EXTRACTION AND ECONOMIC UTILIZATION OF COPPER...
135
EXTRACTION FOR ECONOMIC UTILIZATION OF INDIGENOUS COPPER ORES BY USING HYDROMETALLURGICALTECHNIQUES IFFAT TAHIRA SIDDIQUE QADRI University Registration No. 0499066 UNIVERSITY OF EDUCATION LAHORE 2008
Transcript of EXTRACTION AND ECONOMIC UTILIZATION OF COPPER...
EXTRACTION FOR ECONOMIC UTILIZATION
OF INDIGENOUS COPPER ORES BY USING
HYDROMETALLURGICALTECHNIQUES
IFFAT TAHIRA SIDDIQUE QADRI
University Registration No. 0499066
UNIVERSITY OF EDUCATION LAHORE
2008
14
EXTRACTION FOR ECONOMIC UTILIZATION
OF INDIGENOUS COPPER ORES BY USING
HYDROMETALLURGICALTECHNIQUES
By
IFFAT TAHIRA SIDDIQUE QADRI
University Registration No. 0499066
Submitted in partial fulfillment of the requirements for the degree of
Doctor of Philosophy in Chemistry at the Division of Science and Technology
University of Education, Lahore
Supervisor: Prof. Dr. Izhar-ul-Haq Khan
JUNE 2008
15
DEDICATION
The thesis and the efforts in the way of its completion are dedicated
TO
THE GREATEST TEACHER
BELOVED HOLY PROPHET “MUHAMMAD”
SALLAL LA HO ALY HAY WAALY HI WASLLEM
WHO HAS
ALL THE KNOWLEDGE IN
THE UNIVERSE
iii
16
CONTENTS
S.No PRELIMINARIES Page
1 Title i
2 Bismillah ii
3 Dedication iii
4 Approval Sheet iv
5 Abstract v
6 Acknowledgement vii
7 List of Tables xiii
8 List of Figures xv
9 Symbols, Abbreviations & Acronyms xvi
CHAPTER 1
INTRODUCTION 1
1.1 Copper Ore Deposits in Pakistan 6
1.2 Scope of Study 7
1.3 Significance of the Study 11
1.4 Objectives 12
CHAPTER 2
REVIEW OF LITERATURE 13
2.1 Importance of Hydrometallurgy 14
2.2 Types of Leaching 17
2.2.1 Leaching with Sulfuric Acid 20
2.2.2 Dump and Heap Leaching 21
2.2.3 Pressure Leaching 22
2.2.4 Bioleaching/Bacterial Leaching 25
2.2.5 Catalyzed Leaching 26
2.2.6 Roast Leach Process 27
17
2.3 World Class Hydrometallurgical Leaching Plants 30
2.3.1 MT Gordon Plant 30
2.3.2 Activox Process 30
2.3.3 CESL Process 31
2.3.4 Dynatec Process 31
2.3.5 NSC Pressure Oxidation Process 31
2.3.6 High Temperature Pressure Oxidation Process 32
2.3.7 Nena Tec Process 32
2.3.8 Intec Process 32
2.3.9 Chloride/Sulfuric Acid Leaching Process 33
2.3.10 Canmet Process 33
2.3.11 Cymet Process 33
2.3.12 Dyuval CLEAR Process 33
2.3.13 Dextec Process 34
2.3.14 Elkem Process 34
2.3.15 Cuprex Process 34
2.3.16 Noranda Antlerite Process 35
2.3.17 BHAS Process 35
2.4 Patented Hydrometallurgical Processes 36
2.4.1 Roast-Leach Process 36
2.4.2 Ammonia Leach Process 39
2.4.3 Chloride Based Leach Process 40
2.4.4 Chloride-Sulfate Leach Process 41
2.4.5 Counter Current Leach Process 43
2.4.6 Temperature Pressure Acid Leach Process 43
2.5 Reagents Used in Leach Processes 45
2.5.1 Similarities of Hydrometallurgical Leaching Techniques
(Copper Sulfide Ores)
47
2.6 Solvent Extraction 48
2.7 Kinetics and Thermodynamics Studies 50
18
CHAPTER 3
MATERIALS METHODOLOGY AND EXPERIMENTAL 52
3.1 Materials 52
3.1.1 Equipments 52
3.1.2 Chemical and Reagents 52
3.2 Methods 54
3.2.1 Sample Procurement 54
3.2.2 Sample Preparation 54
3.2.3 Mineralogical Studies 55
3.2.4 Chemical Analysis 55
3.2.5 Chemical Processing 55
3.2.6 Kinetic Studies 57
3.3 Experimental Work 58
3.3.1 Sampling and Comminution Studies 58
3.3.2 Microscopic Analysis of Samples 59
3.3.3 Chemical Analysis 59
3.3.4 Chemical Processing 59
3.3.4.1 Direct Leaching 59
3.3.4.2 Indirect Leaching (Roasting before Leaching) 61
3.3.4.2.1 Simple Roasting before Acid Leaching 62
3.3.4.2.2 Roasting of Sample C with Oxidants before Acid Leaching 63
3.3.4.2.3 Roasting of Sample C with Novel Oxidant before Acid Leaching 64
3.3.4.2.4 Roasting of Sample C with Non-Oxidant Material/Novel Additives 64
3.3.4.2.4.1 Two-Stage Roasting 65
3.3.4.2.4.2 Effect of Combined Oxidant/Additive on the Copper Leaching 66
3.3.5 Leaching of Copper Chalcopyrite Concentrate 66
3.3.6 Amount of Sulfur Analyzed in Different Samples 66
3.3.7 XRD Analyses 67
3.3.8 Kinetic Studies 67
19
CHAPTER 4
RESULTS
4.1 Sampling and Preparation of Analytical Sample 68
4.2 Microscopic Analyses of Samples 69
4.3 Chemical Analysis 70
4.4 Chemical Processing 71
4.4.1 Direct Leaching 71
4.4.2 Roasting before Leaching 71
4.4.2.1 Simple Roasting 72
4.4.2.2 Roasting with Oxidant 79
4.4.2.3 Roasting with Novel Oxidant 81
4.4.2.4 Roasting with Ammonium Compound (Novel Additives) 86
4..4.2.4.1 Two-Stage Roasting 91
4.4.2.4.2 Effect of Combined Oxidant/Additive on the Copper Leaching 95
4.4.2.4.3 Leaching of Copper Chalcopyrite Concentrate 100
4.4.2.4.4 Amount of Sulfur Analyzed in Different Samples 101
102
4.5 XRD Analyses and Microphotography 113
4.6 Kinetic Studies 114
4.6.1 Mechanism of Reaction 116
CHAPTER 5
DISCUSSION AND CONCLUSIONS 118
Discussion 118
Conclusion 124
BIBLIOGRAPHY 125
Appendix A 141
Appendix B
Appendix C Economic Utilization
148
154
Appendix D Patents, Publications 155
20
LIST OF TABLES
Table. No Title of Tables Page
Table 1 Base Metals in Copper Ore Samples 70
Table 2 Effect of Simple Roasting on Copper Leaching in Sample ‘C’ 73
Table 3 Effect of Roasting Time on Copper Leaching at Optimum
Temperature
74
Table 4 Effect of Leaching Time on Leaching of Simple Roasted Ore 75
Table 5 Effect of Sulfuric Acid Concentration on Leaching of Simple
Roasted Ore
76
Table 6 Effect of Solid Liquid Ratio on Leaching of Simple Roasted Ore 77
Table 7 Effect of Particle Size of Simple Roasted Ore on Copper Leaching 78
Table 8 Effect of Oxidant during Roasting on Copper Leaching 80
Table 9 Results of Lower Temperature Oxidant Roasting for long time
period
80
Table 10 Effect of Novel Oxidant and Roasting Time on Copper Leaching 82
Table 11 Effect of Amount of Novel Oxidant Roasting on Copper Leaching 82
Table 12 Effect of Temperature and Time on Novel Oxidant Roasting 83
Table 13 Effect of Particle Size in Novel Oxidant Roasting 84
Table 14 Effect of Different Acids with Different Time on Novel Oxidant
Roasting
85
Table 15 Effect of Novel Additive during Roasting on Copper Leaching 87
Table 16 Effect of Roasting Time on Novel Additive Roasting 88
Table 17 Effect of Leaching Time on Novel Additive Roasting 88
Table 18 Effect of Particle Size on Novel Additive Roasting 89
Table 19 Effect of Temperature and Time on Novel Additive Roasting 90
Table 20 Effect of Two-Stage Roasting on Copper Leaching 91
Table 21 Effect of Particle Size during Two-Stage Roasting 92
Table 22 Effect of Temperature during Two-stage Roasting 93
21
Table 23 Comparison of Copper Leaching at Different Roasting Treatments 94
Table 24 Effect of Combined Oxidant/`Additives on Copper Leaching 96
Table 25 Combinations of Additive at Different Temperatures’ Conditions 97
Table 26 Effect of Particle Size on Roasting with [sample C + NH4Cl +
(NH4)2SO4 ]
98
Table 27 Comparison of Roasting Treatments with Additive NH4Cl and
[NH4Cl + (NH4)2SO4]
99
Table 28 Leaching of Copper Chalcopyrite Concentrate 100
Table 29 Amount of Sulfur Analyzed in Different Samples 101
Table 30 Effect of Time & H2SO4 Concentration on Copper Leaching 104
Table 31 Effect of Liquid Solid Ratio during Sulfuric Acid Leaching 105
Table 32 Effect of Heat during Sulfuric Acid Leaching 106
Table 33 Effect of Time during Sulfuric Acid Leaching under Pressure and
Temperature
107
Table 34 Effect of Sulfuric Acid Concentration under Pressure and
Temperature
108
Table 35 Effect of Oxidizing Acid during Sulfuric Acid Leaching under
Pressure and Temperature
109
Table 36 Effect of Solid Liquid Ratio during Sulfuric Acid Leaching under
Pressure and Temperature
110
Table 37 Effect of Particle Size during Sulfuric Acid Leaching under
Pressure and Temperature
111
Table 38 Effect of Temperature and Pressure during Sulfuric Acid Leaching 112
Table 39 Reaction Mechanism Determining Equations 115
Table 40 Comparison of Activation Energy by using Different Model of
Kinetics
116
Table 41 m-Values for Different Roasting Mixtures to Determine Reaction
Mechanism
117
22
ABSTRACT
The present study based on extraction of copper from indigenous sources in Darosh area
Chitral Khyber Pukhtun Khwa (KPK) and copper concentrate from Saindak (Bluchistan)
of Pakistan was conducted by using hydrometallurgical techniques. The convenient
representative sampling involved collection of high grade ores selected by hand sorting
for studies. The samples were tested by different techniques involving chemical analyses,
roasting with and without oxidants/ additives prior to leaching. The chemical analyses
have shown that on an average these ore samples contained 20 percent Copper, 2.01
percent Iron, 1.54 percent Zinc, 0.074 percent Lead, 2.35 ppm Silver, 1.0 ppm Gold and
20 percent Sulfur. The XRD analysis of the samples has shown the presence of dominant
Chalcopyrite and subordinate amount of Chalcocite, Azurite, Tetrahedrite, Malachite as
other copper minerals. The dominant accessory/ gangue minerals are muscovite and other
carbonates, silicates.
A suitable and cheap method to process the sulfide copper containing chalcopyrite ore
was investigated by thermo-chemical processes. The higher amount of copper leaching
was 70 to 75 percent achieved by prior roasting at 650oC. The effect of various oxidants
to increase copper leaching was also investigated during roasting of sulfide ores sample
at 5000C to 6500C. The oxidants used for the purpose were sodium per sulfate, sodium
chlorate, potassium chlorate, sodium nitrate, and sodium nitrite. The copper leaching was
only 70 percent achieved by roasting with sodium per sulfate at 5000C.
The new method was developed for economic utilization of copper indigenous ores. The
research introducing a series of novel additives used first time in such roasting. The
23
additives have properties of ions and vapors such as nitrate, chlorate, sulfate, phosphate,
acetate, oxalate, carbonate, and ammonia. Some of these ionic species were used only in
aqueous medium as reported in literature. The additive’s dissociation products in the
form of gas or vapors were investigated during roasting in dry and solid medium in
present study.
The novel processes developed envisaged that the vapor-solid metallurgy of complex
copper sulfide ores changes the mineralogy into leachable minerals. The additive roast
leach processes gave about 85 percent to 98 percent copper leaching with low range of
roasting temperature and roasting time. The XRD analysis of additive roasted sample has
shown the formation of sulfur rich mineral of copper as Cubanite. It is prove that the
additive roast leach processes have friendly environment by reducing sulfur dioxide
emission.
The combined effect of additives in two - stage roasting at lower range of temperatures
gave better results with reducing roasting time. The best combination of additive’s
products found was chloride – sulfate of ammonium. This combination is also better in
aqueous medium of direct leach processes as reported in literature. The additive roast
leach processes gave higher percentage of copper leaching with 2.5 percent sulfuric acid
in 1:5 solid liquid ratio within very short time.
The kinetics of roasting processes was studied. The roasting results were compared with
the help of kinetic models of Sharp (1966) to investigate the mechanism of reactions. It
was found the reactions mechanism during the roasting ranges from boundary phase
reaction to diffusion reaction. It was found that roasting with additives has decreased the
activation energy in order to chloride > chloride-sulfate > nitrate > sulfate.
24
CHAPTER 1
INTRODUCTION
The man has been searching facilities and methods instinctively with the divine guidance
for his livelihood since prehistoric times. He used metal for his needs in different ways
and he processed natural resources i.e. used ores to extract metal value. Till now the
technological changes are continuing in the discipline of extractive metallurgy. In the
quest of methods of extraction, the main focus is now shifted to reduce the extent of
damage to the environment. The increasing pressures are to limit the release of toxic
gases, such as sulfur dioxide to the atmosphere; safe disposal of the waste products and
development of low energy consuming processes (Ritchie, 1998).
The world’s economy directly ‘depends’ on the copper, due to its useful physical and
chemical features. It stands third in the context of the world consumption after steel and
aluminum. Although the origin of copper is not exactly known to the human race but it is
estimated that it was discovered in around 9000 BC in the Middle East. The respective
era of history when it was extensively used is named as the Bronze Age 2500 BC-600 BC
(CRN, 2007).
The technologies that are newly emerging are emphasizing on low energy consuming
processes as well as reduction of concomitant emissions. For adjusting to environmental
pressures, the processing of copper ores may be using solutions chemistry to extract
metal values rather than the traditional pyrometallurgy.
25
Men learnt to smelt copper in 3500 BC (Lichtenhan, 2004) and the smelting processes
that are used to refine copper are very old dating around 4500 BC. Present-day Israel,
Egypt and Jordan were some of the earliest locations of copper smelting (Stanczak,
2005). In traditional techniques of pyrometallurgy in which copper ore is mined, crushed,
ground, concentrated, smelted and refined, these processes of copper ores were extremely
energy intensive. For the beneficiation (concentration) purpose copper ore must be
sulfide in nature for efficient flotation. The conventional copper extraction processes
produced the impure molten metallic copper, molten iron oxide and gaseous sulfur
dioxide. The impure copper is then purified by electrolytic purification to pure copper
(Dresher, 2001).
Hydrometallurgy was started to separate the metals from its alloys in China in Bronze
Age (Stanczak, 2005). The hydrometallurgy commonly known as leaching-solvent
extraction-electrowining (L-SX-EW) process was widely adopted in the mid 1980s. This
new copper technology utilizes smelter acids to produce copper from oxidized ores and
mine wastes. The worldwide production of all copper produced by these processes
amounts to 1/5th (Dresher, 2001). These processes are said to have very little
environmental impacts and also cost effective. The hydrometallurgical processes extract
copper from its ores is therefore also more environmental friendly as compared with the
conventional smelting process.
The hydrometallurgical route that employs chemical extraction process (leaching) is
inevitable in certain ore types and offers an alternative to traditional smelting. Leaching
was possible in mid 1885 and remains so today in certain ores due to their unique
mineralogy and applied to ores having both oxides and sulfides copper mineralization
26
(Dresher, 2001a). The different types of leaching are (1) In Situ Leaching, (2) Heap
Leaching, (3) Dump Leaching, (4) Percolation or vat Leaching, (5) Agitation Leaching,
(6) Bacterial Leaching, (7) Acid Leaching, (8) Pressure Leaching, (9) Catalyzed
Leaching, (10) Ammonia Leaching, (11) Chloride Leaching, (12) Sulfate Leaching, (13)
Chloride-Sulfate Leaching and (14) Roast- leach process.
The oxide ores of copper are processed chemically in aqueous solutions of sulfuric acid
usually taking very short time. The leaching of oxide minerals is relatively easy and relies
on dump, vat and agitation leaching using acid solutions. On the other hand sulfide ore
minerals of copper require either pressure (autoclave) or bacterial leaching or heap
leaching or in-situ leaching by using oxidizing media and require much longer time.
These sulfide ores need time consuming long term processes. However the different
copper minerals may need different conditions to extract the copper values.
The copper sulfide ores are treated directly with strong oxidants (ferric chloride) or
subjected to pressure leaching to obtain acceptable leaching. The negative aspects of
direct leaching of chalcopyrite (Padilla, 2003) are that the copper is not extracted
selectively over iron, and the precipitation and disposal of the iron from the leach
solutions represents not only an economic burden for the processes but a pollution hazard
as well. On the other hand roast-leach processes, as pretreatment of chalcopyrite or
copper complex sulfide ores or concentrates, transform the mineralogy to amenable
leaching species.
There are different routes in hydrometallurgical techniques employing prior roasting
before leaching of copper sulfide ores. Examples are roasting with elemental sulfur
(Padilla, 2003) and with sulfur dioxide and oxygen (Huggins, 1977. US-Patent), (Jaquay,
27
1979. US-patent), (Ethem, 1974. US-patent), (Theodore, 1978. US-patent). The other
example is chlorination during roasting with salts or gas (Kanari, 2001), (Kadyrov, 2005.
RU-patent), (Dubrovsky, 1992. US-patent), (Agnew, 1976. US-patent). In some
processes sulfatization and chlorination was carried out simultaneously during the
roasting process (Lippert, 1976. US-patent), (Theodore, 1978. US-patent). Some of the
non-polluting hydrometallurgical processes are based on sulfatizing the copper sulfide
complex ores or concentrates (Minic, 2005), (Norrgran, 1985. US-patent).
The sulfides of copper may be of different nature. The methods for the copper
extraction from refractory sulfide copper ores, ranges from pretreatment of the ores by
oxidative roast, autoclave oxidization or bio-oxidization, chloride leaching, sulfate
leaching and chloride-sulfate leaching which have been studied with a certain degree of
success. The oxidization roast leaching is now termed as the first choice followed
secondly by the temperature pressure leaching (Prasad et al., 1998). To avoid the
emission problems of sulfur dioxide during roasting treatments, attempts have been made
to search for a new pretreatment technique, aimed at sulfur fixation during the roasting of
refractory sulfide copper ores. Barlett, Haung and Haver, Wong initially introduced this
concept in the 1970s to treat the copper sulfide concentrates (Liu, 2000). For the purposes
of fixing sulfur many oxidants as fixing agents were used in the early studies, lime was
also used as the fixing agent, scavenging sulfur into calcium sulfate. Other fixing agents
such as sodium, potassium, calcium, ferrous salts and iron oxide were also tested in
different processes. It was reported that 90-95 % sulfur has been fixed (Liu, 2000).
Norrgran (1985) patented a process for the sulfatization of non-ferrous metal sulfides in
which he used two-stage temperature roasting. First roast of 140 minutes at 420oC and
28
then 210 minutes roasting at 610oC with fixing agent as sodium carbonate. This was
followed by agitation leaching for two hours in water.
Some Leaching Plants of the most common world class processes and plants for
hydrometallurgical extraction of copper from copper sulfide ore or concentrates are as
follows:
MT Gordon Plant by Dreisinger et al. (2002), Activox process by Corrans et al.
(1993), CESL Process by Jones et al. (1998), Dynatec Process by Collins et al. (1998),
NSC Pressure Oxidation Process, Chloride/Sulfuric Acid Leaching Process and High
Temperature Pressure Oxidation by Taylor (2007), Nena Tec Process by Horan et al.
(1996), Intec Process by Everett (1994, 1997), Canmet Process, Elkem Process and
Dextec Process by Zoppi (2000. US-patent), Cymet Process by McNamara et al. (1978),
Noranda Antlerite Process by Stanley et al. (1982), BHAS Process by Lal et al. (1987),
Dyuval Clear Process by Schweiter et al. (1982) and Cuprex Process by Zunkel (1993).
All these processes are related to chloride hydrometallurgy and are very complex
and have many steps. Special equipment is used to resist corrosion. The final copper
product is further processed.
The aim of the present research carried out in this thesis is to investigate the
possibilities of recovering copper from indigenous mono-mineralic copper ores as well as
indigenous complex ores by suitable techno-economical and environmental friendly
method such as by sulfur fixation hydrometallurgical techniques. The principal
advantages for this research would be the avoidance of sulfur dioxide emission and or
solution of H2SO4 marketing problems associated with traditional smelter and roaster
operations.
29
Taylor et al. and Peacey et al. (2007) reviewed in detail the commercial processes related
to leaching plants and chemistry involved in the plant processes. There is very little work
in their review on roasting prior (or pretreatment) to leaching in the hydrometallurgical
techniques. In the present work detailed study of the roast leach processes was
undertaken, which may be applicable to sulfides or refractory minerals. These processes
include the oxidation, chlorination, sulfation, sulfidization that can be performed
separately or in different combinations. The prior roasting or activation of copper sulfides
at low temperature, which may be possible by using activator or additive, is thought to be
very useful approach to extract copper. The pretreatment processes of copper complex
sulfides is recently reported by (Padilla, 2003), (Kadyrov, 2005), and (Kanari, 2001),
(few examples are presented also in Chapter 2).
The present work of leaching by prior roast leach process may also avoid the
corrosion encountered in chloride hydrometallurgy and the emission problems associated
with pyrometallurgy. Different additives as activator would be studied to reduce the
roasting temperature and to enhance the leaching rates in a single step to avoid multi-
stage processes. By this roast leach process, there would be no need of further pressure
leach as reported in (Akcil, 2002) which shown 85% extraction. The research reported in
present thesis gives 85-96% extraction of copper without autoclave pressure leaching.
1.1. Copper Ore Deposits in Pakistan
The copper ore deposits are found in different areas of Pakistan. Among these the
important copper ore deposits are at Saindak, Dalbandin and Chilghzi in Baluchistan and
Dir, Chitral, Waziristan in N.W.F.P (KPK). The ore reserves of most of these deposits are
30
not fully estimated except in Saindak. The exploration work on these deposits was
conducted by various mineral-developing agencies to prove or estimate the reserves in
these areas. The mineralogy and chemical composition of these minerals are oxides,
sulfides and complex sulfides. The deposits are of low to high grade, porphyry, vein type,
hydrothermal and complex in mineralogical nature. The main mineralization of copper
ores in Pakistan is porphyry type, hydrothermal, scarn and subduction related ores in
which the mineralogical composition varies according to the deposition. These deposits
are having mineralogy of sulfide, chalcopyrite and complex minerals; containing copper,
lead, zinc, iron, silver and gold assemblage, such as tetrahedrite and boulangrite. The
exploration work in Pakistan in addition to Saindak area has also shown potential
deposits of copper. No systematic work on these ores has so far been done on the
evaluation, beneficiation and utilization of locally available ores. The oxidized and low-
grade copper ores are mainly unutilized. Developments of economic processes for the
economic utilization of these ores still need investigation (Khan, 2004).
The copper ores samples collected for the study is of different composition i.e.
low or high grade and oxidized or sulfides ores. For mineralogical viewpoint the oxidized
ores contain Malachite, Azurite and Bornite while the sulfide bearing ores contain
dominantly Chalcopyrite, Tetrahedrite and Pyrite.
1.2. Scope of the Thesis Work
The present study was carried out on the minerals samples from indigenous ores deposits
mainly using different chemical leaching techniques for copper extraction. Sample
preparation and comminution study was carried out in PCSIR laboratories Lahore. The
31
extraction methods were studied and developed for extraction of copper in the present
work on indigenous copper ores using different approaches. Two types of
hydrometallurgical techniques such as direct leaching and indirect leaching were
primarily used.
The direct leaching on indigenous copper ore samples from Chitral, Dir and
Saindak were conducted. The leaching parameters such as temperature, pressure,
oxidizing agent and acid concentration were also undertaken.
The indirect leaching was carried out mainly on copper sulfide ores, since it is
known that the oxides minerals were leached easily in acid media whereas the sulfide ore
minerals were not leached easily in acid media without oxidizing agent or at normal
temperature pressure condition. The alternative leaching processes have been developed
for extraction of copper from sulfide ores. The chalcopyrite copper especially presents
problems (Pedilla, 2003) related to the great resistance of this mineral to oxidation during
leaching. On the other hand, the pretreatment of copper sulfide ores and chalcopyrite
concentrate to transform its mineralogy to other species were found more amenable to
leaching. Literature has reported that preheating mixture of copper sulfides and use of
different additives, activators, catalysts, oxidants and reductants at different ranges of
temperature. The subsequent leaching of pretreated ores was performed in various
leaching media. Another technique of hydrometallurgy reported in the literature is prior
sulfatization roasting subsequent to leaching. The process of fixing sulfur of copper
sulfide ore using additives (Norrgan, 1985) during roasting appears a better approach.
This technique was applied to sulfide type ores with different roasting conditions. New
32
and economic collections of additives as activators for sulfatization roasting of copper
sulfide ores were also investigated.
The roasting parameters as temperature, roasting time, effect of additives,
amount of additives were optimized in the present study. The parameters as leaching
time, acid concentration and solid liquid ratio during subsequent leaching were also
studied. The purpose of the present work was to achieve lower temperature of roasting by
using additives in addition to achieve greater yield of copper extraction. The extraction of
copper at high leaching rate was achieved by a novel method. Some activators and
oxidants having catalyst-like effect during roasting were studied to achieve low
temperature in roasting. The two-stage roasting at various temperatures treatments
showed better results, which has a potential to develop as a process. This process may
reduce the roasting time to half of the conventional roasting.
The XRD analyses of sample before and after treatments are presented to show
the changes in the mineralogy. For the analysis of copper and other base metal Atomic
Absorption Spectrometer Z8000 Hitachi was used. The physical changes in texture and
apparent colors of the ore after thermal treatment were observed by microscope. It is
hoped that the result of the present research could be applied to utilize the small, low
grade, complex oxide and sulfide ores by relatively environmentally inert process, which
are not possible by conventional pyrometallurgical techniques.
The present work would have the advantages over the simple hydrometallurgical
techniques performed previously by different workers mainly in solution media. The
present work undertook solid-state reactions of ore and additives to alter the mineralogy
33
so that the roasted product is easily leached. There is very little work of this type reported
in the literature utilizing different additives during roasting of ores.
The chemical extraction methods developed can also be applied to other types of
ores such as complex sulfides ore and on low-grade ore deposits. Such deposits in
Pakistan are mainly unutilized due to absence of work on extraction processes. The need
of today is developing technologies, which are cheap and environmental friendly. Based
on the investigation in the thesis, the author had published a patent in Pakistan. A patent
has been filed on a cheap, environmental friendly and economical process in US patents.
The material processed by roasting process using activators/ additives is also found to
leach instantaneously with little consumption of acid.
34
1.3. Significance of the Research Work
The research will contribute to provide new information and data in the areas of
application of hydrometallurgy for indigenous copper ores utilizing prior roasting with
and without additives (roasting agents), which are not so far investigated. The study
would have importance in the following areas.
The utilization of a number of small deposits and low grade indigenous copper
ores in Pakistan is not being done due to absence of hydrometallurgical and chemical
production technology. The work has immense economic importance, which has been
neglected for a long time with no proper exploration.
The recovery of copper from the ore is increased by using hydrometallurgical
processes, because these processes are time saving and eliminate many steps from
conventional metallurgical processes such as concentration by mineral processing and
pyrometallurgy. The cut off grade in processing is about 0.1-0.2 whereas in
hydrometallurgical processes even the waste can also be processed.
The chemical and metallurgical process studies will be important academically
and also have economic potential. From review of literature it is found that most of the
work was done for extraction of copper by acid leaching with H2SO4, but roasting solid
state reactions with additives are not studied in detailed. Small scale processes based on
hydrometallurgical techniques are economical as these could be based on locally
abundantly available acids for leaching reactions.
35
1.4. Objectives of the Study
1. To study the suitable hydrometallurgical extraction processes of copper.
2. To investigate the optimum conditions of leaching with different reagents in order
to find maximum leachability of the ore.
3. To study the kinetics of developed processes such as prior roasting processes.
4. To study the conventional sulfuric acid leaching processes on local ores and the
effect of temperature, pressure, concentration and catalysts on the rate of reaction.
36
CHAPTER 2
REVIEW OF LITERATURE
The main focus of reviewing of related literature was to find the cheapest and novel
approaches in hydrometallurgical techniques. Different sources of literature studied were
books, journals, online Journals and magazines related to extractive metallurgy,
pyrometallurgy and hydrometallurgy in the libraries of Universities and Pakistan Council
Scientific & Industrial Research (PCSIR) Lahore. The related literature was also
collected from Pakistan Scientific & Technological Information Center (PASTIC)
Lahore. The traditional, classic, conventional and recent processes of the extractive
metallurgies were studied and analyzed. It is impossible to separately distinguish the
method and processes related to hydrometallurgical techniques. Because the addition of
new processes and the combination of old and new processes in the hydrometallurgical
fields are now inter convert terms.
Copper occurs in nature in the form of minerals as oxides, sulfide, silicate and native
form. There are two main types of copper minerals, sulfides that have not been oxidized
in nature and oxides and other minerals that have been oxidized and have little or no
sulfur. The sulfide copper minerals are bornite, chalcocite, chalcopyrite, pyrite and
covellite. Oxide copper minerals include azurite, chrysocolla, cuprite and malachite.
Many copper ores and rocks also have molybdenum, lead, zinc, gold, platinum, nickel,
iron and silver.
37
The following purposes were kept in focus to review the literature.
1- Importance of hydrometallurgy
2- Types of leaching
3- World class hydrometallurgical leaching plants
4- Patented hydrometallurgical processes
5- Reagents used in leaching processes
6- Solvent extraction
7- Kinetic and thermodynamic studies
2.1. Importance of Hydrometallurgy
The reviews give the importance of hydrometallurgy over the other conventional,
classic and traditional extractive metallurgies. According to Eltringham (1997) the ideas
and methods of technology of today’s copper mining were introduced as early as 1865.
The open pit mining, flotation, concentration and the reverberatory smelter for
porphyry copper ores were innovative technologies for copper industry (Biswas et al.
1980). The basic methods of copper production have remained same for 65 years except
leaching- solvent extraction- electrowining (L-SX-EW). The technological innovation in
the copper industry has consisted largely to exploit lower grade ores and continually
reduce the costs of production.
The more efficient, automated equipment have been created to maintain profits
(Eltringham, 1997) and the oxide deposits have been utilized to maximize the production
of copper. Eltringham (1997) proposed that the future success of copper mining lies in
38
the alternative methods of copper extraction as; (i) in situ leaching, (ii) new
hydrometallurgical routes, (iii) mini-smelters.
Yazawa (2003) studied the trends of extractive metallurgy of copper by pyro and
hydrometallurgical methods. Boriskov (2002) studied general characteristics of the metal
recovery from raw materials by hydrometallurgical method. The developments of
hydrometallurgical process were discussed, introduced new technologies and compared
hydrometallurgical recovery of metal with other technologies.
Davenport (1999) postulated that the biggest possible change over the next twenty years
would be complete replacement of smelting / converting by hydrometallurgical
processing. Reviews from the Milton E. Wadsworth International Symposium on
Hydrometallurgy, Parkinson (1993) featured new hydrometallurgical processes that avoid
the emission problems of pyrometallurgical routes from copper and other ores.
The pyrometallurgical processes lack environmental concerns, i.e. generation of
harmful SO2 gas and high energy consumption to recover copper from sulfide copper
ores. The authors directed their attention to the environmental aspect and energy saving,
at copper recovery by using the more advantageous L-SX-EW methods. These can be
designed as closed system and lessen the wastes from the process (Shibata et al. 2000).
Hydrometallurgical methods (L-SX-EW) eliminate many steps in the
conventional metallurgical processing. Hydrometallurgical routes for the extraction of
metals are stated as more simple, rapid, efficient and more economical. These routes have
become more popular to undertake research because of the following reasons; (a) energy
costs are lower, (b) requirements of zero discharge type conditions, (c) the impurities
39
become a problem as concentrating in the copper whether for air or water pollution
smelting processes must be leak proof, (d) hydrometallurgical processes can separate
impurities better, (e) the operating temperatures are lower, (f) sulfur dioxide emitted
during roasting and smelting causes severe air pollution, (g) acid formation to avoid the
sulfur dioxide gas emission.
The powder copper ore can be treated by two methods to extract copper;
1. Copper ore Leached Solution Copper
2. Copper ore Concentrate Leached Solution Copper
The above references show the importance and need of hydrometallurgical
Techniques. Following figure (U.S. 1988) shows both pyrometallurgical and
hydrometallurgical routes.
40
Figure 1. Steps of Pyrometallurgy and Hydrometallurgy
2.2. Types of Leaching
Chemical leaching is a process for chemical dissolution of the desired minerals in
aqueous solutions at different pH. It is possible to separate the compounds of different
metals due to the difference in the dissolution rates. Some oxidative reagents need to be
added to promote leaching. The valuable mineral and other associated minerals are
Comminution
Flotation
Smelting
Converting
Anode refining and casting
Leaching
Precipitation Solvent Extraction
Electrowining
Electro refining
Concentrates
(20.30% Cu)
.
Matte
(98..5% Cu)
Blister
(98.5+ % Cu)
Anodes
(99.5 % Cu)
Cathodes (99.99 + % Cu)
(U.S. 1988. Office of Technology Assessment)
Cement Copper
(85 – 90 % Cu)
Cathodes (99.99 + % Cu)
Pyrometallurgical Route
Sulfide ores
(0.5 – 2.0 % Cu)
Hydrometallurgical Route
Oxides and sulfide ores
(0.3 – 2.0 % Cu)
Pregnant leachate
(20 – 50 % Cu)
Copper electrolyte
(25 – 35 % Cu)
41
dissolved by an appropriate reagent under certain conditions e.g. pH, pressure,
temperature, catalyst. Leaching is known as extraction in the chemical processing
industry. Leaching has a variety of commercial applications, including separation of
metal from ore using acid.
Metallurgical application of leaching is widely used in extractive metallurgy since
many metals can form soluble salts in aqueous media. The leaching is easier to perform
and much less harmful as compared to pyrometallurgical operations because of no
gaseous pollution. The only drawback of leaching is lower chemical rate may be due to
low temperatures of the operation. Zies et al. (1916) studied the sulfides enrichment
products by using cupric sulfate as the enriching agent. The pyrite alters to covellite and
chalcocite. Chalcopyrite changes to covellite and chalcocite. Covellite alters to
chalcocite. Sphalerite and galena change first to covellite and subsequent to chalcocite.
There are a variety of leaching processes, usually classified by the types of
reagents used in the operation. The reagents required depend on the ores or pretreated
material to be processed. A simple acid leaching reaction can be explained by the copper
oxide leaching reaction;
CuO + H2SO4 → CuSO4 + H2O
In the above reaction CuO ore dissolves, forming soluble copper sulfate. In many
cases other reagents are used to leach copper oxides ores i.e. ammonia and ammonium
salts, mixture of acids. The leaching of the oxide material is relatively easy and relies on
dump, vat and agitation using acid solutions.
On the other hand leaching sulfide minerals whether in ore or concentrate requires
a chemical oxidizing agent such as ferric ions (Fe3+). These special ferric ions can be
42
generated by reactions with air. The oxidation can be assisted by either pressure (as in an
autoclave) or more commonly with bacteria. Leaching of sulfides is a more complex
process due to the refractory nature of sulfide ores. It is best described by the following
chemical process;
2CuS + O2 + 2H2SO4 → 2CuSO4 + 2H2O + 2So
This reaction proceeds at temperatures above the boiling point of water, thus
creating a vapor pressure inside the vessel and oxygen is injected under pressure. The
leaching depends on the ore particle size, leaching rate, ore composition, and subsequent
isolation and precipitation techniques.
Taylor (1996) discussed new processes reflect present trends as well as those that
existed in the 1970s. Several of these processes including the Escodida ammonia leach
process, the Intec process, pressure oxidation, bio-oxidation, the Cuprex process, and the
CANMET process were discussed.
Types of leaching described by various authors are;
In Situ Leaching, Heap Leaching, Acid Leaching, Dump Leaching, Agitation Leaching
Bacterial Leaching, Percolation or vat Leaching, Pressure Leaching, Catalyst Leaching
Roasting before Leaching.
Types of leaching described using different leaching agents are as;
Ammonia Leaching, Chloride Leaching, Sulfate Leaching, Chloride-Sulfate Leaching.
43
The more common processes of hydrometallurgy include following leaching
types are used.
2.2.1. Leaching with Sulfuric Acid
A variety of leaching reactions with sulfuric acid under different conditions and
parameters reported were in literature. Hussain et al. (1975) investigated agitation
leaching of Saindak oxidized copper ore using 10-100g/l sulfuric acid and studied the
effect of particle size, acid consumption and liquid-solid ratio. Wilkomirsky et al. (1984)
roasted low grade Cu-Mo concentrates at 650-720oC with 20-60% excess air, followed by
leaching with 2-10% H2SO4 and extracted with a tert-aliph.amine. The molybdenum was
precipitated as ammonium molybdate and copper is recovered by electrolysis or
precipitated with iron. The recoveries are 96% for copper and 84% for molybdenum.
Krushkol et al. (1987) worked on leaching of ores with H2SO4 solution containing
additives is improved and consumption of reagents is decreased, by using phenylamine as
the additive at 0.1-0.2g/l. A copper smelting slime was leached with H2SO4 solution in
the presence of an oxidant. The operation is terminated when the system redox potential
rises sharply (Takewaki et al. 1987).
The effect of temperature (75-97oC), pH (1.21-4.00), oxygen pressure (172-6200
kPa) on the kinetics of chalcopyrite and pyrite leaching by H2SO4 was studied (Lin et al.
1986). Ptitsyn et al. (1987) studied leaching with H2SO4 for four months at 5-35g/l was
investigated at -18 or +20o for copper ore of sulfide, oxidized or mixed types. Recovery
of copper from the mixed or oxidized ores was similar at both temperatures. Panin (2002)
44
leached complex sulfide-oxide ores of copper with sulfuric acid. The residual ore solids
were optimally milled for flotation. The copper recovered about 90% in mixed ore feed
as compared to the only 77.5% by conventional processing.
Medvedev (2002) studied the effect of H2SO4 concentration, temperature, slurry feed,
leaching time, stirring and particle size on copper leaching for the quarts rich ore feed
containing about 1.6% copper sulfide, about 2% copper basic sulfate, and about 3.6%
iron oxide and hydroxides. The maximum recovery of copper was obtained to be 83.8%
in acidic leaching for 6.5 hours at 95ºC temperature with initial sulfuric acid
concentration of 22g/l and the slurry solid liquid ratio of 1:7. Activation energy and order
of the reaction were also determined.
Kutokhov et al. (1988) developed a process for the recovery of copper from an
oxidized low-sulfide concentrates containing 21.8% Cu, 8.0% Fe, 32.1% SiO2, 4.4%
Al2O3, 7.8% S and other components by repeating flotation and leaching with H2SO4.
Mann (1987) described open-pit mining, heap-leaching with H2SO4 of copper ore and
electrowining of copper at the San Manuel mine (AZ, USA).
2.2.2. Dump and Heap leaching
Mc Gregor et al. (2000) described the principle of copper dump leaching had been
understood and applied in Europe and Asia for over 400 years. Copper heap leaching has
developed during the last 15 years. Jose et al. (2000) studied that by dump leaching in
Cananea had successfully improved copper recoveries.
45
Research using surfactant leaching aids to enhance copper recovery was also
discussed (Jenkins, 1994). Witt et al., Lima et al. and Uhire et al. (2000) discussed heap
leaching separately in proceedings of International Conference.
2.2.3. Pressure leaching
Hiroyoshi et al. (2000) studied that ferrous ions are more useful for leaching chalcopyrite
with oxidant ferric ions in sulfuric acid solutions in air at ambient temperatures.
Berexowsky et al. (1999) has developed sulfuric acid with pressure leaching process to
extract copper from the sulfide ore. Rani et al. (1998) conducted pressure and agitation
leaching studies for the extraction of copper and associated metals under varying
parameters by using H2SO4 in complex copper ores. The maximum recovery of copper
obtained at 100oC with 1.87 M H2SO4 and 0.5M HNO3 for 30 minutes.
The leaching process is reported in Cominco Engineering and Mining Journal
(Anonymous, 1998) based on an oxygen-enriched autoclave pressure-leaching operation
following conventional flotation. Carter (2003) had reported that Pressure leaching plant
is an alternative to conventional methods for processing concentrates.
Varga (2000) leached low grade Chalcocite-Covellite-Chalcopyrite ores from the
Cavnic mine in sulfide medium. The parameters such as leaching time, quantity of ferric
sulfate and oxygen as oxidizing agent were modified during the experimentation. Habashi
(2002) reviewed the development of hydrometallurgical technologies especially under
pressure and the latest achievement in performance of non-ferrous metal recovery from
various raw materials.
46
A hydrometallurgical process to extract copper was developed which involves a
sulfuric acid based two stages countercurrent leach process from the sulfide ore
(Berezowsky, 1999).
Dutra (2001) leached the Igarape Bahia gold ore was performed for the extraction of
copper. Temperature, stirring intensity, sulfuric acid concentration and particle size were
selected variables and analyzed.
Deng (2001) presented experimental data related to the effect of temperature,
time, sulfuric acid concentration, chloride additives and oxygen flow rate for leaching of
copper. The sodium chloride additives increased recovery of copper. Cole (2002)
performed chemical processing of metal from sulfide ore in a reactor with oxygen
injection and stirring the slurry. The process was found to be suitable for oxidation
leaching of sulfide ore concentrates for recovery of Gold, Copper, Nickel, Zinc, Lead,
Cobalt, Vanadium and Tungsten.
Marsden (2002) recovered copper from the sulfide ore or concentrate by
following three methods/steps (a) controlled fine grinding (b) pressure leaching at 160-
170ºC with aqueous sulfuric acid solution containing added surfactants or dispersing
agents and (c) conventional copper recovery by solvent extraction and electrowining. The
copper recovery in leaching was about 98% with H2SO4.
Bandyopadhyhay (2003) leached powdered sulfide especially chalcopyrite ore
concentrate with acidic ferric sulfate aqueous solution. The copper was leached in the
presence of silver ions at 2-10 microgram/ml of solution at atmospheric pressure and pH
of 1-2. The process was found to be suitable for leaching of sulfide ore containing 0.92-
3.68% copper. Imamura (2003) recovered copper by high pressure and temperature
47
leaching. The process comprised preparing concentrate from high grade copper ore
mainly containing chalcopyrite by flotation, leaching at high temperature and pressure to
obtain copper leachate containing Ferrous ions and sulfuric acid and then heap/vatt
leaching was performed. Leaching was carried out at 150-220ºC.
Pandey (2003) studied the effect of Sulfuric acid Pressure leaching parameters in
extraction of valuable metals from the sulfide concentrate of Juduguda. The metal was
accompanied by the formation of elemental sulfur that can be recovered as byproduct.
Metal recovery increased by increasing temperature and pressure.
Tsunekawa (2005) mixed the sulfide ore with carbon greater than 1% by weight
and leached with ferrous sulfate aqueous solution containing 1-20g/l iron, pH 1-2.5 and
redox potential of 350-450 mV. Shneerson (2003) processed the sulfide feed material
copper leaching in two steps with sulfuric acid in Autoclave pressurized with oxygen
containing gas. The 1st stage recovery was applied for the limited recovery into solution.
The 2nd stage leaching of the residue was performed by the sulfuric acid at the liquid solid
ratio of 1.5:1-10:1. The copper recovery was about 97-99.5%.
Brewer (2005) tested copper concentrate pressure leaching in laboratory and plant
was taken under Mineral and Metallurgical Processing (Eng.) Phelps Dodge operated a
pressure-leaching vessel for copper concentrate in Arizona. Huang et al. (2007) proposed
technique having four major steps; (1) the acidity adjustment of acidic pressure leaching
solutions, (2) solvent extraction separation of copper by organic reagents XD5640, then
(3) iron in raffinates is selectively removed by high temperature hydrolysis precipitation
in an autoclave, and (4) nickel and cobalt are selectively precipitated by Na2S from the
final solution after recovery of iron.
48
McDonald et al. (2007) compared the reaction condition promoted by various
companies as the Phelps Dodge Placer Dome and Activox processes. The aim was to
improve to understanding of the mechanism and practical issues for competing processes.
McDonald et al. (2007a) compared the reaction kinetic and products for the recovery of
copper from chalcopyrite concentrates under medium temperature conditions at 125-
150oC with in 30 minutes. It was concluded that chloride ion addition and high acid
concentration enhanced copper extraction kinetic and recovery and inhibited the
oxidation of sulfur to sulfate. Chloride ion enhanced the anodic oxidation of mineral
sulfides and the dispersion of molten sulfur.
Padilla et al. (2007) leached sulfidized chalcopyrite with sulfuric acid oxygen
pressure leaching. Oxygen partial pressure was the main variable that controls the
copper/iron selectivity of the leaching. An increase in oxygen partial pressure increased
significantly the rate of copper dissolution but deteriorated the copper and iron
selectivity.
2.2.4. Bioleaching / Bacterial Leaching
Shuey et al. (1999) described Bioleaching is the future for refractory mineral processing.
It is an environmentally friendly, bacterially assisted process that is useful for low grad
heap leaching operations. Acevedo et al. (1993) discussed the feasibility of applying
bacterial leaching with respect to a project involving leaching of copper ores in Chile.
The Bacterial Thin Layer leaching process (BTL) was used in Chile.
A detailed work was done on different types of bacterial leaching by Vasquez et
al. (1997), Kirpishchikov et al. (1982), Mier et al. (1996), Rawliings et al. (1995),
49
Grouddev et al. (1993), Ostrowski et al. (1993), Hazra et al. (1992), Zastrow et al.
(1991), Acevedo et al. (1989), Khalid et al. (1988), Litz (1993), Bruynesteyn (1986),
Agate et al. (1986), Lastra et al. (2002), Holmes (1999). The authors had discussed their
kinetics studies, effect of microorganism, particle size, nutrient, oxygen, temperature,
pressure, pH, organic growth factor, concentration of acid, heap, dump, in-situ leaching,
knowledge on the genetics, physiology and biochemistry of leaching bacteria, all
physical-chemical requirements in relation with process conditions for low, high grade,
oxidized, sulfide ores.
2.2.5. Catalyzed leaching
A method was described by Roche et al. (1987) when copper or uranium ores are crushed
to a particle size that promotes dissolution of >75% of the Uranium, the resulting power
is formed into pellets using a 1st liquor and the pellets are leached with an acid solution
compatible with the 1st liquor. The method is useful for accelerated leaching of copper or
uranium ore.
Seeger (1999) had shown that the oxidation of copper sulfide ore concentrates can
be accelerated in the presence nitrous oxide (NOX) gases as the reaction catalysts,
resulting in the exothermic reaction with formation of sulfur and sulfates. This process is
applied to the catalytic ore leaching in aqueous H2SO4 near normal pressure and ~110o in
a closed container, adding either O2 or air. Investment costs are small compared with
those of the pyrometallurgical processes.
Anderson (2003) had demonstrated a process of nitrogen species catalyze in the
oxidizing pressure-leach system to produce copper via solvent extraction / electrowining.
50
Carranza et al. (2004) leached copper sulfides in two stages with ferric leaching and in
first stage ferric leaching and the unattacked chalcopyrite was leached with silver as
catalyst in 20 hours in second stage.
A process was invented for copper leaching from copper sulfides concentrates.
The pyrite used as catalyst in leaching of chalcopyrite for ferric reduction in order to
eliminate the passiveness of the chalcopyrite surface. In the process maintaining
operating solution potential at suitable level does not materially oxidize the pyrite. The
leaching is carried out in sulfuric acid by oxidization in four hours (Dixon et al. 2005,
US-patent). Okamaete (2005) investigated catalytic effect of six different activated
carbon and coal on the chalcopyrite leaching in sulfuric acid solution at 30oC under
atmospheric. It was noted that except for one activated carbon sample activated carbon
addition enhanced copper extraction but the effect increased by the rank of the coal.
Lorentzen et al. (1996) published a microwave-heated EPA method 3050B for the
leaching of key elements (cadmium, chromium, copper, lead, nickel, zinc) of
environmental importance was tested and compared to conventional hot plate-heated
EPA method 3050B.
2.2.6. Roast-Leach Process
The literature reported that without pretreatment and activation, complex copper sulfides
and chalcopyrite ore has not been leached. The prior roasting at high temperatures
(Bafubiandi et al. 2005) or roasting with additives at low temperatures (Kadyrov et al.
2005, RU-Patent), (Agnew, 1976. US-Patent) has been conducted in literature. The
structure of all type of sulfide minerals is deformed by these activities. The new
51
amenable to leaching species are developed by these pretreatments or roasting activities.
The studies verify Evrard (2001) mineralogical observations that the Chalcopyrite
decomposed due to the formation of discrete particles of Cu2S with a size-range from five
to 20 or more micrometers by controlled desulfurization process at relative low
temperatures (650oC). During roasting the chemistry of chalcopyrite changed. Deer
(1967) explained the X-rays study by Hiller and Probsthain (1956). Heating without
controlled sulfur pressure chalcopyrite passes into a high temperature phase at 550oC.
This shows that the heat treatment of chalcopyrite should be at low temperature less than
550oC for deformation of structure. Deer (1967) stated that when chalcopyrite heated in
air it shows an exothermic DTA peak at 500oC (McLaughlin, 1957; Levy, 1958). A very
few examples of roast-leach process were found and reviewed.
Ziyadanogullari et al. (1999) roasted the oxidized copper ore and converter slag in
different ratios at 600oC for six hours. The amounts 84% copper and 70% cobalt were
determined. The oxidized copper ores, converter slag, FeS and H2SO4 were reacted under
autoclave condition and the solid samples were roasted for sulfatization.
Prasad et al. (1998) reviewed and summaries the work carried out on oxidative
roasting of chalcopyrite and related constituent such as copper sulfides and iron sulfides.
The sulfation roasting process is it suitability for smaller, pocket and complex deposits
with less air pollution due to fixation of sulfur with metal. Parker et al. (1981)
experimented to copper sulfides concentrates on a laboratory scales and oxidative roast at
800-900oC to remove sulfur and reduction of calcine. The dissolution in acetonitrile-
water was achieved.
52
Bafubiandi et al. (2005) used a route of the concentrates-roast-leach-electrowining
process in the hydrometallurgical extraction of zinc and copper from a sulfidic ores.
Optimal roasting conditions were found to be 900oC for three hours and obtained XRD
data, leaching was done with HCl, H2SO4 and HNO3. More then 80% of both were
recovered.
Kanari (2001) investigated the chlorination of two chalcopyrite concentrates with
Cl2 + N2 wider isothermal conditions in temperature range of 20-750oC using boat
experiments. SEM, XRD and chemical analysis were used. The reaction of chlorine with
sulfides is almost completed at about 300oC and over all reaction is exothermic. The
metal chlorides were concentrated and iron and sulfur were volatilized. It is an alternative
route for sulfides treatment without SO(x) emission.
Akcil (2002) used a process of roasting and pressure leaching is an alternative
process. The best results obtained were with a pretreatment by roasting followed by acid
leaching is an autoclave system. The extraction of copper achieved was over 85%.
Nakazawa et al. (1999) developed a modified free energy minimization method to study
the behavior of arsenic during the partial roasting of copper concentrates.
The sulfide minerals are converted into water soluble sulfates by roasting. The
mechanism involved in converting sulfides to sulfates proceeds via an oxide as follows:
MS + 3/2O2 MO + SO2
SO2 + 1/2O2 SO3
MO + SO3 MSO4
M stands for metal.
2CuFeS2 + 13/2O2 2CuO + Fe2O3 + 4SO2
53
CuO + SO3 CuSO4
(Norrgran, Pat. No: US4541993)
2.3. World Class Hydrometallurgical Leaching Plants
The chemical processes and leaching reagents used in the most common world class
hydrometallurgical leaching plants commercially for extraction of copper from copper
sulfide ores/concentrates are as follows;
2.3.1. MT Gordon Plant
Dreisinger et al. (2002) processed chalcocite ore directly. The chalcocite ore was treated
with ferric leach, low pressure oxidation process, after a moderate grind to 75-106
microns. The copper recovered by conventional SX/EW. It was tested in laboratory scale
experiments the processes also applicable to chalcopyrite with fine grinding. The
elemental sulfur caused complications. The process is not tested for precious metals.
2.3.2. Activox Process
Corrans et al. (1993) activated the mineral species. The process is low pressure oxidation
at 110oC and 1000 pka having ultra-fine grinding to 5-15 microns and copper recovered
via SX/EW. The process is applicable to chalcopyrite, chalcocite and gold. The elemental
sulfur is formed which leads difficulty to recover precious metals. The gypsum, goethite
and jarosite are by-products.
54
2.3.3. CESL Process
Jones et al. (1998) developed a process by Tec-Cominco. The process is an oxidizing
pressure leaching using a mixture of cupric sulfate and cupric chloride. The process is
catalyzed by chloride on medium pressure oxidation at about 150oC and 200 psi pressure.
It is regrind to P95 of 45 microns. It is avoided ultra fine grinding but chloride will
require suitable corrosion resistant materials. Besides these sulfur is good by-product.
2.3.4. Dynatec Process
Collins et al. (1998) used additives in leaching process. The process operates at medium
pressure of oxidation at 150oC after a fine grinding to 90% passing 25 microns for
chalcopyrite copper. For chalcocite the temperature is required to 100oC, which is less
than CESL process i. e. 150oC the coal is added to disperse the molten sulfur in dynatec
process. The elemental sulfur is recovered by melting and hot filtration.
2.3.5. NSC Pressure Oxidation Process
This process is based on moderate pressure oxidation at 125-155oC catalyzed with
nitrogen species supplied from sodium-nitrite with particle size of 80% passing 10
microns in autoclaves the process is operated in Sun shine plant and in Montana by Dr.
Corby Anderson (Taylor, 2007).
55
2.3.6. High Temperature Pressure Oxidation Process
The chalcopyrite copper concentrates are leached in high pressure oxidation at
temperature rate of 200-225oC. Sulfur is oxidized to sulfate. In this process no fine
grinding, no catalyst, no elemental sulfur in the autoclaves. The process is operated at in
pilot scale work in Guelb Moughrein Project in Mauritania (Taylor, 2007).
2.3.7. Nena Tec Process
Horan et al. (1996) patented a leaching process. The ferric sulfate leaching is involved in
this process at 80oC and atmospheric pressure with oxygen treatment to maintain an
adequate level of ferric. The feed is finally ground to 20 microns. The leaching
temperature is less than melting point of sulfur and elemental sulfur remains in residue,
which is hurdle in leaching of precious metals. The final side products are goethite,
gypsum and jarosite.
2.3.8. Intec Process
Everett (1994, 1997) presented innovative chemistry of the process is the main
components of the leaching solution are sodium chlorides and bromide and the lixiviant,
BrCl2 is produced by anode reaction in a unique electrowining cell. The process involves
a four stages counter current leach with chloride/bromide solution at 85oC and
atmospheric pressure and air as oxidants in 12 hours. The sulfur is formed in the
56
elemental state. The pilot plant is working in Sydney in 1996. It is very complicated
process.
2.3.9. Chloride/Sulfuric Acid Leaching Process
The leaching is processed with a sodium chloride and sulfuric acid solution with oxygen
treatment at 80-95oC and under atmospheric pressure in 24 hours. The chloride acts as a
catalyst for the oxidation reaction at a relatively coarse grind size (Taylor, 2007).
2.3.10. Canmet Process
The Canmet, Minemet, Recherche, Broken Hill processes combine leaching using CuCl2
at atmospheric pressure, with solvent extraction of copper from chloride solution and
electro-wining from sulfuric baths (Zoppi, 2000. US-patent).
2.3.11. Cymet Process
McNamara et al. (1978) subjected the copper sulfide concentrate to a two stages counter
current leach. The Cymet process (Patent No. US3901776) used a mixture of FeCl3 and
CuCl2 sodium chloride brine solution as the leaching medium. The CuCl obtained is
precipitated from the solution and reduced to copper metal by using a flow of hydrogen
in a fluid bed reactor. The iron was rejected from solution as jarosite and various hydrous
iron oxides during second stage.
2.3.12. Dyuval CLEAR Process
57
Schweiter et al. (1982) operated CLEAR process between 1976 and 1982 and leached
copper concentrate in two stages counter current system at atmospheric pressure. The
Dyuval clear process and the U.S.B.M. process also use leaching medium based on CuCl2
and FeCl3 to procedure a CuCl solution that is electrolyzed in a diaphragm cell. The
CuCl2 is regenerated in anodic compartment. The residue from the first leaching stage
was leached under 145oC and oxygen over pressure to recover the residual copper, using
ferric ion as the principal leaching agent. The copper is deposited in incoherent form is
removed continuously. Copper was deposited as powder in electrowining. The purity of
copper produced in the way is insufficient and further electrolytic purification is required.
The CLEAR plant was closed due to economic reasons and because the product could not
meet LME specification.
2.3.13. Dextec Process
The Dextec process causes anodic dissolution of the copper in the presence of oxidants.
The reaction takes place in a diaphragm cell. The copper sponge containing the impurities
of starting material is deposited in the cathode compartment (Zoppi, 2000. US- patent).
2.3.14. Elkem Process
The process leaches the mixed concentrate of copper with zinc and lead in a ferric and
cupric chloride solution, which operates in counter current. It complicated process and
the metals cannot sale directly without further processing (Zoppi, 2000. US-patent).
58
2.3.15. Cuprex Process
Zunkel (1993) leached copper sulfide concentrates with the solution of sodium chloride
with ferric chloride in two stages, obtaining a CuCl2 solution. The copper is recovered in
a form sponge from this complex process.
2.3.16. Noranda Antlerite Process
Stanley et al. (1982) developed a process for the treatment of chalcopyrite concentrates.
The conversion of chalcopyrite and sulfides copper ore to antlerite, hematite and
elemental sulfur by treatment of cupric sulfate-cupric chloride solution at135-145oC and
200 psi oxygen and leaching of antlerite residue with sulfur acid at pH 2.5 to produce
pregnant electrolyte for copper electrowining.
2.3.17. BHAS Process
Lal et al. (1987) developed a process for treating a copper lead sulfide matte with
oxygenated acid solution containing both sulfate and chloride. In the absence of chloride
the copper extraction was only 30% while it increased to 95% with chloride of more than
10g/l. silver is remained in leaching residue.
All these processes are related to hydrometallurgy and are very complex and
having many steps. Special equipments are used to resist corrosion. The final copper
product is further processed. Taylor et al. and Peacey et al. (2007) and Peters (1992)
reviewed in detail the leaching plants and chemistry involves in their leaching processes
and innovation in hydrometallurgy. But they missed and did not discuss very important
approach of prior roasting to leaching in the hydrometallurgical techniques. The roast-
59
leach process may be avoided the corrosion and emission problems, multi-stage
processes.
2.4. Patented Hydrometallurgical Processes
The patented processes of hydrometallurgy to recover copper from copper containing
materials/ores are dividing according to mode of process/method and use of the reagent
in following categories/ type. The review relate only on chemical/acid based leaching at
different physio-thermochemical conditions, not on bio, insitu, heap, dump-leaching
Roast-leach process
Ammonia leach process
Chloride based leach process
Chloride-Sulfate leach process
Counter-Current leach process
Temperature-pressure leach process
Common factors: Temperature pressure during leaching is common in most cases but
different reagents. The multi-steps leaching is involved in many patents. The sulfuric acid
leaching is mostly used under different parameters.
2.4.1. Roast-Leach Process
60
Huggins et al. (1977-01-18) treated the chalcopyrite ore with sulfur dioxide at 500oC
produced elemental sulfur, magnetite and copper sulfide. The sulfur is removed and
residual sulfides treated with concentrated sulfuric acid at 180-500oC to form copper
sulfate and sulfur dioxide, which is recycled, the solid magnetite roasted at 400-900oC
and leached with dilute sulfuric acid.
Swinkels et al. (1976-06-22) processed the ore/concentrates of sulfides or
thermally treated to activate by non-oxidizing gas at 650-800oC. The acid leach is
conducted as in two stages counter current leach. The ferrous iron in the filtrate from the
first stage is oxidized and hydrolyzed in presence of ammonia and ammonium or alkali
metal compounds for the precipitation of jarosite and production of sulfuric acid. The
residue subjected to an acid oxidization leach for copper sulfate solution.
Pepper et al. (1979-09-25) roasted the material at 620oC for 80% sulfur
elimination. Then roasted pyrite was subjected to an oxidation leach using chloride
solution at 145oC and 150psia.
Jaquay (1979-05-22) charged sulfide concentrate in high speed rotary converter
with oxygen and sulfur dioxide gases at 1000-1650oC, the metal were recovered and
sulfur gas exhausted or converted to sulfuric acid.
Dalvi et al. (1979-09-18) treated copper containing sulfidic material as dead roast
at temperature 750oC and the resulting calcine is leached in sulfuric acid solution at 60-
70oC in 2-3 hours.
Ethem et al. (1974-10-01) modified the Outokumpu process. The concentrate is
roasted in fluidized bed at 580-625oC with mixture of air and sulfur dioxide and then
water leached. The copper is recovered 50% after first roast as copper sulfate, then 85%
61
copper after second roast and 95% after third roast. 75% copper recovered after five
hours roasting.
Theodore (1978-08-29) subjected the material to molten salts NaSO4, NaCl,
KSO4, KCl, at 650 oC to 800 oC with O2 and SO2 in water. The mixture salts coating the
ore particles.
Rastas et al. (2001-08-14) recovered the metals by means of melt and melt coating
sulfation by thermal treatment under oxidizing conditions at 400oC to 800oC. The
recovery of metals completed in nine steps of process. The process is alkali metal sulfate-
sulfuric acid-water solution.
Kadyrov et al. (2005-01-20) used the raw material and ground to 0.2 mm blending
with halogen salts/oxygen salts, roasted at 450-560oC for 1-7 hours at high redox
potential leached with hot water.
Dubrovsky et al. (1992-04-14) chlorination and oxidation of refractory ores was
conducted. Ore and sodium chloride mixture was chlorinated at 450oC for one hour then
oxygenation was introduced for one hour. Then the power was flushed with nitrogen after
cooling leaching with HCl at pH 2.0.
The leaching was done after sulfatizing roasting. The residue subjected to
chlorinating sulfatizing roasting, and leaching the roasted residue with selective solvents
for base metals and precious metals (Lippert et al. 1976-07-13).
Norrgran (1986-09-02) mixed concentrate with sodium carbonate or bi carbonate
and roasted for the sulfatization. The roasted product is mixed with water to separate
these soluble sulfates.
62
Agnew (1976-09-21) explained a method for oxidation of fine divided sulfide
ores or mattes containing iron with one or more elements in the presence of iron chloride
in oxygen containing at atmosphere at a temperature of 220-450 oC during roasting phase
Ralph (1931-09-22) experimented to a specific refractory ore containing silver
and number of base metal and metalloids. The purpose of roasting is first to eliminate the
sulfur to sulfide as in form of sulfuric acid, second to eliminate other volatile compounds
and leaching with hot water. For silver and gold, the ore is subjected to sulfating and then
chlorinating.
Baglin et al. (1983-12-27) presented process comprises (1) smelting the sulfide
with flux as CaO, CaF2 and SiO2 (2) dry grinding 200 mesh (3) leach with 2.06 M
sulfuric acid in first stage for four hours at 95-100oC. The residue was leach in second
stage with Fe2(SO4)3, H2SO4 or FeCl3, HCl for six hours at 70oC.
Norrgran (1985-09-17) mixed the ore 200 mesh with sodium carbonate and
roasted at 420oC for 140 minutes, for second roasting the temperature raised to 610oC for
210 minutes, the ore was rabbled every 30 minutes, cool in furnace. Agitation leaching
was performed with water for two hours.
Frank (1974-02-12) roasted an admixture of an inorganic chloride and sulfide ore
for subsequent extraction of copper values of the ore as water soluble salts. The process
claims to reduce the emission of large quantities of sulfur dioxide and chlorine without
the necessity of using chemical absorbing solutions.
Coffin (1978-10-17) roasted the sulfidic material to produce a calcine. The calcine
mixed with carbonaceous reductant and with at least one halide at a segregation roasting
temperature at 650oC to 700oC.
63
2.4.2. Ammonia Leach Process
Tolley et al. (1982-03-30) treated the ground ore with 200ml of high strength ammonium
carbonate solution contains 20 grams of hydroxylamine hydrochloride for four hours
leaching. This solution was subjected to oxidation at oxygen pressure of five psig for
three hours; other examples used ammonium chloride, carbon monoxide.
Arbiter (1998-08-18) provided both floatation and oxidation leaching for copper
containing material by leaching agents as ferric sulfate, ammonia and ammonium salts
sulfur dioxide sulfuric acid, aeration with intensive agitation in two to four hours.
Arbiter (1998-09-15) invented a process is based on non-selective floatation and
acid leaching with aeration, agitation using ammonia and ammonium sulfite to reduce
copper sulfide to copper ions and sulfide ions in three to four hours leaching time.
Duyvesteyn et al. (1993-01-05) leached the copper sulfide floatation concentrate
with ammonia and ammonium sulfate at 20-40oC about two hours at pH 9-11.
2.4.3. Chloride Based Leach Process
Hyvarinen et al. (1999-12-28) processed ore with the alkali chloride produces copper
oxidule is fed into chloride alkali electrolysis. The reduced copper is in a granular or
pulverous form.
Jones (1999-05-11) leached copper sulfide ore with oxygen pressure chloride
process. The residue used after removing sulfur for recovery of precious metals by
oxidative high temperature pressure chloride leaching.
64
Johnson et al. (1999-06-29) milled the chalcopyrite in a vertical stirred mill to p80
of between two and 20 micron to produce activated copper mineral. The oxidative leach
was at 100oC and 1000 kPa for 75 minutes. The sodium chloride or hydrochloric acid
used for chloride ions.
Han et al. (1999-11-23) combined the material with reactants including water,
bromine/bromide, iodine/iodide, sodium and potassium nitrate, and optionally oxygen
and sulfuric acid to form a charged reaction zone at 150oC and 270 psig then
electrowining, cementation, solvent extraction and chemical precipitation used.
Pepper et al. (1979-09-25) roasted the material at 620oC for 80% sulfur
elimination. Then roasted pyrite was subjected to an oxidation leach using chloride
solution at 145oC and 150 psia.
Kieswetter et al. (1976-04-20) leached the copper material first with ferric
chloride in aqueous solution. The solution passed through sulfuric acid, the HCl and
copper sulfate formed treating with hydrogen gas. Ferric oxide and HCl gone to roasted
step is electrolyzed for hydrogen and chlorine gas generation.
Von (1986-11-12) said process has three broad steps (1) finely divided sulfides
mixed divided chlorides or chloride gas or sulfur chloride (2) heating mixture at to form
chlorides and elemental sulfur (3) convert metal chloride by conventional means.
Demarthe et al. (1980-10-28) stated a method in which lixiviate solution contains
CuCl2 and sodium chloride at 85oC for five hours at pH-2 by 12N hydrochloric acid with
air.
2.3.4. Chloride-Sulfate Leach Process
65
Shaw (2004-06-08) disclosed the leaching of sulfide ore or concentrate having copper,
zinc, silver and gold. The processes include two or more two leach steps. The leach steps
include and oxidative acid sulfate high pressure leaching, ferric sulfate leach, a mix
chloride-sulfate leach.
The invention relates the combined action of chloride ion and sulfate ions with
continuous oxygen injection and vigorous agitation. The chloride ions taken from NaCl
and HCl and sulfate ions produced by H2SO4 (Sawyer et al. 1990-11-20)
Jones (1995-07-11) subjected the sulfide ore or concentrate to first leaching at an
elevated temperature 125-175oC and pressure of oxygen 50 psig-250 psig. The
concentrate was ground to 98% minus 400 mesh and leached with acidic solution of
chloride and bisulfate or sulfate ions to produce insoluble basic copper sulfate, and
leached with sulfuric acid.
Jones (1997-07-08) subjected the sulfide copper ore or concentrate to pressure of
oxygen with chloride acids solution to obtained the insoluble basic copper sulfate salt and
dissolved.
Jones (1997-07-22) leached the sulfide concentrate having base metal like copper,
nickel, zinc, and cobalt and iron. During pressure oxidation the base metals precipitated
as basic salts e. g. basic copper sulfate. The basic salts oxidized to sulfate compounds by
sulfate-chloride leaching under pressure.
Jones (1999-02-23) subjected the ore to oxygen pressure leaching with chloride
and bisulfate or sulfate ion. The residue is used for gold and silver extraction. The sulfur
is removed from residue and then subjected to high temperature pressure leaching with
oxygen and chloride acids solution.
66
Young et al. (2002-10-29) leached copper and nickel from laterite ore with
sulfuric acid-halide-carbon system.
Kruesi et al. (1982-04-13) used microwave energy to selectively heat the copper
compounds, such as sulfidic and oxidic compounds in the ore. The sulfide ore was heated
in the presence of oxygen at 350-700oC. Then ore was heated at 300oC with chlorine or
ferric chloride the leaching was done with sulfuric acid at pH 1.0.
Baczek et al. (1981-03-17) divided ground chalcopyrite into first stream is
leaching with 100 g/l sulfuric acid and 10-30 g/l ferric iron. The copper sulfate solution
combined sulfur dioxide in second stream. Chloride present was about 200-2000 ppm.
The copper sulfides change into sulfate.
2.4.5. Counter Current Leach Process
Zoppi (2000-12-12) leached copper primary and secondary ores in different stages by
counter current leaching processes using leaching regents ferric fluborate and fluboric
acids. Inen (2005-08-16) leached iron copper concentrate in a chloride milieu and counter
currently in several stages.
2.4.6. Temperature Pressure Acid Leach Process
Marsden et al. (2005-05-17) used seeding agents to provide a nucleation site for the
crystallization of solid species which otherwise attend to passive the reactive process of
otherwise encapsulate the metal value during pressure leaching.
67
Marsden et al. (2004-01-03) disclosed a process for recovery of metal by
controlled, superfine grinding to a p98 of less then about 25-micron and temperature from
142-180oC in the presence of surfactants to form copper contained solution.
Marsden et al. (2004-01-20) used pressure leaching operation, need not be
significantly diluted facilitate effective metal recovery using solvent extraction and
electrowining. The temperature range is 210-235oC for the pressure leaching of
concentrated copper sulfide bearing material in oxygen that containing in a seed, agitated
multiple compartment pressure leaching vessel to form product slurry.
Marsden et al. (2003-09-30) used seeding agent in high temperature pressure
leaching at least one portion of residue from said reactive process.
Marsden et al. (2002-09-17) subjected copper containing material to high
temperature pressure leaching in a pressure leaching vessel with seeding agent. The
seeding agent may be a portion of residue.
Shaw (2004-06-08) disclosed the leaching of sulfide ore or concentrate having
copper, zinc, silver and gold. The process includes two or more two leach steps. The
leach steps include oxidative acid sulfate high pressure leaching, ferric sulfate leach, a
mix chloride-sulfate leach.
Collins et al. (1998-03-24) described a process for finely divided sulfidic copper
concentrate of chalcopyrite involving oxidized pressure leach using dilute sulfuric acid
and a carbonaceous additive, at above the melting point of sulfur but below about 200oC.
King (1997-12-16) described a process for recovering copper from a copper
containing material by the pressure oxidation, dilution before solvent extraction and
electrowining.
68
The copper ore concentrate must be 90% minus 325 mesh subjected to vigorous
agitation leaching at 200oC and an oxygen partial pressure of about 150-300 psig in 0.5 to
2.0 hours. In second step the filtrate passes through heap of low grade or waste to reduce
the acidity and increase pH 1.5 to 2.0 by percolation leaching (Jones, 1993-06-29).
Robert (1996-01-16) leached the ore by mixture sulfuric acid and nitric acid at 110-
170oC. During leaching the oxygen gas passed for one hour and the oxygen replaced by
argon for next 30 minutes and temperature raised up to 180oC.
Corrans et al. (1993-08-03) activated the material by fine or ultra-fine in milling
prior to process by methods of oxidative hydrometallurgy. The size of particles reduces to
p80 of 15 micron or less and leaching at 120oC and 1000 kPa.
Dempsey et al. (2003-01-07) used finely divided chalcopyrite concentrate in an
aqueous sulfuric acid solution to form slurry with lignosol or quebracho as surfactants in
pressure vessel at an elevated temperature.
Dixon et al. (2005) patented a process of leaching copper from copper sulfides
concentrated such as chalcopyrite, using pyrite as a catalyst for ferric reduction in order
to eliminate passivation of the chalcopyrite surface. In the process maintaining operating
solution potential as suitable level does not materially oxidize the pyrites. The leaching
is carried out in sulfuric acid by oxidation in four hours.
2.5. Reagents Used in Leach Processes
The reagents from reviewed literature are listed as below:
Gas reagents
Oxygen, Chlorine, Sulfur dioxide, Sulfur trioxide, Argon, Iodine, Bromine, Ammina.
69
Liquid reagents
Sulfuric acid, Hydrochloric acid, Nitric acid, Fluoboric acid
Solid reagents
Oxides: NaNO2, Fe2O3, SiO2, Al2O3, MgO, CaO, K2O, Na2O, TiO2, V2O5, CuO
Chlorides: CuCl2, CuCl, NaCl, FeCl2 FeCl3, MgCl2, KCl, ZnCl2, LiCl, CaCl2
Sulfates: K2SO4, Na2SO4,
Others: Ferric Fluoborate and Ammonium salts.
It was noticed that for the recovery of copper from its ores especially from sulfide
ores by using hydrometallurgical techniques the following two routes were followed; (i)
direct leaching (ii) indirect leaching.
In these routes the following processes or methods were adopted to obtain
concentrated leach solution (1) oxidation (2) sulfation (3) chlorination (4) nitration. These
method were conducted separately or sometime combined or simultaneously. According
to literature review, the roast leach process a few example were seen. In roast leach
process the said methods were used to activate the chalcopyrite or copper sulfide ores
before leaching. These techniques were used for the recovery of all type of metals
affecting by different roasting parameters.
The hydrometallurgical routs are divided into two states reactions as (i) aqueous state
reactions and (ii) solid state reactions. The pretreatments activate the sulfide minerals and
convert those to possible in acid water soluble species or minerals. The purpose of
pretreatments was the same in both states reactions to destroy and disintegrate the sulfide
nature of the minerals by doing (1) oxidation (2) sulfation (3) chlorination (4) nitration.
70
The similarities were found in hydrometallurgical leaching techniques to treating copper
sulfide ores to extract copper from copper ores or concentrate.
2.5.1. Similarities of Hydrometallurgical Leaching Techniques
(Copper Sulfide Ores)
S.No Roast-Leach Direct-Leach
1.
Gas Additives:
O2, SO2, SO3, Cl2 & N2
Gas Additives:
O2, SO2, SO3, Cl2 & N2
2.
Solid Additives:
Inorganic chloride salts
Oxygen containing salts
Solid Additives:
Inorganic chloride salts
Oxygen containing salts
3.
Use in solid state:
The sulfides change into soluble
sulfate/chloride under temperature
pressure conditions
Use in liquid state:
The sulfides change into soluble
sulfate/chloride under
temperature pressure conditions
4.
Process steps
More than two
Process steps
More than two
5. Leach with acid water Obtained Leach solution
6.
7.
Direct Electrowining
Electrowining followed by Solvent Extraction
Pure Copper Metal
71
2.6. Solvent Extraction
The solutions from leaching of low-grade ores are low in copper (<5Kg m3Cu) were
previously recovered by cementation on iron (Moore, 1982). Recently however, organic
solvents have been developed which selectively extract copper ions from these dilute
solutions. The organic solvents can be subsequently stripped of this copper at low pH,
concentrated copper aqueous solutions suitable for electrowinning (Biswass, 1975).
The modified strong extractants used to extract copper from high concentration
aqueous feeds for hydrometallurgical processing of sulfidic copper ores (Maes, C. 2003).
Acorga reagents such as M5640 & PT5050, both of which have better selectivity.
Extensive laboratory scale and commercial operations around the world have
demonstrated that Acorga (ester modified reagent M5640) is the most hydrolytically
stable formulation available (Gohar, 2005). Copper hydrometallurgy usually employs
acidic chelating agents Lix, Kelex, SME and Acorga were used (Alguacil, 1987). Copper
was recovered from dilute aqueous solutions with Acorga M5640 using non-depressive
solvent extraction technique, pH was adjusted at 4.0 stripping was done by 180 g/l H2SO4
solution by (Alguacil, 2002). The solvent extractant 4 (alpha branched) acyl – (3H) –
pyrazol – 3ones is a suitable organic phase with low aromatic hydrocarbons in a metal
recovery from acidic or ammonical leach solution (Campbell, 2003). Acorga P5100 in
Kerosene oil is well for extraction of copper from acid chloride solution and stripping is
completed by dilute HCl (Jena, 2003). Metals are recovered from leach solution by
extraction with organic phase followed by fractional distillation of metal compounds
72
metals are separated from organic phase by heating to distill the organic phase or a metal
compound (Lakshaman, 2005).
Extraction of copper was done with tertiary amines (Khan, 1999). Khan et al.
(2000) described the comparative study with various high molecular weight tertiary
amines to develop hydrometallurgical methods for the extraction and recovery of copper
from different acid media. They limited such as, the effect of acid, extractant and
stripping agents, concentration time of contact and diveroceions were optimized.
Maximum extraction and recovery of copper ion 99% were achieved with 5% solution of
trioctylamine in benzene. The process developed has been made more cost effective by
recycling used solvent.
Agrawal (1982) had published a review for recovery of copper by solvent
extraction. Gonzales et al. (1996) investigated the plants of solvent extraction and
electrowining (SX/EW) in Toqrepala and in Cuajone mine sites. The leaching operation
allows better and more economical resource exploitation.
Zunkel (1993) dveloped the Cuprex Metal Extraction Process (CMEP). The
integrated ferric chloride-based hydrometallurgical process uses mild leaching conditions
and novel solvent extraction and electrowining step to give cathode grade copper from
clean dirty or low-grade copper concentrates.
The history of the evolution of copper hydrometallurgy in mining is reviewed
(Arbiter et al. 1994). A variety of leaching systems, electrowining of porphyry-oxide
copper ores, solvent extraction technique were applied on low-grade ores, tailing and
other wastes. Anonymous in BHP Laboratory (1993) described a novel and patented
73
application of solvent extraction in the copper industry is the Escondida Process
developed by Minerals Laboratory of BHP.
2.7. Kinetic and Thermodynamic Studies
The kinetics and thermodynamics parameters of roasting and leaching processes were
investigated. In the literature many researchers (Strbac, 2006; Zivkovic, 2005; Arslan,
2004; Minic, 2005; Lawson,1992; Syngouna, 1997; Hancock, 1972) has used the
equations for kinetic models defined by Sharp (1966). These kinetic models are used to
determine the reaction rate and calculate the activation energies and thermodynamic
parameters of the reactions in aqueous as well as in solid state. The kinetics model of
Sharp was compared to investigate the mechanism of reactions by using diagnostic
equation. The activation energy, enthalpy, entropy and free energy were calculated and
compared with different kinetic equations of model.
Albery et al. (1988) studied the kinetics and mechanism of the solvent extraction
of copper by using the oxime ligand Acorga P50, with the rotating diffusion cell in both
the extraction and stripping directions. The reaction takes place on the liquid-liquid
interface between the aqueous and heptane phases. A kinetic scheme involving the
sequential addition of two oxime ligands is proposed and all rate constants for the scheme
were measured. Sequences of free energy profile for the extraction and stripping reaction
under industrial conditions are presented.
Smith et al. (1999) investigated the mechanism and kinetics of a leaching reaction
of Raney copper, an active hydrogenation catalyst. The reaction rate was found to be
constant and independent of hydroxide concentration under the reaction conditions used.
74
The activation energy for the reaction was found to be 69 [plus or minus] 7 kJ/mol. The
mechanism of the formation and rearrangement of the copper was a dissolution/
redeposition of copper atoms, with a minor role played by surface and/or volume
diffusion.
Konishi et al. (1999) gave the kinetics of leaching of high-grade chalcopyrite
(CuFeS2) concentrate (38-53mm) by the acidophilic thermophilic bacterium, Acidianus
brierleyi, was studied at 65oC and pH of 1.2 in a batch stirred reactor. The adsorption of
A. brierleyi cells between the sulfide surface and solution was attained within the first 20
min of exposure to the mineral, and the equilibrium adsorption data were correlated with
the Langmuir isotherm. The bioleaching was markedly accelerated in the presence of A.
brierleyi, and greater than 90% leaching copper in the concentrate was achieved within
10 days. Rate data collected under a wide variety of operating variables were analyzed to
determine kinetic and stoichiometric parameters for the microbial chalcopyrite leaching.
Eriksson et al. (1997) investigated the couple defects of flow heterogeneity, primary
dissolution kinetics and equilibrium precipitation / dissolution of secondary copper-
bearing minerals.
It is concluded as the recovery of copper from its ores by using hydrometallurgical
techniques the two routes are used (i) direct leaching (ii) indirect leaching. These
techniques are carried out into two states reaction as aqueous state and solid state
reaction. The pretreatments activate sulfide minerals and convert to those of soluble
species. The additive roasting prior to leaching is mostly used to treat the copper sulfide
ores in the present study.
75
CHAPTER 3
MATERIALS, METHODOLOGY AND EXPERIMENTAL
3.1. MATERIALS
3.1.1. Equipments
Analytical balance, four digits, Sartorius
Atomic Absorption Spectrophotometer Z8000 Hitachi, Japan
X. R. D., D8 Discover and Bruker
Zoom microscope, Olympus SZX7, DF PLAPO 1X-4, Japan
Zoom Digital Camera, Olympus C-740 Ultra zoom 3.2 mega pixel, Japan
pH meter (model # 8520 Hanna Instruments; measurement temperature range 0-100 ºC).
Comminution equipment mineral processing laboratory and pilot plant of PCSIR
Laboratories Lahore.
3.1.2. Chemicals and Reagents
Hydrofluoric acid (AR) HF E. Merk
Hydrochloric acid (AR) HCl E. Merk
Sulfuric acid (AR) H2SO4 E. Merk
Nitric acid (AR) HNO3 E. Merk
Sodium nitrite (Pure) NaNO2 E. Merk
76
Sodium nitrate (Pure) NaNO3 E.Merk
Potassium chlorate (Pure) KClO3 E. Merk
Sodium chlorate (Pure) NaClO3 E. Merk
Ammonium nitrate (Pure) NH4NO3 E. Merk
Ammonium carbonate (Pure) (NH4)2CO3 E. Merk
Ammonium oxalate (Pure) (NH4)2(COOH)2 E. Merk
Ammonium acetate (Pure) (NH4)C2H3O2 E. Merk
Ammonium sulfate (Pure) (NH4)2SO4 E. Merk
Ammonium phosphate (Pure) (NH4)3PO4 E. Merk
Ammonium chloride (Pure) NH4Cl E. Merk
Urea (Pure) Urea E. Merk
Barium chloride (Pure) BaCl2 E. Merk
Sodium thiosulfate (Pure) Na2S2O3 E. Merk
Sodium persulfate (AR) Na2S2O8 E. Merk
Sodium hydroxide (Pure) NaOH E. Merk
Potassium iodide (AR) KI E. Merk
Copper sulfate (Pure) CuSO4. 5 H2O E. Merk
Potassium bromide (AR) KBr E. Merk
Copper powder (AR) Copper metal E. Merk
3.2. METHODS
77
The research involved the valid methods. These are using internationally in
hydrometallurgical techniques which are discussed in literature review;
3.2.1. Sample Procurement
The samples for research studies were prepared from representative and typical ores
having different mineralogical characteristics. The samples of copper ores were from
Kaldam Gol area, Darosh Chitral and Saindak. Copper concentrate was also procured
with the help of mineral exploring agencies (Exploration Division, Director General
Mines & Mineral) working in NWFP (KPK) and Geochemistry section of PCSIR,
Lahore. The samples were also collected from Chitral and Dir areas under HEC Project
No. 870. The samples were representative of both oxidize and sulfide zones. The Darosh
samples were taken from the ore dump at the mine. Trench samples of the ore were also
collected from different areas of oxidized and sulfide zones of copper deposit with the
help of mineral exploring agencies. The flotation concentration of Saindak copper ore
Baluchistan prepared in PCSIR, Lahore was also used.
3.2.2. Sample Preparation
The samples were prepared by crushing and grinding to the required mesh size for
different studies such as for chemical and mineralogical evaluation, XRD, leaching,
roasting, extraction, petrography and stereo microscopy.
3.2.3. Mineralogical Studies
78
The mineralogical studies undertaken for the determination of basic characteristics of the
ore includes petrographic study, ore microscopy and separation of mineral by gravity and
flotation techniques. Mineralogical techniques such as XRD were also used for the
identification of mineral phases and study of the ore texture before and after treatments.
The work also included XRD technique for the qualitative estimation of minerals.
3.2.4 Chemical Analysis
The chemical constituents were determined in copper ore by applying different
conventional, instrumental and chemical analyses methods. The instrumental methods
used for chemical analyses include Atomic Absorption Spectrophotometer, Flame
Photometer and UV-Visible Spectrophotometer. The conventional chemical analyses
(Vogel, 1961) were also conducted. The iodimeteric analyses were used to find the
amount of copper.
3.2.5. Chemical Processing
The study was conducted on indigenous copper ores for extraction of copper. The copper
ore deposits of Pakistan are small and contain low-grade copper and associated metals
and require processing for utilization. According to the related literature review there are
two types of chemical processes used in hydrometallurgy. These processes were
investigated to extract copper from ores e.g. (l) direct leaching, (2) indirect leaching.
Direct Leaching: Leaching is a wet metallurgical process employing different chemicals,
leaching agents and techniques. The technologies were developed for extraction of
79
copper in the world for copper ores. The literature shows that oxide minerals of copper
can be easily leached with mild sulfuric acid at ambient conditions as compared to sulfide
minerals. The acid leaching includes the evaluation of the important parameters affecting
the rate of leaching such as particle size, acid concentrations and solid liquid ratio. The
effect of oxidizing agents, temperature and pressure were also investigated.
Indirect Leaching (Roasting before Leaching): The sulfide minerals of copper are not
easily leached, while its sulfates are soluble in water and its oxides are soluble in dilute
sulfuric acid (Biswas, 1976). The roasting of copper sulfide ores before leaching was also
used for copper extraction. Thus the roasting of sulfides can produce a readily leached
product containing oxide and/or sulfate. The leaching studies were conducted in this
study on the sulfide as well as complex copper sulfides ore after roasting. The various
types of roasting treatment are (1) the simple roasting conducted without using any
reagent during roasting; (2) the roasting conducted using different reagents for oxidation
during roasting e.g. oxidants such as persulfate, sulfate, chlorate, nitrate and nitrite.
Stronger oxidants such as hydrogen peroxides, ozone, Caro’s acid (H2SO5), persulfate,
permanganate, dichromate and manganese dioxide were thought to be of less interests
because of their cost (Peters, 1992), (3) the roasting with additives was also investigated.
(4) Two-stage roasting, (5) roasting using a combination of additives (ammonium
compounds) at various temperatures was also conducted.
3.2.6. Kinetic Studies
80
Chemical kinetics of roasting and leaching reactions was studied and discussed in the
thesis. The work can provide the basis of chemical processing. The kinetic equations
models defined by Sharp (1966) were used to calculate the different parameters of
kinetics.
3.3. EXPERIMENTAL WORK
81
The experimental work conducted was as follows:
3.3.1. Sampling and Comminution Studies
The bulk ore samples of copper containing sulfides and oxides were collected from
different localities Kaldam Gol area in Chitral from N.W.F.P. Pakistan. The Chalcopyrite
concentrate sample of Saindak copper ore Baluchistan was also prepared for roast leach
work.
Preparation of analytical samples: The following procedure of crushing and grinding
was conducted for the preparation of analytical samples. The sample E was prepared by
concentration using flotation in PCSIR Laboratory Lahore.
Procedure: The following procedure was applied to four groups of samples A, B, C and
D. First jaw crusher was used to crush the large pieces of copper ores. The crushed
material was again passed through jaw crusher with opening of one inch size. The
crushed material was ground to ¼ inch size by roll mill. Then material was quartered by
riffle. A portion of quartered material was passed through disc grinder. The material was
divided by riffle. One half of disc grinded material was pulverized by Tema mill, up to
the size of minus 300 to minus 400 mesh. Other half was passed through different sieves
size of 50, 100, 200, 250 and 300 meshes. The analytical samples passing 300 mesh size
were used for the different types of instrumental, mineralogical, physical and chemical
analyses.
3.3.2. Microscopic Analyses of Samples
82
The different mesh size was examined by petrographical microscope. The microscopic
study showed that all samples have large amount of sulfide minerals. The
microphotography of treated copper ore samples was also conducted. These are
confirmed by the study of XRD analysis of treated and untreated. The results are shown
in Figures 2-8 (Appendix-A) and Figures 9-11 (Appendix-B).
3.3.3. Chemical Analysis
The ore samples were completely dissolved by using standard analytical methods. The
chemical constituents were analyzed by using standard analytical methods and Atomic
Absorption Spectrophotometery (Vogel, 1961). The results of chemical analysis are
shown in Chapter 4, Table-1.
3.3.4. Chemical Processing
There are two types of chemical processes used in hydrometallurgical leaching
techniques were investigated to extract copper from ores e.g. (i) direct leaching, (ii)
indirect leaching as roasting before leaching.
3.3.4.1. Direct leaching
Direct leaching of ore samples were conducted by different concentrations of sulfuric
acid in different time periods. The effect of oxidizing agent, temperature and pressure
was studied.
83
Leaching with sulfuric acid: Pulverized copper ore samples marked as A, B, C, D and E
(chalcopyrite concentrate) were taken for leaching studies. In each leaching test the ore
sample was treated with 5.0 percent sulfuric acid in 1:5 solid liquid ratio. The leaching
time was varied as 02, 05, 10, 20 and 30 minutes. The effect of acid concentration,
particle size and solid liquid ratio were also investigated. The samples were analyzed for
copper by using standard analytical methods. It was found that the oxidized ores samples
A & B were not leached to considerable extent, as only 15 percent to 30 percent leaching
was achieved. This may be due to the presence of sulfide minerals in associated with
oxides minerals as these samples contain 15-30 percent oxide minerals. The samples C
(complex copper sulfide ore), D (mixed ores of copper) & E (Chalcopyrite concentrate)
were also not leached by sulfuric acid.
The results obtained are shown in Chapter 4, section-V. The effect of time and
sulfuric acid concentration on copper leaching of samples A, B, C, D and E is shown in
table-30. The effect of solid liquid ratio is shown in table-31 and effect of temperature is
shown in table-32 in Chapter 4, Section-V.
Leaching with sulfuric acid under various temperature and pressure: The pressure
leaching was carried out in Bomb digester using the method described by Rani, et al.
(1998). Weighed quantity of pulverized sample ‘C’ was put in Bomb digester and 10
percent HNO3 and 20 percent H2SO4 were added by using solid liquid ratio of 1:5. It was
placed in boiling water for 30 minutes. After cooling the solution was filtered and
required volume was made. The sample solution was analyzed for extracted copper by
using standard analytical iodimetric methods.
84
The same procedure was repeated for 60 min, 120 min and 180 min. The pressure
leaching of copper sulfide ores had been investigated (Rani et al. 1998). The optimum
conditions given her work on pressure leaching were used for investigated on the present
sample. The leaching of copper was not possible to more than seven percent of the total
amount of copper. Then different parameters were investigated such as acid
concentration, oxidizing agent, different temperature and pressure. The results presented
the effects of time (table-33), sulfuric acid concentration (table-34), oxidizing acid
concentration (table-35), solid liquid ratio (table-36), particle size (table-37), temperature
and pressure (table-38) on copper leaching in Chapter 4, Section-V.
3.3.4.2. Indirect Leaching (Roasting before leaching)
As discuss under methodology section 3.2.4.1 the leaching of complex copper sulfide
ores is difficult as compared to copper oxide ores. It is reported in the literature that
indirect leaching methods based on pretreatment and subsequent leaching of ore could be
an alternative processes for leaching complex copper sulfide ores and chalcopyrite. The
following methods were adopted to leach complex copper sulfide ores.
3.3.4.2.1. Simple Roasting of sample C (complex copper sulfides) and Chalcopyrite
before acid leaching
3.3.4.2.2. Roasting of sample C with oxidants before acid leaching
3.3.4.2.3. Roasting of sample C with novel oxidant before acid leaching
3.3.4.2.4. Roasting of sample C with non-oxidant material (novel additives) before
acid leaching
85
3.3.4.2.1. Simple Roasting before Acid Leaching
The pulverized copper ore samples marked as C (complex copper sulfide), E
(Chalcopyrite) were heated at different time and temperatures. The sample was taken in a
crucible and the covered crucible was placed in Muffle furnace at room temperature.
When the temperature reached 500oC, then crucible was continuously heating for three
hours at this temperature. The crucible was placed in desiccator for cooling. The roasted
material was ground in pestle and mortar. The leaching studies were then conducted on
the roasted material. The weighed roasted material was leached with 2.5 percent H2SO4
by using 1:5 solid liquid ratio for two minute. The amount of leached copper was
determined by standard analytical methods. The same procedure was repeated at 550oC,
600oC, 650oC, 700oC and 800oC. The roasting parameters such as roasting time, roasting
temperature, particle size were investigated. The studies of leaching parameters such as
leaching time, acid concentration, solid liquid ratio were also conducted. The detailed
results show as; the effect of simple roasting of sample C at different time and
temperatures (table-2), roasting time (table-3) and particle size during roasting (table-7)
in Chapter 4, section-I.
Simple roasted ore sample C: The effects of leaching time (table-4), sulfuric acid
concentration (table-5) and solid liquid ratio (table-6) on copper leaching after simple
roasting are discussed in Chapter 4, Section-I.
86
3.3.4.2.2. Roasting of Sample C with Oxidants before Acid Leaching
The roasting was also conducted by mixing ore with different oxidants; the oxygen
containing salts. The oxidants were used during roasting of sample C to see if they
increase the extraction of copper from 72 percent that was achieved by simple roasting.
The following oxidants like NaClO3, NaNO2, NaNO3, KClO3, and Na2S2O8 were
investigated for roasting the ore. The amount of oxidant was 20 percent of the weight of
ore sample C. The mixture of ore and oxidant was taken in a china crucible. The covered
crucible was placed in furnace at room temperature. The temperature of the furnace was
raised to 650oC. The roasting was done for half an hour at 650oC. The roasted material
was cooled and ground. The weighed roasted material was leached with 2.5 percent
H2SO4 by using 1:5 solid liquid ratio for two minutes. The optimized leaching conditions
(4.4.2.1. Section-I) were used for leaching of roasted material and these leaching
conditions were used in subsequent leaching tests. The leaching of copper was not
significantly increased. The roasting with oxidant was carried out at low temperature
500oC as compared to values given in the literature (Norrgran, 1985) for long time
roasting for three hours. The roasted material was cooled and ground. The weighed
roasted material was leached with 2.5 percent H2SO4 by using 1:5 solid liquid ratio for
two minutes. The leached solution was analyzed by standard analytical method. The
particle size, roasting temperature, roasting time and amount of oxidant were varied and
optimized conditions investigated. The results obtained showing the effect of oxidant
during roasting and temperature of oxidant roasting are showing in tables 8 & 9 in
Chapter 4, Section-II.
87
3.3.4.2.3. Roasting of Sample C with Novel Oxidant
before Acid Leaching
Novel oxidant: The study was conducted using new material, which was not investigated
so far for roasting of copper containing sulfides ore. The ammonium nitrate was one such
compound used during roasting of complex copper sulfide ore. This novel roasting
material, which may act as the oxidant, has not been reported in the literature. The results
presented the effect of novel oxidant roasting on copper leaching with roasting time in
table-10, the effect oxidant’s amount in table-11, the effect roasting temperature and time
in table-12, the effect particle size in table-13 and leaching with different acids in
different time in table-14 in Chapter 4, Section-II.
3.3.4.2.4. Roasting of Sample C with Novel Non-Oxidant Material/
Novel Additives
The literature reported roasting additives such as lime, NaCl, (Na)2SO4, (Na)2CO3, and
many other inorganic salts. The roasting of complex copper sulfide ore with novel
additives was investigated. These additives contain ammonia, chloride, sulfate,
phosphate, acetate, oxalate, carbonate and urea. A known percentage of additive by
weight of sample was taken. The sample C of complex copper sulfide ore was used. The
roasting with additive was conducted at temperature 500oC for three hours. The roasted
material was cooled and ground. The roasted material was subjected to leaching. The
additive, which has shown better leaching, was selected for detailed investigations.
88
The different parameters of roasting with novel additives were particle size, roasting
temperature, roasting time, amount of additive. The particle sizes used were of 200, 250,
300 mesh and minus 300 mesh. The effect of roasting temperature on extraction of
copper from complex copper sulfide ore was conducted at 400oC, 450oC and 500oC. The
effect of particle size of ore on roasting with ammonium chloride as additive was also
studied. The results are shown in tables-15, 16, 17, 18 and 19 in Chapter 4, Section-III.
3.3.4.2.4.1. Two Stage Roasting
The special treatment method using two stage roasting was investigated after selecting
different additives, particle sizes, roasting temperatures and roasting times. This
treatment was applied on roasting with additives such as ammonium nitrate, ammonium
chloride and ammonium sulfate. A 10 percent amount of ammonium chloride mixed with
complex copper sulfide ore (sample C) was taken in crucible. The covered crucible was
placed in furnace and heating started from room temperature. When the temperature
reached 350oC, the mixture was roasted for 30 minutes and the temperature was increased
to 500oC and the mixture was again heated for 45 minutes at this temperature. The
weighed roasted material was leached with 2.5 percent H2SO4 by using 1:5 solid liquid
ratio for two minutes. The leach solution was analyzed by standard analytical method.
The tables 20, 21, 22 and 23 show the effects of ammonium chemicals during two-stage
roasting, particle size, temperature and comparison of copper leaching at different
roasting temperatures respectively in Chapter 4, Section-III.
89
3.3.4.2.4.2. Effect of Combined Oxidant/ Additive on Copper Leaching
The combine effect of ammonium nitrate, ammonium chloride and ammonium sulfate
was investigated. The pulverized sample C was mixed with different ratios of combined
additives. The roasting was carried out for 30 minutes at 350oC then temperature was
raised to 500oC, and the material was heated for 45 minutes. The material was leached
with 2.5 percent H2SO4 by using 1:5 solid liquid ratio for two minutes after cooling and
grinding. The leach solution was analyzed by standard analytical method. The results
shown in Chapter 4, Section- IV are the effects of combined additives (table-24),
combinations of additives at different temperatures (table-25), the effect of particle size
during roasting with chloride and sulfate of ammonium (table-26), comparison of
combinations of additive ammonium chloride and chloride & sulfate of ammonium at
different conditions of roasting (table-27) on copper leaching.
3.3.5. Leaching of Copper Chalcopyrite Concentrate
The roasting and leaching conditions were optimized on complex copper sulfide ore
(Sample C). The acceptable conditions were applied on the chalcopyrite concentrate
(Sample E). The results of the behavior of chalcopyrite concentrate towards different
conditions of roasting and leaching are shown in Chapter 4, Section-IV in Table- 28.
3.3.6. Amount of Sulfur Analyzed in Different Samples
The amount of sulfur was determined before and after roasting treatments by
conventional methods (Vogel, 1961). It was found that the amount of the sulfur has
90
remained constant in most of the experiments. That shows the emission of sulfur dioxide
might be controlled. The results shown in Chapter 4, Section- IV as the amount of sulfur
are shown in table-29.
3.3.7. XRD Analyses
The XRD analyses complex copper sulfide samples were conducted. The optimal treated
samples of copper ore were subjected for XRD analyses. The results of XRD analyses
were estimated by using Data Book (1978) for diffraction and Zivkovic (2005). The
results are shown in Chapter 4, Section-IV and Figure-2-8.
3.3.9. Kinetic Studies
Chemical kinetics of roasting reactions was studied and discussed in the thesis. The
kinetic equations models defined by Sharp (1966) were used to calculate the different
parameters of kinetics. The results are interpreted in Chapter 4, section-V, in table-39, 40.
91
CHAPTER 4
RESULTS
Hydrometallurgical routes for the extraction of metals are stated to be simpler, rapid,
efficient and more economical. The hydrometallurgical leaching (chemical processing)
techniques were developed for indigenous copper ores during the research. This part of
thesis presents the results of the different experiments and investigations.
4.1. Sampling and Preparation of Analytical Sample
The samples of copper ores were from Kaldam Gol area, Darosh in Chitral in N.W.F.P.
and Saindak in Baluchitan. The samples were representative of both oxidized and sulfide
zones. The samples for research studies were prepared from representative and typical
ores having different mineralogical characteristics. The sample A has mineralogical
characteristics dominantly of Azurite. The sample A is blue in color with subordinate
amount of Malachite. Sample B is Malachite and is green color with subordinate amount
of Azurite. Sample C is black in color a complex copper sulfide minerals containing
Tetrahedrite with subordinate amount of Galena and Sphelarite. Sample D has the mixed
oxides and sulfides minerals and is composed of samples A, B and C. Sample E is the
Chalcopyrite concentrate sample of Saindak copper ore. The concentrate was heat treated
for drying at about 300oC. The samples were marked as A, B, C, D and E.
Oxidized copper minerals
92
A = Azurite, is major oxide mineral in ore containing 20 to 80 percent of sulfides
ore
B = Malachite, is major oxide mineral in ore containing 30 to 70 percent of
sulfides ore
Sulfide copper minerals
C = Copper containing sulfides ore (Complex Copper sulfides)
D = Mixed ore containing oxides and sulfides samples A, B, C
E = Chalcopyrite concentrate
The sampling and comminution studies of Kaldam Gol area and flotation
concentration of Saindak copper ore were carried out in mineral processing section of
PCSIR laboratories Lahore (please see section in 3.3.1). The analytical samples were
prepared for chemical and mineralogical evaluation, XRD, leaching, roasting, extraction,
petrography and stereo microscopy.
4.2. Microscopic Analyses of Samples
The different mesh size samples were examined by petro graphical microscope. The
microscopic study showed that all samples have large amount of sulfide minerals. The
microphotography of treated and untreated copper ore samples were also conducted.
These are confirmed by XRD study. The results are shown in figures 9-11.
93
4.3. Chemical Analysis
The copper ore samples were dissolved by conventional methods (Vogel, 1961). The
results of dissolved ore solutions were analyzed for base metals by Atomic Absorption
Spectrometric technique. The amount of copper in different experimental batches was
determined by iodimetery titration (Vogel, 1961).
The ore samples were found to contain different base metals. The amount of
copper in the studied sulfide samples ranges from 17.77 percent to 26.17 percent, zinc
ranges from 1.3 percent to 1.72 percent, lead 0.06 percent to 0.35 percent. The oxide
samples of copper ore A and B contain copper ranging from 13.92 percent to 20.98
percent, zinc from 1.28 percent to 1.85 percent, lead from 0.06 percent to 0.13 percent.
The iron in sample A, B, C, D ranges from 1.5 percent to 2.14 percent. However 23.78
percent iron is found in chalcopyrite concentrate sample E. The presence of precious
metals like silver and gold was found very low as silver ranges from 1.7 ppm to 2.84 ppm
and gold from 0.98 ppm to 1.02 ppm in oxide samples and in sulfide samples silver from
1.72 ppm to 3.15 ppm and gold from 0.91 ppm to 1.09 ppm. The results of chemical
analysis for base metals are shown in table 1.
Table 1: Base Metals in Copper Ore Samples
Sample
Copper
%
Zinc
%
Iron
%
Lead
%
Silver
ppm
Gold
ppm
A 20.98 1.85 2.13 0.06 2.84 0.98
B 13.92 `1.28 2.25 0.13 1.70 1.02
C 21.13 1.72 2.14 0.06 3.15 1.09
D 17.77 1.30 1.50 0.35 1.72 0.91
E 26.17 - 23.78 -
94
4.4. Chemical Processing
The chemical processing work was divided into following types as (i) direct leaching (ii)
roasting before leaching. The results are interpreted in different sections in this chapter.
4.4.1. Direct leaching
A series of leaching experiments using different acids were done on different samples of
copper ores containing oxidized and sulfides minerals. The effects of temperatures,
pressures and nitric acid were also studied. The oxidized ores samples were leached
easily with dilute sulfuric acid solutions within short time. But the sulfides minerals in
copper ores samples resisted leaching. The detailed results are presented in section-V of
this chapter.
4.4.2. Roasting before leaching
The sulfides ore samples were subjected to different roasting techniques such as (a)
simple roasting; in which no additive was added during roasting (b) roasting with
oxidant; in this roasting conventional oxidizing salts were used to enhance the leaching
(c) roasting with a novel oxidant; the ammonium nitrate was investigated for roasting of
sulfide ores (d) roasting with novel non-oxidant material or novel additives; a new set of
chemicals containing ammonia, nitrate, chloride, sulfate, acetate, oxalate, urea and
phosphate were studied as additives during roasting. The detailed results are shown in
following sections as:
95
SECTION-I
4.4.2.1. Simple Roasting
In the thesis the simple roasting refer to heating the complex copper sulfide ore in a
furnace at specific temperature. The powder ore of mesh size 300 of complex copper
sulfides ores were taken in a china crucible. The covered crucible is placed in muffle
furnace for roasting at different times and temperatures. The temperature ranges studied
were from 500oC to 800oC. The roasting time was also varied from 60 minutes to 240
minutes. The results of leaching after simple roasting at different temperature and time
were presented in table-2.
It is found from these experiments that the leachability (the amount of copper
leached in a sample by sulfuric acid solutions with or without pretreatment) of complex
copper sulfide ore (Sample C) has increased with increase in roasting time at lower
temperature range from 500oC to 650oC. The leaching of copper was 73 percent within
60 minutes roasting at 650oC but 70 percent copper was leached within 180 minutes at
600oC. When the temperature is raised from 650oC to 800oC the leachability of Sample C
increases to maximum with increase in temperature within 120 minutes roasting time.
However further increase in roasting time beyond 120 minutes the leachability decreased.
The sulfide minerals of copper formed oxides or sulfates on roasting and sulfur
dioxide is released (Biswas et al. 1976). The oxides formed are readily leachable in
sulfuric acid solutions and the sulfates are leachable in water. These changes may occur
at temperature range 500oC to 650oC but when the temperature is raised above 650oC the
96
extraction of copper was decreased. It may be due to the formation of new species of
minerals, which resist to leaching (Evrard, 2001). There may be intermediate minerals or
solid solutions formed on prolonged heating which are difficult to leach.
Optimum roasting temperature: The complex copper sulfides ore (sample C) was
roasted at different temperature for different time periods. The maximum leaching of
copper was achieved at 650oC after 60 minutes roasting. The extraction of copper was
very low 29.25 percent at temperature of 600oC for 60 minutes roasting. But at
temperature of 650oC a marked increase was found in the extraction of copper (up to 73
percent). A slight increase was observed after roasting for 60 minutes above 650oC to
800oC. At 750oC the extraction of copper has decreased to 75.55 percent to 71.4 percent
at 180 minutes roasting. The decrease in extraction of copper after roasting at higher
temperature for long duration may be due to transformation to phases, which resist
leaching. The XRD analyses of simple roasting at 650oC are shown in figure-3, 4.
Table 2: Effect of Simple Roasting on Copper Leaching in Sample ‘C’
S. No. Temperature oC
Copper Leaching (%) at Different Roasting
Temperature & Time
60 min 120 min 180 min 240 min
1 500 08.75 24.52 41.4 58.79
2 550 18.40 38.8 56.15 68.23
3 600 29.25 56.42 70.00 72.2
4 650 73.40 74.25 75.43 70.36
5 700 74.80 75.34 65.00 55.43
6 750 75.55 76.62 58.00 27.2
7 800 71.40 65.92 55.00 25.67
Particle size = 300 mesh, Leachant = 2.5% H2SO4, Solid Liquid Ratio = 1:5 Leaching
Time = 02 min
97
Effect of roasting time on copper leaching at optimum temperature: The Sample C
was roasted at 650oC for time periods of 30, 60, 120, 180, and 240 minutes. The results
indicated that optimum extraction of copper was achieved in 30 minutes by simple
roasting. After thirty minutes no significant difference in extraction of copper was found.
The extraction of copper remained constant after thirty minutes roasting time. The results
are presented in table-3.
Table 3: Effect of Roasting Time on Copper Leaching at Optimum Temperature
S. No. Roasting Time (minutes) Copper Leaching (% )
1 30 72.85
2 60 73.40
3 120 74.25
4 180 75.43
5 240 70.36
Sample C & optimum roasting temperature = 650oC, Particle size = 300 mesh
Leachant = 2.5% H2SO4, Solid Liquid Ratio = 1:5, Leaching Time = 02 min
98
Effect of leaching time on leaching of simple roasted ore: The simple roasted sample
C was leached with 2.5 percent sulfuric acid solution with 1:5 solid liquid ratio for
different time periods i.e. 15, 30, 120, 900, 1800 and 3600 seconds. The optimum amount
of copper extracted was 73 percent in only 30 seconds. Only three percent further
increase in leaching was achieved after 3600 seconds of leaching. It was shown that the
maximum leaching of the complex copper sulfides ore (sample C) was in 30 seconds as
shown by very dilute sulfuric acid. The results are shown in table-4.
Table 4: Effect of Leaching Time on Leaching of Simple Roasted Ore
S. No. Leaching Time (Seconds) Copper Leaching
(percent)
1 15 70.30
2 30 72.55
3 120 72.85
4 900 73.35
5 1800 74.75
6 3600 76.10
Roasted Sample C at 650oC, Roasting Time = 30 min, Leachant = 2.5% H2SO4
Particle size = 300 mesh, Solid Liquid Ratio = 1:5
99
Effect of sulfuric acid concentration on leaching of simple roasted ore: To observe
the leaching behavior of roasted ore, sample C was treated with different concentrations
of sulfuric acid i.e. 0.5, 1.25, 2.5, 5 and 10 percent after roasting. The leaching was done
with solid liquid ratio of 1:5 for a period of two minutes. The results in table-5 indicate
that there is a remarkable difference in the initial rate of copper extraction when using the
acid of 2.5 percent strength. The concentrations higher than 2.5 percent or with 5 percent
do not show a significant increase in the copper extraction. The results show that
maximum copper is leached within two minutes using only 2.5 percent sulfuric acid. It
has been observed that the rate of dissolution of roasted complex copper sulfides ore
increased with increase in the acid concentration. However, the leaching rate becomes
almost constants when the leaching was carried out with an acid concentration higher
than 2.5 percent.
Table 5: Effect of Sulfuric Acid Concentration on Leaching of Simple Roasted Ore
S. No. Sulfuric Acid Conc. (%) Copper Leaching (%)
1 0.50 54.80
2 1.25 59.75
3 2.50 72.85
4 5.00 73.55
5 10.00 73.75
Roasted Sample C at 650oC & Roasting Time = 30 min, Leaching Time = 02 min
Particle size = 300 mesh, Solid Liquid Ratio = 1:5
100
Effect of solid liquid ratio on leaching of simple roasted ore: A series of experiments
were performed using simple roasting of complex copper sulfides ores of mesh size 300.
The concentration 2.5 percent of sulfuric acid was used to determine the effect of solid
liquid ratio on the extraction of copper. The leaching was done for two minutes. The solid
liquid ratio was varied from 1:1 to 1:10. The results indicate that the extraction of copper
was 49 percent at solid liquid ratio of 1:2.5. The further increase of solid liquid ratio from
1:5, only a slight increase in copper extraction was found. From the leaching of copper it
was found that a solid liquid ratio of 1:5 offered a remarkable operating conditions for
simple roasted materials at which 73 percent copper extraction was achieved. The results
are shown in table-6.
Table 6: Effect of Solid Liquid Ratio on Leaching of Simple Roasted Ore
S. No. Solid Liquid Ratio Copper Leaching (%)
1 1:1.0 22.65
2 1:1.5 32.40
3 1:2.5 48.85
4 1:3.5 62.25
5 1:5.0 72.85
6 1:10. 75.76
Roasted Sample C at 650oC & Roasting Time = 30 min, Particle size = 300 mesh,
Leachant = 2.5% H2SO4, Leaching Time = 02 min
101
Effect of particle size of simple roasted ore on copper leaching: A number of tests
were performed to find out a particle size range, which would give maximum copper
extraction. The complex copper sulfides ore was pulverized to obtained samples passing
through 100, 150, 200 and 300 mesh sieves. The leaching tests were conducted at liquid
solid ration of 5:1 using 2.5 percent sulfuric acids solution for two minutes leaching time.
It was observed that leaching action was slow on larger roasted particle size. The effect of
particle size on the copper leaching from complex copper sulfides ores was presented in
table- 7. The finer particle of size minus 300 mesh shows optimum extraction of copper
of roasted materials. Further decrease in size indicates that copper extraction has becomes
constants. It was also found that the roasting of larger particle shows low rate of
extraction of copper.
Table 7: Effect of Particle Size of Simple Roasted Ore on Copper Leaching
S. No. Particle Size (mesh) Copper Leaching (%)
1 -150 +200 45.68
2 -200 +250 68.67
3 -250 +300 72.85
4 -300 +350 73.35
Roasted Sample C at 650oC & Roasting Time = 30 min, Solid Liquid Ratio = 1:5,
Leachant = 2.5% H2SO4, Leaching Time = 02 min
102
SECTION-II
The copper extractions only up to 73 percent could be achieved after simple roasting of
complex copper sulfides ore. The roasting was therefore conducted using different
reagents of increase extraction by oxidation roasting, e.g. oxidants such as persulfate,
sulfate, chlorate, nitrate and nitrite. Stronger oxidants such as hydrogen peroxides, ozone,
Caro’s acid (H2SO5), persulfate, permanganate, dichromate, manganese dioxide were of
even less interests because of their cost economics (Peters, 1992).
4.4.2.2. Roasting with Oxidant
A series of experiments were performed using commonly used oxidants such as sodium
per-sulfate, potassium chlorate, sodium chlorate, sodium nitrite and sodium nitrate, to see
if these enhance the copper extraction from 73 percent. However higher extraction above
73 percent could not be obtained after roasting with these oxidants using roasting
temperature of 650ºC. The sodium per-sulfate and potassium chlorate gave approximately
68 percent copper extraction. The sodium nitrite and sodium chlorate show even less
copper extraction. The sodium nitrate did not show any significance effect. Since they
decreased the copper extraction, it may be possible due to formation of oxidized layer on
particles which resisted leaching. The decrease in copper extraction may not be possible
due to formation of intermediated minerals species, which may not be leached by 2.5
percent sulfuric acid with 1:5 solid liquid ratio within two minutes. The results are
presented in table- 8.
103
Table 8: Effect of Oxidant during Roasting on Copper Leaching
S. No Sample C + Oxidant Copper Leaching (%)
1 Sample C without oxidant 72.85
2 C + Sodium Per-sulfate 68.72
3 C + Potassium Chlorate 68.83
4 C + Sodium Chlorate 47.15
5 C + Sodium Nitrate 11.45
6 C + Sodium Nitrite 51.13
Leachant = 2.5% H2SO4, Solid Liquid Ratio = 1:5, Leaching Time = 02 min
Amount of Oxidant = 20% of sample’s weight, Roasting Time = 30 min
Sample C & Particle size = 300mesh, Roasting Temperature = 650ºC
Results of lower temperature oxidant roasting at longer time period: The effect of
lower temperature was studied. The roasting with oxidant was carried out at low
temperature 500oC as compared to values given in the literature (Norrgran, 1985, US-
patent) for longer time roasting of three hours. The behavior of oxidants, at the
temperature of 500oC shows that the copper extraction was 70 percent with sodium per
sulfates, 69 percent with sodium chlorate and potassium chlorate and 65 percent with
sodium nitrite. The use of oxidants during roasting lowers 150oC of the temperature. The
results are shown in table-9.
Table 9: Results of Lower Temperature Oxidant Roasting for Longer Time Period
S.No Sample C + Oxidants Copper Leaching (%)
1 C + Sodium Per-sulfate 70.15
2 C + Sodium Nitrite 64.53
3 C + Potassium Chlorate 69.85
4 C + Sodium Chlorate 69.90
Leachant = 2.5% H2SO4, Solid Liquid Ratio = 1:5, Leaching Time = 02 min
Roasting Time = 3 hour, Amount of Oxidant = 20% of sample’s weight
Sample C & Particle size = 300 mesh, Roasting Temperature = 500oC
104
4.4.2.3. Roasting with Novel Oxidant
A novel idea developed during investigations was to employ such chemical whose active
dissociation product readily react with sulfide minerals. A number of oxidants were
studied for use as additive during roasting. These chemicals provide different oxides on
heating, which react with sulfide minerals. The study was conducted using new material,
not investigated so far during roasting of sulfides ores. The ammonium nitrate was used
during roasting of complex copper sulfide ore. This novel roasting material as the oxidant
has not been used before by previous workers as no information is available in the
literature.
The novel oxidant (Ammonium Nitrate) was used during roasting for different
time periods. The optimum roasting temperature was found to be much lower i.e. 500oC
the previously obtained. The amount of oxidant required was 10 percent of the sample
weight. The copper extraction achieved was 21 percent, 52 percent, 77 percent and 75
percent after roasting for 60, 120, 180 and 240 minutes respectively. The copper leaching
was increased with increase the roasting time up to 180 minutes; however on further
roasting the copper extraction was decreased. The possible reaction, which changes
mineralogy, was during 180 minutes roasting. It showed favorable leaching with 2.5
percent sulfuric acid 1:5 solid liquid ratio in two minute leaching time. It may be possible
to increase the leaching by changing the roasting and leaching conditions. But it was
found after detailed investigations, that no further significance increased could be
achieved in this study. The results are shown in table-10. The XRD results are shown in
figure-5 and microphotography in figure-9.
105
Table 10: Effect of Novel Oxidant and Roasting Time on Copper Leaching
S. No Roasting Time (min) Copper Leaching (%)
1 60 21.63
2 120 52.15
3 180 77.72
4 240 75.26
Ammonium Nitrate = 10% of sample’s weight, Leachant = 2.5% H2SO4
Sample C & Particle size = 300 mesh, Roasting Temperature = 500oC
Solid Liquid Ratio = 1:5, Leaching Time = 02 min
Effect of amount of novel oxidant roasting on copper leaching: A series of
experiments were performed to study the effect of the amount of oxidant on copper
leaching. The amount of oxidant was varied as five percent, seven percent, 10 percent, 20
percent and 30 percent. The copper extraction obtained in this study show as 28 percent,
59 percent, 77 percent, 71 percent and 62 percent respectively. The optimum amount of
oxidant was found to be 10 percent. The results are presented in table-11.
Table 11: Effect of Amount of Novel Oxidant Roasting on Copper Leaching
S. No Amount of Oxidant (%) Copper Leaching (%)
1 5 28.36
2 7 59.23
3 10 77.72
4 20 71.25
5 30 62.85
Leachant = 2.5% H2SO4, Solid Liquid Ratio = 1:5, Leaching Time = 02 min
Sample C & Particle size = 300 mesh, Oxidant = Ammonium Nitrate
Roasting Temperature = 500oC, Roasting Time = 180 minutes
106
Effect of temperature and time on novel oxidant roasting: The effect of roasting
temperature and time was studied by keeping all other conditions of oxidant roasting the
same. The temperatures of roasting were 400oC, 450oC and 500oC. For the favorable
reaction between ore particle and oxidant, the optimum roasting temperature was found
to be 500oC at 180 minutes roasting time. The results are shown in table-12.
Table 12: Effect of Temperature and Time on Novel Oxidant Roasting
S. No. Temperature oC
Copper Leaching (%) at Different Roasting
Temperature & Time
60 min 90 min 120 min 180 min
1 500 21.63 29.4 52.15 77.72
2 450 10.5 20.8 35.4 52.6
3 400 4.15 13.6 24.8 42.2
Sample C & Particle size = 300 mesh, Leachant = 2.5% H2SO4
Oxidant = Ammonium Nitrate = 10% of sample’s weight
Solid Liquid Ratio = 1:5, Leaching Time = 02 min
107
Effect of particle size in novel oxidant roasting: The effect of particle size on leaching
of copper was studied. Using different particle size of -150+200, -200+250, -250+300
and -300 + 350 mesh, the copper leaching achieved was 78 percent with -200+250 mesh
size keeping other conditions of roasting and leaching the same. It was found that the
oxidant has more significant effect at the particle size of -200+250 mesh range. There
was no significant increase in copper leaching by using 300 mesh particle size. The fine
particle size probably hinders the flow of gaseous product below particles. This decreases
the oxidation of reaction during roasting. It was found that the oxidant roasting had no
need of fine or ultra fine grinding of ore samples. The results are shown in table-13.
Table 13: Effect of Particle Size in Novel Oxidant Roasting
S. No Particle Size (mesh) Copper Leaching (%)
1 -150 +200 72.26
2 -200 +250 77.84
3 -250 +300 77.72
4 -300 + 350 77.63
Roasting Time = 180 minutes, Solid Liquid Ratio = 1:5, Leaching Time = 02 min
Sample C & Roasting Temperature = 500oC, Leachant = 2.5% H2SO4
Oxidant = Ammonium Nitrate = 10% of sample’s weight
108
Effect of different acids with different time on novel oxidant roasting: The
dissolution of roasted material in different acids is shown in table-14 by using one minute
and 180 minutes time. The five percent solution of HNO3, H2SO4, HCl and HCl+HNO3
was used. The copper extraction by leaching achieved was 73 percent, 77 percent, 61
percent, 78 percent respectively for one minute leaching time. When the leaching time
was three hours the copper leaching was 74 percent, 78 percent, 64 percent, and 79
percent respectively. The results show that the copper extraction is greater with dilute
sulfuric acid than the dilute nitric acid. It was found that no significant leaching was
achieved by increasing the acid concentration and time. The results are shown in table-
14.
Table 14: Effect of Different Acids with Different Time on Novel Oxidant Roasting
S. No Acids Copper Leaching (%)
01 min 180 min
1 5%HNO3 73.85 74.23
2 5%H2SO3 77.36 78.56
3 5%HCl 61.59 64.25
4 5%(HCl+HNO3) 78.45 79.23
Oxidant = Ammonium Nitrate = 10% of sample’s weight, Solid Liquid Ratio = 1:5
Sample C & Roasting Temperature = 500oC, Roasting Time = 180 minutes
109
A novel method was developed based on the experiments using roasting with non-oxidant
material / additives. The method is based on solid state reactions of sulfides ores minerals
with additives which generate vapors containing H2O, NH3, HNO3, NOx, HSO4, SOx,
HCl, Cl2 and chloride. A number of solid vapor phase reactions enhance the met stability
of minerals. The over all process not only gave leachable product at lower temperature
and time as compare with roasting temperature of chalcopyrite and other complex copper
sulfides ores given in the literature. The study of novel additive roasting is shown as
follows;
SECTION-III
4.4.2.4. Roasting with Ammonium Compounds (Novel Additives)
A series of experiments were performed with different chemical compounds containing
ammonia/ammonium as an additive for roasting of complex copper sulfide ore. These
chemicals (novel additives roasting) are chloride, sulfate, acetate, carbonate, oxalate,
phosphate of ammonium. The parameters of roasting with ammonium nitrate were
studied (section-II). The novel additive ammonium chloride for roasting was found
suitable, which gave good results. This roasting is called in this thesis as novel additive
roasting for convenience. The result showed up to 85 percent copper extraction. The
additive which has chloride ion, showed better results. It may be due to its action on
complex copper sulfide minerals, which ultimately increases its leachability. The
comparative results of additives are presented in table-15. The XRD and
110
microphotographic results of (S. No. 1) of table-15 are shown in figure-5, 9 and the XRD
results of (S. No. 2) of table-15 are shown in figure-6, 10 respectively.
Table 15: Effect of Novel Additives during Roasting on Copper Leaching
S.No Sample C + Additive Copper Leaching (%)
1 C + Amm. Nitrate 77.72
2 C + Amm. Chloride 85.16
3 C + Amm. Sulfate 53. 62
4 C + Amm. Phosphate 53.23
5 C + Amm. Acetate 38.65
6 C + Amm. Oxalate 36.47
7 C + Amm. Carbonate 27.56
8 C + Urea 26.14
Sample C & Particle size = 300mesh, Roasting Temperature = 500oC
Roasting Time = 180 minutes, Amount of Additive = 10% of sample’s weight
Leachant = 2.5% H2SO4, Solid Liquid Ratio = 1: 5, Leaching Time = 02 min
Effect of roasting time on novel additive roasting: The effect of roasting time was
studied by varying time as 30, 90, 120, 150 and 180 minutes. The extraction of copper
was obtained 53 percent after 30 minutes roasting time. The extraction of copper
becomes constant after 90 minutes roasting. It was found that additive ammonium
chloride decreased the roasting time to half as compared to ammonium nitrate, because
the additive (ammonium nitrate) took 180 minutes for interaction with ore particles. This
novel additive having chloride ion appears to deform the mineral structure of complex
copper sulfide ore in 90 minutes. It may be due to the formation of new mineral phases
during roasting. The results are presented in table-16.
111
Table 16: Effect of Roasting Time on Novel Additive Roasting
S. No Roasting Time (min) Copper Leaching (%)
1 30 52.95
2 90 84.51
3 120 84.77
4 150 84.93
5 180 85.16
Sample C & Particle size = 300 mesh, Roasting Temperature = 500oC
Leachant = 2.5% H2SO4, Solid Liquid Ratio = 1: 5, Leaching Time = 02 min
Additive = Ammonium chloride =10% of sample’s weight
Effect of leaching time on novel additive roasting: The effect of leaching time of
roasted sample with additive ammonium chloride was studied on copper extraction. The
maximum copper extraction was achieved within a 60 to 120 seconds leaching time.
There was no significant change in leaching of copper after 120 seconds. The change of
mineralogy of complex copper sulfide ore after roasting with this additive resulted an
instant leaching after roasting with ammonium chloride additive. The results are shown in
table-17.
Table 17: Effect of Leaching Time on Novel Additive Roasting
S. No Leaching Time (Sec) Copper Leaching (%)
1 60 84.98
2 120 85.16
3 300 85.36
4 900 85.98
5 1800 86.42
6 2700 87.26
Sample C & Particle size = 300mesh, Roasting Temperature = 500oC
Roasting Time = 180 minutes, Ammonium chloride =10% of sample’s weight
Leachant = 2.5% H2SO4, Solid Liquid Ratio = 1: 5, Leaching Time = 02 min
112
Effect of particle size on novel additive roasting: The particle size of -200 +250
showed 96 percent copper extraction. The effect of particle size showed increase in
copper extraction with increase in mesh size. The maximum copper extraction was
obtained with -200 +250 mesh particle size. The decrease was observed in copper
extraction by using fine particle size of 300 mesh and above. The low results indicate the
formation of muscovite (ref. fig. 6, XRD, p-145) and silicate minerals. It is due to
coagulation of finer particles in solid state reaction hinders the leaching. It may be due to
different mineralogical composition in fine particle size. The results are presented in
table-18.
Table 18: Effect of Particle Size on Novel Additive Roasting
S. No Particle Size (mesh) Copper Leaching (%)
1 -150 +200 92.38
2 -200 +250 95.76
3 -250 +300 85.25
4 -300+350 85.16
Leachant = 2.5% H2SO4, Solid Liquid Ratio = 1: 5, Leaching Time = 02 min
Sample C, Roasting Temperature = 500oC, Roasting Time = 180 minutes
Additive = Ammonium chloride =10% of sample’s weight
113
Effect of temperature and time on novel additive roasting: The effect of roasting
temperature was studied by keeping all other conditions of additive roasting the same.
The temperature used was 400oC, 450oC, 500oC. No significant change was found in
copper extraction by lowering the roasting temperature. The favorable reaction between
ore particle and ammonium chloride additive was found to be 500oC and 90 minutes
roasting time. The results are shown in table-19.
Table 19: Effect of Temperature and Time on Novel Additive Roasting
S. No. Temperature oC
Copper Leaching (%) at Different Roasting
Temperature & Time
60 min 90 min 120 min 180min
1 500 52.9 84.5 84.8 85.2
2 450 37.9 64.1 64.2 65.7
3 400 20.3 40.8 49.3 53.1
Ammonium chloride =10% of sample’s weight, Sample C & Particle size =300 mesh
Leachant = 2.5% H2SO4, Solid Liquid Ratio = 1:5, Leaching Time = 02 min
114
4.4.2.4.1 Two-Stage Roasting
The effect of various ranges of lower temperatures was studied during roasting using
novel additives. The process was developed of two stage roasting at various lower
temperatures and shorter roasting time as compared to what is reported in literature
(Norrgran, 1985. US- patent).
The two stage roasting treatment was applied to obtain lower temperatures of
roasting. The additives ammonium nitrate, ammonium chloride and ammonium sulfate
were used during roasting. The roasting and leaching conditions were kept the same. The
roasting temperature of 350oC was kept for 30 minutes and then the temperature was
raised and maintained at 500oC for 45 minutes. This two stage roasting treatment was
found suitable only to roasting with ammonium chloride as it gave better leaching. This
treatment showed 93 percent copper extraction. As compared to it the other additive did
not show better extraction of copper. The additive ammonium nitrate showed only 31
percent and the additive ammonium sulfate showed 26 percent copper extraction. The
results are presented in table-20.
Table 20: Effect of Two-Stage Roasting on Copper Leaching
S. No Sample C+ Additive Copper Leaching (%)
1 C + Amm. Nitrate 31.11
2 C + Amm. Chloride 93.08
3 C + Amm. Sulfate 26.35
Amount of Additive = 10% of sample’s weight, Particle Size = 200 mesh
Leachant = 2.5% H2SO4, Solid Liquid Ratio = 1: 5, Leaching Time = 02 min
Roasting 30min at 350oC.then 45min at 500oC, Sample C
115
Effect of particle size during two-stage roasting: The effect of particle size of sample
during two stage roasting was studied. The maximum copper leaching of 97 percent was
obtained with –200 +250 mesh range of particle size. It may be due to availability of
additive and interactions for long time with ore particles at lower temperature and there
was no need of fine grinding. The results are presented in table-21.
Table 21: Effect of Particle Size during Two-Stage Roasting
S. No Particle Size Copper Leaching (%)
1 -150 +200 93.08
2 -200 +250 97.53
3 -250 +300 84.56
4 -300+350 84.35
Leachant = 2.5% H2SO4, Solid Liquid Ratio = 1: 5, Leaching Time = 02 min
Roasting 30min at 350OC, then 45min at 500OC, Sample = C
Ammonium Chloride = 10% of sample’s weight
116
Effect of temperature during two-stage roasting: The effect of lowering the
temperature of two stage roasting was studied. It was found that the lower temperature
range up to 325oC—330oC at which the copper extraction was 90 percent. The temperature
lowers than 325oC did not show significant copper extraction. The both interaction between ore
particles and additive was in the temperature range 325oC–350oC. The results are presented in
table-22. The XRD analysis of (S. No. 1) of table-22 are shown in figure-7
Table 22: Effect of Temperature during Two-stage Roasting
S. No Two-stage Roasting
Temperatures
Copper Leaching
(%)
1 30min at 350oC. 45min at 500oC.
84.35
2 30min at 325oC. 45min at 500oC
90.34
3 30min at 300oC. 45min at 500oC.
38.68
Leachant = 2.5% H2SO4, Solid Liquid Ratio = 1: 5, Leaching Time = 02 min
Sample = C, Additive = Ammonium Chloride = 10% of sample’s weight
Particle size = 300 mesh
117
Comparison of copper leaching at different roasting treatments: The comparison of
roasted sample with additive ammonium chloride was made between two stage roasting
treatment and single stage for 90 minutes at 500oC. There was no significant difference
was seen in the copper leaching of both treatments. It was found that the two stage
roasting has been much economic process for copper extraction of sulfide ores. The
results presented in table-23.
Table 23: Comparison of Copper Leaching at Different Roasting Treatments
S. No
Acid
Concentration
(%)
Leaching
Time
(Seconds)
Cu-Leaching (%) 30 min at 350
OC.
45 min at 500OC.
Cu-Leaching (%)
90 min at 500OC
1 2.5 % H2SO4 60 83.75 82.89
2 2.5 % H2SO4 120 84.35 84.51
3 2.5 % H2SO4 300 85.39 84.97
4 2.5 % H2SO4 900 85.50 85.02
5 2.5 % H2SO4 2700 86.01 86.78
6 5 % H2SO4 300 86.44 86.35
Sample = C, Additive = Ammonium Chloride = 10% of sample’s weight
Leachant = H2SO4, Solid Liquid Ratio = 1: 5
118
Three novel additives were selected to study the combined effect on copper sulfide ore at
two stage roasting. The combined effect of optimal roasting was also studied at various
temperatures. The results are shown in section-IV as follows;
SECTION-IV
4.4.2.4.2. Effect of Combined Oxidant / Additive on Copper Leaching
The combined effect of the additives on copper leaching was investigated. The effect of
following three additives, ammonium nitrate, ammonium chloride and ammonium sulfate
were studied with amounts of additives. The total amount of additives range was 7
percent to 15 percent. It was found by applying different mixtures of additives that the
ammonium chloride could be reduced to 2.5 percent of ore sample weight for better
leaching. The best combination of additives was selected as chloride and sulfate of
ammonium (S. No: 5, table-24). The combination showed the maximum copper leaching
of 77.75 percent.
Optimum Conditions: The combined effect of additives show better results order, which
are summarized as follows;
NH4Cl +(NH4)2SO4 > NH4NO3+NH4Cl+(NH4)2SO4 = NH4NO3 + NH4Cl + (NH4)2SO4
2.5% + 5% 5% + 2.5% + 5% 5% + 5% + 5%
The optimal combination is NH4Cl + (NH4)2SO4.
The combined action of additives has advantages due to following;
i- Percentage of additive used was less than 10 percent.
119
ii- The amount of required additive NH4Cl is only 2.5 percent. As the chloride ion
has corrosive action during chemical process is in industrial level operations, it
had been minimized by the reducing percentage of chloride by applying this
process.
It was found that the combination of chloride-sulfate in aqueous media of leaching had
shown better copper leaching. The same effect was found in prior roasting with chloride-
sulfate of ammonium additives in subsequent leaching. The solid-state reactions involve
very complex chemistry (West, 1984). The present research has presented the results for
copper extraction from indigenous copper complex sulfide ores by using new chemicals
not reported in the literature during solid state reactions. The results obtained are
presented in table-24 and XRD of (S. NO. 5) of table-24 in figure-8.
Table 24: Effect of Combined Oxidant/Additive on the Copper Leaching
S. No
Sample C + Additives (%)
Amm. Amm. Amm.
Nitrate Chloride Sulfate
Copper
Leaching
(%)
1 2.5 5 2.5 70.39
2 5 5 5 73.07
3 5 2.5 5 73.93
4 5 2.5 - 62.89
5 - 2.5 5 77.75
6 5 - 5 58.05
7 - 10 - 93.08
8 10 - - 31.11
9 - - 10 26.35
Roasting 30 min at 350OC, then 45 min at 500OC, Sample C & Particle Size = 200 mesh
Leachant = 2.5% H2SO4, Solid Liquid Ratio = 1: 5, Leaching Time = 02 min
Additives = Ammonium nitrate, ammonium chloride, ammonium sulfate
120
Combinations of additive at different temperatures’ conditions: The effect of
different combination of additives at different treatment of temperatures was studied. The
combinations of chloride-nitrate and nitrate-sulfate have significantly higher leaching for
three hours roasting at 500oC. The group of chloride-sulfate gave 78 percent copper
leaching on two stage roasting lower temperature. The group of three additives nitrate-
chloride-sulfate was much effective for three hours roasting at 500oC. It gave 99 percent
copper extraction. But this combination gave 73 percent copper extraction at two stage
lower temperature treatment. The treatment of two stage roasting was economically
suitable because the roasting time (1.5hrs) at high temperature (500oC) become half (45
minutes) as compared with two stage roasting time at 500oC. The results obtained are
presented in table-25. The microphotography of (S. No. 2) of table-25 at 500oC for three
hours roasting treatment is shown in figure-11.
Table 25: Combinations of Additive at Different Temperatures’ Conditions
S.
No
Combinations
of
Additives
Cu-Leaching (%) 30min at 350
oC.
45min at 500oC.
Cu-Leaching
(%)3hours at
500oC
1 NH4NO3+NH4Cl 62.89 85.79
2 NH4Cl +(NH4)2SO4 77.75 86.04
3 NH4NO3+(NH4)2SO4 58.05 71.81
4 NH4NO3+NH4Cl+(NH4)2SO4 73.26 98.95
Sample C & Particle Size = 200 mesh, Leachant = 2.5% H2SO4
Solid Liquid Ratio = 1: 5, Leaching Time = 02 min
121
Effect of particle size on roasting with [Sample C + NH4Cl + (NH 4)2SO4]: The effect
of particle size of sample with NH4Cl + (NH4)2SO4 additives on copper leaching was
studied. The additive roasting gave 78 percent copper leaching on particle size of +200
mesh. The roasting of finer particle was conducted the copper leaching obtained less than
large particles of 200 mesh. The results obtained are presented in table-26.
Table 26: Effect of Particle Size on Roasting with [Sample C + NH4Cl + (NH4)2SO4]
S. No Particle Size
(mesh)
Copper Leaching
(%)
1 -150 +200 77.75
2 -200 +250 76.57
3 -250 +300 75.94
4 -300 +350 74.65
Roasting 30 min at 350oC then 45 min at 500oC, Leachant = 2.5% H2SO4
Solid Liquid Ratio = 1: 5, Leaching Time = 02 min
122
Comparison of roasting with additives NH4Cl and [NH4Cl + (NH4)2SO4]: The
comparison of copper leaching was made between roasting with additives NH4Cl and
NH4Cl + (NH4)2SO4. The treatments of roasting were different as described in table-27.
The copper leaching was increased when temperature decrease to 325oC but when the
ore-additive material was roasted at 325oC for 30 minutes the copper leaching was only
17 percent and 14 percent by NH4Cl and NH4Cl +(NH4)2SO4 respectively. It showed the
major changes were during the temperature range of 325oC to 500oC. The results are
presented in table-27.
Table 27: Comparison of Roasting with Additives NH4Cl and [NH4Cl + (NH4)2SO4]
S.No Different
Conditions of
Roasting
Cu-Leaching%
NH4Cl
+ Sample C
Cu-Leaching%
NH4Cl+(NH4)2SO4
+ Sample C
1 30min at 350oC.
45min at 500oC. 84.35 77.75
2 30min at 325oC. 45min at 500oC
90.34 78.45
3 30min at 300oC. 45min at 500oC.
38.68 25.16
4 30min at 325oC. 17.24 14.29
5 3 hours at
500oC
85.16 86.04
6 1.50 hours at
500oC
84.51 74.53
Leachant = 2.5% H2SO4, Solid Liquid Ratio = 1: 5, Leaching Time = 02 min
123
4.4.2.4.3. Leaching of Copper Chalcopyrite Concentrate
A detailed study of simple roasting, oxidative roasting and additive roasting was
conducted then the sample of copper concentrate chalcopyrite was investigated through
the optimal conditions. It was found that the amount of chloride additive was increase to
three times. The copper leaching 99.95 percent was achieved with 30 percent amount of
chloride additive. The copper leaching again decreased with increase the amount of
additive more than 30 percent. The results obtained are presented in table-28.
Table 28: Leaching of Copper Chalcopyrite Concentrate
S. No Different Conditions of
Roasting
%Cu-Leaching
Sample E
1 Sample E + 3 hr at 500oC. 39.79
2 Sample E + NH4Cl (10%) + 3 hr at 500oC
48.47
3 Sample E + NH4Cl (30%) + 3 hr at 500oC
99.95
4 Sample E + NH4Cl (40%)
+ 3 hr at 500oC
65.12
5 Sample E + NH4Cl +
(NH4)2SO4 (5+10%) + 3 hr
at 500oC
98.65
Leachant = 2.5% H2SO4, Leaching Time = 02 min
Sample E, Solid Liquid Ratio = 1: 5
124
4.4.2.4.4. Amount of Sulfur Analyzed in Different Samples
The amount of sulfur was investigated in different treated and untreated samples. For the
study of sulfur dioxide emission, the amount of sulfur was compared in different
experiments are listed in table-29. The sample C has 20 percent amount of sulfur. The
treated samples have no sulfur emissions during roasting as shown by the amount of
sulfur after roasting. The novel additives probably fix the sulfur of sulfides into sulfates
or other mineral species in the reactions. The mineral phase cubanite was also found in
treated sample C with ammonium chloride additive at 500oC for three hours roasting as
shown in XRD analysis Figure-6.
Table 29: Amount of Sulfur Analyzed in Different Samples
S. No Different Conditions of
Roasted Samples
Amount of Sulfur
(%)
1 Untreated Sample C 20.15
2 Sample C + NH4Cl
+ 3 hr at 500oC
20.14
3 Sample C + (NH4)2SO4
+ 3 hr at 500oC
20.16
4 Sample C + NH4NO3
+ 3 hr at 500oC
17.09
5 Sample C+ NH4Cl +
(NH4)2SO4 + 3 hr at 500oC
20.12
6 Sample C + NH4Cl +
(NH4)2SO4 at 350oC &
500oC
20.17
7 Sample C + NH4Cl 350oC & 500oC
20.16
8 Sample E 10.56
9 Sample E + NH4Cl (30%) + 3 hr at 500oC
10.54
125
The literature shows that oxide minerals of copper can be easily leached with sulfuric
acid as compared to sulfide minerals with mild acid at ambient conditions. The acid
leaching includes the evaluation of the important parameters affecting the rate of leaching
such as particle size, acid concentrations and solid liquid ratio. The effect of oxidizing
agents, temperature and pressure were also investigated. The results are shown in section-
V as follows;
SECTION-V
4.4.1.1 Sulfuric Acid Leaching
The copper ore samples were subjected to sulfuric acid leaching at different conditions of
concentration of sulfuric acid, temperature, pressure, oxidizing acid and leaching time.
Leaching of Copper Ores using Sulfuric Acid: The leaching tests were carried out in a
Pyrex beaker with sulfuric acid of known strength. The slurry was agitated with a glass
stirrer and sample was drawn with a pipette attached with a porous plug, after 2, 5, 10, 20
and 30 minutes agitation and preserved for subsequent analysis. The samples were
analyzed by using standard analytical methods. All the leaching tests were carried out at
room temperature (35ºC).
Effect of Acid Concentration and Time: Ore samples passing 300 mesh size were
arbitrarily selected to study the leachability of samples A, B, C, D and E. The leaching
was done with 2.5, 5.0, 7.5 and 10 percent sulfuric acid and at a tentatively selected
126
liquid-solid ration of 2.5:1, for a period of 10 min and agitation speed of 1000 revolution
pr minute. The results indicate that there is a marked difference in the initial rate of
copper extraction when using acids of strength higher than 2.5 percent. Acid strength
higher than 5.0 percent does not show a significant increase in the copper extraction. It is
apparent from the results that maximum copper is leached out within 10 min using even
2.5 percent sulfuric acid. The copper extraction goes as high as 34 percent from oxidized
copper ore samples.
Results given in table-30 indicate that an increase in the initial concentration of
acid is not only unnecessary, but in fact, harmful as it results in an increase in the
dissolution of impurities and thereby higher consumption of the acid per kg of copper
leached. It is, therefore, inferred that no special advantage is obtained by using acid
concentration above 2.5 percent. From the consideration of reaction rate it has been
observed that the rate of dissolution of copper increases with an increase in the acid
concentration. However, the rate becomes almost constant when leaching was carried out
with an acid concentration higher than 5 percent. The results are shown in table-30.
127
Table 30: Effect of Time & H2SO4 Concentration on Copper Leaching
S. No Sample H2SO4
Conc. (%)
Copper Leaching (%)
2 min 5 min 10 min 20 min 30 min
1
A
2.5 15.6 20.8 25.9 29.7 30.3
2 5.0 20.5 24.6 27.9 30.1 31.4
3 7.5 26.12 28 29.8 31.1 32.7
4 10 27.7 29.2 30.8 32.2 33.4
1
B
2.5 11.36 17.3 21.19 24.38 25.68
2 5.0 16.89 21.32 24.8 27.68 28.49
3 7.5 20.12 23.57 26.4 28.36 29.35
4 10 21.17 24.9 27.85 30.21 31.12
1
C
2.5 2.7 3.1 3.4 3.5 3.5
2 5.0 3.8 4.13 4.58 4.9 5.1
3 7.5 4.3 4.7 5.02 5.34 5.6
4 10 4.9 5.1 5.4 5.62 5.7
1
D
2.5 6.83 8.09 9.4 10.7 11.58
2 5.0 7.98 9.21 10.40 11.45 12.45
3 7.5 8.25 9.87 11.25 12.33 13.14
4 10 8.67 10.23 11.73 12.98 13.87
1
E
2.5 2.3 2.81 3.1 3.31 3.4
2 5.0 3.4 3.87 4.2 4.51 4.8
3 7.5 3.9 4.1 4.6 4.72 4.91
4 10 4.3 4.5 4.75 4.8 5.23
Samples = A, B, C, D, E & Particle size = minus300 mesh, Temperature = 35ºC
Effect of liquid-solid ratio during sulfuric acid leaching: A number of tests were
performed using ground ore and 2.5 percent sulfuric acid to determine the effect of liquid
solid ratio by weight on the extraction of copper. The liquid solid ratio was varied from
2:1 to 10:1. The leaching was done for 10 min and the results indicate that increasing the
liquid solid ratio above 3:1 causes only a slight improvement in the copper extraction.
This may be due to that the particle at the beginning are leached superficially, the
concentrated solution enveloping the particle is displaced by the fresh leachant of lower
copper content under the strong agitation conditions. After the surface of the particle is
128
leached a concentration gradient is established at the surface and the interior of the
particle, and the leaching rate becomes diffusion-controlled. The increase in liquid solid
ratio, therefore, will have no marked effect on the leaching rate of the ore as long as the
particle is constantly submerged in the acidic solution. The results are shown in table-31.
Table 31: Effect of Liquid Solid Ratio during Sulfuric Acid Leaching
S. No Liquid-Solid
Ratio
Copper Leaching (%)
A B C D E
1 2.0/1 13.4 11.7 0.8 2.6 0.7
2 2.5/1 25.9 21.9 3.4 9.4 3.1
3 3.0/1 27.4 23.2 3.7 9.9 3.5
4 3.5/1 27.4 23.3 3.9 10.4 3.8
5 4.0/1 27.4 23.6 4.1 10.9 4.1
6 5.0/1 27.9 24.1 4.5 11.2 4.4
7 10.0/1 28.4 24.9 4.9 11.8 4.6
Particle size = minus300 mesh, Temperature = 35ºC, Leachant = 2.5% H2SO4
Leaching Time = 30 min, Samples = A, B, C, D, E
Effect of heat during sulfuric acid leaching: The leaching of samples A, B, C, D and E
was further conducted by heating at temperature of 100ºC. The leaching tests were
performed at a liquid solid ratio of 2.5:1 using 2.5 percent sulfuric acid. The leaching
time was extended up 30 minutes due to observe the leaching action on high temperature.
The oxides and sulfides ore samples did not show the remarkable increase in extraction of
copper. It is evident that the oxide samples were leached totally with mild sulfuric acid at
ambient conditions, while the sulfide ore samples were not leached even at boiling. The
results are shown in table-32.
129
Table 32: Effect of Heat during Sulfuric Acid Leaching
S. No Samples Copper Leaching (%)
2 min 5 min 10 min 20 min 30 min
1 A 21.5 21.68 22.17 22.18 22.25
2 B 17.81 17.92 18.08 18.28 18.49
3 C 3.91 3.98 4.18 4.59 4.61
4 D 8.12 8.21 8.49 9.05 9.45
5 E 3.71 3.75 3.92 4.12 4.33
Leachant = 2.5% H2SO4, Liquid-Solid Ratio = 2.5/1, Samples = A, B, C, D, E
Particle size = 300 mesh, Temperature = 100ºC
Sulfuric acid leaching under pressures and temperatures: A number of tests were
conducted on the leaching. The experiments were conducted in Bomb digester by taking
copper ore sample and different acids under pressure by placing in boiling water. The
different parameters such as acidity, use of nitric acid as oxidizing agent, pressure and
temperature, liquid solid ratio, particle size, and leaching time period were experimented.
The samples were analyzed for copper leaching by using standard analytical methods.
Time required for heat transformation between bomb digester and water bath: The
ore sample was weighed and transferred to the Bomb digester. Then 1.25 ml of each of
10 percent HNO3 and 20 percent H2SO4 were added with solid liquid ratio of 1:5. It was
placed in the water bath already at the boiling temperature but having water level less
than the open end of Bomb digester. A thermometer was passed through the cutting hole
of card and then placed it over open end of the Bomb digester. Temperature was noted at
that time then temperature was noted at several times until it become equal to the water
bath temperature and time required for this transfer of heat was also noted. 15 minutes
were required for the 80ºC temperature rise of the internal system i.e. Bomb digester. It is
130
estimated that heat transfer between internal and external system, 20-25 minutes were
required. It means within 10-15 minutes 70 percent leaching was achieved and the
remaining time is utilized for heat transfer. Each experiment was performed for 30
minutes tentatively.
Effect of time under pressure and temperature leaching: Copper ores samples were
leached with 20 percent H2SO4 and 10 percent HNO3 for different time periods 30, 60,
120 and 180 minutes keeping all other conditions constant. The maximum amount of
leached copper was 67.25, 65.73, 62.58, 64.7 and 46.12 percent in 180 minutes for
Azurite, Malachite, complex copper sulfide ore, mixed ore and Chalcopyrite concentrate
samples respectively. The optimum amount of leached copper in 30 minutes was 60.46,
58.36, 55.25, 56.00 and 35.84 percent for A, B, C, D and E samples respectively. Thus by
increasing the time, rate of copper leaching can be increased. The results are shown in
table-33.
Table 33: Effect of Time during Sulfuric Acid Leaching under
Pressure and Temperature
Sample
Copper Leaching (%)
30min 60min 120min 180min
A 60.46 63.41 66.01 67.25
B 58.36 61.48 64.68 65.73
C 55.25 57.51 61.20 62.58
D 56.01 60.30 63.80 64.70
E 35.84 41.35 45.76 46.12
Solid Liquid Ratio = 1:5, Particle size = 300 mesh, Temperature = 96-100 ºC
H2SO4 conc. = 20%, Oxidizing agent (HNO3) conc. = 10%, Pressure = 70-80 atm
131
Effect of sulfuric acid concentration under pressure and temperature leaching: The
leachability of the copper ores sample was studied by using different concentrations of
H2SO4 as 10, 20 and 30 percent and a tentatively selected liquid solid ratio of 5:1 for a
period of 30 minute at 100ºC. The 10 percent HNO3 was added as an oxidant. The results
indicate that by increasing acid concentration, the leaching of copper increased but up to
20 percent then for 30 percent it was again decreased. For 10 percent H2SO4 the copper
extraction was 38.52, 25.45, 34.45, 32.55 and 28.14 percent for A, B, C, D and E samples
respectively. For 20 percent H2SO4 copper extraction for all the five samples was
maximum i.e. 60.46, 49.2, 52.25, 53.47 and 35.84 percent for A, B, C, D and E samples
respectively. But with 30 percent H2SO4 extraction copper in all cases decreased. Hence
20 percent H2SO4 is most favorable to achieve optimum copper extraction. The results are
shown in table-34.
Table 34: Effect of Sulfuric Acid Concentration under Pressure and Temperature
Sample
Mark
Sulfuric Acid
Concentration (%)
Copper
Leaching (%)
A
10
20
30
38.52
60.46
53.90
B
10
20
30
30.45
49.20
38.25
C
10
20
30
34.45
55.25
44.95
D
10
20
30
32.55
53.47
49.17
E
10
20
30
28.14
35.84
32.14
Sample C, Solid Liquid Ratio = 1:5, Oxidizing agent (HNO3) conc. = 10%
Leaching Time = 30 min, Temperature = 96-100 ºC, Pressure = 70-80 atm
132
Effect of oxidizing acid under pressure and temperature leaching: The ore sample
was leached with H2SO4 and HNO3 as oxidizing agent. The addition of oxidizing agent
increases the leaching of copper. The different concentrations of oxidizing agent were
used as 5 percent, 10 percent and 15 percent, while the other conditions remained the
same. The optimum copper extraction was obtained at 10 percent of oxidizing agent as
38.52, 30.42, 34.45, 32.55 and 28.14 percent for the samples A, B, C, D and E
respectively. The results are shown in the table-35.
Table 35: Effect of Oxidizing Acid during Sulfuric Acid Leaching under
Pressure and Temperature
Sample
Mark
Oxidizing Acid (HNO3)
Concentration (%)
Copper
Leaching (%)
A
15
10
5
41.15
38.52
27.62
B
15
10
5
32.85
30.45
21.34
C
15
10
5
36.52
34.45
22.36
D
15
10
5
35.12
32.55
20.34
E
15
10
5
30.86
28.14
17.95
Leaching Time = 30 min, Temperature = 96-100 ºC, Pressure = 70-80 atm
Sample C, Solid Liquid Ratio = 1:5, Sulfuric acid conc. = 10%
133
Effect of liquid solid ratio under pressure and temperature leaching: A series of
experiments were performed using –300 mesh size with 10 percent HNO3 and 20 percent
H2SO4 to determine the effect of liquid solid ratio which was varied from 2.5:1 to 15:1.
The leaching was done for 30 minutes and copper extraction was determined at the end.
The results given indicate that leaching was 44.20 percent at 2.5:1, 55.25 percent at 5:1,
88.47 percent at 10:1 and 92.73 percent at 15:1 ratio. The results indicate that increasing
the liquid solid ratio increases the copper extraction. It was found that liquid solid ratio of
10:1 offered a reasonable operating condition for –300 mesh size. The results are shown
in the table-36.
Table 36: Effect of Liquid Solid Ratio during Sulfuric Acid Leaching under
Pressure and Temperature
S. No Solid Liquid
Ratio
Copper Leaching
(%)
1 1:2.5 44.20
2 1:5 55.25
3 1:10 88.47
4 1:15 92.73
Sample C & Particle size = -300+350 mesh, H2SO4 = 20%, HNO3 = 10%
Leaching Time = 30 min, Temperature = 96-100 ºC, Pressure = 70-80 atm
134
Effect of particle size under pressure and temperature: A number of tests were
conducted to find out a particle size range which could give maximum copper extraction.
Leaching tests were performed with 100, 150, 200, 300 and 350 mesh size. The liquid
solid ratio was 10:1 using 10 percent HNO3 and 20 percent H2SO4 for a time period of 30
minutes at 100ºC temperature and under > 70 atmosphere pressure. The effect of particle
size on the leachability of copper shows that ore sample of –300 mesh size gives
maximum leaching. The results are shown in the table-37.
Table 37: Effect of Particle Size during Sulfuric Acid Leaching under
Pressure and Temperature
S. No Particle
Size (mesh)
Copper Leaching
(%)
1 -50 +100 12.66
2 -100 +150 18.25
3 -150 +200 45.47
4 -200 +300 81.65
5 -300 +350 88.47
Leaching Time = 30 min, Temperature = 96-100 ºC, Pressure = 70-80 atm
Sample = C, Solid Liquid Ratio = 1:10, H2SO4 = 20%, HNO3 = 10%
135
Effect of temperature and pressure during sulfuric acid leaching: The ore sample
was leached first at room temperature only with ambient condition of pressure. All other
parameters i.e. liquid solid ratio, time period, oxidizing agent, acid concentration and
particle size kept constant. Then the leaching experiment was conducted at room
temperature with pressure by placing lid. The leaching test was conducted at 175¬180ºC
and about 90 atmosphere pressure. It was found from the results that by increasing the
temperature and pressure the leaching increased and optimum results were obtained at
100ºC. The results are shown in the table-38.
Table 38: Effect of Temperature and Pressure during Sulfuric Acid Leaching
S. No
Temperature (ºC)
Copper
Leaching (%)
1 Room Temp. (35) in open beaker 5.82
2 Room Temp. (35) in Bomb digester 15.26
3 Bomb digester in Water Bath (96-100) 55.25
4 Bomb Digester in Sand Bath (180) 86.97
Sample C & Particle size = -300+350 mesh, Solid Liquid Ratio = 1:5
Leaching Time = 30 min, H2SO4 = 20%, HNO3 = 10%
136
4.5. XRD Analyses and Microphotography
The XRD analyses of selected samples were carried out. The XRD analysis of the sample
C has shown the presence of Chalcopyrite, Chalcocite, Azurite, Tetrahedrite and
Malachite as copper minerals. The treated samples show, the formation of oxides and
sulfate of copper. The novel additive may have to fix the sulfur of sulfides into sulfates or
other sulfur rich mineral species. Cubanite is brown in color and formed as a result of the
disintegration of solid solutions (Agol, 1965). The mineral phase cubanite was found by
treating the sample C with ammonium chloride additive at 500oC for three hours roasting,
shown in XRD analysis Figure-6. The results of XRD with treatment are shown in figure
numbers (2-11).
XRD of complex copper sulfide ore without treatment (Sample C) in Figure-2,
XRD of simple roasting of sample C for 30 minutes at 650oC in Figure-3, XRD of simple
roasting of sample C for 60 minutes at 650oC in Figure-4, XRD of roasting of sample C
with ammonium nitrate in Figure-5, XRD of roasting of sample C with ammonium
chloride in Figure-6, XRD of two-stage roasting with ammonium chloride in Figure-7,
XRD of two-stage roasting with combined additive (chloride and sulfate of ammonium)
in Figure-8, microphotography of roasting of sample C with ammonium nitrite at 500oC
for three hours is shown in Figure-9, microphotography of roasting of sample C with
ammonium chloride at 500oC for three hours roasting time is shown in Figure-10 and
microphotography of two-stage roasting with combined additive (chloride and sulfate of
ammonium) at 500oC for three hours is shown in Figure-11.
137
4.6. Kinetic Studies
The kinetic parameters of roasting and leaching reactions were also compared by using
different models of kinetics. These kinetic models are used to determine the reaction rate
and calculate the activation energies of the reactions in aqueous as well as in solid state.
The kinetics model of Sharp (1966) was compared with our experimental results to
investigate the mechanism of reactions by using diagnostic equation. The mechanism of
the solid state reactions was determined by using diagnostic equation as follows;
-ln ln (1- α ) = ln B + m ln t.
The m-value was obtained by plotting the -ln ln (1- α ) versus In t and slope
shows m-value. This value shows the reaction mechanism.
The activation energy was also calculated. It was found that the reactions
mechanism during the roasting has phase boundary reaction as well as diffusion reaction.
The roasting with additives has decreased the activation energy from chloride >
chloride-sulfate > nitrate > simple roasting > sulfate. The model A2 shows randomly
formation of new phase nucleus. It is presented in XRD figures 4-8. The results are given
in table-40. All calculations were made computerized.
138
The table – 39 is derived from Strbac et al. (2006) and Hancock & Sharp (1972) and
Diagnostic equation.
Table 39: Reaction Mechanism Determining Equations
S.
No.
Function Equation Process that define rate
of reaction
m-Values
1 D1 α2 = k.t Linear diffusion 0.62
2 D2 (1- α)ln(1- α) + α = k.t Two-dimensional
diffusion cylindrical
symmetry
0.57
3 D3 [1-(1- α)1/3] 2 = k.t Three-dimensional
diffusion, Spherical
symmetry, Jander’s
Equation
0.54
4 D4 (1-2/3 α)-(1- α)2/3 = k.t Three-dimensional
diffusion, spherical
symmetry Ginstling-
Braunstin equation
0.57
5 F1 -ln(1- α) = k.t Randomly formation of
new phase nucleus,
Avrami equation (1)
1.00
6 A2 [-ln(1- α)]1/2 = k.t Randomly formation of
new phase nucleus,
Avrami equation (11)
2.00
7 A3 [-ln(1- α)]1/3 = k.t Reaction at the phase
boundary, cylindrical
symmetry
3.00
8 R2 1-(1- α)1/2 = k.t Reaction at the phase
boundary, cylindrical
symmetry
1.11
9 R3 1-(1- α)1/3 = k.t Reaction at the phase
boundary, spherical
symmetry
1.07
10 Zero
order
α = k.t 1.24
139
Table 40: Comparison of Activation Energy by using Different Model of Kinetics
S.No Samples Activation Energy (kJ/mole)
D1 D2 D3 D4 F1 A2 A3 R2 R3
1 Simple Roasting of sample C 73 80 87 82 59 33 23 52 54
2 C + Amm. Chloride 53 63 76 67 46 23 15 36 39
3 C + Amm. Nitrate 61 68 77 71 47 26 18 42 43
4 C + Amm. Sulfate 114 120 127 122 72 40 28 68 69
5 C + Amm. (Chloride+Sulfate) 61 71 83 75 49 24 16 39 42
4.6.1. Mechanism of Reaction
The mechanisms of roasting reactions were determined by the computation of
experimental data by using equations of table-39. It was calculated by using Sharp’s
kinetic models. The mechanism of the solid state reactions was determined by using
diagnostic equation as follows; -ln ln (1- α) = ln B + m ln t.
The m-value was obtained by plotting the -ln ln (1- α) versus In t and slope shows
m-value. This value shows the reaction mechanism.
Reaction mechanism for simple roasting of copper sulfides ore: The simple roasting
was carried out at the temperatures range from 500oC to 800oC. The m-value ranges from
1.63 to 0.64 at lower temperature range of 500oC to 600oC. When temperature goes
higher from 600oC to 800oC, the m-values become less than zero. It is found that simple
roasting reaction is from phase boundary to diffusion area at lower temperature range.
The m-values are given below;
Temperature 500oC 550oC 600oC 650oC 700oC 750oC 800oC
m- value 1.63 1.25 0.64 -0.032 -0.37 -0.95 -0.89
Hancock and Sharp (1972) studied the mechanism (m- values) for the decomposition of
BaCO3 at 7380 to 8080C. The m-value was decrease gradually with increase in
140
temperature, different from any value predicted by the commonly used theoretical
equations. In present research the sample is purely natural crude copper sulfide complex
ore. The m-value is also decrease with increase temperature in this research.
Reaction mechanism for additive roasting of copper sulfides ore: The additive
roasting was done at the temperatures range from 400oC to 500oC. The m-values of
different kinetics models were compared to determine the reaction mechanisms of roasted
mixtures. In roasting with ammonium chloride the reaction mechanism starts from R3
model with 1.07 value of mechanism of phase boundary spherical symmetry reaction at
low temperature of 400oC. The mechanism is going through diffusion D1 at 450oC to
randomly formation of new phase nucleus F1 at high temperature of 500oC. The
mechanism of sample C with combined additives [NH4Cl + (NH4)2SO4] roasting shows
in diffusion areas with m-values of 0.7 to 0.57.
The mechanism of roasting sample C with ammonium nitrate is randomly
formation of new phase nucleus shows in model A2. The mechanism of sample C with
ammonium sulfate roasting shows in model A2, A3 models with m-values of 3.3 to 2.
The results of m-values for different roasting mixtures to determine reaction mechanism
are given in table-41.
Table 41: m-Values for Different Roasting Mixtures to Determine
Reaction Mechanism
Sample 400oC 450oC 500oC
C+ NH4NO3 2.37 1.76 1.64
C+ NH4Cl 1.07 0.69 0.77
C+ (NH4)2SO4 3.30 2.49 2.02
C+ [NH4Cl + (NH4)2SO4] 0.73 0.64 0.57
141
CHAPTER 5
DISCUSSION AND CONCLUSION
DISCUSSION
The copper ore deposits are found in different areas of Pakistan. The exploration work on
these deposits was conducted by various mineral developing agencies to prove or
estimate the reserves in these areas. The mineralogy and chemical composition of these
minerals are oxides, sulfides and complex sulfides. The deposits are of low to high grade,
porphyry, vein type, hydrothermal and complex in mineralogical nature. The main
mineralization of copper ores in Pakistan is porphyry type, hydrothermal, scarn and
subduction related ores in which the mineralogical composition varies according to the
deposition. These deposits are having mineralogy of sulfide, chalcopyrite and complex
minerals; containing copper, lead, zinc, iron, silver and gold assemblage, such as
tetrahedrite and boulangrite. No systematic work on these ores has so far been done on
the evaluation, beneficiation and utilization of locally available ores. The oxidized and
low grade copper ores are mainly unutilized. Developments of economic processes for
the economic utilization of these ores still need investigation.
After detailed review of literature related to hydrometallurgical processes, the
aims of present research carried out in this thesis to investigate the possibilities of
recovering copper from indigenous copper ores. These ore samples were collected from
Chitral, Dir and Saindak areas. Comminution study was carried out in PCSIR laboratories
142
Lahore. Two types of hydrometallurgical techniques studied (1) direct leaching which is
a wet metallurgical process and (2) the indirect leaching (roasting before leaching) was
carried out for copper indigenous ores. The roasting before leaching is an alternative
leaching processes and have been developed for extraction of copper from sulfide ores.
Indigenous copper ore from Chitral, NWFP in Pakistan, were investigated for chemical
and hydrometallurgical processing and evaluation using mineralogical and chemical
methods. It was found that on an average these ore sample contained 20 percent copper,
21 percent iron, 1.54 percent zinc, 0.074 percent lead, 2.35 ppm silver and 1.0 ppm gold.
The sulfur content in the ore on average is 20 percent. The leaching results showed that
copper oxidized ores containing Azurite and Malachite are amenable to leaching in low
concentration of sulfuric acid. But the sulfide ores are not leached easily. The leaching of
copper sulfide ore samples in this study is conducted and results described below.
The copper extraction of about 80 percent was achieved by using 20 percent
sulfuric acid with 10 percent nitric acid at approximate temperature of 100oC and
pressure of 70-80 atmosphere with solid liquid ratio of 1:10 and 30 minutes leaching
time.
Alternative pretreatment processes of sulfide ores as roasting were developed and
are given as follows;
The sulfide ores was subjected to simple roasting at different temperature and
time without any additive. The maximum leaching results were obtained at 650oC for 30
minutes roasting. The 72 percent leaching was obtained with dilute sulfuric acid
solutions.
143
Some conventional oxidants, oxygen containing salt were used. The oxidants were
sodium persulfates, sodium chlorate, sodium nitrate, sodium nitrite, potassium chlorate.
The advantage is found the reduction of temperature range from 650oC to 500oC.
However roasting time was increased to three hours.
The ammonium nitrate was used during roasting. It gives 77 percent copper
extraction after three hours roasting at 500oC. The amount of ammonium nitrate is 10
percent of ore sample weight.
A non-oxidizing set of chemicals was used during roasting of copper sulfide ore.
Some chemicals containing chloride, nitrate, acetate, oxalate, phosphate, sulfate of
ammonium and urea were used during roasting. The ammonium chloride is found very
useful additive. It gives 85 percent copper leaching at 500oC for 90 minutes roasting.
The ammonium chloride additive was used in the two-stage roasting. The ore and
additive mixture was subjected to roasting for 30 minutes at 350oC then temperature was
raised to 500oC and the material was roasted 45 minutes at 500oC. The roasted material
was grinded and leached. The leaching of copper was above 96 percent in very short time
with dilute sulfuric acid solution.
The ammonium chloride, ammonium nitrate, ammonium sulfate were selected for
the study of combined effect of additives. The best combination was chloride-sulfate
during roasting. It is noticed here that this combination was also better for in aqueous
medium of leaching processes according to literature.
A detailed study of simple roasting, oxidative roasting and additive roasting was
conducted then the sample of copper chalcopyrite concentrate was investigated through
the optimal conditions. It was found that the amount of additive was increase to three
144
times. The copper leaching again decreased with increase the amount of additive more
than 30 percent.
The unique and common findings of the research work are use of novel
additives such as ammonium chemicals. These chemicals are used first time in solid
state as roasting additives. The more suitable among ammonium chemicals are found
ammonium chloride, ammonium nitrate and ammonium sulfate. These ammonium
chemicals generate the reactive ions in the form of vapor or gaseous species as NH3, HCl,
NO2, N2, HNO3, NO, N2O, H2O, HSO4. The decomposition of different chemicals is
inferred as follows:
Ammonium chloride sublime at 340-350oC by reaction (Cotton, 1962)
NH4Cl(s) NH3(g) + HCl(g)
Ammonium salts having oxidizing anions decompose on heating, with oxidation
of ammonia to NO2, N2 and HNO3 with the contamination of NO. The melting point of
Ammonium nitrate is 190oC and boiling point is 210oC, may decompose according to
equations;
NH4NO3(s) NH3(g) + HNO3(g)
NH4NO3(s) N2(g) + 2H2O(g) + 1/2O2(g)
NH4NO3(s) N2O(g) + 2H2O(g) (contaminated with NO)
The gas N2O decomposes at above 600oC.
2N2O(g) 2N2 (g) + O2 (g) (Heslop et al. 1963)
The decomposition of Ammonium sulfate is according to equation on 280oC as
(NH4)2SO4(s) NH3(g) + HSO4(g)
145
The ammonium chemicals produce NH3 at low temperature. It is possible that NH3
oxidizes and gives NO, nitric oxides at or above 500oC. It may be possible in the
presence of copper, ammonia oxides and converts into NO.
4NH3 + 5O2 4NO + 6H2O (Heslop et al. 1963)
NO + NO N2O2
N2O2 + Cl2 2NOCl
The NO reacts with metal in ammonia, so it may be possible that formation of
CuNO for short time as an intermediate phase.
N2O4 + HCl NOCl + HNO3
The chemistry of ammonium compound may show variety of complex reactions
in solid state reactions. These additives may produce the NH3, H2O, Cl2, HCl, HSO4-,
HNO3, SO4=, NO3
- and NOx species which react as vapor solid phase reactions. The
ammonia changes into NO or NOx species. Nitrogen oxide species may react with ores
sulfides minerals in ores (Anderson, 2002).
2MS + 4NO+ 2M+2 + 2So + 4NO
According to these equations or reaction mechanism the additives first break
down into possible species at lower temperatures. These ions or gaseous species may
attack the ore particles. The interstitial penetration of ions may take place into the
structure of ore particles (West, 1984). There is very complex mechanism of the solid
state reactions. The additives may interact as catalysts.
These novel additives change mineralogical phase into leach able minerals. The
cubanite mineral is formed in some reactions, which is high in sulfur as compared to
chalcopyrite, covellite, chalcocite. It is shown in Figure-6 of XRD analysis in Appendix-
146
A. It is brown in color shown (Agol, 1965) micro photographically is shown in Figure-10
in Appendix-B.
2CuFeS2 CuFe2S3 + CuS
(Additive roasting)
The dissociated products NH3, H2O, Cl2, HCl, HSO4, HNO3, SO4= and NOx
species of these additives behave as catalyst during roasting treatments. This finding is
compared with previously reported findings (Dunn, 1989) of the recovery of copper
improved after catalytic roasting at temperature 450-500oC.
The chemical reactions involve solids and liquefication do not occur, the progress
of chemical change is a complex process which depends on the area and defects in the
structure of the contact areas between the reactant solid and the product (Galwey, 1967).
The XRD analyses of selected sample were carried out. The XRD analysis of the
samples has shown the presence of Chalcopyrite, Chalcocite, Azurite, Tetrahedrite and
Malachite as copper minerals. The treated samples show, the formation of oxides and
sulfate of copper. It is also found that the formation of sulfur rich minerals of copper.
Cubanite brown in color is formed (Agol, 1965).
The kinetics of roasting processes was investigated. The kinetic models (Sharp,
1966) are used to determine the reaction rate, mechanism of reaction and calculate the
activation energies of the reactions. It was found the reactions mechanism during the
roasting has phase boundary reaction as well as diffusion reaction.
It is probable that roasting with additives has decreased the activation energy from
chloride > chloride-sulfate > nitrate > simple roasting > sulfate.
147
CONCLUSION
The copper oxidized ores containing Azurite and Malachite are found amenable to
leaching in low concentration of sulfuric acid. But the copper sulfide ores containing
chalcopyrite and complex sulfide minerals are not leached easily. A suitable and cheap
method to leach the sulfide copper containing chalcopyrite ore for the economic
utilization is developed. The optimum of copper leaching achieved by prior roasting was
70 to 75 percent.
A series of novel additives during roasting was investigated having properties of
ions and vapors such as nitrate, chloride, sulfate, phosphate, acetate, oxalate, carbonate
and ammonia. It was found that these ions or group of ions have similar effect as reported
in the direct leaching of hydrometallurgical techniques for copper ores. The developed
processes envisaged that the vapor-solid metallurgy of complex copper sulfide ores
change the mineralogy into leachable minerals. The roast leach processes gave about 85
percent to 98 percent copper leaching with 2.5 percent sulfuric acid in 1:5 solid liquid
ratio within very short time. The additive containing chloride ion showed more leaching
than nitrate or sulfate species and decreases the roasting time to half. The combined
effect of additives studied at two stage roasting showed the best combination of additives
found was chloride-sulfate of ammonium.
The kinetics of roasting processes showed that the reactions mechanism during
the roasting ranges from phase boundary reaction to diffusion reaction. It was found that
roasting with additives has decreased the activation energy in order to chloride >
chloride-sulfate > nitrate > sulfate.
