RECOVERY OF PARTICULATE GOLD BY COAL-OIL AGGLOMERATES

A Thesis by

JUANCHO PABLO S. CALVEZ B. Sc. Metallurgical Engineering

Submitted for the degree of

Master of Science

School of Chemical Engineering & Industrial Chemistry Faculty of Engineering UNIVERSITY OF NEW SOUTH WALES

March 1998 11

CERTIFICATE OF ORIGINALITY

I hereby declare that this submission is my own work and to the best of my knowledge, it contains no materials previously published or written by another person, nor material which to a substantial extent has been accepted for the award of any other degree or diploma at UNSW or any other educational institution, except where due acknowledgment is made in the thesis. Any contribution made to the researchby others, with whom I have worked at UNSW or elsewhere, is explicitly acknowledgedin thethesis.

I also declare that the intellectual content of this thesis is the product of my own work, except to the extent that assistance from others in theproject's design and conception or in style, presentation andlinguistic expression is acknowledged. lll

ABSTRACT

Fundamental and applied studies were undertaken to evaluate the underlying ' principles by which gold is recovered by coal-oil agglomerates. The utilization of coal-oil agglomerates as a potential environmentally safe alternative to conventional methods of recovering gold has been patented over a decade ago but has not gained its ground commercially in the mining industry. This study then was conducted to further improve the understanding of the process.

Experiments were conducted in a 1.5 liter baffled glass vessel agitated by four­ axial blade impellers. Each experiment was carried out in two phases: first, the formation of agglomerates and second, the agitated mixing of agglomerates and ground gold bearing samples. Carbonaceous materials used for agglomeration were bituminous coal, lignitic coal and graphite while the agglomerating liquids were diesel oil and kerosene. Gold bearing samples employed for the fundamentals studies were synthetic mixtures of cyanide leaching residues (or pure silica) and pure gold powders while the materials for the applicability tests were real gold bearing samples - copper concentrates and amalgamation tails - from the Philippines.

Tests to determine the effects of different variables - oil:coal ratios, agglomerate:ore ratios, pH, impeller speed, particle suspension, coal particle size, ore particle size - on gold recovery were conducted. The effects of sulfides were also evaluated. The mechanism by which gold are collected by the agglomerates was investigated by the use of scanning electron microscope. IV

Results of the fundamental studies revealed that rate of gold recovery was increased by increasing agglomerate:ore ratio (AOR) or decreasing oil:coal ratio (OCR). At AORs of 1 and 0.13, 99.8% and 99.3% gold recoveries were attained after 15 and 120 minutes, respectively. Increasing agglomerate:ore ratio correspondingly increases the number and surface area of agglomerates available for gold capture. At 30 minutes mixing time, gold recoveries at OCRs of 0.269 and 0.403 were 99% and 91 %, respectively. Decreasing the amount of oil relative to constant weight of coal (decreasing OCR) produced smaller agglomerates thereby increasing their number and available surface area.

At pH range of 4 - 12, there was no significant difference in gold recoveries while at pH of 2, recovery was adversely affected specifically at limited mixing time. Whereas superficial organic contamination or oxidation influences gold hydrophobicity, there may have been an initial cleaning effect of sulfuric acid (pH regulator), that lessens the hydrophobicity of the gold surfaces. By extending mixing time from 30 to 60 minutes, however, gold recovery was increased considerably due to the ease by which gold could be re-contaminated.

Increasing impeller speed increased gold recovery due to enhancement of inter­ particle collisions and gold particle suspension. The final recovery at 300 rpm was notably low at 64% as compared to over 90% attained at 500 rpm and above. Agglomerate strength and stability as dictated by the size of coal particles, the type and amount of oil, and the type of carbonaceous material also influenced gold recovery. With weaker agglomerates, formed by using coarser coal particles, lighter oils, over-critical amount of oil or less hydrophobic carbonaceous agglomerating material, gold recovery was adversely affected when V

fine agglomerate particles (possibly loaded with gold) were abraded from the main agglomerates and were lost with the tailings.

At up to 5% sulfides in the feed, gold recovery was not significantly altered. While the agglomerates mostly recovered chalcopyrite, pyrite and galena, sphalerite, due to the oxidized nature of the sample, gave poor recovery. Increasing pH lowered pyrite recovery that at pH of 12 virtually no pyrite was recovered.

Experiments with real gold bearing samples revealed that increased liberation size and impeller speed raised gold recovery. Too much sulfides had an adverse effect on gold recovery.

Microscopic examination of loaded agglomerate showed that gold particles penetrated within the sub-surface up to 60 µm depth. Vl

ACKNOWLEDGMENT

I would like to convey my heartfelt appreciation and gratitude to the following persons and entities, without them, this study would not have been a reality:

To Associate Professor Tam Tran, my supervisor, for his dedicated guidance and support during the whole course of my research.

To the Australian Agency for International Development (AusAID), for granting me a scholarship to pursue this MSc degree.

To Dr. Patrick Wong, my former co-supervisor, for his technical support during the initial phase of this study.

To the staff of Centre for Minerals Engineering, Jin Song and Ling Lau, for their invaluable assistance and to my fellow postgraduate students, for helping me cope up with the rigors of laboratory works.

To the personnel and staff of the Metallurgical Technology Division of the Philippine Mines and Geosciences Bureau, for the unselfish help they accorded me during the conduct of my experiments in the Philippines.

To the Filipino Students Society of UNSW, for sharing with me the ups and downs of being away from home. vii

To the Couples for Christ - Singles for Christ Ministry here in Australia, for supporting me spiritually and sharing with me the joys of being together in fellowships and service.

To my family and friends, for their prayers and words of encouragement that inspired me to do my best in hurdling this another milestone of my career.

And most especially to our Good Lord, the Almighty King, for always being there - watching and guiding me in all my endeavors. Vlll

TABLE OF CONTENTS

i, Page

ABSTRACT 111

ACKNOWLEDGMENT VI

LIST OF FIGURES Xlll

LIST OF TABLES XVI

LIST OF APPENDICES XVIll

LIST OF PUBLICATIONS lX

CHAPTERl INTRODUCTION 1

CHAPTER2 LITERATURE REVIEW 5 2.1 INTRODUCTION 5 2.2 GOLD MINERALOGY 5 2.3 GOLD RECOVERY PROCESSES 7 2.3.1 Cyanidation Process of Gold Extraction 9 2.3.1.1 The Chemistry of Cyanidation 9 2.3.1.2 Treatment of Gold Ores by Cyanidation 10 2.3.1.3 Methods of Recovering Gold from Cyanide Leach Liquor 11 2.3.1.3.1 Zinc Cementation 11 2.3 .1.3 .2 Carbon Adsorption Process 12 2.3.1.3.3 Resin Technology 16 2.3.1.3.4 Ion Flotation 18 IX

2.3.2 Non-Cyanide Lixiviants for Gold Dissolution 20 2.3.2.1 Thiourea 21 2.3.2.2 Thiosulfate 22 2.3.2.3 Thiocyanate 23 2.3.2.4 Halides 24 2.3.3 Gravity Concentration 25 2.3.4 Froth Flotation 27 2.4 AGGLOMERATION OF FINE COAL WITH OIL 32 2.5 COAL-OIL AGGLOMERATION FOR GOLD RECOVERY 36 2.5.1 Patents for Coal-Oil Agglomeration Process 37 2.5.1.1 The CGA Process 38 2.5.1.2 The CARBAD Process 39 2.5.1.3 The Bateman Process 40 2.5.2 Application of Coal-Oil Agglomeration on Gold Recovery 41 2.6 THEWETTABILITYOFGOLDBYWATER 45

CHAPTER 3 SUMMARY OF LITERATURE REVIEW AND RESTATEMENT OF OBJECTIVES 48

CHAPTER 4 RESEARCH METHODOLOGY 52 4.1 MATERIALS 52 4.1.1 Gold-Bearing samples 52 4.1.1.1 Synthetic Mixtures 52 4.1.1.1.1 Gold Powder - Cyanidation Residue Synthetic Mixture (Blend A) 52 4.1.1.1.2 Gold Powder-High Purity Silica Mixture (Blend BJ 53 4.1.1.2 Real Gold-Bearing Samples 54 4.1.2 Carbonaceous Materials 54 4.1.2.1 Bituminous Coals 54 X

4.1.2.2 Graphite 55 4.1.2.3 Lignitic Coal 56 4.1.3 Agglomerating Liquids 56 4.1. 4 Chemical Reagents 56 4.2 EQUIPMENT AND INSTRUMENTATION 57 4.2.1 Experimental Apparatus 57 4.2.2 Sample Preparation Equipment 58 4.2.3 Analytical Equipment 58 4.2.4 Scanning Electron Microscope 59 4.3 EXPERIMENT AL DESIGN 59 4.3.1 General Procedures 59 4.3.1.1 Agglomeration 60 4.3.1.2 Gold Recovery Stage 60 4.3.2 Preliminary Experiments 61 4.3 .3 Experiments to Determine the Effects of Various Parameters in Gold Recovery 62 4.3.3.1 Agglomerate : Ore Ratios 62 4.3.3.2 Oil: Coal/Graphite Ratios 63 4.3.3.3 pH 63 4.3.3.4 Particle Suspension 64 4.3.3.5 Coal Particle Size 65 4.3.4 Experiments to Determine the Effects of Sulfides 66 4.3.5 Laboratory Experiments Conducted in the Philippines 68 4.3.5.1 Tests on Real Gold Bearing Samples 68 4.3.5.2 Tests on Philippine Lignitic Coal 70 4.3.6 Determination of the Extent of Gold Attachment/ Penetration in the Agglomerate 71 4.4 ANALYTICAL PROCEDURES 72 X1

4.4.1 Analysis of Heads and Metallurgical Products 72 4.4.2 Ash Determination 74 4.5 SIMULTANEOUS PULP AND AGGLOMERATE SAMPLING PROCEDURE 74 4.6 METAL RECOVERY CALCULATIONS 75

CHAPTERS RESULTS AND DISCUSSION 77 5.1 RESULTS OF PRELIMINARY EXPERIMENTS 77 5.2 EFFECTS OF AGGLOMERATE:ORE RATIOS 80 5.3 EFFECTS OF OIL:COAL RATIOS 83 5.3.1 Effects of Oil:Coal Ratios on Agglomerate Size, Surface Area and Number 84 5.3.2 Effects of Agglomerate Surface Area and Size on Gold Recovery 86 5.3.3 Summary of Effects of Different Oil:Coal Ratios on Gold Recovery 88 5.3.4 Effects of Oil:Graphite Ratios on Gold Recovery 89 5.4 EFFECTS OF pH ON GOLD RECOVERY 90 5.5 EFFECTS OF PARTICLE SUSPENSION 92 5.6 EFFECTS OF COAL PARTICLE SIZE ON AGGLOMERATION AND GOLD RECOVERY 95 5.7 EFFECTS OF SULFIDES ON GOLD RECOVERY AND RECOVERY OF SULFIDES BY COAL-OIL AGGLOMERATES 96 5. 7 .1 Effects of Varying Amounts of Sulfides on Gold Recovery 96 5.7.2 Sulfides Recovery by Coal-Oil Agglomerates 97 5.7.3 Effects of pH on Gold and Sulfides Recovery 99 5.8 RESULTS OF LABORATORY EXPERIMENTS CONDUCTED IN THE PHILIPPINES 103 5.8.1 Tests on Real Gold Bearing Samples 103 Xll

5.8.2 Tests on Philippine Lignitic Coal 106 5.8.3 Summary of Results on Tests Conducted in the Philippines 107

CHAPTER 6 DETERMINATION OF THE EXTENT OF GOLD ATTACHMENT/PENETRATION IN 108 '£HE AGGLOMERATES 6.1 PHOTOGRAPHS OF AGGLOMERATE SURF ACES 109 6.2 SEM IMAGES OF AGGLOMERATE SURF ACES 112 6.3 SEM IMAGES OF AGGLOMERATE CROSS-SECTIONS 115

CHAPTER 7 CONCLUSIONS AND RECOMMENDATIONS 119 7.1 CONCLUSIONS 119 7.2 RECOMMENDATIONS FOR FUTURE WORK 123

REFERENCES 125

APPENDICES 134 Xlll

LIST OF FIGURES

Chapter 2 Pages

Figure 2.1 : Generalized Flowsheet for Gold Recovery by Cyanidation- Zinc Cementation Process 13

Figure 2.2: Generalized Flowsheet for Gold Recovery by Cyanidation- Carbon Adsorption Process 15

Figure 2.3 : Schematic Diagram of a Continuous Ion Flotation Unit 19

Figure 2.4 : Typical Philippine Flowsheet for Small-Scale Gold Recovery Operation 28

Figure 2.5: Flotation Cells - Conventional Cell and Jameson Cell 30

Figure 2.6: The States of Bonding Liquid Saturation 35

Figure 2. 7 : The CARBAD Mechanism 40

Figure 2.8: Use of Sessile Drop of Bubble in Contact Angle Determination 46

Chapter 4

Figure 4.1 : Schematic Diagram of Experimental Set-up 57

Figure 4.2 : Schematic Diagram of Agglomerate Preparation for Scanning Electron Microscopy 73

Figure 4.3 : Simultaneous Pulp and Agglomerate Sampler 75

Chapter 5

Figure 5.1 : Effect of Mixing Time on Gold Recovery 79

Figure 5.2: Effect of Impeller Speed on Gold Recovery (Variable Agglomerate Size) 79

Figure 5.3 : Effect of Agglomerate:Ore Ratio on Gold Recovery (15 min) 81 XIV

Figure 5.4: Effect of Agglomerate:Ore Ratio on Gold Recovery (120 min) 82

Figure 5.5: Effect of Mixing Time at Different AOR on Au Recovery 82

Figure 5.6 : Effects of Oil/Kerosene Ratios on Agglomerate Surface Area and Size 86

Figure 5.7: Effect of Agglomerate Size on Gold Recovery 87

Figure 5.8: Effect of Available Surface Area on Gold Recovery 87

Figure 5.9: Effect of Oil:Coal Ratios on Gold Recovery 89

Figure 5 .10 : Effect of Oil:Graphite Ratios at Different Impeller Speeds and Mixing Times on Gold Recovery 90

Figure 5 .11 : Effect of pH on Gold Recovery 92

Figure 5.12: Effect of Impeller Speed on Gold Recovery (Constant Agglomerate Size) 93

Figure 5.13 : Effect of Impeller Elevation on Gold Recovery 94

Figure 5.14: Effects of Varying Amounts ofSulfides on Gold Recovery 97

Figure 5.15 : Recovery of Sulfides by Coal-Oil Agglomerates 98

Figure 5.16: Effects of pH on Gold Recovery at 2% Sulfide Matrix 100

Figure 5.17: Effects of pH on Sulfides Recovery 101

Figure 5.18 : Photograph of Agglomerates with Chalcopyrite 101

Figure 5.19: Photograph of Agglomerates with Pyrite 102

Figure 5.20: Photograph of Agglomerates with Galena 102

Figure 5.21 : Photograph of Agglomerates with Sphalerite 103

Figure 5.22 : Recovery of Gold from Flotation-Sluice Concentrate at Different Agglomerate:Ore and Oil:Coal Ratios. 105 xv

Figure 6.1 : Photograph of Agglomerate at 15 Minutes Mixing Time 109

Figure 6.2: Photograph of Agglomerate at 30 Minutes Mixing Time 110

Figure 6.3 : Photograph of Agglomerate at 45 Minutes Mixing Time 110

Figure 6.4 : Photograph of Agglomerate at 60 Minutes Mixing Time 111

Figure 6.5 : Photograph of Agglomerate at 120 Minutes Mixing Time 111

Figure 6.6: Typical EDS Pattern for Gold Analysis 112

Figure 6. 7 : SEM Image of Agglomerate Surface at 15 minutes 113

Figure 6.8: SEM Image of Agglomerate Surface at 15 minutes (Higher Magnification) 114

Figure 6.9: SEM Image of Agglomerate Surface at 60 minutes 114

Figure 6.10 : SEM Image of Agglomerate Surface at 120 minutes 115

Figure 6.11 : SEM Image of Agglomerate Cross-Section at 60 minutes (#1) 116

Figure 6.12 : SEM Image of Agglomerate Cross-Section at 60 minutes (#2) 117

Figure 6.13 : SEM Image of Agglomerate Cross-Section at 120 minutes (#1) 117

Figure 6.14: SEM Image of Agglomerate Cross-Section at120 minutes (#2) 118 XVI

LIST OF TABLES Pages Chapter 2

Table 2.1 : List of Gold Minerals 7

Chapter 4 s

Table 4.1 : Summary of Test Conditions for Preliminary Experiments 61

Table 4.2: Experiments to Determine the Effects of Agglomerate:Ore Ratio Variations 62

Table 4.3 : Experiments to Determine the Effect of Varying Oil: Coal/Graphite Ratios 63

Table 4.4 : Experiments to Determine the Effects of pH 64

Table 4.5 : Experiments to Determine the Effects of Particle Suspension 65

Table 4.6 : Experiments to Determine the Effects of Coal Particle Size 66

Table 4. 7 : Experiments to Determine the Effects of Sulfides on Gold Recovery 67

Table 4.8: Experiments to Determine the Effects of pH on Gold and Sulfides Recovery 67

Table 4.9: Effects of Oil: Coal Ratios on Flotation-Knelson Concentrates 68

Table 4.10: Effect of Grinding and Impeller Speed on Amalgamation Tails 69

Table 4.11 : Effect of Agglomerate:Ore and Oil:Coal Ratios on Flotation-Gravity Concentrate 69

Table 4.12 :Experiments on the Efficiency of Philippine Coal on Synthetic Mixture 70

Chapter 5

Table 5.1 : Results of Preliminary Experiments 78

Table 5.2: Effects of Agglomerate:Ore Ratio (15 min) 81 XVll

Table 5.3: Effects of Agglomerate:Ore Ratio at Extended Mixing Time 81

Table 5.4 : Effects of Variable Diesel Oil/Kerosene:Coal Ratios on Agglomerate Size and Surface Area, No. of Agglomerates and Gold Recovery 85

Table 5.5: Effect of Oil:Coal Ratios on Gold Recovery at Extended Time and Higher Impeller Speed 88

Table 5.6: Effect of Oil:Graphite Ratios at Different Impeller Speeds and Mixing Times on Gold Recovery 90

Table 5.7: Effect of pH on Gold Recovery 91

Table 5.8: Effects of Impeller Speed on Gold Recovery 93

Table 5.9: Effect of Impeller Elevation on Gold Recovery 93

Table 5.10: Effect of Coal Particle Size on Gold Recovery 96

Table 5.11 : Effects of Varying Amounts of Sulfides on Gold Recovery 97

Table 5.12 : Recovery of Sulfides by Coal-Oil Agglomerates 98

Table 5.13 : Effects of pH on Gold Recovery at 2% Sulfide Matrix 99

Table 5.14: Effects of pH on Sulfides Recovery 100

Table 5 .15 : Recovery of Gold from Flotation-Knelson Gravity Concentrate at Different Oil:Coal Ratios 104

Table 5.16 : Recovery of Gold from Flotation-Sluice Concentrate at Different Agglomerate:Ore and Oil:Coal Ratios. 105

Table 5.17 : Effect of Grinding and Impeller Speed on Gold Recovery from Amalgamation Tails 106

Table 5.18 : Agglomeration and Gold Recovery Using Philippine Lignitic Coal 106 XVlll

LIST OF APPENDICES

Page

Appendix A MINERALOGY OF CYANIDE LEACHING RESIDUE (BLEND A MATRIX) 135

Appendix B SCREEN ANALYSIS OF GOLD BEARING AND COAL SAMPLES 140

Appendix C COAL ANALYSES 144

AppendixD SCREEN ANALYSES AND SURF ACE AREA CALCULATIONS OF AGGLOMERATES AT DIFFERENT OIL/KEROSENE:COAL RATIOS 145

Appendix E ANALYTICAL PROCEDURES 150 X1X

LIST OF PUBLICATIONS

• J. P. S. Calvez, P. L. M. Wong and T. Tran, Use of Coal-Oil Agglomerates for Particulate Gold Recovery submitted to Minerals Engineering, March 1998

• J.P. S. Calvez, P. L. M. Wong and T. Tran, Recovery of Particulate Gold by Coal-Oil Agglomerates submitted to Chemeca '98, March 1998 1

CHAPTERl

INTRODUCTION

The search for new technologies to recover gold from various types of ores has been a preoccupation of metallurgists and mineral processing engineers for several decades now. Some of the advances in the field include processes that could treat low and marginal grade ores, and those that minimize environmental impact. The use of cyanide however is still the predominant method until now since its inception more than a century ago. Earlier researches were centered on improving the cyanide-based process, most notably of which was the use of activated carbon in adsorbing the leached gold in lieu of the older process of utilizing zinc to precipitate the gold from the pregnant solution. This development was responsible for the minability of low to marginal grade ores. Next to the use of cyanide, another most predominant method still conducted in underdeveloped areas is the use of mercwy in recovering free gold especially in small-scale mining. These two gold recovery processes however use toxic chemicals - cyanide and mercwy - which unless handled properly, have deleterious effects on the environment.

The increasing global concern over environmental protection has broadened the area of research in gold recovery. Although researches focusing on the 2 improvement of the cyanide-leach process are still being conducted, numerous studies that take environmental preservation into consideration are being undertaken. These studies aimed at either to lessen the impact of cyanide in the tailings (i.e. destruction of cyanide) or to avoid using cyanide or mercury at all. Among the non-cyanide leach options, several lixiviants are worth mentioning; these include thiourea, thiosulfate, thiocyanate, chloride, iodide and bromide. While the use of these leachants have technical and environmental advantages over cyanidation, the economics of operations have prevented them from taking off industrially. On the non-leaching processes like flotation and gravity concentration, a promising technology known as coal-oil agglomeration has caught the attention of some researchers because of its environmental and technical viabilities.

The coal-oil agglomeration process for gold recovery became known to the mining industry in the mid-80s after the first related patent was approved. This process involves either the attachment of gold to oil-agglomerated coal or the simultaneous agglomeration of gold with coal and oil. Although the use of oil to recover gold was already introduced in the early 1900's, it is only this late that the potential of oil-coal agglomerates is being investigated as one of the alternatives to the conventional gold recovery processes. A decade has passed since that first patent was issued and various tests and researches had been conducted but the process has not gained its commercial acceptance so far. This study was undertaken to take a closer look at the coal-oil agglomeration process, the variables that affect both agglomeration and gold recovery, and its application to the industry. 3

The objectives of this study are:

1. To determine the underlying fundamentals by which the gold is recovered by coal-oil agglomerates, and '

2. To test other factors that may further explain the principles of gold recovery by coal-oil agglomeration process.

To achieve these aims, synthetic and real gold-bearing samples were prepared and subjected to laboratory bench-scale tests to evaluate different variables that affect gold recovery by coal-oil agglomerates. The extent of gold capture and/or penetration in the agglomerates was determined through the use of electron scanning microscope. The effect of sulfides on gold recovery was likewise evaluated.

A comprehensive review of related literature is presented in Chapter 2. This includes conventional and non-conventional gold recovery processes; oil agglomeration technology; patents and applications of coal-oil agglomeration for gold recovery; and hydrophobicity/wettability of gold. The summary of this review is enumerated in Chapter 3, which also includes the re-statement of objectives.

Chapter 4 describes in detail the experimental procedures, analytical methods, materials, intrumentation and apparatus used in the study.

The experimental results on the effects of variables namely; agglomerate:ore ratio, oil:coal ratio, pH, particle suspension and coal particle size on 4 agglomeration and gold recovery are presented in Chapter 5. The outcome of the laboratory testworks conducted on Philippine gold bearing samples and coal is also included in this chapter together with the effects of sulfides on gold recovery.

The extent of gold particle penetration/attachment in the agglomerate is discussed and illustrated in Chapter 6. SEM images and photographs of agglomerate surfaces and cross-sections at different loading times are shown.

Chapter 7 gives the summary of findings and conclusions in the study and introduces some recommendations for future work. References and appendices follow this fmal chapter. 5

CHAPTER2

LITERATURE REVIEW

2.1 INTRODUCTION

Gold, from the earliest times, has been considered as the most valuable of all metals as it possesses very important properties that are not common to other metals. Its richness of color and untarnishable brightness have made gold in demand for decorative and ornamental purposes (Eissler, 1896). Its scarcity, durability and attractiveness led to its early use as a medium of currency or exchange for goods and purposes. Today, gold is extensively used in electronics, heat shield and aerospace industries owing to its valuable properties such as chemical inertness, malleability, high electrical conductivity, high reflectivity and high specific gravity (Paterson, 19~0).

2.2 GOLD MINERALOGY

Gold occurs in nature as a native element, an alloy, a telluride (possibly a tellurate) or a compound (with sulfur, antimony or sellenium), the most 6 common of which are native gold and electrum. (see Table 2.1 for the list of gold minerals) An electrum is only a variety of native gold containing 20% or more silver and has a similar crystal structure and optical properties with native gold (Gasparrini, 1983). Gold occurrence is usually associated with most of the common rock-forming sulfide minerals like pyrite, arsenopyrite, stibnite and chalcopyrite as very small inclusions or ionic substitutions (Paterson, 1990). In oxidized environment, gold is commonly hosted by magnetite and secondary iron oxides. It could also be found in association with; uranium minerals (as fme grains totally enclosed in uranium minerals); silicates and carbonates (enclosed or distributed along grain boundaries); carbonaceous materials (associated with graphite or other carbonaceous matter forming fine disseminated particles of native gold) and; sand and gravel (placer deposits - as loose detrital grains in sediments of variable mineralogical composition). The occurrence of gold in host minerals is very varied and can be summarized as follows (Gasparrini, 1983):

• gold distributed in fractures or at the border of the same mineral

• gold distributed along the border between grains of different minerals (e.g. two sulfides, a sulfide and a silicate, a sulfide and an oxide)

• gold totally enclosed in the host mineral 7

Table 2.1 : List of Gold Minerals

Native Elements, Alloys and Metallic Compounds Tellurides

Native Gold Au Calaverite AuTe2 Electrum (Au, Ag) Montbrayite (Au, Sb)Te3 Curpoauride (Au, Cu) Petzite ( antamokite) Ag3AuTe2 Porpezite (Au, Pd) Muthmannite (Ag, Au)Te Rhodite (Au, Rh) Sylvanite Au, Ag)Te4 Iridic Gold (Au, Ir) Kostovite AuCuTe4 Platinwn Gold (Au, Pt) Nagyagite Pb5Au Bismuthian Gold (Au, Bi) (Te, Sb)4 Ss-s Gold Amalgam (Au2Hg3)(?) Sulfide Maldonite (Au2Bi) Auricupride (AuCu3) Uytenbogaardite Palladiwn Cuproauride ( Cu, Pd)3Au2

Antimonide Selenide

Aurostibite F ischesserite

(Paterson, 1990)

2.3 GOLD RECOVERY PROCESSES

Gold, being very widely distributed in nature both in native or combined forms, is extracted through various recovery techniques from simple gravity concentration to most complex biochemical schemes since time immemorial until today. The commonly used processes include :

1. Gravity concentration (using jigs, tables, spirals, etc.) 2. Amalgamation (with mercury) 8

3. Flotation (as free particles or contained m base metal sulfide concentrates) 4. Pyrometallurgy (in the smelting and refining of base metal ores and concentrates) ' 5. Hydrometallurgy ( direct cyanidation, cyanidation with carbon adsorption, heap-leach, and chlorination-leach) 6. Combination methods (gravity or flotation concentration followed by cyanidation or roasting of flotation concentrates followed by cyanidation)

With the depletion of high-grade and free-milling ores, various efforts are being made to develop new technologies to process low grade and refractory ores (Bhappu, 1990). Heap leaching, for one, is being utilized to recover gold from mill tailings containing as little as 1 ppm (Habashi, 1987) or low-grade oxide ores (Bhappu, 1990). For treating refractory ores where gold is locked in pyrite crystals or in organic matter (thus having a poor response to conventional cyanidation), pre-treatment methods such as roasting, high temperature and high pressure oxidation and biological oxidation are done prior to cyanidation (Bhappu, 1990). With the ever growing concern over environmental protection, efforts are also undertaken to develop alternative processes to cyanidation. Non-cyanide lixiviants such as thiourea, thiosulfates, thiocyanate, chloride, bromide, and iodide are being tried and developed. In the non-leaching side, coal-oil agglomeration may be considered as a potential process. 9

2.3.1 CYANIDATION PROCESS OF GOLD EXTRACTION

2.3.1.1 The Chemistry of Cyanidation

Cyanidation is the process of dissolving metallic gold in an aerated dilute cyanide solution. The process was first commercially employed in New Zealand in 1889 (Davis, 1991) and after more than a century, is still the most common method of extracting the metal today. The overall dissolution of gold in dilute cyanide solutions may be represented by the classic Elsner's equation:

(1)

This equation proceeds through intermediate product H2O2 as follows (Bhappu, 1986):

= 2Au(CN)2- + 2OH- (3)

The above equations indicate that the dissolution process can be considered as an electrochemical process, in which oxygen takes up electrons at one part of the metallic surface (the cathodic zone) while the metal gives them up at another (the anodic zone) (Habashi,1966).

Clean gold surfaces dissolve at a rate of 3.25 mg per sq. cm. per hour at optimum cyanide concentrations (about 0.05% NaCN), thus a 44 µm gold 10

particle will dissolve in 13 hours and a 149 µm gold particle will dissolve in 44 hours. In practice, coarse particles (> 149 µm) found in placer deposits and vein type ores are usually recovered by gravity concentration prior to cyanidation (Bhappu, 1990).

2.3.1.2 Treatment of Gold Ores by Cyanidation

The treatment of gold ores by cyanidation generally consists of percolation or agitation leaching with dilute cyanide solution generally less than 0.3% sodium cyanide. With the agitation leach, the ore is usually ground to 80% passing 75 µm and leached in mechanical agitators at 40 - 45 % pulp densities with adequate aeration ( effected by compressed air injection in the slurry) to provide enough oxygen essential in the leaching step as shown in equations (1) and (2). Lime is added in the cyanide pulp to prevent hydrolysis and to neutralize any acidic constituents present in the ore. Besides maintaining pH of 10 - 11, lime (which could also be substituted with sodium hydroxide or sodium carbonate), also helps in the decomposition of bicarbonates in mill water, improvement of settling rates in the counter-current decantation (CCD) thickeners and improvement in extraction rates for certain types of ores such tellurides and ruby silver (Bhappu, 1990). After leaching, which normally takes 20 -30 hours, the dissolved gold (more than 90% recovery) is commonly recovered by either zinc precipitation or carbon adsorption.

Cyanidation of low-grade (0.03 to 0.10 oz. per ton) oxidized ores through heap leaching has been given considerable interest for the last decade 11

primarily because gold values are located in the fracture fillings and cyanide solution is able to contact them at coarser sizes ranging from the run-of­ mine ore down to crushed product (minus 3/8"). In actual plant operation, the uncrushed or crushed ore, with or without agglomeration is placed in an impervious surface and percolation leached over a period ranging from 30 to 150 days or more. The length of leaching is dependent on the size of the ore/agglomerate, height of heap and the mineralogy of the valuable minerals as well the host rock. In general, gold recovery from 60 - 80 % is achieved. Like the pregnant liquor in the agitation leach, gold is recovered by either zinc cementation or adsorption into activated carbon (Bhappu, 1990).

2.3.1.3 Methods of Recovering Gold from Cyanide Leach Liquor

Gold in cyanide solutions are commonly recovered by either zmc cementation or carbon adsorption. Several recovering techniques, however, are being developed, two of which - resin technology and ion flotation - will be discussed.

2.3.1.3.1 Zinc Cementation (Merrill-Crowe Process)

The recovery of gold from the pregnant solution by zinc cementation/ precipitation has been practiced since the introduction of cyanidation. The process, which is commonly known as the "Merrill-Crowe Process" after the two scientists, C.W. Merrill, who developed the process in 1897 and T.B. Crowe, who introduced de-aeration prior to actual cementation in 1918, (Wan, 1990) is governed by the reaction (Jha, 1984): 12

2Au(CN)2- + Zn° = 2Au0 + Zn(CN)/- (4)

In practice, the leached pulp is passed through a series of countercurrent decantation (CCD) thickeners to initially separate the pregnant solution ' from the tailings. The overflow of the last thickener is passed through a filter and then to a clarifier to obtain clear pregnant solution. The solution is then de-aerated followed by the introduction of zinc. After cementation, the precipitated gold is filtered and finally smelted into gold bullion. Figure 2.1 shows the generalized flowsheet for gold recovery by cyanidation-zinc cementation process.

2.3.1.3.2 Carbon Adsorption Process

The recovery of gold from cyanide leach solution by carbon adsorption was patented as early as 1894 but did not gain much attention until the 1960s. The advancement of effective methods of eluting gold from the carbon which made possible the regeneration and recycling of activated carbon was responsible for the widespread use of carbon starting in the 1980s (Wan, 1990). At present, the carbon adsorption process commonly known as carbon-in-pulp (CIP) and carbon-in-leach (CIL) has become standard practice for gold recovery. It has become the predominant recovery method in use worldwide, accounting to about 44% of world production followed by solid-liquid separation/zinc precipitation at 30% (Marsden, 1993). In contrast to the Merill-Crowe process which requires efficient solid-liquid separation, the CIP/CIL involves the direct introduction activated carbon (usually of 6 - 28 mesh in size) into the cyanided pulp. In the CIP process, 13

RUN-OF-MINE GOLD ORE

' i CRUSHING.,______,1

' I HEAP PREPARATION I GRINDING! tailings ...------. -.--HEAPLEACIDNG AGITATED LEACHING

leach liquor COUNTERCURRENT tailings DECANTATION leach 11quor

FILTRATION

--t CLARIFICA TIO1""'·~1 .------

I DE-AERATior~ pregnant solution CN solution ZINC CEMENTATIO~1---- for re-use gold precipitate

SMELTING

GOLD BULLION

Figure 2.1 : Generalized Flowsheet for Gold Recovery by Cyanidation-Zinc Cementation Process 14

activated carbon is introduced after leaching through a series of carbon adsorption tanks where carbon moves in countercurrent with the pulp. In the CIL method, carbon is added during leaching of the ore wherein the carbon moves in countercurrent with the pulp in a series of agitated ' cyanidation tanks. The CIL method is practiced in carbonaceous ores to prevent the adsorption of gold cyanide complex on the organic matter which then reports to the tailings. By simultaneous leaching and adsorption, any gold dissolved is immediately retained by the carbon and not by the carbonaceous matter in the ore (Habashi, 1987).

After the gold is adsorbed, the loaded carbon is separated from the pulp by simple screening. The loaded carbon is then stripped-off of its gold by elution with hot caustic cyanide solution using either the Zadra method or the AARL (Anglo-American Research Laboratories) technique. The Zadra process involves circulating the eluate (1% NaOH, 0.2% NaCN, 98°C) continuously between the elution column and the electrowinning cell for 48 hours. The AARL technique separates the elution part from gold-winning wherein the carbon is presoak in 1% NaOH and 5% NaCN followed by elution with fresh water at 100 to 120°C for 8 to 12 hours. The fresh eluate is accumulated in storage and subsequently circulated between the electrowinning cell and the storage cell (Wan, 1990; Stanley, 1990). Electrowinning is conducted using steel wool cathodes at 2.5 to 3.5 volts per cell and amperage consistent with 30 to 40 % current efficiency (Bhappu, 1990). The loaded cathodes are smelted to produce gold bullion. Figure 2.2 illustrates a generalized flow diagram of cyanidation-carbon adsorption process. 15

RUN-OF-MINE GOLD ORE

I I CRUSHING! •• I HEAP PREPARATION I GRINDING ! tailings l ' ' HEAP LEACHING AGITATED LEACHING AGITATED LEACHING/ CARBON ADSORPTION I CIS CIP CIL I CARBON ADSORPTION I

to tailings/solution tank I I SCREENING I loaded carbon carbon for recycling ELUTION I

pregnant solution

ELECTROWINNING

gold cathode I SMELTIN

Figure 2.2: Generalized Flowsheetfor Gold Recovery by Cyanidation-Carbon Adsorption Process 16

The activated carbon after repeated use/re-cycling is reduced of its activity

due to the presence of CO3=, CaCO3, silica, and flotation reagents such as xanthates, frothers, and hwnic acid. The pores of the activated carbon are blocked by the presence of silica or iron as fine quarts, clay or calcine and ' severe deactivation are caused by frothers and xanthates. In order for the carbon to be utilized further, carbon reactivation such as acid washing to remove most of the organic salts and thermal regeneration to restore carbon activity are practiced.

2.3.1.3.3 Resin Technology

The recovery of precious metals from pregnant liquors by ion-exchange using resin-in-pulp (RIP) or resin-in-column (RIC) has been gaining attention over the past years. The process is similar to the carbon adsorption method wherein strong or weak-base resins are utilized to adsorb gold and other metals either directly from the cyanided pulp (RIP) or from pregnant solution (RIC). The stripping stage, however, does not require elevated temperatures and pressures to desorp precious metals from loaded resins (Bhappu, 1986). It is reported that resin technology could eliminate the disadvantages of carbon-adsorption process such as slow kinetics, carbon fouling due to salt precipitation and regeneration problems. The disadvantages of using resins, on the other hand includes size, density and mechanical stability problems (Wan, 1990). 17

The controlling equations involved in the resin technology are presented as follows : (Bhappu, 1986)

1. Adsorption stage : (R is an anion exchanger)

---+ RCl + NaAu(CN)2 +-- RAu(CN)2· + NaCl (5)

2. Stripping stage:

---+ RAu(CN)2· + NRtSCN +-- RSCN + NRtAu(CN)2 and, (6)

3. