role of copper ion catalysis and the second term accounts for the fact that the rate will drop as

the solids concentration goes to zero.

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The kinetic equations suggest that the amount of copper in solution at any given time is

very important to therate and that lower copper concentrations will cause slower leaching rates.

Ifthe copper in solution is depleted, the rate is expected to go to zero. This is the anticipated

slow leaching behavior observed in the plant known as the "slow cook" condition.

In this preliminary model, the rate constants were treated as fixed values because of the

unavailability of additional data. These values probably change due to changing conditions

during leaching.

Figures 2.4-2.11 show some sample outputs of the model generated from various initial

concentrations ofacid and copper in solution. It can be seen that the slow cook condition occurs

only when there is a depletion of copper in solution.

Figure 2.4 shows the model output from a case when there are acid and copper in solution

at a level which will cause the reactions to proceed as described in case (i) above. The output can

be split into 3 sections. In the first section, according to reaction 1, acid is being consumed to

leach copper from Cuand the copper concentration is increasing in solution. When the acid is

depleted, at around the 16-minute mark, the process proceeds according to reaction 2 and

produces basic copper sulphate from CuxS and CuS04. The copper concentration decreases as

shown. Eventually, when all the Cu^S is transformed to CuS, the leaching proceeds according

to reaction 4 until all the CuS is leached. Figure 2.5 is the corresponding oxygen consumption

rate curve. This shows that the initial rate of reaction rises very quickly as the amount of copper

in solution rises and the rate drops as the copper is depleted via reaction 2. The reaction rates

never drop to low levels because there is always a lot of copper in solution available for

catalysis. The complete leach time is approximately 5 hours and is close to the times observed

in plant operations for a normal leach.

Figures 2.6 and 2.7 represent case (ii) in which the there is no acid available at the

beginning of the leach and the initial copper concentration is approximately 100 g/L. The

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320 -

280 -

240 -

^ 20 0 -c

o

a

s

160 -

c -

120 -

880 -

40 -

0 -

Concentration Profiles Predicted by Model

_ \\\

\C u

xs

V\V\\\ 2Cu(OH)2.CuS04

- Y/ \///;

Cu

N

W - H 2 S 0 41 1 1 I 1 1 1 I

_ _ " " ~ r

40 80 120 160 200 240 280

Time (min)

Figure 2.4 Distribution ofspecies as predicted by model for case (i) conditions.

0.04

0.035

Oxygen Flowrate Predicted by Model

28 0

Time (min)

Figure 2.5 Oxygen flow rate as predicted by model for case (i) conditions.

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model predicts a higher amount of basic copper sulphate formation via reaction2 until all Cu,S

is converted to CuS. The kinetics are good in this case because the copper in solution never

drops below 50 g/1.

Figures 2.8 and 2.9 shows the effect ofexcess acid in solution at the beginning of leach.Because there is still acid present after all the CuxS has been converted to CuS, sulphur

formation is expected via reaction5 and no basic copper sulphate is produced. Near the end of

the leach process, the solubility of copper sulphate is exceeded resulting in the precipitation of

CuS04.5H 20.

The model was operated under conditions which would cause a copper deficiency during

operation. This was done by setting a high copper to sulphur ratio in the feed solids and less

acid in the feed solution. The resulting output is shown in Figures 2.10 and 2.11. The acid and

copper in solution drop very quickly and the result is a severe drop in the leaching rate. It is also

important to note that a significant amount of basic copper sulphate solids is produced. This

high amount of solids could cause a significant drop in gas-liquid mass transfer although the

model does not incorporate the effect of solids loading into the gas-liquid mass transfer rate.

The total leach time is predicted to exceed 25 hours.

The model predicts slow leaching conditions under copper depleted conditions but the

model requires verification. Firstly, the sequential reaction chemistry proposed in the model

needed to be verified through experimental work. Secondly, better kinetic relationships need to

be developed because the model used an empirical relationship for the rate of reaction. The

objective of this thesis project was to obtain experimental data to improve the understanding of

the leaching process and eventually to develop a better model.

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co034=C

oo

280

240

200

160

120

80

Concentration Profiles Predicted by Model

ri

i

\

I \ C"xS/ ^

2Cu(OH)2.CuS04rii

\

I \ C"xS/ ^/ \ // " " ^ " Cu

1 \,

100 200

Time (min)

300 400

Figure 2.6 Distribution ofspecies as predicted by model forcase (ii) conditions.

Oxygen Flowrate Predicted by Model

200

Time (min)

Figure 2.7 Oxygenflowrate as predicted by model for case (ii) conditions.

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320 -i280 -240 -

i 200 -

ritratiot

160 -ncei

120 -8

80 -40 -0 -

Concentration Profiles Predicted by Model

\CujSCu

CuS04.5H20

Time (min)

Figure 2.8 Distribution ofspecies as predicted by model for case (iii) conditions.

0.05 Oxygen Flowrate Predicted by Model

0.04c!o 0.03 -sf 0.02

0.01

0 I i iii ii ii ii i ii i iiiii -0 20 40 60 80 100 1 20 140 160 180 200Time (min)

Figure 2.9 Oxygen flow rate as predicted by model for case (iii) conditions.

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co

sc

500

450

400

350

300

250

200

150

100

50

0

Concentration Profiles Predicted by Model

\ i

I

^ 2Cu(OH)2.CuS04

C U x S

Cu'TfrTrTrT

1

500 1000T I I 1

1500

Time (min)

Figure 2.10 Distribution ofspecies as predicted by model for case (iv) conditions.

Oxygen Flowrate Generated by Model0.04 -n

0.035 -

.g 0.03 -

| 0.025 -

| 0.02 -

o

c 0.015 -

I

I" 0.01 -0.005 -

0 " I I I \ 1 I I I I I I I I I 1

0 500 1000 1500

Time (min)

Figure 2.11 Oxygen flow rate as predicted by model for case (iv) conditions.

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2.3 Scope of the Study

A review of the history of the second stage leached has shown that a number of variables

have been considered in trying to improve the leaching process. Clearly, the dificulties with the

process have not been satisfactorily overcome. The most recent attempt at analyzing theproblem was the UBC screening model discussed earlier.

The present study was designed based on the results of the UBC screening model and the

earlier workdone at INCO labs. The main objetives ofthis study are:

1. to understand the reaction chemistry and check if the reactions are indeed

sequential as suggested in the UBC screening model.

2. to identify any intermediate copper sulphide products formed.

3. to investigate chemical kinetics of copper leaching for the first stage residue cake.

4. to provide a basis for further work to improve the present mathematical model or

develop a new mathematical model.

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2.4 Literature Review

2.4.1 Copper Sulphides -Chalcocite to Covellite

One of the most important economic copper minerals, chalcocite (CujS), is produced

by the reduction of CuS0 4 solutions descending from oxidation zones of copper rich

deposits in the earth [14]. Covellite (CuS), similarly, is usually found as an oxidation

product of chalcocite or other primary copper sulphides like chalcopyrite as a zone of

secondary enriched copper deposit. The oxidation of chalcocite does not lead to the direct

formation of covellite. The decomposition process produces many intermediate sulphides

such as djurleite (Cu196S), digenite (Cu, 76., g^S), blue-remaining covellite (Cu,,. 14S) and

covellite (CuS). Potter [15] has shown the existence of numerous intermediate phases andprovided free energy data for these phases. These phases are Cu ^S , Cu, o S, Cu, 76 5S,

Cu, 4S, Cu, ,S and CuS. Figure 2.12 shows the crystal structures of chalcocite, covellite and

digenite.

Chalcocite (hexagonal) Digenite (cubic) Covellite

Figure 2.12 Crystal structures ofcopper sulphides relevant to this study [14].

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In chalcocite and djurleite, the sulphur species are arranged in a hexagonal close

packed structure [14]. The copper ions are located near the triangular faces of the tetrahedral

sites as opposed to being in the center of the tetrahedral interstices. The structure of digenite

is a cubic close packed arrangement of the sulphur species and the copper ions are located

off-center in the tetrahedral interstices. One eighth of the tetrahedra are unoccupied [14].

Natural digenite has 1 at.% iron to maintain a stable solid solution composition but a stable

iron-free "low digenite" called anilite (Cu^S) is also found.

The structure of covellite is more complex as can be seen in Figure 2.12. The base

structureconsists of three layers of hexagonal close packed sulphur species. The copper ions

occupy the centers of the equilateral triangles and the centers of tetrahedral sites in the

layers.

2.4.2 Leaching of Chalcocite and Covellite

The leaching of chalcocite has often been reported as a two-step process according to

the following reaction sequence:

Cu** -> CuS+ Cu1++ 2e~

CuS->Cu2+ + S+ 2e~

A variety of leaching processes have been investigated in the laboratory and classified

according to the leaching steps and the types ofreagents used. The most common oxidizing

agents are:

Ferric sulphate in acid

Ferric chloride in acid

Oxygen in sulphuric acid

Oxygen in ammoniacal solutions

Nitric acid

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There are other oxidizing agents but they have not received much attention, mostly because

oftheir cost and commercial availability.

Sullivan [16] studied the leaching chemistry of chalcocite in acidified ferric sulphate

solutions using bottle roll tests at temperatures below 50 C The dissolution process was

found to occur in two steps. The first step was rapid until about 50% copper dissolution and

the second step, oxidation ofCuS, was relatively slow. The two reactions involved in the

leaching process were reported as:

CujS+ Fe2(SOJ3 -> CuSOA + 2FeSOA + CuS

CuS+ Fez(SO,\ -> Cu504 + IFeSOA + S

The study showed that the dissolution rate was independent of the strength offerric

sulphate provided sufficient reagent was present. At a constant ferric concentration, the rate

was also independent of acid strength. Various particle size fractions in the range of

-10-mesh to 200-mesh were also studied. Although there is considerable difference in

surface area per unit weight in this range, there was almost no difference in the leaching

rates (ie. time taken to dissolve a given amount of copper) observed, provided that the

particles were open to solution attack. The rate of dissolution was greatly affected by

temperature. For example, 73% copper dissolution required 1, 5 and 15 days at 50 C, 35 C,

23 C respectively. Sullivan reported no information on intermediate copper sulphide

phases.

