High Temperature Roasting of Sulphide Concentrate and Its Effect on the Type of Precipitate Formed
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
Table 4.4.4.2: Dissolution of elements during hot acid leaching8l
Table 4.4.4.3: Table showing percentages of elements remaining after hot acid leaching. 82
Table 4.5.1: Percentage Extraction in 5M H 2SO4 and 10% (w/v) Mn02 83
Table 4.5.2: Percentage Extraction in 7M H 2SO4 and 10% (w/v) Mn02. 85
Table 4.5.3: Percentage Extraction in 5M H 2SO4 and 20% (w/v) Mn02. 85
Table 4.5.4: Percentage Extraction in 7M H 2SO4 and 20% (w/v) Mn02 85
Table 4.5.5: Percentage Extraction in 5M H 2SO4 and 10% (w/v) Na2S2O8 . 86
Table 4.5.6: Percentage Extraction in 7M H 2SO4 and 10% (w/v) Na2S2O8. 87
Table 4.5.7: Percentage Extraction in 5M H 2SO4 and 20% (w/v) Na2S2O8. 88
Table 4.5.8: Percentage Extraction in 7M H 2SO4 and 20% (w/v) Na2S2O8. 88
Table 4.5.9: Percentage Extraction of residue in 7M H 2SO4 and 20% (w/v) Mn02 89
Table 4.6.1.1: Table summarising the amount of iron (in %) remaining from precipitation at 80 °C and filtration after 24 hours. 94
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 HNO3 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) Cu2S + 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 —> 0.87ZnSO4 + 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 (2.3.1- AH°(685°C) = 156.3 Kcal 7)
ZnO.2ZnSO4 + Fe2O3 -+ ZnO.Fe203 + 2ZnSO4 (2.3.1-8) AH° (685°C) = —1.79 KCal
CuFeS2 + 3.7502 —> CuSO4 + 0.5Fe203 + SO2 (2.3.1-9) AH° (685° C) = — 309.21 KCal
CuFeS2 + 3.7502 —> 0.5(CuO.CuSO4 )+ 0.5Fe203 + 1.5S02 (2.3.1-10) AH° (685°C) = — 412.0 KCal
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 —> Zn2+ + SO4- + H2O (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.Fe2 0 3 + 4H 2S0 4 —> ZnSO4 + Fe2 (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 (2.4.2-3) RT K is a constant E is activation energy R is the gas constant T is absolute temperature
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 = (2.5-1)
Where 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 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+ OH Fe(OH)2+ "4-111' Fe OH - FiVVH\e
A O H O H n+ nH + Ne/ Ne/ n/2 \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.
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 Nil Nat, K+, NH 4,1- (= R) Nil required
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/dm3, 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 Dowfroth % (Mt) (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 in Figure 4.2.2. Only the phases of interest (sphalerite, chalcopyrite and pyrite) were identified in the spectra.
ROM
CONCENTRATE
10 40 50 70 Two Theta Degrees
Fig 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 % S after % S after % S after (°C) 2 Hrs 3 Hrs 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 minimum amountofoxygen thatcouldbeprovidedbythefurnace. have beenlowerifthe furnacesweresupplyingenoughoxygen. Inthisstudy,wereliedonthe to 800 roasting hasbeendone.Itisassumedthat thesulphurcontentateachroastingconditioncould roasting conditions,800 significantly asthetemperatureincreases. More sulphurisremovedwhenchangingfrom700 with respecttotemperatureandtime.The removalofsulphurisalsonotedtobedecreasing temperatures. Theslopesofthegraphscan beusedtocomparethekineticsofreactions Figure 4.3.1.1showstheamountofremainingsulphurwhenroastingwasdoneat different Figure 4.3.1.1: ° C thanwhenroastingatthesametemperature for2,3and4hours.Attheoptimised % S Retained 700 Percentage sulphurremainingatvaryingroastingconditionsobservingthe ° C for3hrs,thereisonly2.92%sulphurremaining, thusarelativegood 750 Roasting Temperature( 800 51 850 temperature effect. 900 ° C) 950 1000 ° C 0 1 1 1 1 0 1 2 3 4 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 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)
1530
Cap entrate 1707
577
A
703°C
800°C
470 900°C
200
0 10 27 37 40 57 70 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
53 900°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.
1600
- CuFeS2 1200 - ZnS ZnSnS Concentrate 800 - CuFeS2 ZnS FeS2 400 - FeS2 0 ZO ZF ZO 700°C,2hrs
a) ZO ZF ZO 700°C,3hrs
300- re 3 ZF ZO ZF ZF Z ZO 200 - / 700°C, 4hrs 100 -
0 #V1"40/1L. 10 20 30 40 50 70 Two theta degrees
Figure 4.3.2.2: Effect of roasting time on the formation of zinc oxide and zinc ferrite at 700°C.
54 10 30 40 Two theta degrees
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 Pyrit Sphalerite Chalcopyri Quartz Chlonochlore Zinc oxide Muscovite Zinc Wille identific e ZnS te CuFeS2 Si02 (mg,Fe)6 ZnO KAI2(Si3A1)10 Ferrite mite ation FeS2 (Si,A1)4010(0F1)13 (OH, F)2 Zn Fe2O4 Zn2Si 04 ROM 88 696 166 556 187 275 506 2625 527 3155 1492 619 Concent 104 1004 190 400 116 166 rate 587 4642 1629 466 1636 415 R.C.900° 132 187 219 84 C 4hrs 689 1042 2235 524 R.C.800° 571 1741 2391 946 C 4hrs R.C.800° 1016 1933 2480 631 C 3hrs R.C.700° 780 1371 2785 865 C 4 hrs R.C.900° 328 428 384 44 C 2hrs R.C.900° 151 306 585 64 C 3hrs magneti 70 567 146 412 169 132 cs Non- 67 506 174 416 174 70 magneti cs Iron AMORPHOUS precipita te X — lab = A, X = lab B, X = lab C R.C. - ROASTED CONCENTRATE.
59 420 Zinc Ferrite 400 - 380 - • 360 - Zincite
340 - P' Cc 320 - a) - 300 -
280 -
260 - A Willemite 240 - • 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 Zn),Fe3_,(04, Cu.Fe3..04 and/or Cui..Zn.Fe204, may have overlapping Mossbauer parameters for a range of compositions (x values).
U)0 — E (/) C cu
(T) OC 0.95 -
• -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 r, IS/Fe QS Bhf *Abundance Component Linewidth (mm/s) (mm/s) (T) % (mm/s) Concentrate Chalcopyrite 0.32 0.25 -0.01 34.9 21(2)
61 Pyrite 0.32 0.30 0.62 - 60(2) Chlinochlore 0.38 0.99 2.64 - 6(1)
Fe2+ 0.40# 0.40# 0.61 4 - 13(2) Fe3+
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.
Spectral F, Linewidth IS/ Fe QS Bhf *Abu ndance % SAMPLE Component (mm/s) (mm/s) (m m/s) (T)
Hematite 0.34 0.36 -0.2 51.2(2) 42(2) Franklinite, 0.40 0.33 0.45 36(2) Zn-rich ferrites 700° C Ferrites 0.32 0.36 0.14 10-50 22 (2)
Hematite 0.32 0.37 -0.2 51.3(2) 23(2) Franklinite, 0.42 0.33 0.45 46(2) Zn-rich ferrites 800° C Ferrites 0.32 0.26 -0.15 10-50 31(3)
Hematite 0.30 0.39 -0.2 51.3(2) 6(2) Franklinite, 0.45 0.33 0.48 59(2) Zn-rich ferrites 900° C 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 These ferritesaresaid tobemagnetic,dependingontheamount ofzincinthespinel.The in thesampleroasted at700 magnetism collapses astheamountofzincincreasesinspine!. Similartotheassignments not beidentifiedintheXRDpatterndueto overlapwiththezincoxideandferritepeaks. sextets wasassignedtotheotherZn-orCu-depleted ferrites.Thesecopperferritescouldalso franklinites intheirdifferentformsasdidthe Mossbauer.InFigure4.3.3.2,thedistributionof resulted intheformationoflargeamounts offranklinite.TheXRDhowevercouldnotgivethe by theXRDbecausetheyoverlapedwith thoseforzincoxideandferrite.Itspresence crystillanity (acomparativelynarrowIinewidth F-0.34mm/s)itspeakscouldhardlybeidentified sextet andthecentraldoubletrepresentingfranklinite.Thehematite,thoughshowing good 900 Figure 4.3.3.2showstheMossbauerspectraofcalcinesroastedfor2hrsat700 Figure 4.3.3.2: ° C.The 700 . Transmission ° 0 0 C spectrumwasfittedwithtwocrystallinecomponents.Thehematitebeingthe -8 800 Mossbauer spectraofcalcinesroastedfor2hrsatdifferenttemperatures(700 11 ° C, and900 -6 ° C, the800 ° -4 C). Velocity (mm/s) -2 ° C calcinesamplehad thesextetassignedtohematite '3111. 0 64 • lb 10 0 14 0 1•
e t 0. 2
4 900 700 ° ° C C 6 8 0 II ° C, 800 ° C and ° C, 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 Abundance Abundance Abundance component (%) (%) (%) 700°C 800°C 900°C Hematite 42(2) 23(2) 6(2)
Franklinite (zn- 36(2) 46(2) 59(2) rich Ferrites)
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 2O3 ZnFe2O4 (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 60
55
50
45
40
35 dance 30
25 Abun
% 20
15
10
5 I I I 1 I 700 750 800 850 900 Temperature ( °C)
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
Spectral F, Linewidth IS/Fe OS Bhf Abundance % SAMPLE Component (mm/s) (mm/s) (mm/s) (T)
Hematite 0.30 0.37 -0.2 51.4(2) 49(2) 700° C Franklinite, 0.42 0.34 0.44 - 34(2) Zn-rich ferrites Other/mixed . Ferrites 0.30 0.32 -0.07 10-50 17 (2)
Hematite 0.41 0.37 -0.25 47(1) 12(2) Franklinite, 0.48 0.32 0.42 - 48(2) 800° C Zn-rich ferrites Other/mixed Ferrites 0.32 0.33 0.09 10-50 40(2)
Hematite 0.34 0.36 -0.17 51.0(2) 9(2) Franklinite, 0.48 0.33 0.51 - 52(2) 900° C Zn-rich ferrites Other/mixed 0.30 0.35 -0.02 10-50 39(3) Ferrites
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 was insignificant. the sulphurremoval,changeinamount ofsulphurfrom800 to beless,ascomparedthatwhentheroasting wasdonefor2hours.Asitobservedin change intheamountofhematiteformedat 800 formed at900 shows moreclearlyatthehightemperature (900 hematite isseentobedisappearingasthetemperatureincreased.Thesextetfor ferrites 700 Figure 4.3.3.4showsthedisappearanceofhematiteastemperatureisincreased from Figure 4.3.3.4: ° C to900
Transm ission ° ° C. Theeffectoftemperatureisseenatthe4hoursroastingperiod.sextetfor C forbothroastingperiodsisalmostequal ascanbeseeninTable4.3.3.6.The -8 M6ssbauer spectraofthecalcinesamples,roastedfor4hrsatdifferenttemperatures -6 -4 t ;;;;;; Velocity (mm/s) tttttttttttt -2 68 0 ° C, comparedtothatformedat900 ° C). Theamountofferritesandfranklinite 2 4 ° C to900 6 ° C (4hourroasting) 8 ° C isnoted Table 4.3.3.6: Comparison of component abundances (4 hrs) Spectral Abundance Abundance Abundance (%) component . (%) (%) 700°C 800°C 900°C Hematite 49(2) 12(2) 9(2)
Franklinite (zn- 34(2) 48(2) 52(2) rich Ferrites)
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.
Spectral F, IS/Fe OS Bhf Abundance % SAMPLE Component Linewidth (mm/s) (mm/s) (T) (mm/s)
Hematite 0.32 0.37 -0.2 51.3(2) 23(2) Franklinite, 0.42 0.33 0.45 - 46(2) Zn-rich 2hrs ferrites Other/mixed 0.32 0.26 -0.15 10-50 31(3) Ferrites
Hematite 0.33 0.36 -0.2 50.8(2) 28(2) Franklinite, 0.41 0.32 0.46 - 40(2) Zn-rich 3hrs ferrites Other/mixed 0.32* 0.37 0.08 10-50* 32(3) Ferrites*
Hematite 0.41 0.37 -0.25 47(1) 12(2) Franklinite, 0.48 0.32 0.42 - 48(2) Zn-rich ferrites 4hrs Other/mixed 0.32 0.33 0.09 10-50 40(2) Ferrites
70 •._O ca 4 h rs U)
• a) CC
-8 -6 -4 -2 0 2 4 6 8 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 Abundance Abundance Abundance (%) (%) component (%) 2hrs 3hrs 4hrs Hematite 23(2) 28(2) 12(2)
Franklinite (zn- 46(2) 40(2) 48(2) rich Ferrites)
Other/mixed 31(3) 32(3) 40(2) Ferrites
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.
heat ZnSO4 + 02 ZnO + Sgas (4.3.3 -5)
heat ZnO + Fe203 Zn O Fe2 03 (4.3.3-6)
heat CuSO4 + 02 CuO Sgas (4.3.3-7)
heat CuO + Fe203 CuO • Fe203 (4.3.3 - 8)
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 HCI H2SO4 HNO3 (Hrs) % 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- ion t
80- trac
Ex 75- %
70-
65-
60 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 4.4.3 H2SO4 Neutral leach
100 -
95 -
90 -
85 - ion t 80 -
trac 75 - Ex
cY0 70 -
65 -
60 -
55 , I I i i . I . 1 I I 0 1 2 3 4 5 6 7 8 9 Time (Hrs)
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 4.4.4 HNO3 Neutral leach
100 -
95 -
90 -
85 - ion 80 - tract 75 - % Ex 70 -
65 -
60 - 1 I i I , I i 2 3 4 5 6 7 8 9 Time (Hrs)
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 , I g i • 1 0 1 2 3 4 5 6 8 9 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 90- • A ♦ t-___--:____--- . 85- /* 80- / c 75- N _ , . / -° 70- ♦ 65- 72 HCI • A 60- HNO3A H2S0 55 I I i , I . I . I I 0 1 2 3 4 5 6 7 8 9 Time (Hrs)
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 HCI H2SO4 HNO3 (hrs)
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 and theresultsareshowninTable4.5.2 was increasedto7Mwhiletheconcentrationforoxidisingagentkeptconstant(10`Yow/v) sphalerite (thoughtoalesserextent)atthesametime.Theconcentrationforleaching acid be extracted.StartingwiththeMn02alsoseemedtoattackbothchalcopyrite andthe pyrite. Theextractionhoweverwasslowestandafter6hours,only47%Cu44%Zn could chalcopyrite sincehigherpercentagescouldbeobservedifitwascombinedwiththat fromthe chalcopyrite orfromthepyrite.Itseemedthoughthatironintosolutionis fromthe The dissolutionofironwasalsoobservedbutnotprovenifitextracted fromthe More copperthanzincseemstobeextractedwhentheMn02isusedfirstasanoxidising agent. Figure 4.5.1
% Extraction 20 - 10 25 - 15- 30 - 35 - 40 45 - 50
0
Percentage Extractionin5MH 1 2 Time (Hrs) 3 4 2 SO 84 4 and10%(w/v)Mn02. 5 6 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 Table 4.5.5: which mineralwouldbeattackedfirst.Table4.5.5andFigure4.5.3showtheresultsobtained. The concentratewasalsoleachedusingNa2S2O8asastartingoxidisingagentto determine were added,increasedsimultaneously. time. Thetotalextractionwashoweverhigherthanwhenboththeacidandoxidising agents Though theextractionrateisnotedtoincrease,zincandcopperweredissolvedat thesame Time (Hrs) % Extraction 0.5 6 4 2 Percentage Extractionin5MH2SO4and10%(w/v)Na2S2O8. Figure 4.5.2 1 2 48 Zn 43 36 29 Percentage Extractionin7MH Time (Hrs) 3 % Extraction 4 86 Cu 12 11 8 5 5 2 SO 6 4 and20%(w/v)Mn02. Fe 8 8 6 3 Table 4.5.6: concentration, aswouldbeexpected.TheresultsareshowninTable4.5.6. The acidconcentrationwasincreasedto7Mwhilekeepingtheoxidisingagentconstant at concentrations weretoolowtobeextractedfromboththechalcopyriteandpyrite. 10%w/v. Thepercentageextractionwasnotedtoincreasewithan inacid degree thanchalcopyrite(seeFigure4.5.3).Ironisalsonotedtobeextracted but extraction ofzincpreferableovercopper,theuseNa2S2O8attackssphalerite to ahigher Though Figure 4.5.3 Time (Hrs) % Extraction 0.5 6 4 2 it couldbeseenthatthe5Mofconcentrationleachingacidisnotoptimumfor 0 Percentage Extractionin7MH2SO4and10% (w/v)Na2S2O8. Percentage Extractionin5MH 1 2 54 49 Zn 43 31 Time (Hrs) 3 I % Extraction 4 1 2 SO 87 Cu 15 13 9 6 4 and10%(w/v)Na2S208. 5 1 6 i Fe 12 11 8 5 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 Table 4.5.9: The resultsareshowninTable4.5.9andFigure4.5.5. analysed, leachedwiththesameconcentrationofacidbutusingMn02asanoxidising agent. The residuefromthe7MH2SO4and20%(w/v)Na2S2O8was,afterbeingdigested and and leavingmostofthechalcopyritepyritenotattacked. be leached.ItishoweverseenthatstartingwiththeNa2S2O8providedapossibilityto leach zinc again nottheoptimumconditionsforleachingsinceeven70%oftargeted zinccould percentages arenotedtoincrease.The7MH2SO4and20%(w/v)Na2S2O8(Figure4.5.4.) were The increaseintheconcentrationsalsogavesametrendextractionexcept thatthe Figure 4.5.4 Time (Hrs) % Extraction 0.5 6 4 2 Percentage Extractionofresiduein7MH2SO4 and20%(w/v)Mn02. Percentage Extractionin7MH i Zn 2 15 12 11 4 Time (Hrs) 3 % Extraction 4 2 89 SO Cu 41 37 28 19 4 and20%(w/v)Na 5 6 2 17 Fe 17 15 11 S 2 O 8 . much pyritehasbeen attackedinanyoftheoxidisingagentsused. characterisation techniques beusedfortheresidue.Thiswill allow thedeterminationofhow extraction percentages andthesourceofirondissolved, itisrecommendedthatother feasible ifthemineralwithhigherpercentage (sphalerite)wasleachedfirst.Toconfirmthe was alsonotedtobedissolvedwellafterthe sphaleritehadbeenremovedfromtheconcentrate. In thisconcentratewithahigherpercentage ofsphalerite,theselectiveleachingwasmore here thatchalcopyritecanalsodissolvein H2SO4 mostlyinthepresenceofMn02.Chalcopyrite choice oftheoxidisingagentplayedamajorroleindeterminingwhichmineralsdissolved first The selectiveleachingofthissulphideorewaspossibleinthepresencesulphuric acid.The into thesolution.Thoughleachingofchalcopyrite iscommonlydoneinHCI,itwasobserved 4.5.1 to whentheinitialconcentratewasstartedwithMn02asanoxidisingagent. preference oftheleachingisnotedtobehigheraftersomezinchadbeenleachedas compared rich inzincandmostofthewasstillremainingresidue.Thedifference inthe residue digestionwerestillhigherthanthepercentageofcoppersinceconcentrate wasvery Mn02 seemedtofavourthedissolutionofchalcopyrite.Thepercentageszinc afterthe
c/0 Extraction Conclusions Figure 4.5.5% 0 1 2 Percentage Extractionin7MH2SO4and20%(w/v)Mn02. Time (Hrs) 3 4 90 5 6 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 pH Time in hours added 30 60 2 3 4 5 24 mins mins 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 PH Time in hours 3 6 24 added 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 The cationsfromthefigureaboveseemtoplaynomajorroleonrateandefficiency of remaining fromboththeNaOHandNH4OH. precipitation. Afterthesamedurationofprecipitation,therewasanalmostequalamount ofiron 4.6.1.1: pregnant solution.
`)/0 Fe Remaining 20 - 30 - 10 - 40 - 50 - 60 0 0
The comparisonofNaOHandNH4OHusedinprecipitatingironfromanHCI 5 10 Time (Hrs) 97 15 20 +— N N NH Na Na Na H H + + ÷ 4 4 4 ,pH2.5 ,pH2.0 ,pH1.5 + + + , , ,pH1.5 pH2.5 pH2.0 25 Figure the HCImeaningthatthereismoreironremaininginH2SO4comparedto in theHCI The effectofthe2acidscanbeobservedinFigure4.6.1.2.precipitation has been pregnant solutionwhenprecipitatedatthesamepH. compared inthe2acids.WhenvaryingpH,valuesforironH2SO4arehigher thanin Figure 4.6.1.2:
% Fe Remaining 20 30 40 10 50 60 70 0 0 The comparisonofHCIandH NaOH. 5 10 Time (Hrs) 98 15 2 SO4 pregnantsolutionsinprecipitatingironusing —•— H —•— HCI,pH2.0 • 20 H H HCI,pH2.5 HCI,pH1.5 2 2 2 SO SO SO 4 4 4 ,pH2.0 ,pH1.5 ,pH2.5 25 Figure 4.6.1.3alsoshowstheinsignificanteffectonrateandefficiencyofprecipitating iron. pregnant solution. 4.6.1.3:
`)/0 Fe Remain ing 20 30 40 10 50 60 70 0 0 The comparisonofNaOHandNH4OHusedinprecipitatingironfromaH2SO4 5 I 10 Time (Hrs) 99 15 1 —•— NH —v— 20 I A +— NH *— Na - Na NH Na + + + 4 4 4 ,pH2.5 ,pH2.0 ,pH1 .5 + + + , , pH2.0 , pH2.5 pH 1.5 25 I Figure efficiency ofironremoved. The changeoftemperaturefrom80 pH. It canbeseenthatdifferentacidsdohaveaneffectontheamountofironprecipitated atagiven Figure 4.6.1.4:
°/0 Fe Remain ing 20 30 10 40 50 60 70 0 0 The effectofH NaOH. 5 2 SO 4 andHNO ° C to95 10 3 Time (Hrs) ° intheamountofironremainingfromprecipitatingwith 100 C wasfoundtohavelittleeffectontherateand 15 —v— HNO —•— H +— HNO A 20 - H HNO H 2 2 2 SO SO SO 3 3 3 4 4 4 ,pH2.5 ,pH2.0 ,pH1.5 ,pH2.5 ,pH2.0 ,pH1.5 25 4.6.3 XRD characterisation of precipitates.
