High Temperature Roasting of Sulphide

Concentrate and its Effect on the Type of

Precipitate Formed.

A dissertation submitted to the School of Mines, Faculty of Engineering,

Technikon Witwatersrand, Johannesburg, South Africa, for the fulfilment of

the degree of

MAGISTER TECHNOLOGIAE: EXTRACTION METALLURGY.

By

Fortunate Magagula

Supervisor: Dr. A.F. Mulaba-Bafubiandi

Department of Metallurgy, Technikon Witwatersrand,

Johannesburg

December 2002.

Declaration

I Fortunate Magagula, hereby declare that this dissertation is my own unaided work. It is being

submitted to the Technikon Witwatersrand for the degree MAGISTER TECHNOLOGIAE

Extraction Metallurgy. It has not been submitted before by myself or any other person to any

institution for any degree or examination.

Author's signature Date20421 eiq

ii

Acknowledgements

I sincerely thank my supervisor, Dr Mulaba for his academic guidance and logistical support and

for making this study a success. I would also like to thank the Technikon Research Committee

for the financial support. I thank the University of the Witwatersrand for allowing me to use their

facilities. I am so grateful to my colleagues for their moral support. I would like to extend my

gratitude to my husband for loving and motivating me throughout this study.

iii

Dedication

To my husband who made everything possible and bearable. Thank you for your tireless

support.

iv

Abstract The most commonly used route in the hydrometallurgical extraction of zinc and copper is the

roast-leach-electrowin process. During the roasting process, the concentrate is subjected to

either relatively low temperatures (partial roasting) or high temperatures (to achieve dead

roasting) to produce a calcine that will be leacheable to extract zinc and copper. The resulting

calcine contains zinc and copper in a form of oxides (ZnO, CuO), sulphates (ZnSO4, CuSO4)

and ferrites ((Zn,Cu 1-x, Mx)0Fe203) or Zn,CuFe2O4) in the case of partial roasting. In the case

of dead roasting, mostly the oxide forms are produced but in most cases ferrites will form as

well.

The means of avoiding the ferrites completely have not yet been achieved. Attempts in the past

had only been focusing on either partial roasting or dead roasting without actually finding the

optimum roasting conditions to minimise the ferrite formation. In this study the main objective

was to identify optimised conditions for roasting, i.e. the possibility of producing these ferrites in

minimum amounts as compared to the targeted zinc/copper oxides. Optimised roasting

conditions were achieved in this study on a Zinc-copper ore from Maranda mine, where the

sulphur removal test was used to ensure a dead roasting. This was done by analysing the

amount of sulphur remaining after each roasting condition. Characterisation of the calcine has

been done using the XRD and the Mossbauer spectroscopy. More zinc oxide than zinc ferrite

was obtained at conditions of 800 °C for 3 hours as per the XRD analyses. The sulphur removal

test however, showed a dead roasting at 900 °C (2% remaining sulphur) and this is attributed to

the inadequate (not designed as in industry) supply of oxygen by the laboratory furnaces used.

The precipitation of iron from the three acids (HCI, H2SO4 and HNO3) was done using NH4OH

and NaOH. The Mossbauer and XRD characterisation techniques were used, where the XRD

characterisation showed different spectra of the precipitate attributing to different compounds.

The results of the precipitates from the optimised roasting conditions are those precipitates that

are not commonly found in industry. The effect of the acids and the cations showed goethite to

be formed from H2SO4 and HNO3, with NH4+ and Na+ respectively.

The possibility of the selective leaching of the concentrate has been investigated. This

eliminates the roasting process completely and thus provides a possibility of leaving the pyrite

(FeS2) in the residue and thus minimising the amount of iron to be handled. Selective leaching

has been done using Mn02 and Na2S208 in the presence of H2SO4. It was observed that

starting with Mn02 as an oxidising agent does not achieve good selective leaching results

between the sphalerite and the chalcopyrite. It was however possible to preferable leach

sphalerite over chalcopyrite with the use of Na 2S2O8 as a starting oxidising agent. So the choice

of the oxidising agent plays a role in selectively leaching different minerals.

The optimised roasting conditions at high temperatures resulted in some type of precipitates,

(mohrite, ferrihydrite and akaganeite) that are not commonly formed in industry. Jarosite, which

is the most common precipitate formed in industry, could not be precipitated. Goethite was also

fcund to be present.

vi

Table of Contents

Declaration ii

Acknowledgements iii

Dedication iv

Abstract

Table of Contents vii

List of Tables

List of Figures xiii

List of Abbreviations xvi

CHAPTER 1 — INTRODUCTION 17

CHAPTER 2 - LITERATURE REVIEW 20

2.1 Introduction 20

2.2 Iron in Sulphide ores 20

2.3 Roasting of sulphide ores 21

2.3.1 Partial roasting 21

2.3.2 Dead roasting 24

2.4 Leaching 25

2.4.1 Conventional leaching of sulphide ores 25

2.4.2 Selective leaching of sulphide ores 25

2.5 Precipitation 28

2.5.1 Hydrolysis of Iron in Aqueous Media 29

2.5.2 Iron precipitate products 32

2.6 Mossbauer 37

vii

Chapter 3 — EXPERIMENTAL 38

3.1 Research Method 38

3.2 Materials 38

3.2.1 Ore and its origin 38

3.3 Reagents and Apparatus 39

3.3.1 Flotation Reagents 39

3.3.2 Leaching Reagents 39

3.3.3 Apparatus 40

3.4 Experimental Procedures 40

3.4.1 Flotation procedures 40

3.4.2 Roasting procedure 40

3.4.3 Sulphur determination 40

3.4.4 X Ray Diffraction (XRD) 42

3.4.5 Mossbauer Effect Spectroscopy (MES) 42

3.4.6 Leaching procedures 43

3.4.7 Selective Leaching procedures 43

Chapter 4—RESULTS AND DISCUSSION 45

4.1 Introduction 45

4.2 Flotation Results 45

4.3 Roasting Results 49

4.3.1 Sulphur removal 49

4.3.2 X-Ray Diffraction (XRD) 52

4.3.3 Mossbauer Results 60

4.4 Leaching Results 74

4.4.1 Neutral leaching 74

4.4.2 HCI Neutral leach 75

4.4.3 H2SO4 Neutral leach 76

4.4.4 HNO3 Neutral leach 77

4.4.5 CONCLUSION 82

4.5 Selective leaching 83

4.5.1 Conclusions 90

viii

4.6 Precipitation results 92

4.6.1 Procedure 92

4.6.2 Results 94

4.6.3 XRD characterisation of precipitates 101

4.6.4 Mossbauer characterisation of precipitates. 102

Chapter 5 — Conclusions 104

5.1 Recommendations 105

References 106

Appendices 111

ix

List of Tables

Table 2.2.1.1: Comparison of Iron Precipitation Processes 35

Table 3.1.1.1: Mineral and metal abundances in the Run Of Mine. 39

Table 4.2.1: Results showing mass recovery with respect to variation of modifiers types and ratios 46

Table 4.2.2: Table showing zinc and copper recoveries 47

Table 4.3.1.1: Percentage sulphur from varying roasting conditions. 50

Table 4.3.2.1: XRD spectral Intensities of phases at different ... roasting conditions. 57

Table 4.3.3.2: Hyperfine interaction parameters of the components in the concentrate sample. 61

Table 4.3.3.3: Hyperfine interaction parameters components in calcines roasted for 2 hrs at different temperatures .

63

Table 4.3.3-4: Comparison of component abundances (2hrs) ... 65

Table 4.3.3.5: Hyperfine interaction parameters of the components in the calcines, roasted for 4 hrs at different temperatures 67

Table 4.3.3.6: Comparison of component abundances (4 hrs) .. 69

Table 4.3.3.7: Hyperfine interaction parameters of the spectral components in the spectrum of calcines roasted at 800°C

for different durations. 70

Table 4.3.3.8: Comparison of component abundances (800 °C). 72

x

Table 4.4.1.1: Table showing neutral leaching results 74

Table 4.4 4.1: Percentages of elements remaining after neutral leaching. 81

Dissolution of elements during hot acid leaching8l

Table showing percentages of elements remaining after hot acid leaching. 82

Percentage Extraction in 5M H 2SO4 and 10% (w/v) Mn02 83

Percentage Extraction in 7M H 2SO4 and 10% (w/v) Mn02. 85

Percentage Extraction in 5M H 2SO4 and 20% (w/v) Mn02. 85

Percentage Extraction in 7M H 2SO4 and 20% (w/v) Mn02

85

Percentage Extraction in 5M H 2SO4 and 10% (w/v) Na2S2O8 .

86

Percentage Extraction in 7M H 2SO4 and 10% (w/v) Na2S2O8 .

87

Percentage Extraction in 5M H 2SO4 and 20% (w/v) Na2S2O8 .

88

Percentage Extraction in 7M H 2SO4 and 20% (w/v) Na2S2O8 .

88

Percentage Extraction of residue in 7M H 2SO4 and 20% (w/v) Mn0 2

89

Table summarising the amount of iron (in %) remaining from precipitation at 80 °C and filtration after 24 hours. 94

Table 4.4.4.2:

Table 4.4.4.3:

Table 4.5.1:

Table 4.5.2:

Table 4.5.3:

Table 4.5.4:

Table 4.5.5:

Table 4.5.6:

Table 4.5.7:

Table 4.5.8:

Table 4.5.9:

Table 4.6.1.1:

xi

Table 4.6.1.2: Results summarising the amount of iron (in %) remaining from precipitation at 95 °C and filtration after 24 hours 96

Table 4.6.3.1: Hyperfine interaction parameters of the spectral components in the spectrum of samples. 103

Table A: Hyperfine interaction parameters at room temperature of candidate Fe-bearing phases that may occur in the samples

111

Table B: Hyperfine interaction parameters at room temperature of candidate Fe-bearing phases that may occur in the precipitate samples

112

xii

List of Figures

Figure 2.5.1.1: Schematic representation of the hydrolysis-precipitation process 30

Figure 4.2.1: Effect of conditioning time on mineral recoveries. 46

Figure 4.3.1.1: Percentage sulphur remaining at varying roasting conditions.observing the temperature effect 51

Figure 4.3.1.2: Percentage sulphur remaining at varying roasting conditions observing time effect 52

Figure 4.3.2.1: XRD graph showing the effect of roasting temperature on the formation of zinc oxide and zinc ferrite.... 53

Figure 4.3.2.3: Effect of roasting temperature on the formation of zinc oxide and zinc ferrite at 800 °C 55

Figure 4.3.2.5: XRD spectral Intensities of the zinc ferrite and ... zinc oxide at 700°C. 58

Figure 4.3.2.6: Graph showing intensities of Zinc ferrite and zinc oxide (zincite) at 800 °C. 60

Figure 4.3.3.1: Mossbauer spectr um of the concentrate from the zinc ore sample. 61

Figure 4.3.3.2: M6ssbauer spectra of calcines roasted for 2hrs at different temperatures (700 °C, 800°C, and 900°C). 64

Figure 4.3.3.3: Graph showing the effect of temperature on the amount of phases formed during roasting for 2 hours. 66

Figure 4.3.3.4: Mossbauer spectra of the calcine samples, roasted for 4hrs at different temperatures 68

Figure 4.3.3.5: Mossbauer spectrum of calcine samples roasted at 800°C for different durations. 71

Figure 4.4.2.1: Dissolution of the metals in HCI showing different extraction rates. 75

Figure 4.4.3.1: Dissolution of the metals in H 2SO4 showing different extraction rates. 76

Figure 4.4.4.1: Dissolution of the metals in HNO 3 showing different extraction rates. 77

Figure 4.4.4.2: Dissolution of copper in the three leaching acids. 78

Figure 4.4.4.3: Dissolution of Zinc in the three leaching acids ... showing the effect of each acid. 79

Figure 4.4.4.4: Dissolution of iron in the acids showing the effect of each acid. 80

Figure 4.5.1 Percentage Extraction in 5M H2SO4 and 10% (w/v) Mn02 . 84

Figure 4.5.2 Percentage Extraction in 7M H 2SO4 and 20% (w/v) Mn02 86

Figure 4.5.3 Percentage Extraction in 5M H 2SO4 and 10% (w/v) Na2S2O8. 87

Figure 4.5.4 Percentage Extraction in 7M H 2SO4 and 20% (w/v) Na2S2O8. 89

Figure 4.6.1.2: The comparison of HCI and H2SO4 pregnant solutions in precipitating iron using NaOH. 98

Figure 4.6.1.4: The effect of H 2SO4 and HNO3 in the amount of iron remaining from precipitating with NaOH 100

Figure 4.6.2.1 XRD spectra for precipitates where iron precipitation was optimum. 101

xiv

Figure 4.6.3.1: Mossbauer spectrum of precipitates. 102

RV

List of Abbreviations

XRD X-Ray Diffraction

AAS Atomic Absorption spectroscopy

MS Mossbauer Spectroscopy

WN Weight to volume

mV Millivolts

CP Chemical purity

ROM Runoff mine

KV Kilovolts

mA Milliamps

MES Mossbauer Effect Spectroscopy

QS Quadrupole splitting

Bhf Magnetic field

IS Isomer shift

xvi

CHAPTER 1 INTRODUCTION

1.1 Aims and objectives

The aim of this project was to optimise the roasting conditions i.e. to find the conditions that

produce ,minimal ferrites after the roasting process; to characterise precipitates that result from

the pregnant solution after leaching of calcine from the optimised roasting conditions; to

establish / find out the effect of high temperature roasting on the type of precipitate formed and

to investigate the possibility of selective leaching.

1.2 Overview

Roasting may be used to prepare sulphide concentrates for subsequent pyrometallurgical or

hydrometallurgical operations. For pyrometallurgical processing, the usual purpose of roasting is

to decrease the sulphur content to an optimum level for smelting to a matte. Partial (oxidizing)

roasting is accomplished by controlling the access of air to the concentrate; a predetermined

amount of sulphur is removed, and only part of the iron sulphide is oxidized, leaving the copper

sulphide (for example) relatively unchanged. Total, or dead, roasting involves the complete

oxidation of all sulphides, usually for a subsequent reduction process. (For hydrometallurgical

extraction, roasting forms compounds that can be leached out.)

Iron plays an important role in the production of non-ferrous metals. It is the fourth most

common element in the composition of the earth after oxygen, silicon, and aluminium, and the

second most common metal after aluminium. Mainly because of this abundance, iron may be

present as an essential constituent of the ore or gangue, as a solid solution, or may be mixed

with the ore in the form of various iron minerals (Ozberk and Minto, (1986)). Though it may be

present as an essential constituent in other ores, it is also regarded as an impurity in many non-

ferrous metals.

There have been a number of other processes on the subject of iron control, most of which were

not practiced further because of their disadvantages. To name a few, Stein and Spink, (1990)

made some developments for partial oxidation roasting of zinc concentrates, which afforded a

solution to the ferrite problem. In this process, complete avoidance of zinc ferrite formation can

be attained with resultant higher overall recoveries of zinc than are presently achieved via the

17

conventional dead-roast-leach-electrowinning process. However this results in some zinc being

undigested in the leach residue and has to be recycled in the roasting circuit.

In this study the behaviour of iron compounds in a concentrate is being observed as the

concentrate is subjected to high temperature roasting. Roasting was done at different roasting

conditions. This was to determine the ferrite formation and to find conditions that favour the zinc

oxide formation over the ferrites. Conditions that promote ferrite formation over the zinc oxide

must therefore be scrutinized to keep it as minimal as possible. Optimization of the roasting

process has been investigated by studying the calcine using the XRD technique in conjunction

with the sulphur removal test and M6ssbauer spectroscopy. The ore used was a zinc-copper

ore from Maranda mine in the Murchison Greenstone belt in South Africa.

As it will be stated in the literature review, studies have been made on partial roasting where the

temperatures are kept low; this study focuses on the effect of high temperature roasting. The

resulting calcine is leached and iron precipitated from the pregnant solution. The precipitated

iron using different bases (NH4OH and NaOH) has been studied to see the effect of these

cations. The possibility of selective leaching has also been studied as an alternative to the

roast-leach route.

The behaviour of individual sulphide minerals can aid in the understanding of selective leaching

of a zinc-copper sulphide ore where pyrite exists. As it eliminates the roasting step, it thus

reduces the energy consumption and also the purification step will be cheaper since less iron

will have to be managed. Selective leaching has been done in this study using Mn02 and

Na2S208 as oxidising agents in the presence of H2SO4. Though the optimum concentrations

have not been established, the increase in concentrations has shown increase in the

percentage extraction. The choice of the oxidising agent has been observed to play a role in

selectively leaching each mineral over the other. Sphalerite was found to be leached first if

Na2S2O8 was used first than when Mn02 was used.

The thesis is divided into five chapters. The first chapter introduces the work, what it focuses on,

its aims and problem statement. Chapter two gives the literature review. The main theme in

chapter two is to outline what has been done previously in the field of iron removal and why the

other approaches on the subject of iron control have been unsuccessful. In the third chapter, a

methodology is outlined on the experiments conducted. This includes experiments from the as

18

received ore, how the liberation size was achieved, the concentration process, roasting

experiments and the characterisation techniques of calcines, leaching experiments, iron

precipitation experiments and the selective leaching experiments.

Chapter four gives a detailed report on the results and their discussion. Some mini conclusions

are stated at the end of some discussion of results. Chapter five gives a summary, conclusions

and recommendations on the project.

1.3 Problem Statement

Iron is invariably associated with most minerals. Its presence in significant amounts results in it

locking a significant amount of the desired metal in a ferrite form, which is produced during

roasting. The roasting process has to be optimised to minimise the production of ferrites.

19

CHAPTER 2 - LITERATURE REVIEW

2.1 Introduction

The most commonly used route in the zinc-copper industry is the roast-leach-electrowin

process. The roasting conditions depend on industry preference and especially the advantages

that each industry would prefer over the other. Partial roasting and dead roasting have been

tried and are practiced by some industries. The presence and amount of iron in the concentrate

also plays a major role in determining the choice of the roasting process where the products

from each roasting process differ from each other. The cost of the iron removal process is

dependent on the roasting process being used. Selective leaching is another alternative that if

properly done, would eliminate the iron removal process.

2.2 Iron in Sulphide ores

Zinc occurs in nature mainly as the sulphide (ZnS), which is mineralogically known as

sphalerite. Various iron minerals generally accompany the occurrence of zinc in these sulphide

deposits, (Elsherief, 1999). Conventional zinc concentrates i.e. beneficiated zinc ores; typically

contain 5 — 10 percent iron. The iron commonly associated with the zinc concentrates can be

present as either a replacement for zinc in sphalerite (ZnFeS2) or marmatite (Zn,Fe)S which is a

variety of sphalerite or as separate minerals such as pyrite, pyrrhotite, or chalcopyrite. This

makes the disposal of iron an integral part of the design and operation of zinc refineries,

(Dutrizac, 1987).

Deposits of metallic copper have been mined in several parts of the world. Currently, however,

copper is found naturally as simple or complex sulphides or as compounds such as hydroxides

or carbonate produced from sulphides by local weathering. There are some copper sulphides of

economic importance that are associated with sulphides of Iron. They are chalcopyrite (CuFeS2)

and Bornite (Cu5FeS4) with copper contents of 34.6% and 63.3% respectively, (Ozberk and

Minto, 1986). The roasting of the concentrate produces some ferrites which at times are

combined with the zinc in the case where the concentrate contains sphalerite and chalcopyrite.

Sphalerite present in the ore usually reports in the concentrate (though in small amounts).

20

When leaching the calcine, besides copper, oxides of iron present in the ore are also leached. It

is therefore necessary to exercise control over the amount and strength of the acid to be used

for leaching to attain maximum copper and minimum iron extraction. When copper oxide

minerals are leached, sulphuric acid has been found to be about five times the weight of the

dissolved copper, (Ozberk and Minto, 1986)

2.3 Roasting of sulphide ores 2.3.1 Partial roasting

Stein and Spink, (1990) made some developments for partial oxidation roasting of zinc

concentrates, which affords a solution to the ferrite problem. In their process, complete

avoidance of zinc ferrite formation was attained with resultant higher overall recoveries of zinc

than were presently achieved via the conventional dead-roast-leach-electrowinning process.

In their process, the iron was maintained in its 2 + state throughout the roast by a controlled set

of roast operating conditions. However this resulted in some zinc being undigested in the leach

residue and had to be recycled in the roasting circuit. Since the zinc sulphide ore contained

significant amounts of iron, there was a formation of zinc ferrite in the roasting of the

concentrate. During the dead roasting of copper concentrates the following reactions occur:

CuFeSO4 + Heat —› Cu2S + FeS + SO 2 (2.3.1-1)

Cu2 S + 02 --> Cu2O + SO2 (2.3.1-2)

FeS + 02 —> FeO +S02 (2.3.1-3)

During roasting of the zinc-copper concentrates the zinc is tied up in ferrites and also silicates

according to the following reactions:

Fe2O3 ZnO --> ZnO • Fe2O3

(2.3.1-4)

SiO2 + ZnO --> ZnO • SiO 2 (2.3.1-5)

Chen and Cabri, (1993), studied the sulphation roasting in which the significant differences from

the conventional dead roast are the roaster operating temperatures, the method of gas cooling

and cleaning, the recycling of solutions to the roaster for thermal decomposition, also the

absence of an iron removal stage such as jarosite or goethite. In this process the temperature is

kept at

21

675 °C and the following exothermic reactions can take place:

0.87ZnS + 0.13FeS + 1.9702 —> 0.87ZnSO 4 + 0.64Fe203 + 0.13502

AH°(685°C) = 177.80 Kcal

(2.3.1-6)

.0.87ZnS + 0.13FeS + 1.8202 --> 0.29(ZnO 2ZnSO 4 )+ 0.64Fe203 + 0.4202

AH°(685°C) = 156.3 Kcal

7)

(2.3.1-

ZnO.2ZnSO4 + Fe2O3 -+ ZnO.Fe203 + 2ZnSO4

AH° (685°C) = —1.79 KCal

CuFeS2 + 3.7502 —> CuSO4 + 0.5Fe203 + SO2

AH° (685° C) = — 309.21 KCal

CuFeS2 + 3.7502 —> 0.5(CuO.CuSO 4 )+ 0.5Fe203 + 1.5S02

AH° (685°C) = — 412.0 KCal

(2.3.1-8)

(2.3.1-9)

(2.3.1-10)

The highly exothermic character of the above reactions and the high oxidation states result form

the heat being evolved from the reacting particles at a faster rate than it can be dissipated into

the surrounding gas phase. This effect increases the temperature of the reacting particle over

the bulk temperature of the bed (685 °C) and hence the local displacement of the

thermodynamic equilibrium from zinc sulphate to the oxy-sulphate (ZnO.2ZnSO4) can occur.

This oxy-sulphate can react readily with hematite formed from the iron present in sphalerite and

/ or pyrite to produce zinc ferrite.

Avoidance of the zinc ferrite would result in the production of zinc oxide. In the case of copper, it

would results in the production of copper oxide. The roasted ore (calcine) would then be

leached with H2SO4. The principal reactions occurring during the leaching would be:

C U0 + H2SO4 Cu 2÷ + SO4- + H2O

(2.3.1-11)

ZnO + H2SO 4 —> Zn 2+ + SO4- + H 2O

(2.3.1-12)

22

Toho zinc's Annaka Refinery developed the Waelz process as a means of zinc refining, (Ozberk

and Minto, 1986). The main feature of this process is that zinc is vaporised and recovered at the

same stage and the clinker sold to cement companies as a source of iron. This eliminates the

problem of iron residual disposal.

One method that was tried by Van Niekerk and Begley, (1991) is the so-called residue fuming,

which depends on decomposition of zinc ferrite and formation of zinc sulphate by roasting at

high temperatures with sulphuric acid. The reaction is probably represented by the following

equation

ZnO.Fe 2 0 3 + 4H 2S0 4 —> ZnSO 4 + Fe 2 (SO 4 ) 3 + 4H 2 0

(2.3.1 -1 3)

The sulphate roasting process was developed in Akita refinery Materials Corporation, (Dutrizac,

(1987)). In this process the iron in the zinc concentrate is removed as Fe203 to the final residue,

which can be safely stockpiled without the possibility of the re-solution of the remaining heavy

metals.

Gula et al. (1992) developed an ion exchange process using diphonix® resin for the control of

iron from a cobalt solution to replace the bleed stream process.

Another area of considerable activity is the development of a solvent extraction route for iron for

integration into zinc hydrometallurgy to replace conventional iron removal by precipitation as

jarosite, goethite or hematite. Solvent extraction has however not yet been successfully used

commercially in the primary zinc industry (Dutrizac and Harris, 1996).

Lakshmanan et al. (1992) studied the separation of iron from acid leach solutions containing

zinc, by solvent extraction using N-alkyl hydroxamic acids. However, solvent extraction is being

used commercially for iron removal, particularly when chloride hydrometallurgy is used.

The extraction of copper from its sulphide ores also involves some iron removal processes.

Copper sulphide concentrates generally contain valuable impurities such as silver, gold,

platinum, selenium and tellurium, which because of the small quantities in which they report to

iron-dominated residues, are totally uneconomical for hydrometallurgical recovery processes.

23

Several hydrometallurgical processes have, however been developed to treat copper sulphide

concentrates (Ozberk and Minto, 1986).

In the roast-leach-electrowin process, iron is removed from the leach residue after

neutralization. In the Arbiter Ammonium-oxygen leach process, iron is removed after leaching

the concentrate with ammonia and oxygen and before the copper is electrowon, (Ozberk and

Minto, 1986).

The chloride leaching and copper electrowinning process involves two stages of iron removal.

Iron present in the gangue minerals does not dissolve in a chloride medium and remains in the

leach residue. Iron contained in chalcopyrite, marmatite and pyrrhotite, dissolves with copper

and is removed from the spent electrolyte after electrowinning of copper in a diaphragm cell

(Ozberk and Minto, 1986).

A special mechanical aid leach was developed at Great Falls and at Eitrheim to dissolve the

ferrites in the leach residues and to precipitate the iron as basic iron sulphate with calcine and

ground lime-bearing rock, (Dutrizac, 1987). A method of residue treatment, called red roasting,

which, is in fact, a sulphating roast, worked well, but the subsequent water leach to recover the

zinc sulphate seriously affected the sulphate balance in the leaching circuit.

2.3.2 Dead roasting

The calcine obtained by high temperature (800 °C and above) roasting of the sulphide

concentrate contains predominantly zinc oxide, minimal zinc sulphate, and about 10-15 percent

zinc ferrite. One of the problems of the hydrometallurgical methods of zinc extraction is the

ferrite formation during the high temperature roasting (800 °C and above) of the sulphide

concentrate. The removal of iron was a major difficulty for the industry for many decades and

was responsible for low overall zinc recoveries.

The zinc ferrite locks in about 10 — 15% of the zinc originally present in the concentrate and is

not dissolved in dilute acid and combines with residues. The efficient recovery of zinc metal

requires rejection of iron residue in a form that minimises the zinc entrainment. Recovering zinc

from these residues can be achieved by leaching with hot sulphuric acid, but under such

24

conditions a substantial portion of iron also dissolves. The reaction for the dissolution of zinc

ferrite is shown below:

ZnO Fe203 + 4H2SO4 --> ZnSO4 + Fe(SO4 )3 + 4H20 (2.3.2-1)

The maximum leach temperature was restricted to below the melting point of the elemental

sulphur (119 °C) because the molten element sulphur, formed in the leach, coated the partially

reacted metal sulphides and thus limited zinc extraction (Van Niekerk et al, 1991).

The amount of zinc ferrite formed is directly proportional to the amount of iron present in the

concentrate. This implies that the higher the iron level in the concentrate the more important and

costly its removal is. The formation of zinc ferrite during roasting must be minimised to avoid

unnecessary re-treatment of residues to recover the zinc. It is therefore important and

necessary to optimize the roasting conditions that will result in less zinc ferrite formed.

2.4 Leaching Leaching, in general, is a process whereby a solid and a liquid chemically react and all part of

the solid is dissolved as a soluble species in the liquid solvent. The leaching of sulphides can be

conducted by a number of leaching methods. In this study, the conventional and the selective

leaching were used. The hydrometallurgical processing of complex concentrates represents an

ecologically attractive alternative with respect to classical pyrometallurgical technologies.

2.4.1 Conventional leaching of sulphide ores

Conventional leaching is mostly done after the roasting process. The calcine is in the form of an

oxide and thus easy to be leached in neutral sulphuric acid, which is normally used in industry.

The leaching acid in this case is aimed to attack all the oxides. Some ferrites do not dissolve in

the neutral leach and are therefore leached at rather intense conditions (high temperatures and

stronger acid concentration). This also attacks the ferrites simultaneously.

2.4.2 Selective leaching of sulphide ores The overall recovery of the metal and the difficulty of separating it from impurity metals are

generally governed by the efficiency selectivity of the leaching process (Lakshmanan et al,

1992).

25

Leaching is described as a heterogeneous (that is solid/liquid) reaction, as opposed to a

homogeneous reaction with all the reactants in a single phase (e.g. all in solution as soluble

species). Leaching can occur in a number of ways and various models or descriptions are used

to describe these ways (Junea et al, 1996). There are basically two types of leaching reactions:

those in which an oxidised metal compound is dissolved in a reagent solution:

CuO H2SO4 CuSO4 + H2O

(2.4.2-1)

Those in which the metal or compound must be oxidised during leaching:

CuS + H2SO4 + 2 02 —> CuSO4 + H2O + S

(2.4.2.2)

The oxidative leaching of metal sulphides is a complex process involving a number of possible

chemical reactions in parallel and series. In addition, sulphide minerals display a variety of

complex crystal structures, with replacement metal atoms often present and variations from the

ideal stoichiometry. This diversity of character between sulphide minerals and even between

samples of the same sulphide makes it almost impossible to predict accurately their behavior

during oxidative leaching, (Junea et al. 1996).

The overall leach can be limited by:

Mass transport of reactants and products in solution

In general, adequate transport of chemical species in solution can be achieved by suitable

mixing equipment.

Diffusion of reactants and products.

Films of reaction products can often form on the surface of sulphide minerals undergoing

oxidative leaching. The most common film forming materials are iron oxide, sulphur and

insoluble sulphates. The formation of an iron oxide film is largely dependent on the pH value of

the solution. With pyrite at a pH above 3.0, iron oxide will usually precipitate on the surface.

Sulphide sulphur also occurs during oxidative leaching and can transform in a variety of

oxidation states. The temperature affects the nature of the elemental sulphur. Below 118 °C the

sulphur is usually porous and the film does not inhibit the rate of oxidation. Above 118 °C (which

26

is close to the melting point of sulphur) the sulphur forms a very effective barrier against further

rapid oxidation. In the case of chalcopyrite, inhibition of the oxidation by a film of sulphur can

occur at a temperature below 118 °C, especially in the sulphate system.

c. Chemical reactions at the surface

In systems with adequate mass transfer and where no surface films inhibit the leaching, the rate

can be controlled by the heterogeneous reactions at the sulphide surface. Rates controlled by

chemical reactions with relatively high activation energies are significantly enhanced by

increasing temperature according to the Arrherius relationships: (Junea et al, 1996)

Rate a K exp RT

K is a constant

E is activation energy

R is the gas constant

T is absolute temperature

(2.4.2-3)

For many oxidative leaching systems, the rate will be more than double for every 10 °C rise in

temperature. Ferric sulphates can leach the sulphide according to the following mineral reaction.

(Metallurgical test work and research report, Australia, 1989)

MS + Fe(SO4 )3 —> MSO4 + 2FeSO4 + S (2.4.2-4)

Sulphide minerals display great variations in their response to leaching. Dutrizac and Palencia,

(2002), studied the effect of the iron content in sphalerite on its rate of dissolution in ferric

sulphate and ferric chloride media. They observed that the leaching rate increased in a linear

manner with increasing iron content in both cases. A study was also done on the dissolution of

chalcopyrite in ferric sulphate and ferric chloride media. Rates were observed to be faster in the

chloride system. The activation energy was found to be about 42KJ/mol and 75 KJ/mol in the

ferric chloride and ferric sulphate respectively. Leaching was found to be independent of the iron

concentration in the sulphate system, as opposed to the leaching of sphalerite in the same

media (Dutrizac, 1981).

27

A study on hydrochloric leaching of a complex zinc sulphide ore was investigated with the

objective of obtaining selective dissolution. Only electrolysis under an applied potential of A300

mV permitted selective dissolution (Elsherief, 1999). Dissolution of chalcopyrite in a

hydrochloric acid medium in the presence of manganese dioxide was studied. Chalcopyrite did

not dissolve independently but underwent oxidative dissolution in the presence of manganese

dioxide (Devi et al, 2000).

Oxidative leaching of chalcopyrite with dissolved oxygen and / or with ferric ions was modeled

and found to be promoted by high concentrations of ferrous ions in sulphuric acid solutions,

containing cupric ions. The ferrous ions were released from chalcopyrite together with cupric

ions during the leaching. This ferrous promoted chalcopyrite leaching and an auto-catalytic

process and no supply of a promoter was necessary. This process could avoid running cost

increases while improving copper extraction rates (Naoki et al, 2000).

Godocikova et al. (2002) did a study on the structural and temperature sensitivity of chloride

leaching of copper, lead and zinc from a complex CuPbZn sulphide concentrate, which was

mechanically activated. They observed that mechanical activation influenced the leaching

kinetics and recoveries of copper and zinc. Their suggested order of structural sensitivity was

galena>chalcopyrite>sphalerite, in accordance with the temperature sensitivity.

The understanding and application of the leaching behavior of individual sulphide minerals can

be used to device and optimise more efficient and selective leaching processes for the

sulphides in South African mining industry. This could be an alternative to the roast-leach-

electrowin process and the problem of iron removal with costs being minimized.

2.5 Precipitation

The precipitation process mostly governs the removal of iron in a filterable form. Precipitation

commonly involves the addition of a solution containing a precipitating agent to an aqueous

solution of the desired metal. The colloidal chemistry involved is quite complex, such that the

28

procedures to attain the desired gel structure in the precipitate have been developed through an

empirical research (Rao, 1992 and Schwertmann, 1995).

According to Schwertmann (1991) and Skoog et al, (1992), a precipitate forms through the

process of nucleation and crystal growth, and the size of the particles formed is being

determined by the rates at which these two processes occur. The initial step in the formation of

a colloidal suspension is nucleation (Mullin, 1961). Further precipitation then involves a

competition between additional nucleation and growth of the existing nuclei. The rate of

nucleation has been shown to increase exponentially with increasing relative super saturation

(Nielson, 1964).

Saturation is the extent by which the solubility limit has been exceeded. The relative super

saturation is given by the following equation (Nielson, 1964).

Re lative Super Saturation =

Where

(2.5-1)

Q is the concentration of the solid at any instant

S is the equilibrium solubility.

In contrast to the nucleation reaction, the rate, of particle growth is only moderately enhanced

by a high relative super saturation. Thus, when the relative super saturation is high, nucleation

is the major precipitation mechanism and this results in formation of a large number of small

particles. When the relative super saturation is low, the particle growth predominates and a

small number of large particles are produced. When the two mechanisms are present at almost

equal rates, particles with a broad particle size distribution are produced.

2.5.1 Hydrolysis of Iron in Aqueous Media

The study of the precipitation of a metal hydrous oxide, M203.xH2O, from solution is full of

experimental and conceptual difficulties related to multiple reaction pathways that occur

simultaneously and are difficult to separate. The hydrolysis of ferric ions can however be

29

OH

Fe(OH) 2+ "4-111' Fe OH -

A

described in a simpler approach as a series of hydrolytic polymerisation reactions involving

deprotonation of the original hexa-aqua, Fe • (H 2 0)63+ ion to form hydroxo- and oxo-species

(Blesa, 1989).

In general hydrolysis of ferric solutions is readily induced by addition of a base. Doumsa and de

Bruyn. (1979) studied the hydrolysis and precipitation of ferric oxide and oxy-hydroxide from a

Fe(NO3)3 solution. They elected Fe 3+ concentrations, temperature and ionic strength as

important variables during the homogeneous based titration of the acidified solutions. Using

titration methods coupled with optical density measurements, they distinguished between four

stages in the hydrolysis-precipitation process. These are (i) hydrolysis to monomers and

dimmers; (ii) reversible, rapid growth of small polymers; (iii) slow formation of large polymers;

and (iv) precipitation of a solid phase. The schematic representation of these stages is shown in

Figure 2.5.1.1.

— 6 +

FiVVH\e

O H O H n +

Ne/ Ne/

\1 \117 \ n/2

Figure 2.5.1.1: Schematic representation of the hydrolysis-precipitation process (Doumsa and de

Bruyn, 1979).

Doumsa and de Bruyn, (1979) investigated the role of chloride ions on the formation of Fe 3+

oxyhydroxides by homogenous injection of an alkali (NaOH) into acidified Fe 3+ solutions with

Cl/Fe ratios varying between 0-12.8 and found that chloride is an important constituent of the

polynuclear particles formed during the early stages of titration (pH<2). With an increase in the

amount of base added, this anion was found to be largely displaced by hydroxyl ions.

nH + n/2

30

An investigation was done on the mechanism of formation of the Fe3+ oxyhydroxides and oxides

from the hydrolysis of Fe3+ salt solution at elevated temperature (Music et al. 1994). The

hydrolysis of Fe3+ ions in nitrate and chloride solutions was found to commence with the

formation of simple goethite complexes. This process was followed by the formation of

polymeric species from the monomers. It has been shown that the goethite polymers in the

nitrate solutions do not have nitrate ions in the polymer chain, whereas the polymers formed in

chloride solutions contained some chloride ions in the place of the OH - ions.

Murphy et al, (1976 a, b, c) observed that the nature of the anion does not affect the initial

products of iron polymerisation, but affects the subsequent ageing process. The hydrolysis of

iron from a solution containing sulphate anions has also been studied (Doumsa and de Bruyn,

1979). In this study it was noted that the SW,± /Fe ratio in the solution had a marked effect on

the titration curves obtained by plotting the pH of the solution against the amount of the base

added. A similar study was done by Music et al, (1994) where he the formation of a (FeSO 4 )+

complex in a solution of ferric sulphate was observed to suppress the polymerisation process

and the formation of oxyhydroxides and oxides. Instead basic Fe (III) sulphates were formed.

Cornell et al, (1989) in their review, stated that the addition of sufficient base to give OH/Fe ratio

greater than 3 immediately lead to precipitation of poorly ordered ferric hydroxide of which the

degree of ordering depended on the method of preparation and the time of ageing. These

precipitates resembled the mineral ferrihydride (Fe5HO8.4H20) and showed some similarity to

the oxyhydroxide core of ferritin. Thermodynamic data indicated that ferrihydrite is unstable and

with time transforms into more stable, crystalline oxides such as a-FeO(OH) and a-Fe203, which

form via different mechanisms. Conditions that promote coagulation of particles of ferrihydrite

favour formation of a-Fe203, whereas the formation of a-FeO(OH) proceeded and most readily

at pH values that promoted dissolution of ferrihydrite.

The master variable that governed the reaction products was the pH of the solution. As the a-

Fe0(OH) nucleated and grew in the solution, it became more susceptible to the effect of

solution variables than did a-Fe203 which formed within a solid phase. Other factors that

influenced the composition of the reaction products are temperature of ageing, suspension

concentration, ionic strength and pre-treatment of either ferrihydrite or the ferric solution

(Atkison et al, 1968 and Cornell et al, 1989).

31

The breakthrough for the precipitation of iron in a filterable form came in 1964 with the discovery

of a method to precipitate iron in the form of jarosite. Other methods to precipitate iron from

sulphate solutions soon followed and were put into practice. They are the goethite process,

hematite process, conversion process, basic iron sulphate precipitation and the direct leaching

of the sulphide concentrate. All these have permitted the recovery of 10-15% extra zinc, which

was previously lost in leach residues with the ferrites. The dissolution of the ferrites in

concentrated H2SO4 however leads to; (Murphy et.al . 1976a, b, c)

An increase in the impurity level, especially iron and extensive purification would be

required before electro winning;

High acidity, which must be neutralised so that electro winning would be feasible.

This led to the solution of precipitating the iron in a crystalline form easy to filter. This effect

resulted in a poor dissolution of the zinc since only the zinc oxide fraction dissolved during

leaching. It was known that ferrites would dissolve in strong acid solution at temperatures close

to boiling point, but the combined dissolution of iron and zinc in sulphuric acid led to the problem

of separating the two metals. Much research work followed in an attempt to precipitate the iron

in a filterable form.

The hydrolysis of iron (III) ions in aqueous solutions has received considerable attention through

the years. A large number of studies have been devoted to the behaviour of iron oxide

suspensions in solution in aqueous media. A review of the literature, which deals with both

naturally occurring minerals and synthesised iron oxides, makes it clear that the method of

preparation and subsequent treatment of the oxide may change drastically its surface activity.

2.5.2 Iron precipitate products

Genin et al. (1991) studied the mechanism of oxidation of ferrous hydroxide in sulphated

aqueous media and the importance of the initial ratio of the reactants. Using M6ssbauer

spectroscopy, coupled with direct recording of the pH and the electrode potential, they

concluded that the factor R, which is the ratio of the initial concentration of Fe 2+ and S0 42- to

OW ions, has a remarkable influence on the end products as well as the initial products. Genin

32

et al, (1991), also concluded that the presence of certain anions like CO3 - , NOT, CI Br and

SO4- has an influence, depending on the pH of the solution, and the oxidation mechanism of

the ferrous hydroxide.

Besides the jarosite form, which came to existence in 1964, other methods to precipitate iron

from sulphate solution soon followed. Veille Montagne of Belgium developed the goethite

process (Davey and Scott, 1976). Bryson et al. (1994) studied the batch precipitation of goethite

from sulphate solution containing ferrous and zinc cations. In this process ferric ion is reduced

to ferrous by adding zinc sulphide concentrate. Air was then injected in the hot solution at a pH

2 — 2.5 to oxidise and precipitate crystalline ferric oxide hydroxide, a-FeOOH, known as

goethite.

Many of the parameters governing the removal of iron from leach liqueurs by the goethite

process have been identified and investigated. Davey and Scott, (1976) studied the two

variations of the process, namely, the Veille Montagne and the electrolytic zinc procedures.

Both procedures were examined in sulphate and chloride liquors. They came up with some

advantages of the goethite process over the other processes and the main ones being its

superiority with respect to iron removal down to low levels and its ability to function without

added alkali metals.

The hematite process was developed independently by Dowa Mining Company of Japan and by

Ruhr-Zinc of Germany (Dutrizac, 1987). In this process, iron is precipitated predominantly as

iron oxide (Fe203) in the presence of oxygen at approximately 180 °C and a total pressure of 1.8

— 2.0 Mpa. The other process is the direct leach of zinc sulphide concentrate at an elevated

temperature and in the presence of oxygen, which was developed by Sherritt Gordon (Van

Niekerk and Begley, 1991). The last being the process in which iron is precipitated as a basic

sulphate and was developed by Zincor in Springs (South Africa), (Van Niekerk and Begley,

1991).

Murphy et al 1976. a, b, c), studied the ferric hydroxypoly-cations formed in partially neutralised

ferric solutions. They studied the growth of precipitate phases in hydrolysed ferric chloride, ferric

nitrate and ferric perchlorate solutions via electron microscopy and X-ray diffraction techniques.

33

Knight and Sylva, (1974) also studied the induction times before precipitation of iron oxides from

nitrates, perchlorates and chloride over a wide range of initial OH/Fe ratios.

Over the years much focus has been put on the chemistry of the hydrolysis of iron into a

filterable form with much emphasis on the following properties of the precipitate:

Filterable and washable,

Must not incorporate sought after metal values,

Other impurities should be controlled,

Process should yield a precipitate that is ideally marketable,

Stable final product.

Gordon and Pickering, (1975) gave an account of the precipitation processes, the comparison

being based on plant practice. Table 2.2.1.1 below is a comparison of the processes which are

most commonly used in South African companies with, some variations in the precipitation

temperature.

34

Table 2.2.1.1: Comparison of Iron Precipitation Processes Gordon et al. (1975)

Process

Goethite Jarosite Hematite

Operating

Conditions

pH 2 — 3.5 < 1.5 Up to 2% H2SO4

Temperature 70 — 90 °C 90 — 100 °C -,,,, 200 °C

Anion Any SO4- only SO4 only

Added cations

required

Nil Nat, K+, NH 4, -1 (= R) Nil

Product

Compound formed a- and fl-FeOOH Rfe3(SO4)2(OH)6 a-Fe203

(a-Fe203)

Cationic impurities Medium Low (apart from R) Low

Anionic impurities Medium High Medium

Filterability Very good Very good Very good

Fe in filtrate < 1 g/L (often < 1 —5 g/L 3 g/L

0.05 g/L)

Studies have been conducted to define the precipitation parameter, to explain the structural

relationship and to determine the extent of impurity incorporation. Most of these were carried

over synthesised iron oxides. Current practice focuses on the elimination of iron from process

streams via aqueous hydrolysis of ferric ions (Mullin, 1961). Davey and Scott (1976) and

Dutrizac (1987) concluded that the conditions employed for the forms of iron to be precipitated

from solution depend on the solution conditions. Murphy et al, (1976 a, b, c), observed the

growth of precipitate phases in hydrolysed ferric nitrate, chloride and perchlorate solutions via

electron microscopy and X-ray diffraction (XRD).

35

The Zincor basic ferric sulphate process is another method used to control iron. In this process

iron is precipitated as an easily filterable basic sulphate, achieved by careful pH control as

follows

Fe(SO 4 ) 3 + 2Zn(OH)2 —> Fe03.S03 .H20 + 2ZnSO4 (2.5.2-1)

This process is used at Zincor in Springs, (Van Niekerk and Begley, (1991))

The residue sulphate roasting process is being carried out at the Akita Zinc Refinery of

Mitsubishi Metal Corporation (Ozberk and Minto, (1986)). In this process, hot acid leach residue

from the zinc plant is subjected to flotation, where silver is removed. The first stage flotation

tailings are subjected to a sulphonation roast at 650-700 °C to decompose zinc ferrites as

follows:

ZnO.Fe203 + SO3 ---> ZnSO4 + Fe203 (2.5.2-2)

Zinc recovery from the roast is accomplished by leaching with a weak acid from the main zinc

circuit, and iron is discarded as Fe203 in the final residue.

The Grassroots Process was developed by Sherritt Gordon (Dutrizac, 1987). This process is

used when an entire roaster section has to be replaced and involves two processes. The first

process involves a neutralisation autoclave leach where most of the iron is precipitated, and the

solution from this leach is then subjected to a further neutralisation step where more iron is

precipitated. Solids from the first-stage are further leached for zinc recovery. The second

pr6cess makes use of a high-acid pressure leach, which leaches zinc and iron. Ferric iron and

concentrated zinc solution are then passed to a low-acid first-stage leach where iron is

precipitated conventionally.

An intensive study has also been conducted on the factors that affect the kinetics of nucleation

and growth and purity of iron precipitates. Bryson and Teriele, (1994) concluded that

contamination of the precipitate with zinc is shown to be approximately proportional to the zinc

concentration in solution. Also, from measuring particle size distribution, he concluded that

nucleation occurs as a result of weak outgrowths being dislodged from growing particles. As a

36

consequence he predicted that to produce relatively large particles, low stirrer speeds should be

used.

Also tried on the field of precipitation, was the removal of iron from a chloride leach solution.

(Filippou and Choi 2002). It does not yet constitute a well established technology, as it happens

with iron removal from a sulphate solution. Two techniques were studied (i) oxidative hydrolysis

of FeCl2 with and without any seed at 95°C, (ii) hydrolysis of FeCI3 by controlled crystallization

with and without any seed at 95 °C. Akaganeite (13-Fe00H) was the main crystalline phase

identified in the product obtained by controlled crystallization of FeCI3 and goethito (a-FeOOH)

was the product obtained upon oxidative hydrolysis of FeCl2. (Filippou and Choi (2002)).

2.6 M6ssbauer

57Fe M6ssbauer effect spectroscopy (MES) is a very useful technique for monitoring the ore

dressing of minerals especially the iron bearing minerals. 57Fe Mossbauer effect spectroscopy is

a (nuclear) y-radiation resonance technique involving the recoil-free emission of probing

radiation by the source and the subsequent recoil-free absorption of this radiation by 57Fe nuclei

in the absorber. It is therefore extremely sensitive to the local surroundings of 57Fe atoms; these

are present in all iron bearing compounds. The local surroundings of the iron nuclei are primarily

constituted by the electronic structure of the iron atom, the compositional make-up of

neighbouring atoms and by neighbouring defect structures. It can be used for establishing the

oxidation state of iron (ferrous and ferric) and quantifying the abundance's of different iron

compounds within the same sample. Iron phase abundances as low as five percent may be

detected with high accuracy. The spectral parameters that serve as fingerprints in conventional

transmission MES are the isomer shift (6), quadrupole splitting (A) and internal magnetic field

(Bhf). They serve to characterise the chemical state of Fe in the solid phase.

M6ssbauer spectroscopy only detects Fe-bearing phases. Each spectrum may be comprised of

a superposition of sub-spectra (sextets, doublets, singlets or even a distribution of sextets)

which represent different Fe-bearing phases in the sample. Its application in this project will be

in analyses and monitoring the phase transformation of the compounds listed in the Table A in

the Appendix.

37

Chapter 3 EXPERIMENTAL

3.1 Research Method

As mentioned in the aims and objectives, a calcine with minimum ferrites had to be produced.

The method to be followed was first to produce a concentrate that would be used in the roasting

process. A description of the preparation of this concentrate is outlined in section 3.4.1. The

roasting process was done (see section 3.4.2) and its optimisation was achieved through the

characterisation of the calcine by the XRD (see section 3.4.4), MES (section 3.4.5) and the

sulphur test (section 3.4.3.2). To obtain the precipitates, the calcine produced from the

optimised roasting conditions was leached with HCI, H2SO4 and HNO3. Precipitation was

performed using NH4OH and NaOH.33

3.2 Materials

3.2.1 Ore and its origin

A zinc — copper ore was received from Maranda mine, which is located in the Murchison

Greenstone Belt. The Murchison Range is situated in the north- eastern Transvaal along the

Drankensberg escarpment. The zinc — copper mineralization occurs in a series of volcanogenic

massive — sulphide deposits

Some representative rock chips were investigated for petrographic analysis. The ore minerals

found were mainly sphalerite, pyrite, chalcopyrite and pyrrhotite, silica, traces of lead (not

determined whether a sulphide or carbonate) while magnetite occurs in minor amounts.

The composition of the ore was confirmed from mineralogical analysis and chemical analysis.

The origin sample was composed of 39% sphalerite, 3.75% pyrrhotite, 7% pyrite, 8.35%

magnetite, 0.027% Pb, 23% chalcopyrite and the remainder being gangue. The minerals are

fine and intergrown with each other. The sphalerite was present as fine inclusions in the ore and

the sample showed some form of metamorphism due to compaction.

Table 3.1.1.1 gives a summary of the ore mineralogy prior to treatment of the ore and the

mineralogy of the host rock, however the detailed mineralogy of the representative samples is

given in the appendices. There is a difference between the percentages of the three main

minerals as determined by the microscopy, and the percentages as determined by XRD

38

analysis. The reason for this is due to the approximation used in the microscopy and also the

presence of the gangue minerals.

Table 3.1.1.1: Mineral and metal abundances in the Run Of Mine.

Minerals Microscopy(%) XRD(%)

Sphalerite 39 - 35

chalcopyrite 23 18

pyrite 7 5

Elements Chemical analysis (%)

Zn 16.5

Cu 2.78

Fe 12.57

3.3 Reagents and Apparatus

3.3.1 Flotation Reagents

The pH of the slurry was adjusted using lime. Copper sulphate was used as a depressant.

Xanthate SNPX was used as a collector and Dowfroth 400 was used as a frother.

3.3.2 Sulphur test Reagents

The reagents for the sulphur test were Hydrogen peroxide, concentrated Nitric acid (density:

1.42), Sodium Carbonate, silica, concentrated hydrochloric acid (density: 1.18), Barium Chloride

Solution: 100/dm 3, concentrated Ammonium Hydroxide Solution (density: 0.90) and Methyl red

Indicator Solution: 1g/dm 3 in ethanol. (W.C. Lenahan and R.de L. Murray —Smith, 1986)

3.3.3 Leaching Reagents

The three leaching acids that were used and their results compared were the Hcl 32% CP,

H2SO4 90% CP and HNO3 55% CP. Two oxidising agents were used, MNO2 and Na2S2O8.

39

3.3.4 Apparatus

The stereo microscope was used for the macro analysis and description of the samples. The

reflected optical microscope was used for optical identification of opaque minerals. A 2L

WEMCO flotation cell was used for the flotation process. Analyses were done using an X-ray

diffraction (XRD) equipment (PW1830 Philips generator equipped with a CuK a cathode) and

57Fe Mossbauer spectroscopy.

3.4 Experimental Procedures

3.4.1 Flotation procedures

Various flotation tests were done with the sample mass ranging from 200g-600 g. Water was

added to make up the volume and the slurry was agitated at 1500rpm to wet all the particles.

The pH was then adjusted to 9.5 using lime.

The concentrations of the modifiers (Copper sulphate, Xanthate SNPX and Dowfroth 400) were

varied from test to test. The conditioning time was also varied between 5-10 minutes after

addition of each modifier. The concentrate was collected after every 2 minutes and the

procedure repeated to get 5 samples. The products were dried in a laboratory oven at low

temperature (<50 °C). The tailings were analysed and scavenged for pyrite and sphalerite where

necessary.

3.4.2 Roasting procedure

The concentrate was placed in ceramic dishes (15 cm diameter) lined with silica. About 25 g of

sample was put in one dish to make a layer as thin as possible. The roasting was done in a

laboratory furnace and the temperatures were varied from 700 °C to 1000°C. The duration of the

roasting was also varied (2, 3 and 4 hours) until the optimum roasting time was found.

3.4.3 Sulphur determination

3.4.3.1 The principle

The method used to determine the sulphur content was by means of gravimetric measurement

of the sulphur (W.C. Lenahan and R. de L. Murray —Smith, 1986). Sulphide and other forms of

40

sulphur are oxidised to sulphate by addition of bromine and 1, 1, 2-trichloro-1, 2, 2-

trifluoroethane or by addition of hydrogen peroxide. I chose to use hydrogen peroxide in my own

experiments. Decomposition is completed with a mixture of nitric acid and hydrochloric acids

and effected by fusion with sodium carbonate. Silica is dehydrated with hydrochloric acid and,

after filtration the sulphate is determined as barium sulphate by precipitating with barium

chloride.

3.4.3.2 Procedure

A mass of 0.5g of sample was transferred to a 250 cm 3 beaker. A reagent blank determination

was carried out with each batch of samples. 20 cm 3 of hydrogen peroxide was added. The

beaker was swirled and allowed to stand for 10 minutes. 5 cm 3 of nitric acid and 15 cm 3 of

hydrochloric acid were added. The mixture was allowed to stand for a further 5 minutes after

which it was evaporated on the hot plate slowly to dryness. The same procedure was repeated

three times and then 50 cm 3 water was added. The mixture was boiled and filtered (the paper

and residue was discarded). 400 cm 3 of water was added together with about 0.2 cm 3 of methyl

red indicator solution. Ammonium hydroxide solution was added drop wise to the first

permanent red colour indicator solution. The solution was boiled and while hot, 15 cm 3 of warm

barium chloride solution was added. Hydrochloric acid was added until the indicator solution

turned pink and the iron precipitate dissolved. The solution was boiled and the precipitate

allowed to settle overnight.

The precipitate was filtered and the paper put in a crucible and transferred into a muffle furnace

and ignited at 900°C. The mass of the barium sulphate was then determined by weighing the

residue and the % S was calculated as follows

%S = 0.1374x100x( a ID ) c

(3.3.3.1)

Where

a = mass of barium sulphate

b = mass of blank

c ..mass of sample taken

m = average mass of samples

0.1374 is the factor by which barium sulphate is converted to sulphur.

41

3.4A X Ray Diffraction (XRD)

All XRD patterns were collected using a Phillips PW 1830 diffractometer with an anode

consisting of a Cu element (Current (I) = 1.54A). The diffractometer was operated at a generator

voltage of 40KV and a current of 20mA with the goniometer scanning 28 values from 10 ° to 70°

at a scan rate of 1.0s/ step (step size = 0.020 ° 28). The sample to be analysed was ground in a

ceramic mortar and pestle and pressed into a disk of 1cm radius.

3.4.5 Mineralogical Analysis 10 samples were submitted for polishing in the microscopy laboratory. The samples were

studied under the microscope to determine the occurrence of the minerals and their

percentages.

3.4.6 Mossbauer Effect Spectroscopy (MES)

3.4.6.1 Procedure

Samples were received in powdered form and MOssbauer spectra were recorded in the normal

transmission geometry. An Austin Associates K3 constant acceleration motor was used to scan

the velocity range of interest, with a 57Co(Rh) source (-10 mCi) at room temperature. The

absorbers (calcine samples) were thoroughly mixed with an inorganic buffer to form a

homogeneous disk of uniform thickness in copper powder-clamp sample holders, with sample

thicknesses of about 30 - 60 mg/ cm 2. The velocity calibration was obtained by measuring the

spectrum of a 25 mm thick Fe foil at room temperature. All samples were recorded on a different

spectrometer, which gave a minimum linewidth of 0.32 ± 0.2 mm/s for the 25-mm thick

calibration Fe-foil at room temperature.

Each spectrum was recorded for about 30 - 60 hours duration. Each spectrum was collected in

1024 channels for a duration that ensures at least 500 000 counts in each channel. Prior to

fitting, each spectrum was folded to remove geometrical base-line curvature. Each spectrum

was analysed using the non-linear least squares fitting program NORMOS-90 (distributed by

Wissenschaftliche Elektronik GmbH-Germany). In the analysis each spectrum was fitted with a

six-line pattern, doublet(s) and/or singlet(s). The nuclear hyperfine interaction parameters

comprising the isomer shift IS, quadrupole splitting QS, magnetic hyperfine field Bhf, and

linewidth, were extracted from the least-squares fits to the data.

42

The integrated area under each sub-spectrum yielded at. % of iron in that particular phase. The

cases of a distribution of magnetic hyperfine fields (sextets) have been used to represent the

possibility of a range of ferrite compositions present in the sample, eg. ZnxFe3-x04, CuxFe3-x04

and/or Cu1_xZnxFe204 with a range of x values. Alternatively samples might have been

inhomogeneous with microscopic regions of the sample having Zn- or Cu-rich phases and other

regions being Zn- or Cu-depleted. The Bhf parameter distribution is meant to reflect the

distribution of Fe-local environments in the ferrites (from nearest-neighbour and next-nearest

neighbour effects), whereas parameters like QS for the distribution are considered to be similar

for each component of the distribution. Therefore, detailed features of the Bhf distribution may

not be entirely reliable, but the absorption area associated with the distribution is.

In the roasted samples, typically a magnetic field distribution has been used to model possible

cases of Zn or Cu- depleted ferrites Zn„Fe3,04, Cu xFe3_x04 and/or Cu1_xZnxFe204 with x < 0.5.

Whereas a central doublet has been used to represent Zn-rich ferrites such as Franklinite,

ZnxFe3,04 (x -1) and/or Zn xCu1_xFe204 (x > 0.5). A tabulation of the hyperfine interaction

parameters and abundances of each phase or site is presented in the appendices.

3.4.7 Leaching procedures

The leaching was done in a 1L-laboratory beaker. The density of the slurry was kept at 20%

solids. An agitation leaching was performed on a hot plate/magnetic stirrer. The leaching was

first done at different temperatures (80 °C - 95°C) and for different durations (1-6 hrs) until the

optimum leaching conditions were established. Selective leaching was done at 90°C and the

oxidising agents used were Mn02 and Na2S2O2. In the conventional leaching, the temperature

was varied between 80°C to 95°C with the use of HCL, H2SO4 and HNO3 as leaching acids.

3.4.8 Selective Leaching procedures

Selective leaching of the concentrate was done at 90 °C for duration of 6 hours and sulphuric

acid was used for the leaching process. The oxidising agents used were Mn0 2 and Na2S2O8

and were used due to availability and cost, otherwise manganese dioxide (Mn02) would have

been preferred because of its high oxidation power. The concentrate was first leached in the

presence of Na2S2O8 and the residue after 6 hours in the presence of Mn02. The leaching

43

temperature was kept at 90 °C; samples were taken at 30 mins, 2 hours, 4 hours and 6 hours

intervals for AAS analysis to determine the percentage dissolution.

All the results for the outlined experimental work are outlined in chapter 4. Their interpretation

and discussions are also included in the same chapter. Some mini conclusions are also stated

in some of the sections.

44

Chapter 4-RESULTS AND DISCUSSION

4.1 Introduction

The results in this chapter cover mostly the roasting process and the precipitation process.

Though most industries use the roast-leach-electrowin route, little focus has been done on the

competitive formation of the targeted zinc oxide over the zinc ferrite. Little work has been done

to study the possibility of selective leaching on this particular ore and the results on the

efficiency of extraction are presented.

From the mineralogical analysis, it could be seen that a good liberation of the sphalerite occured

if particle sizes were at about 80 tim. Most of the chalcopyrite was very fine and of the order of

50-70 vtrn and locked in the gangue, in what is called a chalcopyrite disease (Lenahan and

Murray-Smith, 1986). The milled products of -75 vim were studied under the microscope and

showed a very good liberation for both the chalcopyrite and the sphalerite. The flotation was

therefore done on the -75pm fraction. The 'reporting of the chalcopyrite to the tails could not be

avoided much since some of it was in very fine grains and liberating it would result in production

of fines that cannot be floated.

Only a few results are displayed on the selective leaching of the sulphide concentrate. In the

selective leaching, oxidation was done using Mn02 and Na2S2O8. In some experiments, Mn02

was used first and in the other set of experiments, Na2S2O8 was used first and then the residue

leached later with Mn02.

4.2 Flotation Results

Bulk flotation experiments (10) were carried out and the results (average) are shown in Table

4.2.1, each point being a representation of the average point in the 10 experiments.

Optimisation of the conditioning time was done and after a period of 10 minutes there was no

substantial change on the percentage mass recovery of the sphalerite and chalcopyrite. This

observation is shown in Figure 4.2.1. Each data points in Figure 4.2.1 is also averaged for 10

experiments.

45

100

80

0 60 - ED.

40 - ca E 20 - 0

0

0 5 10 15 20 25 time(mins)

Figure 4.2.1: Effect of conditioning time on mineral recoveries.

The optimisation of the flotation was then done at a conditioning time of 10 minutes. The

flotation reagents used at different concentrations were;[CuSO4 (30-60 g/t), Xanthate (30-60g/t)

and Dowfroth (25 and 30g/t)]. As can be seen from Table 4.2.1, the highest mass recovery was

achieved with a condition of 50 g/t Xanthate and 30 g/t Dowfroth. Copper sulphate as a

depressant was also used but it was observed that doses more than 30 g/t did not have any

significant impact on the mass recovery of the concentrate.

Table 4.2.1: Results showing mass recovery with respect to variation of modifier types and

ratios.

Sample Mass (g) CuSO4 (0) Xanthate

(Mt)

Dowfroth

(Mt)

%

Recovery

1 600 60 60 30 79.9

2 600 50 30 30 78.6

3 500 50 60 25 60.1

4 200 50 50 25 67.6

5 200 30 40 25 62.7

6 200 60 30 25 54.7

7 200 30 60 25 69.8

8 200 60 50 25 68.0

9 200 40 50 30 89.1

10 200 30 40 30 83.4

11 200 40 30 30 76.3

12 200 40 40 30 80.9

46

The highest mineral recovery was obtained with conditions for sample 9 where the

concentrations are 40 g/t CuSO4, 50g/t Xanthate and 30 g/t Dowfroth, as shown in Table 4.2.1.

The concentrate was collected at intervals of 2 minutes to compare the mass recovery to the

metal at a given time to enable optimum the collecting times of the concentrate.

It is to be noted that the highest mass recovery of the floated sample did not necessarily mean

optimum conditions for metal recovery. The concentrates were digested and the metal

concentrations analysed using AAS. The results are shown in Table 4.2.2.

Table 4.2.2: Table showing zinc and copper recoveries.

Sample % Zinc recovery % Copper recovery

1 85.5 73.0

2 87.0 74.2

3 79.3 70.4

4 72.7 69.8

5 77.9 69.9

6 91.2 80.2

7 89.4 87.7

8 87.1 88.6

9 89.3 80.0

10 94.3 86.3

11 91.7 83.7

12 96.7 88.1

Conditions for sample 9 gave the highest concentrate (mass) recovery (Table 4.2.1) but

analysis showed that the percentages for the minerals were a bit lower than when using the

conditions for sample 12 (40g/t CuSO4, 40g/t xanthate and 30g/t Dowfroth) (see Table 4.2.2).

The rest of the concentrate was then prepared using the conditions for sample 12 (40g/t CuSO4,

40g/t xanthate and 30g/t Dowfroth) which showed a good metal recovery for both the zinc and

the copper. The lesser amount of mineral recovery in the conditions for sample 9 is attributed to

the excess of Xanthate (50g/I as compared to 40g/I in sample 12).

The concentrate was further analysed using the XRD technique to determine the efficiency of

the flotation process. The comparison of the ROM and concentrate XRD-diffractogram is shown

47

ROM

CONCENTRATE

40 50

Two Theta Degrees 10 70

Fig

in Figure 4.2.2. Only the phases of interest (sphalerite, chalcopyrite and pyrite) were identified in

the spectra.

ure 4.2.2: XRD Spectra showing the efficiency of the flotation process.

From Figure 4.2.2, extra peaks attributed to the gangue were observed at 20 lower than 27 ° for

the diffractogram obtained for the ROM sample. Muscovite (2A =13), chlinochlore (20 = 25 °) and

other gangue materials (20 = 20° and 28°) were observed in high amounts in the diffractogram.

A very weak peak, which is characteristic of a silicate was observed at 20 :.--- 22 ° in the sample,

(Nakbanpote, 2000). This indicates almost an absence of a silicate peak and suggests a

dolomite characteristic of the rock. Pyrite, observed at 20 =33.082 °, 56.277° and 37.105° are

also noted to be decreasing in intensity in the diffractogram obtained for the concentrate. The

48

sphalerite and chalcopyrite peaks, in terms of their intensities, have increased significantly in the

diffractogram of the concentrate. Some peaks attributing to the gangue minerals were still

observed in the concentrate diffractogram but in very small amounts, showing how efficient the

flotation process has been.

4.3 Roasting Results

4.3.1 Sulphur removal

All the roasting experiments were done in the laboratory furnaces. It was however difficult to

simulate the industry conditions because of the following limitations:

Though the roasting at industry is conducted at about 930-960 °C, and done in a fluidised

bed, it was difficult to simulate the fluidised bed roaster.

The roasting duration could not be exactly like in industry since the roasting duration is

dependent on the size of the roasting bed and the feed to the roaster so the laboratory

roasting was done on ceramic dishes coated with silica and the concentrate layer was

about 0.5 cm thick. It was not possible to address the bed roaster because of the kind of

furnaces used in the laboratory and the oxygen supply could not be altered. Oxygen was

supplied through the back of the furnaces by the built in air conditioner.

The aim was to obtain a dead roasting to be able to study the resulting products using the

laboratory furnaces. This was to optimise the conditions that favour mostly the zinc oxide

formation over the undesired zinc ferrite. The sulphur test was done to analyse the amount of

sulphur remaining at the different roasting conditions and to ensure a dead roasting. The

roasting temperatures were varied from 700 °C to 1000°C for durations of 2, 3 and 4 hours. .

Table 4.3.1.1 shows the amount of sulphur remaining after roasting under the different

conditions.

49

Table 4.3.1.1: Percentage sulphur from varying roasting conditions.

TEMPERATURE

(°C)

% S after

2 Hrs

% S after

3 Hrs

% S after

4 Hrs

700 10.03 9.10 8.10

800 5.52 2.92 2.17

900 3.17 2.22 2.00

1000 1.86 1.13 1.00

Feed 17.17 17.17 17.17

While optimising the roasting conditions, the effect of time was also studied. The samples were

subjected to different temperatures for different durations to see if time had any effect on the

resulting calcine. Roasting the samples at the same temperature for a longer duration was

found to have a minimal effect, compared to varying the temperature. Figure 4.3.1.1 shows the

comparison between the effect of time and that of temperature.

50

% S

Re t

ain

ed

700 750 800 850 900 950

1000

Roasting Temperature ( °C)

Figure 4.3.1.1: Percentage sulphur remaining at varying roasting conditions observing the

temperature effect.

Figure 4.3.1.1 shows the amount of remaining sulphur when roasting was done at different

temperatures. The slopes of the graphs can be used to compare the kinetics of the reactions

with respect to temperature and time. The removal of sulphur is also noted to be decreasing

significantly as the temperature increases. More sulphur is removed when changing from 700 °C

to 800°C than when roasting at the same temperature for 2, 3 and 4 hours. At the optimised

roasting conditions, 800°C for 3hrs, there is only 2.92% sulphur remaining, thus a relative good

roasting has been done. It is assumed that the sulphur content at each roasting condition could

have been lower if the furnaces were supplying enough oxygen. In this study, we relied on the

minimum amount of oxygen that could be provided by the furnace.

51

3 4

0 1 1 1 1 0 1 2

Time (Hrs)

Figure 4.3.1.2: Percentage sulphur remaining at varying roasting conditions observing time effect

Figure 4.3.1.2 shows the resulting sulphur remaining after roasting was done for different

temperatures and durations.

4.3.2 X-Ray Diffraction (XRD)

The calcines were also characterised using XRD. The aim of using the XRD was to identify the

phases formed and to determine the conditibns that favoured the formation of zinc oxide over

the undesired zinc ferrite. Furthermore the disappearance of the sulphides at the different

roasting conditions was also of interest. A semi-quantitative analysis was also extrapolated from

the XRD results and the resulting phases of interest were determined to be hematite, zinc

ferrite, zinc oxide and copper oxide.

Figure 4.3.2.1 shows the appearances of new peaks as a result of roasting. Peaks which were

present in the diffractogram of the concentrate are seen to be disappearing after subjecting the

concentrate to the roasting. All the samples shown in the figures were roasted for the same

duration (2hrs). Formation of the zinc oxide and zinc ferrite phases are seen to be dependent on

the roasting temperature. Peaks that are seen at 700 °C are also observed at the higher

temperatures but with different intensities. The zinc ferrite is seen to be dominating in the data

52

27 37 40 57 70

1530

1707

577

Cap entrate

A

703°C

800°C

470

200

0 10

900°C

obtained for roasting at 700 °C than it is at 800°C and thus the competition of these two phases

(zinc ferrite and zinc oxide) is noted to vary with temperature. The dominance of zinc ferrite

seems to diminish at 800 °C (28 = 30°) and the zinc oxide peak (28 = 36°) is larger as the

temperature is increased even further to 900 °C after which the zinc ferrite again starts to

compete with zinc oxide. Another observation is that the phases are seen to be (a bit)

amorphous at 700°C in comparison to that formed at 800 °C and 900°C. (see Figure 4.3.2.1)

Two Theta Degrees

Figure 4.3.2.1: XRD graph showing the effect of roasting temperature on the formation of zinc oxide

and zinc ferrite.

NB: ZO = Zinc Oxide, ZF = Zinc Ferrite, scales for 700°C, 800°C and 900°C are the same, and

all roasted for 2hrs.

Peaks for hematite are seen at 28 = 33.21 °, 35.37°, and 54.12° and for copper ferrite at 28 =

35.37°, 62.42° and 30.03°. The peaks for copper ferrite are very difficult to identify because they

overlap with those of zinc ferrite and some of the zinc oxide peaks, thus they could not be

identified. Hematite is observed to be decreasing as the temperature is increased from 700 °C to

53

a) ZO ZF ZO

300- re 3 ZF ZF Z ZO

200 -

100 -

ZF ZO

/ 700°C, 4hrs

#V1"40/1L. 0 10 20

1600

1200 -

800 -

400 -

0

ZnS

CuFeS2 ZnS

FeS2

- CuFeS2 ZnS nS

FeS2

ZO ZF ZO 700°C,2hrs

700°C,3hrs

30 40 50

Two theta degrees 70

Concentrate

900°C. Pyrite in the spectra from the concentrate is observed though in very small amounts at

28 = 33.08°, 56.27° and 37.10°. It is observed to be disappearing as the concentrate is

subjected to heat. More peaks are observed in the spectra for 700 °C and are attributed to the

sulphides and some sulphates that have formed as a result of the partial roasting, occurring at

that temperature.

Figure 4.3.2.2: Effect of roasting time on the formation of zinc oxide and zinc ferrite at 700°C.

54

30 40

Two theta degrees

10

Figure 4.3.2.3: Effect of roasting temperature on the formation of zinc oxide and zinc ferrite at 800°C.

Figure 4.3.2.2 shows the resulting phases at 700 °C. The most intense peaks observed in the

diffractogram are those for the franklinite. More Zinc ferrite (franklinite) than zinc oxide is

observed to be formed at this temperature. The diffractogram in Figure 4.3.2.3 shows the

roasting products at 800 °C. The two competing phases are seen to be in almost equal amounts.

More significant is that at 800 °C for 2 hours, the zinc oxide dominates the zinc ferrite in

abundance.

55

Two theta degrees

Figure 4.3.2.4:

Effect of roasting temperature on the formation of zinc oxide and zinc ferrite

at 900°C.

As the temperature was increased even further, the zinc ferrite was again noted to be more

significant than the zinc oxide. In terms of the percentages of the phases, the duration of

roasting time was observed to have a minimum effect on the formation of phases, if the

temperature was fixed.

56

Table 4.3.2.1: XRD spectral Intensities of phases at different roasting conditions.

Intensity

Phase 700°C 800°C 900°C

2hrs 3hrs 4hrs 2hrs 3hrs 4hrs 2hrs 3hrs 4hrs

ZnFe204 300 . 265 254 390 385 405 500 555 550

ZnO 127 155 148 395 396 410 493 525 520

Zn2SiO4 75 105 87 250 235 250 250 256 275

Si02 424 210 200 200 675 433 255 142 210

From table 4.3.2.1 it can be seen that at low temperatures, the ratio of zinc ferrite (ZnFe2O4) to

zinc oxide (ZnO) is seen to be more (300:127) at 700 °C for 2 hours, i.e., the formation of zinc

ferrite (franklinite) seems to dominate that of zinc oxide which is the most important phase. The

production of the two competing phases seems to be affected by temperature. The ratio of zinc

ferrite to zinc oxide produced decreases as the temperature is increased to 800 °C. At about

900°C zinc ferrite is again dominating over the zinc oxide. This means that roasting at low

temperatures might be costly because very little zinc will be leached and thus the residue will

have to be recycled (taken back to the furnace for roasting). On the other hand very high

temperatures would result in more zinc ferrite formed, also detrimental to production rates.

The other phase affected by temperature is willemite (Zn2SiO4). Its formation is noted to

increase with the roasting temperature and the production of willemite does not seem to have

any known effect on the efficiency of zinc extraction since it can easily be dissolved during the

neutral leaching. Its production can however be avoided by careful control of the quartz added

to the concentrate during flotation. The hematite and copper ferrite peaks could not be easily

identified since they overlap with those for zinc ferrite and/or zinc oxide.

57

2.0 2.5 3.0 3.5

4.0

Roasting Time (Hrs)

Figure 4.3.2.5: XRD spectral Intensities of the zinc ferrite and zinc oxide at 700°C.

From Figure 4.3.2.5, the conditions seem to favour the production of zinc ferrite over that of zinc

oxide. The intensity is observed to decrease with an increase of roasting time for the zinc ferrite,

yet the opposite is observed for zinc oxide. More zinc oxide is observed as the roasting time is

increased to 3 hrs after which no further increase in observed but a decline. Roasting at 700 °C

is therefore seen not to be the ideal temperature for roasting this particular sulphide ore, though

it is difficult to predict with confidence from only three data points.

Table 4.3.2.2 overleaf is a summary of the semi-quantitative analysis of the XRD. The figures

from that table were obtained from the intensity peaks of the XRD spectra and were converted

to area by multiplying by the full length at maximum height.

58

► Ie 4.3.2.2: A semi- quantitative XRD analysis

Sample

identific

ation

Pyrit

e

FeS2

Sphalerite

ZnS

Chalcopyri

te CuFeS2

Quartz

Si02

Chlonochlore

(mg,Fe)6

(Si,A1)4010(0F1)13

Zinc oxide

ZnO

Muscovite

KAI2(Si3A1)10

(OH, F)2

Zinc

Ferrite

Zn Fe2O4

Wille

mite

Zn2Si

04

ROM 88

506

696 166 556 187 275

2625 527 3155 1492 619

Concent

rate

104 1004 190 400 116 166

587 4642 1629 466 1636

415

R.C.900°

C 4hrs

132 187 219 84

524 689 1042 2235

R.C.800°

C 4hrs

571 1741 2391 946

R.C.800°

C 3hrs

1016 1933 2480 631

R.C.700°

C 4 hrs

780 1371 2785 865

R.C.900°

C 2hrs

328 428 384 44

R.C.900°

C 3hrs

151 306 585 64

magneti

cs

70 567 146 412 169 132

Non-

magneti

cs

67 506 174 416 174 70

Iron

precipita

te

AMORPHOUS

X — lab = A, X = lab B, X = lab C

R.C. - ROASTED CONCENTRATE.

59

420

400 -

380 -

360 -

Zinc Ferrite

• Zincite

Willemite A

280 -

260 -

240 -

340 - P'

C c 320 -a)

- 300 -

• I I I 1

2.0 2.5 3.0 3.5

4.0

Roasting Time (Hrs)

Figure 4.3.2.6: Graph showing intensities of Zinc ferrite and zinc oxide (zincite) at 800°C.

At 800°C (Figure 4.3.2.6) the two phases of interest are produced in almost equal quantities,

though the zinc ferrite is still seen to be the dominating phase. The increase of the roasting time

at 800°C does not seem to have much effect on the production of both zinc oxide and zinc

ferrite. The willemite is still produced in small quantities and the increase in roasting time does

not seem to have any effect on its production.

4.3.3 Mossbauer Results

4.3.3.1 Introduction

Mossbauer spectroscopy only detects the Fe-bearing phases present in a sample. Each

spectrum may be comprised of a superposition of sub-spectra (sextets, doublets, singlets or

even a distribution of sextets) which represent different Fe-bearing phases in the sample. The

integrated area under each sub-spectrum yields relative percentages of iron in that particular

phase.

Table A in the appendices has the M6ssbauer parameters from the literature of most candidate

Fe-bearing phases that might be present in the samples investigated, as suggested by XRD

results. MOssbauer spectroscopy alone cannot easily distinguish between these ferrites, i.e.,

60

0 U)

— E (/) C cu

(T) OC

0.95 -

•

Zn),Fe3_,(04, Cu.Fe3..04 and/or Cui..Zn.Fe204, may have overlapping Mossbauer parameters for

a range of compositions (x values).

-10 -5 0 5

10 Velocity (rrirn)

Figure 4.3.3.1: Mossbauer spectrum of the concentrate from the zinc ore sample.

The overall fit to the spectrum is the solid line through the data points represented as open

circles. The spectrum is fitted with a sextet (six-line pattern) and three doublets. The sextet

represents chalcopyrite, the intense central doublet pyrite and the other two doublets

chlinochlore.

Table 4.3.3.2: Hyperfine interaction parameters of the components in the concentrate sample.

SAMPLE Spectral

Component

r, Linewidth

(mm/s)

IS/Fe

(mm/s)

QS

(mm/s)

Bhf

(T)

*Abundance

%

Concentrate Chalcopyrite 0.32 0.25 -0.01 34.9 21(2)

61

Pyrite 0.32 0.30 0.62 - 60(2)

Chlinochlore

Fe2+

Fe3+

0.38 0.99 2.64 - 6(1)

0.40# 0.40# 0.61 4 - 13(2)

From the integrated area under the sub-spectrum. Statistical error indicated in parenthesis.

# These parameters were fixed to allow for a sensible fit consistent with reports on clinochlore

The pyrite with 60 % abundance with respect to the other iron compounds in the concentrate

was also identified by the XRD (Table 4.3.3.2). All the other phases with lower abundances

identified here could also be discerned by the XRD. The chlinochlore, though could be identify

by XRD, did not show any different oxidation states (ferrous and ferric) as given by the

Mossbauer.

62

Table 4.3.3.3: Hyperfine interaction parameters components in calcines roasted for 2 hrs at

different temperatures.

SAMPLE

Spectral

Component

F, Linewidth

(mm/s)

IS/ Fe

(mm/s)

QS

(m m/s)

Bhf

(T)

*Abu ndance

%

700° C

Hematite 0.34 0.36 -0.2 51.2(2) 42(2)

Franklinite,

Zn-rich ferrites

0.40 0.33 0.45 36(2)

Ferrites 0.32 0.36 0.14 10-50 22 (2)

800° C

Hematite 0.32 0.37 -0.2 51.3(2) 23(2)

Franklinite,

Zn-rich ferrites

0.42 0.33 0.45 46(2)

Ferrites 0.32 0.26 -0.15 10-50 31(3)

900° C

Hematite 0.30 0.39 -0.2 51.3(2) 6(2)

Franklinite,

Zn-rich ferrites

0.45 0.33 0.48 59(2)

Ferrites 0.30 0.45 0.19 10-50 35(3)

The percentage abundances of the hematite, Franklinite and ferrites in the above Table 4.3.3.3,

are summarised in Table 4.3.3.4. This was basically to show the relationship between these 3

phases as the temperature is increased and hematite decreasing to the favour of ferrites and

franklinite.

63

. Tra

nsm

iss i

on

•

900 °C

0 11 0

'3111.0 e t0.

0 II

700 °C lb 10

14 0

1•

-8 -6 -4 -2 0 2

4 6 8

Velocity (mm/s)

Figure 4.3.3.2: Mossbauer spectra of calcines roasted for 2hrs at different temperatures (700 °C,

800°C, and 900°C).

Figure 4.3.3.2 shows the Mossbauer spectra of calcines roasted for 2 hrs at 700 °C, 800°C and

900°C.The 700°C spectrum was fitted with two crystalline components. The hematite being the

sextet and the central doublet representing franklinite. The hematite, though showing good

crystillanity (a comparatively narrow Iinewidth F-0.34 mm/s) its peaks could hardly be identified

by the XRD because they overlaped with those for zinc oxide and zinc ferrite. Its presence

resulted in the formation of large amounts of franklinite. The XRD however could not give the

franklinites in their different forms as did the Mossbauer. In Figure 4.3.3.2, the distribution of

sextets was assigned to the other Zn- or Cu-depleted ferrites. These copper ferrites could also

not be identified in the XRD pattern due to overlap with the zinc oxide and zinc ferrite peaks.

These ferrites are said to be magnetic, depending on the amount of zinc in the spinel. The

magnetism collapses as the amount of zinc increases in the spine!. Similar to the assignments

in the sample roasted at 700 °C, the 800°C calcine sample had the sextet assigned to hematite

64

and the central doublet to franklinite. There was also a distribution of sextets which were

assigned to other Zn- or Cu-depleted ferrites. The sample for 900°C also shows the same

features. The only difference lies in the percentage abundances of the components as shown in

Table 4.3.3.4.

Table 4.3.3-4: Comparison of component abundances (2hrs)

Spectral

component

Abundance

(%)

Abundance

(%)

Abundance

(%)

700°C 800°C 900°C

Hematite 42(2) 23(2) 6(2)

Franklinite (zn-

rich Ferrites)

36(2) 46(2) 59(2)

Ferrites 22(2) 31(3) 35(3)

The hematite decreases as the temperature was increased from 700 °C to 900°C. The peaks

were seen to be becoming less sharp and disappearing as the temperature increased. The

decrease in the hematite was due to the reaction of the zinc oxide/ copper oxide and the

hematite, to produce ferrites according to the following reactions:

ZnO + Fe 2O 3 ZnFe2O 4 (4.3.3.1)

CuO + Fe2O3 —> CuFe2O4 (4.3.3.2)

Table 4.3.3.4 shows the percentage abundances of the iron bearing components. Hematite is

seen to be decreasing (42% at 700 °C to 6% at 900°C), while the ferrites are increasing as the

temperature is increased. The ferrites are magnetic and the magnetism in the calcine seems to

be decreasing as the temperature is increased. Figure 4.3.3.3 shows the graph of the hematite

decreasing and the ferrites forming.

65

I I I 1 I

700 750 800

850

900

Temperature ( °C)

% A

bun

dan

ce

60

55

50

45

40

35

30

25

20

15

10

5

Figure 4.3.3.3: Graph showing the effect of temperature on the amount of phases formed during

roasting for 2 hours.

The hematite is seen to decrease as more ferrites are being formed with an increase in

temperature. More hematite is produced at lower temperatures and then gets converted to

ferrites as it reacts with the zinc oxide. The formation of Franklinite is also shown to be favoured

by the increase in temperature. This is so as most of the iron from the decreasing hematite is

converted to the franklinite.

66

Table 4.3.3.5: Hyperfine interaction parameters of the components in the calcines, roasted for

4 hrs at different temperatures

SAMPLE

Spectral

Component

F, Linewidth

(mm/s)

IS/Fe

(mm/s)

OS

(mm/s)

Bhf

(T)

Abundance

%

700° C

Hematite 0.30 0.37 -0.2 51.4(2) 49(2)

Franklinite,

Zn-rich ferrites

0.42 0.34 0.44 - 34(2)

Other/mixed .

Ferrites 0.30 0.32 -0.07 10-50 17 (2)

800° C

Hematite 0.41 0.37 -0.25 47(1) 12(2)

Franklinite,

Zn-rich ferrites

0.48 0.32 0.42 - 48(2)

Other/mixed

Ferrites 0.32 0.33 0.09 10-50 40(2)

900° C

Hematite 0.34 0.36 -0.17 51.0(2) 9(2)

Franklinite,

Zn-rich ferrites

0.48 0.33 0.51 - 52(2)

Other/mixed

Ferrites

0.30 0.35 -0.02 10-50 39(3)

The summary of the above Table 4.3.3.5 is in Table 4.3.3.6. This shows .the percentage

abundances of the hematite and ferrites as the temperature are increased. Hematite is noted to

be decreasing as the temperature is increased from 700oc to 900oC and the zinc ferrites are

seen to be increasing. The iron from the hematite is reacting with the zinc as the temperature is

increased to form the zinc ferrites. The same is observed with the other ferrites, including

copper ferrite.

67

Tra

nsm

issi

on

-8 -6 -4 -2 0 2 4 6 8

t ;;;;;; tttttttttttt

Velocity (mm/s)

Figure 4.3.3.4: M6ssbauer spectra of the calcine samples, roasted for 4hrs at different temperatures

Figure 4.3.3.4 shows the disappearance of hematite as the temperature is increased from

700°C to 900°C. The effect of temperature is seen at the 4 hours roasting period. The sextet for

hematite is seen to be disappearing as the temperature increased. The sextet for the ferrites

shows more clearly at the high temperature (900 °C). The amount of ferrites and franklinite

formed at 900°C for both roasting periods is almost equal as can be seen in Table 4.3.3.6. The

change in the amount of hematite formed at 800 °C, compared to that formed at 900 °C is noted

to be less, as compared to that when the roasting was done for 2 hours. As it was observed in

the sulphur removal, the change in the amount of sulphur from 800 °C to 900°C (4 hour roasting)

was insignificant.

68

Table 4.3.3.6: Comparison of component abundances (4 hrs)

Spectral

component .

Abundance

(%)

Abundance

(%)

Abundance

(%)

700°C 800°C 900°C

Hematite 49(2) 12(2) 9(2)

Franklinite (zn-

rich Ferrites)

34(2) 48(2) 52(2)

Other/mixed 17(2) 40(2) 39(3)

Ferrites

69

Table 4.3.3.7:Hyperfine interaction parameters of the spectral components in the spectrum of

calcines roasted at 800 °C for different durations.

SAMPLE

Spectral

Component

F,

Linewidth

(mm/s)

IS/Fe

(mm/s)

OS

(mm/s)

Bhf

(T)

Abundance

%

2hrs

Hematite 0.32 0.37 -0.2 51.3(2) 23(2)

Franklinite,

Zn-rich

ferrites

0.42 0.33 0.45 - 46(2)

Other/mixed

Ferrites

0.32 0.26 -0.15 10-50 31(3)

3hrs

Hematite 0.33 0.36 -0.2 50.8(2) 28(2)

Franklinite,

Zn-rich

ferrites

0.41 0.32 0.46 - 40(2)

Other/mixed

Ferrites*

0.32* 0.37 0.08 10-50 * 32(3)

Hematite 0.41 0.37 -0.25 47(1) 12(2)

Franklinite,

Zn-rich

ferrites

0.48 0.32 0.42 - 48(2)

4hrs

Other/mixed

Ferrites

0.32 0.33 0.09 10-50 40(2)

70

-8 -6 -4 -2 0 2 4 6 8

4 h rs •._ ca

O

U)

• a)

CC

Velocity (mm/s)

Figure 4.3.3.5: Mossbauer spectrum of calcine samples roasted at 800 °C for different durations.

From Figure 4.3.3.5, the roasting of the concentrate at 800 °C was found to be interesting from

the XRD results. Although a possible phase transformation was observed in the small decrease

of hematite and ferrite phases from 2 hours to 3 hours, no clear trend in the evolution of phases

at 800°C for 2, 3 and 4 hours was observed. The two phases of interest (zinc oxide and zinc

ferrites) detected by the XRD were found to be competing at this temperature. In the other

temperature regions, it was evident that more franklinite was formed than zinc oxide.

71

Table 4.3.3.8: Comparison of component abundances (800°C)

Spectral

component

Abundance

(%)

2hrs

Abundance

(%)

3hrs

Abundance (%)

4hrs

Hematite 23(2) 28(2) 12(2)

Franklinite (zn-

rich Ferrites)

46(2) 40(2) 48(2)

Other/mixed

Ferrites

31(3) 32(3) 40(2)

As it can be seen from Table 4.3.3.8, the hematite content increased from 2 hours to 3 hours

and decreased again at 4 hours. The reverse was observed for the ferrites. This means that the

conditions for minimum ferrites can be established at roasting for 800 °C, if only the correct time

can be found. The magnetic properties of the calcine followed the same pattern. It increased

from 2 hours to 3 hours roasting and then decreased' at 4 hours roasting.

In all the roasting processes carried out, it is evident that it was a dead roasting because of the

absence of the parameters attributed to sulphide from the M6ssbauer data when compared to

those that were observed by Bandyopadhyay et al, (2000). These parameters are tabulated in

Table A in the Appendices.

The formation of hematite as a sulphide concentrate is subjected to heat due to the reaction of

ZnS, CuFeS2 and FeS2 at lower temperatures can be explained by the following proposed

reaction scheme:

CuFeS2 Fe2 (SO4 )3 Fe203 (4.3.3-3)

72

The FeS2 produced during partial oxidation will also be further oxidised and can result in ferrite

(copper ferrite) formation when reacted with the Cu 20. In the case of zinc, zinc sulphide will

react with pyrite in the following reaction:

ZnS 2FeS2 + 702 —ZnSO4 + Fe203 + S2

(4.3.3-4)

This is again a result of partial oxidation. Further roasting will result in the production of ZnO or

CuO and that will react with hematite to produce zinc and copper ferrites respectively.

ZnSO4 + 0 2

ZnO + Fe 203

heat ZnO + Sgas (4.3.3 -5)

(4.3.3-6)

(4.3.3-7)

(4.3.3 -8)

heat

heat

ZnO Zn Fe 203 2

CuO S gas CuSO4 + 02

CuO + Fe203 heat CuO Fe20 3 •

73

4.4 Leaching Results

The calcine to be leached was then roasted at the optimised roasting conditions (800 °C for 3

hours). It was leached with three different acids (Hcl,H2SO4 and HNO3) to compare their

efficiencies and rates. Analysing the pregnant solution hourly, using atomic absorption

optimised the leaching conditions. Neutral leaching was first done and the residue analysed

after which a hot acid leaching was then done.

4.4.1 Neutral leaching

The leaching was done in the three different acids (Hcl,H2SO4 and HNO3)and the results are

shown in Table .4.4.1.1 below.

Table 4.4.1.1: Table showing neutral leaching results.

Time

(Hrs)

HCI H2SO4 HNO3

% Cu % Zn % Fe % Cu %Zn % Fe % Cu %Zn % Fe

1 69.3 63.2 67.5 63.4 57.8 61.2 65.7 60.5 62.9

2 76.2 66.7 71.3 72.3 61.3 63.7 74.2 63.7 66.7

3 81.3 70.1 74.4 77.9 66.9 66.1 78.6 69.4 69.5

4 86.1 76.5 77.6 81.2 72.6 73.5 84.9 72.9 76.2

5 90.6 82.3 80.1 87.7 81.3 77.6 88.3 80.6 79.3

6 98.4 88.1 81.1 90.5 86.7 79.9 92.4 87.9 80.6

7 96.6 86.4 80.2 95.7 85.4 81.3 95.8 87.8 82.2

8 97.7 87.6 79.6 98.3 86.0 78.8 98.5 89.2 81.8

74

4.4.2 HCI Neutral leach

100-

95-

90-

85-

80-

75-

70-

65-

60

% E

x tra

ctio

n

0

2 3 4 5 6

7

8

9

Time (Hrs)

Figure 4.4.2.1: Dissolution of the metals in HCI showing different extraction rates.

From figure 4.4.2.1, the leaching of copper seems to be the fastest and best among the three

metals. About 70% of copper is leached within the first hour. The optimum leaching time is

observed to be six hours after which there is no significant increase in the amount of metal

leached. About 98% of copper is leached in the final stage. Zinc and iron remained in the

residue after the final stage (after 6 hrs) of the neutral leach and only about 80% and 78% had

been leached respectively.

75

6 I i i . I . 1 2 3 4 5

Time (Hrs)

7 9 8

cY0 E

x tra

ctio

n

100 -

95 -

90 -

85 -

80 -

75 -

70 -

65 -

60 -

55 , I 0 1

I I

4.4.3 H2SO4 Neutral leach

Figure 4.4.3.1: Dissolution of the metals in H 2SO4 showing different extraction rates.

From Figure 4.4.3.1, the sulphuric acid also shows a similar pattern to that observed in the Hcl

except in the case of copper.. A large amount (about 20%) of the zinc and iron remained in the

residue after 8 hrs while a lot of copper was leached up to almost 100 % at 8 hrs. The

dissolution of zinc (85%) is however not as high as in the case of the hydrochloric acid (89%)

leaching. Iron is leached less (80%) compared to the Hcl (85%) after 6 hrs.

76

6 7 8 9

% E

x tra

ctio

n

100 -

95 -

90 -

85 -

80 -

75 -

70 -

65 -

60 -

I i I , I 2 3 4 5

Time (Hrs)

1 i

4.4.4 HNO3 Neutral leach

Figure 4.4.4.1: Dissolution of the metals in HNO3 showing different extraction rates.

From figure 4.4.4.1, dissolution of the metals in nitric acid is seen to be lower than that in

hydrochloric acid yet higher that that in the sulphuric acid. The effect of the ferrite is also

assumed to be locking some part of the zinc and copper.

77

1 2 0 3 8 9 , I g i • 1

4 5 6

Time (Hrs)

Figure 4.4.4.2: Dissolution of copper in the three leaching acids.

The extraction of copper in hydrochloric acid was noted to increase from about 70% in the first

hour to just under 100% in the eighth hour. The leaching was almost complete at the sixth hour

and no increase in the extraction was visible. The rate of the leaching was found to be 4.25% /h.

The extraction of copper in sulphuric acid was a bit lower in the first hour (about 60%) as

compared to that in hydrochloric acid. It still took place even after the sixth hour. Least copper

has been leached after six hours in the sulphuric acid (90%). The final percentage extraction

after eight hours was almost 100%. The rate of dissolution is seen to be about the same as that

obtained with HCI.

About two thirds of copper was extracted in the first hour, and after eight hours almost all the

copper was extracted. The rate of copper extraction is again about the same as for HCI and

H2SO4.

78

/*

• A ♦

90-

85- t___---:---____.

80-

75- c N _

° , . - 70-

65-

60-

/

♦72

HCI • A HNO3A

H2S0

I I i , I . I . 1 2 3 4 5 6

Time (Hrs)

55

/

0 I I 7

8 9

Figure 4.4.4.3: Dissolution of Zinc in the three leaching acids showing the effect of each acid.

The dissolution of zinc in the HNO 3 increased from about 63% in the first hour to about 88% in

the sixth hour. This is seen to be the optimum time for the leaching since there is no more

increase in the % extraction afterwards. The rate of leaching is found to be 3.92 and is lower

than that of copper in the same acid (4.25). This could be due to the fact that some of the zinc

cannot dissolve in the neutral leaching due to it being locked in the ferrite. This also results in

the final percentage extraction being around 88.1%.

Less zinc (60%) has dissolved in the H2SO4 in the first hour when compared to that in HCI (70).

the optimum leaching time is again observed to be 6 hours (86.7%), after which there is no

more increase in the % extraction. The rate of dissolution is however seen to be higher (4.59)

than that of HCI (3.92). The final % extraction is around 86% and can also be due to the locking

in of the zinc in the ferrites that do not dissolve in neutral leach.

In the first hour in the HNO 3 the dissolution of zinc is more (60%) than that in H2SO4 (58%) but

less than in HCI (70%). The final rate of dissolution is however faster (4.58) than the final rate of

79

HCI (3.92). Even here some zinc is still locked in the ferrite and cannot dissolve in the neutral

leaching. The final % extraction is a little bit more (around 89%) than in the other two acids.

84

82

80

78 -

76 -

74 -

N 72 -

70 -

68 -

66 -

64

62

60 - I • i •

0

2 3 4 5 6

7

8

9

Time (Hrs)

Figure 4.4.4.4: Dissolution of iron in the acids showing the effect of each acid.

The leaching of iron is found to be optimum at six hours. About 80% of the total iron could be

extracted with HCI after six hours of leaching. This is the minimum extraction when compared to

that of copper and zinc. This could be due to the conditions of the solution that could force some

of the iron to be precipitated and thus settle down with the neutral leach residue. However, most

of it is attributed to the ferrites that have not yet been dissolved. The rate of iron dissolution is

found to be 1.8. This is also the slowest leaching rate of iron among the three acids.

The rate of dissolution of iron (about 3.) is also noted to be slowest among the other metals

leached in the same acid. Iron could also be precipitated here, as the final percentage extraction

was only 78%, most was believed to be locked in the ferrite. The dissolution of iron in sulphuric

acid was the fastest among the three acids that have been compared.

80

4.4.4.1 Neutral leach digestion

About 0.5g of sample from the neutral leach was dried and digested using aqua regia. The

solution was diluted to 250m1 in a volumetric flask and was analyzed using the AAS. The results

obtained are as follows.

Table 4.4 4.1: Percentages of elements remaining after neutral leaching.

Leaching acid %Cu %Zn %Fe

H2SO4 1.24 4.01 8.23

HCI 1.73 3.63 9.01

HNO3 1.51 3.87 7.76

4.4.4.2 Hot Acid Leaching

The residue from the neutral leach was taken for hot acid leaching. The procedure for leaching

was the same as that for the neutral leach, except that the conditions were rather intense with

temperatures of 95 °C and the acid had a concentration of 50g/L. The leaching was done for six

hours. The samples were also analysed at one-hour intervals using the atomic absorption

spectrophotometer. The results are shown below.

Table 4.4.4.2: Dissolution of elements during hot acid leaching.

Time

(hrs) HCI H2SO4 HNO3

Cu(ppm) Zn(ppm) Fe(ppm) Cu(ppm) Zn(ppm) Fe(ppm) Cu(ppm) Zn(ppm Fe(ppm)

1 565.2 1831.7 1759.4 569.2 1833.4 1782.3 567.6 1830.1 1778.2

2 565.7 1832.8 1761.1 571.2 1839.4 1783.5 569.4 1837.6 1780.1

3 566.0 1833.1 1762.9 574.5 1840.7 1783.9 569.1 1838.3 1781.6

4 567.9 1833.9 1764.3 574.4 1844.3 1785.6 571.9 1841.7 1783.6

5 568.1 1834.2 1766.9 576.3 1846.2 1786.8 573.2 1843.4 1785.2

6 568.3 1834.9 1767.6 577.3 1846.9 1787.1 573.9 1843.9 1785.7

81

4.4.4.3 Hot Acid Leaching digestion

The final residue was again digested using aqua regia. The atomic absorption results are shown

below.

Table 4.4.4.3: Table showing percentages of elements remaining after hot acid leaching.

Leaching acid %Cu %Zn %Fe

H2SO4 0.21 0.376 0.48

HCI 0.27 0.145 0.36

HNO3 0.24 0.194 0.45

4.4.5 CONCLUSION

From the results, it can be seen that the HCI leaching results in a faster leaching rate than in the

case of HNO3 and H2SO4. At the final stage there were however some of the desired metals

remaining in the residue. • The rate of leaching at the hot acid leach was rather slow and

eventually negligible at around the fifth and sixth hour of leaching. The remaining metal in the

digested residue from the neutral leach could probably be a result of zinc ferrite locking and

precipitation of iron. The residue was discarded with some part of the pregnant solution

containing the already dissolved metals. The recovery with HCI was however much better than

HNO3 and H2SO4.

82

4.5 Selective leaching

Much of the work done on selective leaching was focused on the use of HCI. Ferric

sulphate/chloride have dominated the possible leaching agents in the presence of Mn02. In this

work the use of HCI was avoided because of the following considerations.

It requires corrosion resistant materials

It is more expensive than H2SO4

It demands special care for environmental pollution.

Chloride systems are usually non-selective and all the sulphides would be attacked and

thus introducing the iron problem again.

The use of ferric chloride/sulphate was also avoided because of the introduction of ferrous iron,

which causes difficulties with downstream processing due to iron impurities. For recycling, the

ferrous iron has to be re-oxidised to ferric iron and these expenses would then discredit the

selective leaching process if it were to be considered in terms of costs.

The following results show the leaching of zinc and copper from the sulphide concentrate using

H2SO4 and two oxidising agents (Mn02 and Na2S2O8). The choice of the two oxidising agents in

this project was mainly due to limited funds in the project. Other oxidising agents could have

been used if costs were not an issue.

Table 4.5.1: Percentage Extraction in 5M H2SO4 and 10% (w/v) Mn02. Time (Hrs) % Extraction

Zn Cu Fe

0.5 21 30 12

2 34 40 16

4 40 43 18

6 44 47 22

83

% E

xtr

act i

on

50

45 -

40

35 -

30 -

25 -

20 -

15-

10

0 1 2

3 4

5

6

Time (Hrs)

Figure 4.5.1 Percentage Extraction in 5M H 2SO4 and 10% (w/v) Mn02.

More copper than zinc seems to be extracted when the Mn02 is used first as an oxidising agent.

The dissolution of iron was also observed but not proven if it was extracted from the

chalcopyrite or from the pyrite. It seemed though that the iron into solution is that from the

chalcopyrite since higher percentages could be observed if it was combined with that from the

pyrite. The extraction however was slowest and after 6 hours, only 47 % Cu and 44 %Zn could

be extracted. Starting with the Mn02 also seemed to attack both the chalcopyrite and the

sphalerite (though to a lesser extent) at the same time. The concentration for the leaching acid

was increased to 7M while the concentration for the oxidising agent was kept constant (10`Yow/v)

and the results are shown in Table 4.5.2

84

Table 4.5.2: Percentage Extraction in 7M H2SO4and 10% (w/v) Mn02. Time (Hrs) % Extraction

Zn Cu Fe

0.5 26 34 13

2 41 45 18

4 45 48 21

6 50 53 23

The oxidising agent (Mn02) concentration was then increased to 20% w/v while keeping the

acid concentration constant (5M). The extraction of the metals was seen to increase with an

increase in oxidising agent present than it did with an increase in the acid concentration. The

results in Table 4.5.2 are compared with those in Table 4.5.3.

Table 4.5.3: Percentage Extraction in 5M H2SO4 and 20% (w/v) Mn02. Time (Hrs) % Extraction

Zn Cu Fe

0.5 29 36 15

2 44 47 19

4 47 50 21

6 52 55 24

Both the acid concentration and the oxidising agent concentration were increased

simultaneously. It was noted that this simultaneous increase of the reagents promoted the

percentage extraction even more than when each reagent was increased separately. The

results are shown in Table 4.5.4.

Table 4.5.4: Percentage Extraction in 7M H2SO4 and 20% (w/v) Mn02. Time (Hrs) % Extraction

Zn Cu Fe

0.5 31 39 22

2 46 49 28

4 51 56 33

6 55 59 35

85

% E

x tra

ctio

n

1 2 3 4

5

6

Time (Hrs)

Figure 4.5.2 Percentage Extraction in 7M H 2SO4 and 20% (w/v) Mn02.

Though the extraction rate is noted to increase, the zinc and copper were dissolved at the same

time. The total extraction was however higher than when both the acid and oxidising agents

were added, increased simultaneously.

The concentrate was also leached using Na2S2O8 as a starting oxidising agent to determine

which mineral would be attacked first. Table 4.5.5 and Figure 4.5.3 show the results obtained.

Table 4.5.5: Percentage Extraction in 5M H2SO4 and 10% (w/v) Na2S2O8.

Time (Hrs)

% Extraction

Zn Cu Fe

0.5 29 5 3

2 36 8 6

4 43 11 8

6 48 12 8

86

% E

x tra

ctio

n

1 2 3 4 I 1 1 i

Time (Hrs)

Figure 4.5.3 Percentage Extraction in 5M H 2SO4 and 10% (w/v) Na2S208.

Though it could be seen that the 5M of concentration leaching acid is not the optimum for

extraction of zinc preferable over copper, the use of Na2S2O8 attacks the sphalerite to a higher

degree than chalcopyrite (see Figure 4.5.3). Iron is also noted to be extracted but

concentrations were too low to be extracted from both the chalcopyrite and the pyrite.

The acid concentration was increased to 7M while keeping the oxidising agent constant at

10%w/v. The percentage extraction was noted to increase with an increase in acid

concentration, as would be expected. The results are shown in Table 4.5.6.

Table 4.5.6: Percentage Extraction in 7M H2SO4 and 10% (w/v) Na2S2O8. Time (Hrs) % Extraction

Zn Cu Fe

0.5 31 6 5

2 43 9 8

4 49 13 11

6 54 15 12

0 5 6

87

The oxidising agent concentration was then increased to 20%(w/v) while keeping the acid

concentration at 5M. The effect of increasing the oxidising agent was more significant in the

percentage extraction than when increasing only the acid concentration if the results in Table

4.5.6 are compared with the results in Table 4.5.7

Table 4.5.7: Percentage Extraction in 5M H2SO4 and 20% (w/v) Na2S 2O8. Time (Hrs) % Extraction

Zn Cu Fe

0.5 39 9 6

2 46 11 10

4 51 14 11

6 57 16 13

Both concentrations for the acid and oxidising agent were increased subsequently. This

improved the percentage extraction to 61`)/0 for zinc whilst chalcopyrite was still not attacked

much. (See Table 4.5.8.)

Table 4.5.8: Percentage Extraction in 7M H2Satand 20% (w/v) Na2S 2O8. Time (Hrs) % Extraction

Zn Cu Fe

0.5 43 13 7

2 51 14 11

4 57 17 13

6 61 19 16

88

% E

xtr

acti

on

i

2 3 4

5

6

Time (Hrs)

Figure 4.5.4 Percentage Extraction in 7M H 2SO4 and 20% (w/v) Na2S2O8 .

The increase in the concentrations also gave the same trend in the extraction except that the

percentages are noted to increase. The 7M H2SO4and 20 % (w/v) Na2S2O8 (Figure 4.5.4.) were

again not the optimum conditions for the leaching since not even 70% of the targeted zinc could

be leached. It is however seen that starting with the Na2S2O8 provided a possibility to leach zinc

and leaving most of the chalcopyrite and pyrite not attacked.

The residue from the 7M H2SO4 and 20% (w/v) Na2S2O8 was, after being digested and

analysed, leached with the same concentration of acid but using Mn02 as an oxidising agent.

The results are shown in Table 4.5.9 and Figure 4.5.5.

Table 4.5.9: Percentage Extraction of residue in 7M H2SO4 and 20% (w/v) Mn02. Time (Hrs) % Extraction

Zn Cu Fe

0.5 4 19 11

2 11 28 15

4 12 37 17

6 15 41 17

89

c/0 E

x tra

ctio

n

0 1 2 3 4

5

6

Time (Hrs)

Figure 4.5.5 % Percentage Extraction in 7M H2SO4 and 20% (w/v) Mn02.

Mn02 seemed to favour the dissolution of chalcopyrite. The percentages of zinc after the

residue digestion were still higher than the percentage of copper since the concentrate was very

rich in zinc and most of the zinc was still remaining in the residue. The difference in the

preference of the leaching is noted to be higher after some zinc had been leached as compared

to when the initial concentrate was started with Mn02 as an oxidising agent.

4.5.1 Conclusions

The selective leaching of this sulphide ore was possible in the presence of sulphuric acid. The

choice of the oxidising agent played a major role in determining which minerals dissolved first

into the solution. Though the leaching of chalcopyrite is commonly done in HCI, it was observed

here that chalcopyrite can also dissolve in H2SO4 mostly in the presence of Mn02. Chalcopyrite

was also noted to be dissolved well after the sphalerite had been removed from the concentrate.

In this concentrate with a higher percentage of sphalerite, the selective leaching was more

feasible if the mineral with higher percentage (sphalerite) was leached first. To confirm the

extraction percentages and the source of the iron dissolved, it is recommended that other

characterisation techniques be used for the residue. This will allow the determination of how

much pyrite has been attacked in any of the oxidising agents used.

90

Though the optimum conditions for the leaching acids and oxidising agents were not attained, it

was observed that as the concentration of oxidising agent was increased, the percentage

extraction also increased and the same was observed with an increase in the acid

concentration. The choice of oxidising agent had also been found to play a role in determining

which metal will go into solution first.

91

4.6 Precipitation results

Many of the parameters governing the removal of iron from acid leach liquors by the various

processes have been identified. In the jarosite formation, the effect of the cation added on the

resulting precipitate has been studied. In this study, the effect of the cation and the type of

leaching acid used have been studied on the rate and efficiency of iron removed.

From the analysis of the concentrate the constituents of the ore and their percentage

occurrence could be determined. The iron content, which is of interest, was shown to be about

21percent. A synthetic solution was first made based on the results. This was done in order to

optimise the precipitation conditions of iron before being implemented on the pregnant solution.

4.6.1 Procedure

250 ml of the pregnant solution was introduced in a beaker. The beaker was tight-closed using a

plastic to minimise any loss due to evaporation of solution caused by increase in temperature

and to minimise oxidation. The solution was placed on a hot plate/magnetic stirrer and stirred at

about 360 rpm. The iron was then precipitated at pH ranging from 1.5 to 2.5 using NH4OH and

NaOH. The temperature was varied between 80 °C and 95°C. The experiment was run for about

24 hours with the same conditions maintained.

About 3 ml of the solution was removed and filtered at 30 minutes time intervals for the first hour

and then at 1 hour-intervals for the next 4 hours and then analysed after 24 hours. The filtered

solutions were taken for AAS to determine the content of iron from time zero up to the last hour.

The results are shown in Table 4.6.1.1.

The solution remaining after the 24 hours was taken and filtered; the filtrate was then subjected

under the same conditions and some more iron precipitate starts forming. It was noted that the

more dilute the iron was in the solution, the more difficult its removal was (poor settling

properties).

In all the acids it was noted that not all the iron could be removed from the solution. This could

be due to the period of precipitation, which might need to be lengthened. It also could be that

the solution was saturated and the iron in solution was at equilibrium with the iron hydroxides

already formed. The color of the second stage precipitate changed from dark brown to very light

92

brown. Raising the pH to 3 resulted in some of the iron precipitating but was difficult to filter and

also taking too long to settle. This would entrain most of the zinc and copper and thus result in

lower recoveries.

The pregnant solution from the neutral leach indicated 443 ppm, 429 ppm and 372 ppm of iron

in HCI, HNO3 and H2SO4 respectively as per the atomic absorption analyses. The iron from the

neutral leach could not be removed using the precipitation. Addition of the bases resulted in

change of colour, which indicated some precipitate, but the solution could not be filtered. From

the hot acid leach, the solution had 1767.6 ppm, 1734.9 ppm and 1517.8 ppm of iron with

respect to HCI, HNO3 and H2SO4.

93

4.6.2 Results

Table 4.6.1.1: Table summarising the amount of iron (in %) remaining from precipitation at

80°C and filtration after 24 hours.

Solution Cation

added

pH Time in hours

30

mins

60

mins

2 3 4 5 24

HCI Na+ 1.5 58.8 50.7 42.9 37.6 34.9 28.8 6.89

HCI Na+ 2.0 47.3 44.8 35.3 28.4 22.9 19.4 2.3

HCI Na+ 2.5 46.5 42.9 34.8 26.4 21.6 16.2 2.2

HCI NH4+ 1.5 57.3 48.4 42.4 35.6 31.9 24.9 8.5

HCI NH4+ 2.0 55.6 46.8 40.9 34.8 29.5 22.8 3.8

HCI NH4+ 2.5 56.9 46.9 40.7 35.2 31.8 24.3 3.5

H2SO4 Na+ 1.5 67.6 61.3 54.8 47.7 42.5 34.7 7.3

H2SO4 Na+ 2.0 62.3 56.6 49.8 42.9 31.9 20.2 5.47

H2SO4 Na+ 2.5 64.6 59.5 53.2 46.2 39.2 26.5 3.7

H2SO4 NH4+ 1.5 65.7 59.8 52.5 45.8 30.4 23.5 5.9

H2SO4 NH4+ 2.0 63.9 58.3 49.9 43.7 27.9 22.6 6.0

H2SO4 NH4+ 2.5 66.3 60.8 50.6 46.5 28.3 23.2 4.9

HNO3 Na+ 1.5 71.7 64.2 57.3 51.5 44.8 32.4 14.4

HNO3 Na+ 2.0 65.3 59.8 52.5 45.7 38.5 25.3 16.6

HNO3 Na+ 2.5 68.5 64.3 58.9 49.8 42.7 36.4 14.7

HNO3 NH4+ 1.5 70.6 62.9 55.5 50.3 44.2 27.4 18.2

HNO3 NH4+ 2.0 69.4 60.9 54.3 50.2 43.6 24.8 16.0

HNO3 NH4+ 2.5 69.9 61.5 54.4 51.5 43.8 25.4 17.4

As can be seen from Table 4.6.1.1, the removal of iron was noted to be more affected by the

type of acid used than it was by the type of cation added. Time was also found to have an effect

on the amount of iron removed. This is observed after the first hour, the sixth and the twenty-

fourth hour. The precipitation of iron from the HCI solution seems to be the fastest and even

more efficient than for the H2SO4 and the HNO3 in solution and can be due to the type of iron

compound formed. In all the cases the precipitate was formed but the rate and the type of

precipitate was in each case different. The strength of the solution also played a major role on

94

the type of compound formed. Leaving the precipitation process for a longer time (24hrs)

resulted in more iron precipitating from the solution and can be filtered even better than when

the solution was only precipitated for a shorter time (1hr). In agreement to this, Murphy et al,

(1976) observed that the process of precipitation depends on the solution conditions. In all the

solutions they studied (Ferric nitrate, ferric chloride and ferric perchlorate) they found that some

rods (sticks like) formed.

The precipitation of iron from the nitric acid was the worst of all the three acids. About 11-14% of

the iron remained in the solution. This could be due to the type of the precipitate forming. The

rods that formed in the precipitate were too small to filter and they remained in the solution. The

solutions were analysed using the AAS and thus the total amount of iron was analysed. Murphy

et al, (1976) discovered that in a nitric acid solution, as the precipitation time was increased, the

rods formed did not increase in length or in width but they remained the same. This makes them

float in the solution as they are too small and can thus not be filtered.

95

Table 4.6.1.2: Results summarising the amount of iron (in %) remaining from precipitation at

95°C and filtration after 24 hours.

Solution Cation

added

PH Time in hours 3 6 24

HCI Na+ 1.5 53.6 45.5 38.9 32.4 29.7 23.6 4.89

HCI Na+ 2.0 42.1 39.6 30.1 23.2 17.9 14.2 2.3

HCI Na+ 2.5 41.3 37.9 29.8 21.2 16.4 11.0 2.2

HCI NH4+ 1.5 52.1 43.2 37.2 30.4 26.7 19.9 8.5

HCI NH4+ . 2.0 50.4 41.6 35.7 29.6 24.3 17.8 2.8

HCI NH4+ 2.5 51.7 41.9 35.5 30.0 26.6 19.1 2.5

H2SO4 Na+ 1.5 62.4 56.1 49.8 42.5 37.3 29.5 4.3

H2SO4 Na+ 2.0 57.1 51.4 44.6 37.7 26.9 15.0 2.4

H2SO4 Na+ 2.5 59.4 54.3 48.0 41.0 34.0 21.3 1.7

H2SO4 NH4+ 1.5 60.5 54.8 47.3 40.6 25.2 18.3 2.9

H2SO4 NH4+ 2.0 58.7 53.1 44.9 38.5 22.9 17.4 3.0

H2SO4 NH4+ 2.5 61.1 55.6 45.4 41.3 23.1 18.0 1.9

HNO3 Na+ 1.5 66.5 59.0 52.1 46.3 39.6 27.2 11.4

HNO3 Na+ 2.0 60.1 54.6 47.3 40.5 33.3 20.1 13.6

HNO3 Na+ 2.5 63.2 59.2 53.7 44.6 37.5 31.1 11.7

HNO3 NH4+ 1.5 65.4 57.8 50.2 45.1 39.0 22.2 15.2

HNO3 NH4+ 2.0 64.2 55.7 49.1 45.0 38.4 19.6 13.0

HNO3 NH4+ 2.5 64.9 56.3 49.2 46.3 38.8 20.2 14.4

Table 4.6.1.2 above is a summary of the amount of iron remaining from the precipitation at 95 °C

and filtration after 24 hours. It shows that there is no significant effect of the cations on the rate

of precipitation.

96

)̀/0 F

e R

ema

inin

g

0 10 15

Time (Hrs)

5 20 25

Na÷,pH1.5

Na+,pH2.0

Na+,pH2.5

NH4+ ,pH1.5

N H4+ , pH2.0

+— N H4+ , pH2.5

60

50 -

40 -

30 -

20 -

10 -

0

Figure

4.6.1.1: The comparison of NaOH and NH4OH used in precipitating iron from an HCI

pregnant solution.

The cations from the figure above seem to play no major role on the rate and the efficiency of

precipitation. After the same duration of precipitation, there was an almost equal amount of iron

remaining from both the NaOH and the NH4OH.

97

% F

e R

ema

inin

g

70

60

50

40

30

20

10

0

HCI,pH1.5 —•— HCI,pH2.0

HCI,pH2.5 —•— H2SO4 ,pH1.5

• H2SO4,pH2.0 H 2SO4 ,pH2.5

0

5

10 15

20

25

Time (Hrs)

Figure 4.6.1.2: The comparison of HCI and H 2SO4 pregnant solutions in precipitating iron using

NaOH.

The effect of the 2 acids can be observed in Figure 4.6.1.2. The precipitation has been

compared in the 2 acids. When varying pH, the values for iron in the H2SO4 are higher than in

the HCI meaning that there is more iron remaining in the H2SO4 compared to that in the HCI

pregnant solution when precipitated at the same pH.

98

`)/0 F

e R

ema

inin

g

70

60

50

40

30

20

10

0 I 1 I I 10 15

Time (Hrs)

5 0 20 25

Na+ ,pH1 .5

*— Na+ ,pH2.0

A- Na+ ,pH2.5

—v— NH4+ , pH 1.5

—•— N H 4+ , p H2.0

+— NH 4+ , pH2.5

Figure

4.6.1.3: The comparison of NaOH and NH4OH used in precipitating iron from a H2SO4

pregnant solution.

Figure 4.6.1.3 also shows the insignificant effect on the rate and efficiency of precipitating iron.

99

°/0 F

e R

ema

inin

g

20 25

H2SO4 ,pH1.5 —•— H2SO4 ,pH2.0

A- H2SO4,pH2.5 —v— HNO3,pH1.5

HNO3,pH2.0 +— HNO3,pH2.5

70

60

50

40

30

20

10

0

0 5 10 15

Time (Hrs)

Figure 4.6.1.4: The effect of H 2SO4 and HNO 3 in the amount of iron remaining from precipitating with

NaOH.

It can be seen that different acids do have an effect on the amount of iron precipitated at a given

pH.

The change of temperature from 80 °C to 95°C was found to have little effect on the rate and

efficiency of iron removed.

100

akaganeite akaganeite

60 20 30 40 50

goethite

morhit goethite

H2SO4+NH4+,pH2.5 morhite

HCI+Na+,pH2.5

INP4WAW1/44/$0-64.4"4.40,0ilw,ve.woinf*A-.440.,,,*".. vvrov'"N

1800 -

1600 -

1400 -

1200 -

1000 -

800 -

600 -

400 -

200 -

0 10 70

400

350 -

300 -

250 -

200 -

150-

100-

50

4.6.3 XRD characterisation of precipitates.

Two Theta degrees

Figure 4.6.2.1 XRD spectra for precipitates where iron precipitation was optimum.

Figure 4.6.2.1 shows the spectrum of the different precipitates resulting from the optimised

conditions for precipitation in the different acids. The calcine used for the leaching was from the

experiments where the laboratory roasting conditions for dead roasting were the best optimum.

From the spectra it can be seen that the products are all different as expected and this could be

attributed to the effect of the acid or the cation used in the bases for precipitation. The

precipitates were characterised using the XRD and were found to be too amorphous in the

precipitate from H2SO4 and NH4OH, however Goethite [FeO(OH)] and Mohrite

{(NH4)2[Fe(H20)6](SO4) 2} were identified by the XRD. Jarosite and other iron hydroxides and/or

oxides could not be identified as expected in this study. These products (Jarosite and other iron

hydroxides and/or oxides) are mostly precipitated in industry where the roasting conditions are

different from those used in these experiments. From the sample of the HCI and NaOH,

101

akaganeite (13-Fe0OH) was identified by the XRD; this is an iron oxide chloride Hydroxide. Only

goethite could be identified by XRD from the HNO 3 and NaOH precipitate sample. Further

identification was done with MS.

4.6.4 Mossbauer characterisation of precipitates.

The Mossbauer results gave similar spectra for the characterised precipitates and showed two

doublets with almost similar properties in all the precipitates from the 3 acids.

1.00

0.95-

0.90-

0.85-

0.80 -

0.75-

(.0

a)

1) cc toi 1.00 0.99 0.98 0.97 096 0.95

-4 -2

Velocity (rrm's)

...

2

4

Figure 4.6.3.1: M6ssbauer spectrum of precipitates.

The overall fit to the spectrum is the solid line through the data points represented as open circles. The

spectrum is fitted with a minimum number of doublets (two), to obtain reasonable linewidth values.

102

Table 4.6.3.1: Hyperfine interaction parameters of the spectral components in the spectrum of samples.

SAMPLE

Spectral Component

F, Linewidth

(mm/s)

8/Fe (mm/s)

QS (mm/s)

Bhf (T)

Abundance %

H2so4+NR4ox pH 2.5

Quad-1 0.36 0.37 0.56 - 67(2)

Quad-2 0.39 0.37 0.96 - 33(2) HCI+NaOH

pH2.5 Quad-1 0.36 0.36 0.59 - 75(2) Quad-2 0.33 0.37 1.0 - 25(2)

HNO3+NaOH Quad-1 0.40 0.34 0.62 - 76(2) pH 1.5 Quad-2 0.33 0.35 1.0 - 23(2)

From the integrated area under the sub-spectrum. Statistical error indicated in parenthesis.

From the XRD pattern (Figure 4.6.2.1), goethite, mohrite and akaganeite were identified from

the precipitates even though the H2SO4 and NH4OH precipitate sample represented poorly

crystalline or amorphous phases. It would be expected however, that some jarosite,

schwermannite or ferrihydrite would be identified from these samples. The hyperfine interaction

parameters obtained from MS (See Table 4.6.3.1) differ from those given in literature (see Table

B in the appendices). None of the obtained parameters resemble any form of jarosite or

schwermannite. Ferrihydrite could be identified from all the precipitates. Some properties of

schwermannite are seen but the isomer shift differs from literature values (Table B in

appendices). This could be due to the method of precipitation which is believed to have an

effect on the parameters for precipitate identification, Young Yu et al, (2002). These two phases

are believed to be difficult to differentiate. Bernhard et.al , (2001) tried to discriminate ferrihydrite

and schwermannite and concluded that their Fe/S mol ratios can be used for the discrimination.

They categorised the stability of these phases according to pH zones. The oxidation zone at

(pH1.5-3.5, high sulphate contents), a neutralisation zone (below pH 3.5.5) and an underlying

primary zone with near neutral pH. Schwermannite and jarosite were found to be more stable in

the oxidising zone and ferrihydrite in the next underlying zone. Schwertmann et al 1995

observed that ferrihydrite is stable at less acidic conditions (pH>4) than schwermannite (pH 3-4)

and jarosite (pH<3). The stability and crystallinity of these iron hydroxides (ferrrihydrite and

schwermannite) are dependent on the method of preparation (Young Yu et al, (2002). Their

crystallinity properties differ from that observed in nature. According to Young Yu at al, (2002),

Goethite is believed to be formed from the transformation of ferrihydrite and schwermannite

after 72 hours passed from onset of the synthesis, which is what happened in this case.

103

Chapter 5 Conclusions

The optimum roasting conditions can be obtained by roasting at the right temperature for the

right duration. From the characterisation of the resulting calcine by the sulphur test, XRD and

MES it is possible to identify the right conditions where more zinc oxide is produced than zinc

ferrite. From the sulphur test, the optimum roasting conditions were obtained at 800 °C for three

hours. The furnaces used however had a poor supply of oxygen to blow out the sulphur and

enhance oxidation of the sulphide. The XRD results indicated the dominance of zinc oxide over

zinc ferrite at 800°C for three hours. Almost equal amounts of these two competing phases were

observed at 800°C. The calcine, further studied with MS, shows that more zinc ferrite was

formed as temperatures were increased higher than at 800 °C. MS however did not show any

relation of the zinc ferrite to the zinc oxide.

From the precipitation results, the cations Na+ and NH4 were found to have no effect on the rate

of precipitation than did the acids which the iron was precipitated from. It was observed that

precipitation was slowest from nitric acid and more efficient from sulphuric acid.

The characterisation of the precipitates with XRD showed different spectra. The jarosite that

was expected could not be identified with either XRD or Mossbauer but geothite, mohrite and

akaganeite could with XRD. Ferrihydrite was found by the MS to be present in all the

precipitates. These precipitates are not commonly found with the present industry conditions for

roasting.

There is not yet a proper route of dealing with or avoiding the zinc ferrite if roasting is done on

the concentrate, but can be minimised by optimising the roasting conditions. High temperature

roasting at optimised conditions can provide a solution to the ferrite problem without having to

treat the residue (as in the case of partial roasting). However, selective leaching can be an

alternative to the iron problem by leaving a substantial amount of iron in the residue (in the case

where the ore contains a lot of pyrite). The only iron to be dealt with will be that from the

chalcopyrite in the case of selectively leached copper.

Selective leaching has also been tried on this particular ore and was found to be feasible if

Mn02 and Na2S2O8 were used as oxidising agents in the presence of H2SO4. Though optimum

104

conditions have not yet been established in the use of these two oxidising agents, it was

observed that the higher the concentrations the more the percentage extraction increases. The

choice of the oxidising agent plays a role in selectively leaching the minerals present in the ore.

The use of Na2S2O8 first in the leaching has been found to attack sphalerite better than it does

to chalcopyrite. Other oxidising agents can be used at optimised conditions and if a feasibility

study was to be done, it could be possible to compare the partial roasting, dead roasting and

selective leaching, to find which route is cheaper in dealing with the iron problem.

5.1 Recommendations

Roasting products from industry must be used and compared with laboratory roasting

products and a feasibility study can be conducted.

The selective leaching route has to be studied further in South African ores with the use

of different oxidising agents including oxygen.

Residue from the leaching must be characterised by either XRD and/or M6ssbauer to

complement the AAS results and determine there the iron comes from in the case where

the ore contains pyrite and chalcopyrite.

A feasibility study must be done to prove the cheaper way of dealing with the iron

problem among the partial roasting, dead roasting and selective leaching routes.

The kinetics of the ferrite formation need to be studied intensively at the 800 °C roasting

temperature.

The characterisation of residue after each selective leaching is necessary.

105

References

Atkison, R.J. Posner, A.M. and Quirk, J.P. (1968). Neutralization of iron solutions. J Inorg

Nucl. Chem., 30, 2371.

Bandyopadhyay, D., Singru, R.M and Biswas, A.K. (2000). Technical Note Study of the

Roasting of Chalcopyrite Minerals by 57Fe Mossbauer Spectroscopy. Minerals

Engineering, 13(8-9), 973-978.

Blesa, M.A. and Matijevic, E. (1989). Hydrolysis of Iron. Adv. Colloid and Interface. Sci., 29,

173.

Bryson, A.W., TeRiele, W.A.M. (1994). Factors that affect the kinetics of nucleation and

growth and the purity of geothite precipitates produced from sulphate solution. Iron

Control in Hydrometallurgy. Eds. Dutrizac, J. E., Monhemius, A. J.

Chen, T.T. and Cabri, L.J. (1993). Mineralogy, Geochemistry, Geology Recovery. Canadian

Institute of Mining and Metallurgy. 23.

Cornell, R.M., Giovandi, R. and Schneider, W. (1989). Effect of solution conditions on the

proportion and morphology of goethite formed from ferrihydrite. Clays and Clay minerals.

33, 424.

Davey, P.T., Scott, R.T. (1976). Removal of iron from Leach Liquors by the Goethite

Process. Hydrometallurgy, 2, 25-33.

Devi, N.B., Madhuchhanda, M. Sririvasa, Rao K., Rath, P.C. and Paramguru, R.K. (2000).

Oxidation of chalcopyrite in the presence of manganese dioxide in HCI medium.

Hyrdometallurgy, 57, 57-76.

Dold, B. Fontbote, L. and Wildi, W., (2001), Selective metal retention by ferric

oxyhydroxides and oxyhydroxide sulfates in mine tailings and influence of climate to

element mobility. Earth Science Department, University of Geneva, Switzerland.

106

Doumsa, J. and De Bruyn, P.L. (1979). Hydrolysis-Precipitation studies of iron solutions-

Model for hydrolysis and precipitation from Fe (III) nitrate solutions. Journal of colloid and

Interface Science, 56(3) 526-539.

Dutrizac, J.E. (1981). Dissolution of chalcopyrite in ferric sulphate and ferric chloride

media. Metallurgical transaction B (Process Metallurgy). 12B(2) 371-378.

Dutrizac, J.E. (1987). An overview of iron precipitation in hydrometallurgy, in

crystallisation and precipitation. Proceedings of the International Symposium, Eds.,

Strathdee GL, Klein MO Melis LA, Pergamon, Oxford, 121-126.

Dutrizac, J.E. and Harris, G.B. (Eds). (1996). Solvent Extraction for iron control in

hydrometallurgy. Iron Control and disposal Pro. 2nd Int. Symp. on iron Control in

Hydrometallurgy, Ottawa, Canada' 331-356.

Dutrizac, J.E. and Palencia,( 2002). Effect of iron content of sphalerite on its rate of

dissolution in ferric sulphate and ferric chloride media. Hydrometallurgy, 26(2), 211-232.

Elsherief A.E. (1999). Selective dissolution of zinc from a complex sulphide ore by

controlled-potential HCI leaching. IMM Mineral Processing and extractive Metallurgy,

section C.

Filippou, D. and Choi, Y. (2002). Iron Removal from Chloride leach Solutions. Noranda

Inc-Technology, Pointe-Claire, Quebec, Canada, 3-8.

Genin, J.M.R. and Olowe, A.A. (1991). The mechanism of oxidation of ferrous hydroxide

in sulphated aqueous media: Importance of the initial ratio of reactants. Corrosion

Science, 32(9), 965-984.

Godocika, E., Balaz, P. and Boldizarova, E., (2002). Structural and Temperature sensitivity

of the chloride leaching of copper, lead and zinc from a mechanically activated complex

sulphide. Hydrometallurgy. 65, 83-93.

107

Gordon, A.R. and Pickering, R.W. (1975). Improved leaching technologies in the

electrolytic zinc industry. Metallurgical transaction, 6B(2), 45-53.

Gula, M.J., Dreisinger, D.B., Montijo, J.C. and Young, S.K. (1992). The ion exchange of iron

in copper electrolyte streams. Iron control and disposal, 315-329.

Hiroyati, N., Miki, H., Hirajima, T. and Tsunekana,M., (2000). A model for ferrous-promoted

chalcopyrite leaching. Hydrometallurgy. 57, 31-38.

Junea, J.M., Singh, S. and Bose, D.K. (1996). Investigation on the extraction of

molybdenum and rhenium values from low grade molybdenum concentrate.

Hydrometallurgy. 41, 201-209.

Knight, R.J. and Sylva, R.N, (1974), Precipitation in hydrolysed iron (II) solutions. Journal

of Inorganic Nuclear Chemistry.35, 591-597.

Lakshmanan, Rathie, V.I.N. and Monzyk, B., (1992), Evaluation of N-alkyl hydroxamic acid

for selective iron separation from zinc process liquors. Iron control and disposal, 357-364.

Lenahan, W.C. and Murray-Smith, R.de. (1986). Assay and Analytical Practice in the South

African Mining Industry, 156-160.

Mullin, J.W. 1961, Crystallization. London Butterworths,71-73.

Murphy, P.J., Posner, A.M. and Quirk, J. (1976a). Characterisation of partially neutralised

ferric perchloride solutions. Journal of colloid and Interface Science, 56(2), 298-319.

Murphy, P.J., Posner, A.M. and Quirk, J.P. (1976b). Characterisation of partially

neutralised ferric nitrate solutions. Journal of colloid and Interface Science, 56 (2), 270-283.

Murphy, P.J., Posner, A.M. and Quirk, J.P. (1976c). Characterisation of partially neutralised

ferric chloride solutions. Journal of colloid and Interface Science, 56(2), 284-297.

Music, S., Vertes, A.Z., Simmons, G.W. and Czak6-Nagy, I. (1994). Mossbauer

Spectroscopy of Frozen Solutions. J. Colliod and Interface Sci. 85, 287.

108

Nakbanpote, W., Thiravetyan,P. and Kalambaheti,C. (2000). Preconcentration of gold by

rice husk ash. Minerals Engineering, 13(4), 391-400.

Nielson, A.E. (1964). Kinetics of Precipitation, Pergamon Press, New York, 43-44.

Ozberk, E. and Minto, R. (1986). Impact of Iron Control on Process Selection. Iron Control

in Hydrometallurgy, Symposium. Dutrizac, J.E. and Monhemius, A.J. Toronto, Canada, 1 and

12-54.

Rao, V.U.S., Stiegel, G.J., Cinquegrane, G.J. and Srivastara, R.D. (1992). Processing of

complex Ores. Proceedings of Metallurgical Society of the Canadian Institute of Mining. 12,

83.

Schwertmann U., Bigham, J.M and Murad, E. (1995). Iron Oxides in the Laboratory:

Preparation and Characterisation. European Journal of mineralogy. 7. 547-552.

Schwertmann, U. and Cornell, R.M. (1991). Iron Oxides in the Laboratory:. Preparation

and Characterisation. VHC Publishers, N.Y.13-19

Skoog, D., West, D. and Holler, F. (1992). Fundamentals Of analytical Chemistry, 6th Ed.

Saunders College Pub.197-201.

Stein, J.Y and Spink, D.R. (1990). A New Approach for Iron control in Zinc Concentrate

Processing. Hydrometallurgy, 2,17-21.

Stevens, J.G., Kasanov, A.M., Miller, J.W., Pollak, H. and Li, Z. (1998). Mossbauer Mineral

Handbook. M6ssbauer Effect Data Center.

Toni Kaskiala, (1999), Metallurgical Testwork and Research, 12 Aitkein way, Kewdale,

Australia, 6105.

Van Niekerk, C.J and Begley, C.C. (1991). Zinc in South Africa. Journal of South African

Institute of Mining and Metallurgy, 91, 233-248.

109

Young Yu J, Park M. and Kim J. (2002). Solubilities of synthetic schwermannite and

ferrihydrite. Geochemical journal, 36(2), 119-132.

110

Appendices

Table A: Hyperfine interaction parameters at room temperature of candidate Fe-bearing

phases that may occur in the samples (Stevens et al, (1998)).

Spectral

Component

Chemical Formula 6/Fe

(m m/s)

QS

(m m/s)

Bhf

(T)

References

Chalcopyrite# CuFeS2 0.36 0.03 34.2 I

Pyrite FeS2 0.25(1) 0.62(1) - II

Hematite a-Fe203 0.36(1) -0.20(1) 51.1(2) III

Clinochlore* (Mg,Fe)6(Si,A1)4010(0F1)8 : - IV

Fe2+ 1.13(5) 2.61(5) -

Fe3+ 0.40(5) 0.61(5)

Franklinite-ZnFe2O4 0.350(6) 0.333(1) - V

Zn-ferrites® 0.40(2) 0.14(4) 48.0(5) III, VI

0.75(2) 0.08(4) 44.0(5) (x=0.2)

ZnxFe(3-x)04 0.40(2) 0.21(4) 46.4(5) (x=0.4)

0.71(2) 0.15(4) 41.4(5)

0.49(2) 0.08(4) 42.0(5) (x=0.6)

0.62(2) 0.09(4) 35.0(5)

Cu-Zn 0.38(2) -0.6(1) 52(1) VII (x=0)

Ferrites® 0.24(2) 0.0(1) 48(1) CuFe2O4

ZnxCui-x Fe204 0.34(2) 0.2(1) 47(1) (x=0.2)

0.22(2) 0.3(1) 47.5

0.31(2) 0.0(1) 44(1) x=0.4

0.22(2) 0.1(1) 45(1)

relaxation spectru -- x=0.5

m

0.34(2) 0.43(4) x=0.6

0.25(2) --

0.37(2) 0.41(4) -- x=0.8

0.21(2)

111

Spectral

Component

Chemical Formula 6/Fe

(mm/s)

QS

(mm/s)

Bhf

(T)

References

0.36(2) 0.38(4) x=1.0

ZnFe2O4

Cu-Ferrite®' CuFe2O4 049 -0.17 53.8

0.37 0.0 50.3 VIII

Ferric Fe2(SO4)3 0.47

Sulphate

Ferrous FeSO4 1.25 2.88

Sulphate

# No errors given in this article

* Clinochlore has both Fe2 + and Fe3+ oxidation states.

Where two sets of parameters (eg, Bhf values) are indicated per sample, the first set

normally refers to the octahedral B-site, and the other set to the tetrahedral A-site in the

spinel ferrite structure.

4 Parameters for the tetragonal phase. Cubic phase is also magnetic at RT.

Table B: Hyperfine interaction parameters at room temperature of candidate Fe-bearing phases

that may occur in the precipitate samples.

RT , 300 K LHe , -5 K

IS/Fe QS Bhf (T) IS/Fe QS Bhf(T)

mm/s mm/s mm/s mm/s

Jarosite-N H4 0.43 1.22 - - - 48.0

( 1 0.40) (1.15)

Jarosite-Na 0.43 1.20 - - - 47.0

(0.40) (1.05)

Jarosite-K 0.43 1.24 - - - 47.0

(0.40) (1.15)

Jarosite-H30 0.43 1.00 - - - -

112

(0.40) (1.00)

Jarosite-Pb (0.40) (1.15)

Jarosite-K 0.45

(IS/?)

0.42

1.05

1.08

46.0

Zn0.2 Fe2.8 04 0.42

0.75

0.14

0.08

48(5)

44(5)

Zn0.4 Fe2.6 04 0.40

0.71

0.21

0.15

46.4(5)

41.4(4)

Zn0,6 Fe2.4 04 0.49

0.62

0.08

0.09

42(5)

35(5)

Franklinite ZnFe2O4 0.35(1) 0.33 - ? -0.1

0.17

50.2

50.3

Ferrihydrite

5Fe203.9H20

0.35 0.71 0.32 0.02 49.2

0.34 0.71 0.34 0.02 48.4

0.33

0.34

0.87

0.54

0.33 -0

-0

-0

50.8

48.4

44.4

0.34 0.62 0.47 -0 49.8

Schwertmannite

Fe16016(OH)12(SO4)

2

0.36 0.64 0.49 -0.41 45.6

113

0.33

0.39

0.68

0.68

-0.03

-0.33

44.9

45.0

Pyrite Pyrite 0.25 0.62

Table C: Petrographic analysis of polished sections.

No.1 OBSERVATION UNDER STEREOMICROSCOPE: (MACRO DESCRIPTION OF SAMPLE)

A lot of metallic yellow mineral

Colourful gangue inside the mineral

Dark grey mineral which cannot be scratched

OBSERVATION UNDER REFLECTED LIGHT: (OPAQUE MINERALS ONLY)

MAJOR (APPROX. VOL.%) ACCESSORY (APPROX.VOL%)

Chalcopyrite =70% Sphalerite = 15%

Pyrrhotite = 15%

INDIVIDUAL DESCRIPTION OF EACH MINERAL PRESENT, HOW IT OCCURS AND ITS CONDITION

Chalcopyrite

Quite coarse grained

Has some gangue inside it

:. Has some cracks

Pyrrhotite

Has some cracks

:• Medium grained

Sphalerite

:- Medium grained

:. Some part intergrown with sphalerite

A small amount of the sphalerite is diseased with chalcopyrite

Metallurgical implications

the shalerite, liberation will be difficult. the sphalerite. However most of the

Due to the very fine chalcopyrite in Some of the chalcopyrite will go with chalcopyrite is coarse.

114

No.2 OBSERVATION UNDER STEREOMICROSCOPE (MACRO DESCRIPTION OF SAMPLE)

Silicates intergrown with some pinkish mineral

Metallic yellow mineral very fine inside the gangue

The sample seems to be taken across a fissure or fault, which is not mineralised.

OBSERVATION UNDER REFLECTED LIGHT: (OPAQUE MINERALS ONLY)

MAJOR (APPROX. VOL.%) ACCESSORY (APPROX.VOL%)

Pyrite = 20%

Sphalerite =20%

chalcopyrite =1%

INDIVIDUAL DESCRIPTION OF EACH MINERAL PRESENT, HOW IT OCCURS AND ITS CONDITION

Chalcopyirte .

:- Very fine grained (about 20 i_tm)

:• Interlocked in the gangue

Pyrite

4'• Medium grained

❖ Interlocked in gangue

Sphalerite

:- Intergrown with the gangue, grains of less than 100pm

Metallurgical implications

tails due to it being intergrown with be broken up by the fault and are

that would be reporting to the

The sphalerite might be lost with the thw gangue.The minerals seem to complexely intergrown with the minerals tailings.

115

No.3 OBSERVATION UNDER STEREOMICROSCOPE (MACRO DESCRIPTION OF SAMPLE)

Pinkish-grey mineral interlocked in a yellow mineral

A grey mineral which is intergrown in with the gangue

Complex intergrowth of minerals

OBSERVATION UNDER REFLECTED LIGHT: (OPAQUE MINERALS ONLY)

MAJOR (APPROX. VOL.%) ACCESSORY (APPROX.VOL%)

Sphalerite = 40%

Chalcopyrite =25%

Pyrite = 20%

Pyrrhotite = 5 %

INDIVIDUAL DESCRIPTION OF EACH MINERAL PRESENT, HOW IT OCCURS AND ITS CONDITION

Chalcopyrite

Medium to course grained

lntergrown with sphalerite

:• The medium grained has some pyrrhotite in it

Pyrite

Medium to coarse grained

Interlocked in chalcopyrite

Interlocked in gangue

:• Often coated in cpy therefor loss of cpy to tailings

Sphalerite

Coarse grained

lntergrown with the gangue has fine chalcopyrite in it

Pyrrhotite

:. Fine grained

Interlocked with pyrite

Metallurgical implications

the zinc concentrate and also in the Some of the chalcopyrite will be in tailings.

No.4 OBSERVATION UNDER STEREOMICROSCOPE (MACRO DESCRIPTION OF SAMPLE)

116

♦ Metallic yellow mineral interlocked in a greyish-brown mineral

Fine pinkish mineral interlocked in greyish mineral

The metallic minerals are quite sparsely distributed in this sample with a lot of silicate (gangue) mineral

OBSERVATION UNDER REFLECTED LIGHT: (OPAQUE MINERALS ONLY)

MAJOR (APPROX. VOL.%) ACCESSORY (APPROX.VOL%)

Sphalerite = 40%

chalcopyrite = 20%

Pyrite = 20%

Pyrrhotite = 5%

INDIVIDUAL DESCRIPTION OF EACH MINERAL PRESENT, HOW IT OCCURS AND ITS CONDITION .

Chalcopyrite

Fine to medium grained (100iim)

•:- Interlocked in sphalerite

Pyrite

Coarse to fine (sometimes +1mm)

Interlocked in sphalerite

Sphalerite

Coarse grained

lntergrown with the gangue has fine chalcopyrite in it

Pyrrhotite

Very fine grained

Interlocked in sphalerite

There may be some fine magnetite

Metallurgical implications

since it is finely intergrown with pyrite might report to the zinc

so should be liberated with little spl

The sphalerite will be difficult to liberate the gangue. The chalcopyrite and concentrate. The pyrite is fairly coarse but maybe some CU

117

No.5 OBSERVATION UNDER STEREOMICROSCOPE (MACRO DESCRIPTION OF SAMPLE)

Brass metallic mineral intergrown with a dark grey mineral

Metallic yellow mineral on the edges of the gangue

OBSERVATION UNDER REFLECTED LIGHT: (OPAQUE MINERALS ONLY)

MAJOR (APPROX. VOL.%) ACCESSORY (APPROX.VOL%)

Pyrite = 65% Sphalerite = 10%

Chalcopyrite = 10%

INDIVIDUAL DESCRIPTION OF EACH MINERAL PRESENT, HOW IT OCCURS AND ITS CONDITION

Chalcopyrite

Fine grained

Interlocked in sphalerite

lntergrown with gangue

Quite intricate or complicated locking

Pyrite

:. Course grained (±20011)

Sphalerite interlocked around pyrite

Has some cracks

Some medium grained pyrite is interlocked in chalcopyrite

:. some pyrite contains sphalerite inclusions

:. Zn in tailings

Sphalerite

Coarse grained

:. lntergrown with chalcopyrite

Locked in some of the gangue

May float with sph

Fine grains interlocked in pyrite

Metallurgical implications

so a lot of gangue may report with cpr and pyrite.

It will be difficult to liberate the sphalerite, with the zinc concentrate. Zn will report

No.6 OBSERVATION UNDER STEREOMICROSCOPE (MACRO DESCRIPTION OF SAMPLE)

Grey mineral interlocked with a metallic yellow mineral and a brass

118

coloured merallic mineral

♦ Gangue inside the grey mineral

OBSERVATION UNDER REFLECTED LIGHT: (OPAQUE MINERALS ONLY) MAJOR (APPROX. VOL.%) ACCESSORY (APPROX.VOL%) Sphalerite = 40%

Chalcopyrite = 30%

Pyrite = 25%

Pyrrhotite = 5%

INDIVIDUAL DESCRIPTION OF EACH MINERAL PRESENT, HOW IT OCCURS AND ITS CONDITION Chalcopyrite

:• Coarse grained and fine (20 - 3000

Interlocked in sphalerite (fairly complex locking)

Some cpy is embedded within the gangue as fine inclusions

:• cu in silicate tailings

Pyrite

Course grained (-10011)

Interlocked in sphalerite and in chalcopyrite but simple locking

Has some fine sphalerite in it and so Zn in pyrite which will be in tailings

Sphalerite

Coarse grained and should be easy to get 80% liberation at -74µ

Intergrown with chalcopyrite

Intergrown with gangue

Pyrrhotite

Medium to course grained(±50 - 100p) Intergrown with silicate that looks like mica Interlocked in sphalerite along simple curved contacts should be liberated

Metallurgical implications

intergrown along curved contacts with the and chalcopyrite will report to the zinc of sphalerite ±70% should be liberated.

Loss of Zn to tailings. The sphalerite it minerals, so an amount of pyrrhotite concentrate but quitr large coarse grains

119

No.7 OBSERVATION UNDER STEREOMICROSCOPE (MACRO DESCRIPTION OF SAMPLE)

Grey mineral with patches of yellowish-white mineral

Gangue intergrown with the grey mineral

OBSERVATION UNDER REFLECTED LIGHT: (OPAQUE MINERALS ONLY)

MAJOR (APPROX. VOL.%) ACCESSORY (APPROX.VOL%)

Sphalerite = 40%

Pyrite = 50%

Chalcopyrite = 2%

INDIVIDUAL DESCRIPTION OF EACH MINERAL PRESENT, HOW IT OCCURS AND ITS CONDITION

Chalcopyrite

:. Fine grained (2011m)

Interlocked in gangue and sphalerite

Pyrite

:• Fine to coarse often cubic (100gm)

Intergrown with sphalerite along straight contacts

:•

Interlocked in sphalerite wher sphalerite is the main mineral and pyrite is enclosed

:- Some fine grain gangue associated (501.tm)

Sphalerite

:• Coarse grain

:• Has some fine grained gangue included

Intergrown with pyrite as described.

Metallurgical implications

sphalerite which is intergrown with contacts and therefore the liberation

This polished section shows coarse 1001..tm grains of pyrite along straight will be good.

120