Thomas et al. [17] examined the kinetics of dissolution of synthetic chalcocite and

digenite in acid ferric sulphate solutions using a rotating sintered disc technique. The study

found that digenite and chalcocite dissolved at similar rates. The dissolution was reportedto occur in stages where chalcocite is progressively converted into djurleite, digenite,

blaubleibender covellite (also known as blue-remaining covellite) and normal covellite. The

covellite is transformed to elemental sulphur according to the second of two equations

above. The rate of dissolution was found to be directly proportional to ferric ion

concentration between the ranges tested (0.025 M - 0.2 M). This ferric ion concentration

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dependency was not found in the studies by Sullivan. This difference is probably due to the

differences in experimental methodology. The rate was also found to depend significantly

on the temperature.

Mao and Peters [13] studied the leaching of chalcocite underautoclave conditions and

found the same two-stage leaching behavior reported in ferric sulphate leaching. During the

first stage, 50% of the copper is extracted and up to 100% of the chalcocite is converted to

covellite. The first stage is also separated into two steps which involves the initial

conversion of chalcocite to digenite and then subsequent conversion of digenite to covellite.

A leaching model to explain the leaching kinetics was based on three parts shown in Figure

2.13. The first step is a shrinking core kinetics model, the second includes particle break-up

and third is the effect of elemental sulphur morphology on kinetics. The leaching process is

described as a mixed-potential electrochemical model in which the first stage kinetics are

predominantly cathodically controlled. The presence of iron in solution leads to higher

leaching rates and decreases sulphur oxidation during the second stage. The second stage

kinetics are explained by the passivation of covellite by oxygen leading to a high mixed

potential. The Evans diagram in figure 2.14 shows a schematic of the applicable polarization

curves. Depassivation occurs in the presence ofFe2+

ions where the process operates at point

D and leads to a higher exchange current (leaching rate) and a lower mixed potential.

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First Step: Stage I LeachingCathodic:

02+4H+

+4e -^WiO (Cu & Cu^ Surface)

Anodic:CUTS -> CKgS + 0.2Cu2+ + OAe (Cu^S only)Volume Reduction: 7.6%

Second Step: Stage I LeachingCathodic:02 + 4H

+ + 4e

Anodic:CU2S- CulJSS+ 0.2Cw

2+

+OAe (CiijS only)

Cumulative Volume Reduction:24.4%

2H20 (CuS Surface)

Stage II LeachingCatnodic:02+4H

++4e^2H20 (CuS Surface)

Anodic:CuS-*Cu2++S + 2e

CuS+ 4H20 ->Cu+

+SOl'+ 8/T + Se

Cumulative Volume Reduction:43.6%(or more depending on degree of sulphuroxidation)

Figure 2.13 Leaching morphology for a chalcocite particle (a) 0-20% copper extraction

Ob) 20-50% copper extraction (c) 50-100% copper extraction [13].

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Otfluston - l \limited Otygen i ^hi Fww i \

0!

n _ A - B First SleO- Stag I teacftatgC Secofic Sm . Slage I teachingD Slags II Leaching (Fa * present)

Stage 11 Leaching(Fe' absent]

Legl-Cabin* Cumrtt [lenemalk)

Figure 2.14 Evans diagram of applicable polarization curves during oxygen pressure

leaching of chalcocite [13].

Oxygen pressure leaching experiments in an iron containing solution done by

Chmielewski and Charewicz [18] show that the partial pressure of oxygen governs the

process rate. The oxygen increased the kinetics mainlyby oxidizing the ferrous iron to ferric

iron, the main leaching agent, and not directly by interaction with the copper mineral.

King et al. [19] studied the leaching of chalcociteby acidic ferric chloride solutions and

found the same two stage leaching process observed in other studies. The first stage of the

reaction, to approximately 50% copper dissolution, was complete in less than 4 minutes at

temperatures between 40C and 80C.However, the second stage of leaching was strongly

affected by temperature as the kinetics were much more rapid at higher temperatures. The

apparent activation energy, Ea, for the first stage and second stage was 3.43 kj/mol and

101-122 kj/mol respectively. This difference in Ea was attributed to a difference in the

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charge transfer process between the two reactions. The first stage is probably controlled by

diffusion of copper ions in the particles and the second stage is chemically controlled by the

reaction ofS2' ions in CuS to form sulphur.

The first stage of the process proceeded more slowly with the larger particle sized

material due to a larger diffusion layer for the copper to travel through in the solid. The

second stage was less affected by particle size. There was no effect ofacid (HCl) strength on

the leaching rate. An increase in ferric concentration up to 0.25 M produced an increase in

the dissolution rate of the first stage but was independent of concentration beyond this

point. The addition of ferrous chloride to ferric chloride solutions showed the exact same

net increase that would have been found ifthe same amount offerric chloride had been used

instead. Use of ferrous chloride alone (ie. no ferric chloride) produced slow leaching.

Particle size fractions tested were 425-600 um, 150-300 um, and 75-106 um. These size

fractions were similar to the ones Sullivan studied. There was almost no difference observed

in the leaching rates, a result similar to that ofSullivan. Particle sizes of1.18-1.70mm and

2.36-4.76 mm produced much slower leaching rates. These were attributed to a larger

diffusion distance required by the copper to travel in the solid.

The most interesting observations of this study were based upon the X-ray diffraction

data. The results show that a whole range of intermediate non-stoichiometric copper

sulphide phases are formed as copper is leached out of the solid matrix. The basic chalcocite

crystal structure does not change until the copper level is below Cu 1 - g 9 1S which is outside the

digenite stoichiometric range. There appears to be a similar behavior as digenite transforms

to covellite. This could be caused by some local areas and particles becoming more depleted

ofcopper and achieving compositions at which phase transformations occur before others.

This would explain the mixtures ofphases observed.

The leaching of chalcocite and covellite was studied by Grizo et al. [20] at pH values

between 0.7 and 2 in the presence ofsulphuric acid and ferric sulphate. They divided the

leaching process into three stages. The three stages were identified by changes in kinetics

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from linear to non-linear and backto linear dissolution rates. This is unlike previousstudies

which divided the stages according to the formation of intermediate species such as digenite

and covellite. They do, however, suggest that during the second stage, leaching of

chalcocite, digenite and covellite is occurring in parallel. The activation energies were found

to progressively increase in each stage but the rise in activation energy is higher between the

second and the third stage. This increase in activation energy suggests a change in

mechanism from diffusion control to chemical kinetics control. An increase in particlesize

decreases the rate at which copper is leached but does not affect the kinetic mechanism in the

three stages. The leaching data on various particle size fractions were also found to show

that the second stage was controlled by the diffusion ofspecies through a product layer.

Cheng and Lawson [21] investigated the leaching ofsynthetic chalcocite and covellite

in oxygenated acidic sulphate-chloride solutions. The leaching was described in terms ofa

shrinking core model with the rate being surface chemical reaction controlled in the first and

second stages. The late second stage was accompanied by pore diffusion control. Elemental

sulphur formation on the surface of the particles was found to retard the dissolution rate

during covellite leaching.

Thomas and Ingraham [22] studied the kinetics of dissolution ofsynthetic covellite in

aqueous acidic ferric sulphate solutions via a rotating sintered disk technique in the

temperature range 25 to 80C. They identified two rate controlling steps. The first, below

60 C, was surface chemical reaction controlled and the second, at higher temperatures, was

solution transport controlled. The respective activation energies were 92 kj/mol and 33

kj/mol. The leaching rate was directly proportional to the ferric concentration below 0.005

M but not sensitive to higher ferric sulphate concentrations.

Dutrizac and MacDonald [23] also studied the dissolution of synthetic CuS and

high-grade natural covellite in the temperature range 25 to 95C in acidified ferric sulphate

solutions. They found little difference in the leaching rate between natural and synthetic

covellite. Other leaching observations were similar to those of Thomas and Ingraham [22].

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2.4.3 Electrochemical Studies

A number of researchers have investigated the electrochemical dissolution of copper

sulphide ores and mattes in order to develop a process for the direct electrorefining ofthese

materials. Etienne [24] studied the electrochemical aspects ofaqueous oxidation of copper

sulphides using rotating diskanodes of digenite and chalcocite. The results showed that the

rate of chalcocite oxidation was under the control of diffusion of the cupric ions through the

solution in the pores of the covellite-sulphur product layer. Etienne also explained the cause

oflarge overpotentials observed by many researchers that occurred at some time into the

electrolysis process. She theorized that the polarizations causing the overpotentials were

due to the precipitation of copper sulphate in the pores formed from the leaching of copper

out of the matrix. This blocked the transfer ofcurrent to the reactive surface siteswhere the

electrolyte contacts the solid surface thus setting up a high electrical resistance.

Biegler and Swift [25] also investigated the dissolution of copper sulphide anodes and

the results supported Etienne's theory for the cause of the polarization. They also used other

electrolytes besides copper sulphate and found that the time at which polarization occurs is

directly related to the time at which the hmit of solubility of the copper salt in the solution is

reached. They also noted that the structure of the product layer is poorly understood and

further investigation of the non-equilibrium products formed during dissolution would be

required to understand the leaching process.

The study of the mechanism of the anodic dissolution ofCU2S was performed in the

presence ofsulphuric acid under galvanostatic and potentiostatic conditions by Winand et

al. [26]. In all cases, a layer of digenite, Cu : gS, was found to form on the surface according to

the following reported reaction.

SCu^S - 5CulxS+Cu2+ + 2e

A concentration gradient of copper was observed through the digenite layer. This digenite

layer stays at a constant thickness after the Cu 1;1S layer appears on the surface. The reaction

in this step was reported as

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3C,5 -> 4Cu1AS+Cu2+

+2e

Ifthe anodic potential is low during the electrolysis process, the next reaction proceeds as

follows

WCuLlS - llCu2+

+10.S+2r?

However, if the current density is sufficiently high to achieve a sharp increase in anodic

potential or the potential is kept high, the reaction path is given by the following two

reactions

lOCulAS -> \QCuS+ Cu2+ + 2e

followed by

CuS-+Cu2+

+S+2e

Furthermore, sulphate is also found to be formed to some extent at high anodic potentials

according to the following reactions.

CuS+AH20 - Cu2+ + S02-+W++8e

It must finally be noted that the formation ofa copper sulphate precipitate on the

surface was stated not to be the cause ofthe sharp rises in anodic potential observed in the

experiments because the calculated current density for such a film on the surface was a

factor of ten higher than the currentdensities used in the study.

McKay [27] studied the anodic decomposition ofcopper-rich mattes using particulate

electrodes. Anodic decomposition ofsynthetic chalcocite was defined as a three-stage

process according to the following reactions:

1. Cu^ -+Cu2_xS+xCu2+ + 2xe~, 1.75 < (2-x) < 1.83

2. Cu2_xS-+Cu2_yS+ {y-x)Cu2

*+2(y-x)e-, 0.7

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Stage 1 is associated with crackformation along grain boundaries and in stage 2, the grains

begin to deteriorate as copper is depleted from them. Total-bed polarization occurs because

of the deterioration of the electrode and a reduction in conducting reaction interfaces

between the electrolyte and the solid surface due to sulphur formation in stage 3.

The electrochemical dissolution of copper sulphides was investigated by Hillrichs et

al. [28,29,30] through cyclic voltammetric methods. The anodic dissolution ofQ . S in

sulphuric acid pointed to three factors controlling the current density. These are, (a) solid

state diffusion through the CuS product layer, (b) pore diffusion in the product layerand (c)

resistance polarization due to CuS04 precipitation in the pores formed during dissolution.

The formation of a thin metastable copper oxide layer was also thought to affect the

dissolution of CuS. Further studies [29] confirmed the formation of this metastable,

non-stoichiometric copper oxide/hydroxide layer.

MacKinnon [31] investigated the anodic dissolution of chalcocite using a fluidised-bed

anode method. The intermediate formation of "blue-remaining" covellite (CuuS) was

observed. The dissolution process became inhibited after about 50% copper removal and

was accompanied by increased oxygen evolution on the platinum current distributor. This

inhibition was inferred to be caused by sulphur formation on the surface of the reactive sites.

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2.4.4 E h -pH Relationships and Phase Systems

Potential-pH diagrams show chemical equilibrium relationships for aqueous systems.

The plots are generated from hydrolysis and oxidation reduction reactions. Figures 2.15a

and 2.15b are E-pH diagrams for the Cu-S-H20 system at 25 and 100 "C generated at unit

activity for all species [32].

Figures 2.16a-b and 2.17 [32] are thermal precipitation diagrams to show shifts in

solution-solid equihbria with respect to temperature and pH.

Figures 2.18a-b [33] are phase diagrams for the Cu-S system to show the stability

ranges for the various species. This phase diagram does not show some of the metastable

phases that have been observed by many researchers.

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Figure 2.16a-b Thermal precipitation diagrams for the CuSO 4-H 2S0 4-H 20 system [32].

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Figure 2.17 Thermal precipitation diagram for the CuSO^HjSCvHjO system at

100-CforaCiiJt = V . [ 3 2 ] .

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J20Cv,1i

r

1000 IN

8 0 0 -

r-x>l) is very fast compared to the

leaching ofCuS because the leach times to 20% consumption were less than 30 minutes and

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the remaining 4 hours of leaching were devoted mostly to CuS dissolution. The lower than

expected amount of basic copper sulphate observed can only be explained as result of the

reactions not occurring sequentially.

The most important result was that there was no slow leaching behavior exhibited in

the laboratory experiments. This result is consistent with other experiments done under

normal leaching conditions previously at INCO labs.

4.1.2 Iron and Arsenic

The iron is slowly leached into solution during the reaction time (Figure 4.6).

However, most of the iron in solution precipitates rapidly with the basic copper sulphate to

a stable level by the 25% oxygen consumption point. The level ofiron in the precipitate

begins to decrease slightly by the end of the leach process probably due to some iron being

resolubulized into the leach solution.

The role of iron in the leaching process is very important. As discussed in the

literature review section, iron was found to increase the leaching rates in most studies

primarily by acting as a charge carrier between the oxygen and the copper sulphide

particles. Iron is also thought to have an important affect on the precipitation behavior ofbasic copper sulphate. The experiments with no iron in solution showed some very

interesting behavior. The resulting basic copper sulphate slurry was much more viscous

and the agglomerate size of the precipitate was finer (see Figure 4.7). The rate of oxygen

consumption also slowed down considerably at approximately 15% oxygen consumption

and was very sensitive to stirring speeds. This suggests that gas-liquid mass transfer is

severely affected by the increased viscosity.

Arsenic in the solids is initially leached very quickly to the 5% oxygen consumption

point but then precipitates out to report to the releach solution (Figure 4.8). All the arsenic

that is leached out steadily for the remaining leaching time reports to the releach solution

and not to the leach solution. There appears to be an error in the assay ofthe releach solution

at the 5% point because the mass balance at this point does not add upto the total amount of

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Distribution ofIronAsa Function of Oxygen Consumption

3.0

3

$ 2.0

ID 1J

2

Leach Sol'n

Releach Sol'n

Solids

0 20 40 60 80 100

Percent Oxygen Consumption (%)

Figure 4.6 Distribution ofiron during leaching.

arsenic in the system. The releach solution assay is thought to be incorrect because the

arsenic is expected to be in the leach solution before it can precipitate. The other reason that

this point is thought to be in error is because the results of the experiment with no iron in

solution show that arsenic first reports to the leach solution and then to the precipitate

beyond the 5% oxygen point (see assay results in Appendix C).

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(a) With Iron-xl 500 (c) Without Iron-xl 500

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Distribution ofArsenicAs a Function of Oxygen Consumption

3.0 |

0 20 40 60 80 100

Percent Oxygen Consumption (%)

Figure 4.8 Distribution of arsenic during leaching.

4.1.3 Nickel and Cobalt

The leaching behavior ofnickel and cobalt is very similar (Figures 4.9 and 4.10). Both

metals leach very quickly and stay in the leach solution without any appreciable amount of

precipitation. Nickel leaches quickly in the presence ofacid and reaches a steady state by

20% oxygen consumption. Cobalt also leaches quickly in the presence ofacid but continues

to leach slowly until the end of the process. Neither metal is thought to affect the leaching

process significantly at the levels at which they are present in the system.

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1.0

Distribution ofNickelAsa Function of Oxygen Consumption

Percent Oxygen Consumption (%)

Figure 4.9 Distribution of nickel during leaching.

Distribution ofCobaltAsa Function ot Oxygen Consumption

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4.1.4 Additional Observations of Part A Experiments

The experiments done with excess arsenic in solution or using the "copper depleted"

cake yielded no unusual leaching behavior. Only the experiments done with high Cu/S

cakeand

no iron in the electrolyte showed significant differences in leaching behavior.

The experiments done with a high Cu/S ratio cake showed significant sensitivity to

agitation speed and the slurries were higher in viscosity due to more basic copper sulphate

being formed. The higher viscosity observed was not measured but was just based on visual

observations in comparison to other experiments.

The effect of no iron in the original electrolyte was discussed earlier and reported as

having a significant effect on the viscosity probably as a result of the finer precipitateformation. After six hours of leaching time, there was 4 times as much unleached material

as expected with copper sulphides still appearing in the x-ray diffraction results (see Table

4.1). The slow leaching in these experiments appears to be more a function ofhigh viscosity

and therefore mixing/gas-liquid mass transfer rather than chemistry. It must be noted that

the "no-iron" condition is almost impossible in the plant because there is always significant

levels ofiron present in the spent electrolyte and entrained in the first stage cake liquor.

4.2 Part B: Kinetic Experiments

The part B experiments can be divided into 2 sections; those performed with and without

acid in solution. The two types of experiments must be considered distinct because they take

different reaction paths in the leaching of chalcocite. For this reason, the originally suggested

factorial design analysis was not carried out on the results. The experiments done with acid in

the initial solution are not very representative of the leaching path in the plantbecause the acid

is present during the whole leaching time in these experiments. The leaching rates that are of

interest are the ones in which the acid is depleted early in the process.

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The leaching rate versus percent oxygen consumption plots areprovided inAppendix D

for the acid and no-acid experiments. These graphs were generated by subtracting the blank

experiments (no solids added) from the actual experiments to obtain net rates. The "loops"

formed at the initial part of the plot are caused by the subtraction of the two runs.

The experiments that begin with acid in solution proceed initially via reaction 1 where

Cu xS is converted to CuS. This is followed by the leaching ofCuS via reaction 5. However,

reaction 4 must occur to a limited degree because some of the sulphur is oxidized. The chemical

analyses of the leach residues contained approximately 90% sulphur and the x-ray diffraction

results show very strong elemental sulphur patterns. The degree ofsulphur oxidation is shown

in Table 4.2 for the various runs performed with acid. It should be noted that those experiments

with no copper and/or iron in solution showed higher levels of sulphur oxidation. This is

consistent with the work done by Mao and Peters [13]. They found that the presence ofiron

lowered the levels of sulphur oxidation. They observed 90.6% elemental sulphur in their

residues, values very similar to the ones observed in these experiments. All of the sulphur was

found to oxidize in the experiments done with no acid in the solution.

The times taken to 10,15 and 20% oxygen consumption are given in Table 4.3. They show

very clearly that the initial leaching rates are very fast where CuxS is being converted to CuS.

The time to 20% oxygen consumption is usually less than 3 minutes and is a very short duration

relative to the total reaction time of most ofthese experiment of 2.5-3 hours. This initial rapid

leaching of chalcocite is consistent with the work of Mao and Peters [13] as well as other

researchers.

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Table 4.2 Degree ofsulphur oxidation at various initial acid, copper and iron concentrations.

Cu/S [HjSOJ [Cu] [Fe] % Sulphur

Ratio (g/D (g/U (g/D OxidizedMedium 60 40 5 7.6

Medium 60 40 0 15.9

Medium 60 0 0 31.6

Medium 60 0 5 21.4

Medium 60 20 2.5 16.0

Medium 60 20 2.5 10.4

Medium 120 0 2.5 23.9

Medium 120 40 2.5 1.6

Medium 120 20 5 9.6Medium 120 20 0 28.4

Low 60 20 0 28.5

Low 60 0 2.5 26.6

Low 60 20 5 15.3

Low 60 40 2.5 15.4

In the experiments with no acid in the initial batch recipe, the material is expected to leach

via reaction 2 followed by reaction 4. The reactions are not expected to be entirely sequential as

the results ofpart A have indicated. In these experiments, all of the sulphur was to oxidize. The

times to 20% oxygen consumption are very fast in the no-acid experiments also.

The total leaching times in these experiments were much shorter than part A. This is most

likely due to the fact that the agitation speed was higher and the pulp density was lower

contributing to better mixing and much higher gas-liquid mass transfer rates.

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Table 4.3 Time taken to leach to 10, 15 and 20% oxygen consumption at various initial

acid, copper and iron concentrations.

Cu/S [HjSOJ [Cu] [Fe] Time (s) to %consumptionRatio ( g / U ( g / u (g/D 10% 15% 20%

Medium 0 20 5 90 142 305

Medium 0 0 2.5 100 175 334

Medium 0 40 2.5 80 105 186

Medium 0 20 0 162 332 499

Medium 60 40 5 78 92 107

Medium 60 40 0 89 110 143

Medium 60 0 0 84 110 174

Medium 60 0 5 95 109 122Medium 60 20 2.5 115 129 146

Medium 60 20 2.5 90 108 129

Medium 120 0 2.5 168 284 306

Medium 120 40 2.5 83 98 116

Medium 120 20 5 100 116 137

Medium 120 20 0 90 113 141

Low- 0 20 2.5 230 399 559

Low 60 20 0 137 250 640Low 60 20 2.5 89 105 125

Low 60 20 5 92 110 136

Low 60 20 2.5 132 152 180

4.3 Part C: Leaching CuS in the presence of basic copper sulphate

Figures 4.11 and 4.12 show the oxygen consumption rates at regular intervals during the

leaching ofCuS in the presence of basic copper sulphate. The rate of oxygen consumption at

low copper concentrations, between 1 and 10 g/L, appears to be slower. A minimum point

probably exists somewhere in this range but it is not possible to depict without more data. This

dip in rate is apparent in both the experimental conditions of iron and no-iron in solution.

Beyond the minimum point, the rate of oxygen consumption (CuS leaching) generally appears

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to be copper catalyzed and increases with increasing copperin solution.

Oxygen Consumption RateAlVarious Copper Concentrations

0.0 I 1 1 1 1 : 1 1 1 ' 0.00 20 40 60 80

Initial Copper Concentration (g/L): [Fe]=0.25

Figure 4.11 Oxygen consumption rate at various initial copper concentrations with initial

[Fe]=0.25 g/L. Each curve represents a different level of oxygen consumption.

Comparing the leaching rates of experiments done with and without iron in solution, it

appears that iron in solution increases the leaching rate. Appendix F contains plots ofrates at

various oxygen consumption points with and without iron in solution. There is a crossover of

rates between the 0 and 10 g/L Cu points.

Table 4.4 and Figure 4.13 show that increases in the initial copper concentration or iron

concentration results in an increase in the oxidation-reduction potential (ORP). This is

consistent with the Nernst equation, e.g.:

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Figure 4.12

Oxygen Consumption RateAt Various Copper Concentrations

20 40 60

Initial Copper Concentration (g/L): [Fe]=0

10%20%30%

40%50%60%70%80%

Oxygen consumption rate at various initial copper concentrations with initial

[Fe]=0 g/L. Each curve represents a different level ofoxygen consumption.

Cu2+ + e'^Cu+

R T [Cu+

1EL = E - I n -

1

nF [Cu2+]

A higher solution potential will tend to increase the leaching rate by imposing a higher

exchange current on the mineral. Figure 4.14 shows a schematic of an Evans E>iagram of the

possible polarization curves during CuS leaching. The reversible potential ofCuS leaching as

shown on Figure 4.14 at 388 Kis approximately 0.21 V. The actual ferric and cupric polarization

curves will be a result of the mixed potential caused by both iron and copper in solution. Theiron in solution is reported to be easily oxidized from ferrous to ferric in the presence of copper

as the cupric-cuprous couple is thought to catalyze the oxidation of the ferrous species [36].

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The increase in the leaching rate observed at higher copper concentrations and higher iron

concentrations is now more understandable. The increase in the cupric/cuprous, ferric/ferrous

ratios or a net increase of the cupric or ferric concentrations will increase the mixed solution

potential and as a result increase the anodic current (leaching rate).

The dip in the solution potential, and consequently the leaching rate, observed at

approximately 10 g /L copper concentration was not explainable. This point is not likely to be

in error because it occurs in both the experiments (ie. [Fe]=0 and [Fe]=0.25 g/L).

Table 4.4 Potential and pH measurements of the leach slurry after leaching

and the approximate temperatures at which they were measured.

Initial Initial pH ORP (mV) Temperature CO

[Cu] (g/L) [Fe] (g/L)

0 0.25 2.84 350 80.1

1 0.25 2.79 353 77.9

10 0.25 2.53 319 81.5

40 0.25 2.51 491 84.6

80 0.25 2.43 499 83.0

0 0 2.78 314 85.9

1 0 2.68 360 82.9

10 0 2.50 321 84.0

40 0 2.68 445 86.9

80 0 2.43 449 85.1

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>E

550.0

500.0 -

g 450.0 -

ocug 400.0 -

o350.0

300.0

Solution Potential

at Various Copper Concentrations

10 20 30 40 50 60Initial Copper Concentration (g/1)

80

Figure 4.13 The effect ofinitial copper concentration on the measured ORP at [Fe]=0 and

[Fe]=0.25 g/L.

Evans Diagram Schematic ofApplicable

PolarizationsCurves During CuS Leaching

1.0

0.8

0.6

o

> 0.4

0.2

0.0

-0.2

Increasing ferric or cupric ion

concentration

CuS+ 4HjO-- Cu + SO*+ 8H++ 8e_

Log I - Galvanic Current (Schematic)

Figure 4.14 Evans diagram ofapplicable polarization curves during pressure leaching of

CuS (schematic).

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4.3.1 Discussion ofORP Measurements

Since there is a small amount ofsolids present in the slurry when the ORP is measured,

the value obtained is really a "mixed potential" because it is affected byboth the solids

present and the prevailing solution conditions. However,since the solids concentration is

very low in the part C experiments, the measured ORP value is much closer to the solution

potential. A simple calculation shows that the solid potential changes very little as a result

of the change in copper concentration.

The equilibrium reaction is:

Cu2+ + SOl~ + 8/Y+ + &e CuS+ 4H20

This potential will vary as a function of[Cu2+]:

ie RT 1' E = E- 2.303log

a

Cu**

a

*-a

+

And assuming: asol- = aCuit

2.303(8.314) (298.15), 2 2.303(8.314) (298.15) E = E +

8 965001l 0 g a

(96500)p H

E = E+ 0.0148 \ogaCult - 0.059\6pH

Therefore, the "solid potential" will vary by -15 mV per order ofmagnitude change in

copper concentration.

The cathodic reaction is:

02 + 4ht+4e -*2H20

However, it is also possible for other reactions to catalyze the reaction:

Fe3+

+ e -> Fe2+

Cu2+

+ e -> Cu

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The observed changes in the ORP measurements are large (Table 4.4). However,

according to the calculation above, the "solid potential" can only change by -15 mV per

order of change in the Cu 2 + concentration. It is therefore more likely that the observed

change in the ORP measurement is a result of the Cu 2 + /Cu + and/or Fe 3 +/Fe 2 + couples. An

increase in potential ofthese couples increases the rate of electrochemical dissolution and

hence increases the rate ofleaching. This observation is important since it indicates that the

levels of copper and iron in solution are important as catalysts and. may be serving as

surrogate oxidants.

4.4 Comparison ofLeaching Rates - Part A and Part C

A comparison of leaching rates between part A and part C experiments was performed todetermine an empirical relationship between them (see Table 4.5). The main difference between

these experimental section was the pulp density and agitation speeds. The comparison ofrates

is done at 60and 70% oxygen consumption for the following reasons:

1. The copper concentrations are similar in these experiments.

2. The predominant material (solids) remaining is CuS.

3. The flow rates are extremely stable and are not affected by initial transients.

4. There is no manual oxygen flow control, as was practiced in the early stages of the

part A experiments4.

Ifthe ratio ofrates is taken between these two parts, it can be seen that the rate of leaching

is 4.5 and 4.2 times as high in the 60% and 70% runs respectively in the part C experiments. This

lends further credence to the fact that the rates must be dependent upon agitationand/or pulp