400
350 - goethite goethite 300 - morhit 250 - H SO +NH4+,pH2.5 200 - 2 4 150- morhite
100-
50 akaganeite akaganeite
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 20 30 40 50 60 70 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)
cc1) toi 1.00 0.99 0.98 0.97 096 0.95 -4 -2 2 4 Velocity (rrm's)
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. Spectral F, 8/Fe QS Bhf Abundance Component Linewidth (mm/s) (mm/s) (T) SAMPLE % (mm/s) H2so4+NR4ox Quad-1 0.36 0.37 0.56 - 67(2) pH 2.5 Quad-2 0.39 0.37 0.96 - 33(2) HCI+NaOH Quad-1 0.36 0.36 0.59 - 75(2) pH2.5 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 Chemical Formula 6/Fe QS Bhf References Component (m m/s) (m m/s) (T) 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 Chemical Formula 6/Fe QS Bhf References Component (mm/s) (mm/s) (T) 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 (10.40) (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 1.05 46.0 (IS/?) 1.08 0.42
Zn0.2 Fe2.8 04 0.42 0.14 48(5) 0.75 0.08 44(5)
Zn0.4 Fe2.6 04 0.40 0.21 46.4(5) 0.71 0.15 41.4(4)
Zn0,6 Fe2.4 04 0.49 0.08 42(5) 0.62 0.09 35(5)
Franklinite ZnFe2O4 0.35(1) 0.33 - ? -0.1 50.2 0.17 50.3
Ferrihydrite 0.35 0.71 0.32 0.02 49.2 5Fe203.9H20
0.34 0.71 0.34 0.02 48.4
0.33 0.87 0.33 -0 50.8 0.34 0.54 -0 48.4 -0 44.4
0.34 0.62 0.47 -0 49.8
Schwertmannite 0.36 0.64 0.49 -0.41 45.6 Fe16016(OH)12(SO4)
2
113 0.33 0.68 -0.03 44.9 0.39 0.68 -0.33 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 Due to the very fine chalcopyrite in the shalerite, liberation will be difficult. Some of the chalcopyrite will go with the sphalerite. However most of the 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% chalcopyrite =1% Sphalerite =20% 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 The sphalerite might be lost with the tails due to it being intergrown with thw gangue.The minerals seem to be broken up by the fault and are complexely intergrown with the minerals that would be reporting to the 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% Pyrrhotite = 5 % Chalcopyrite =25% Pyrite = 20% 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 Some of the chalcopyrite will be in the zinc concentrate and also in the 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% Pyrrhotite = 5% chalcopyrite = 20% Pyrite = 20% 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 The sphalerite will be difficult to liberate since it is finely intergrown with the gangue. The chalcopyrite and pyrite might report to the zinc concentrate. The pyrite is fairly coarse so should be liberated with little spl 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 It will be difficult to liberate the sphalerite, so a lot of gangue may report with the zinc concentrate. Zn will report with cpr and pyrite. 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% Pyrrhotite = 5% Chalcopyrite = 30% Pyrite = 25% 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 Loss of Zn to tailings. The sphalerite it intergrown along curved contacts with the minerals, so an amount of pyrrhotite and chalcopyrite will report to the zinc concentrate but quitr large coarse grains of sphalerite ±70% should be liberated.
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% Chalcopyrite = 2% Pyrite = 50% 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 This polished section shows coarse sphalerite which is intergrown with 1001..tm grains of pyrite along straight contacts and therefore the liberation will be good.
120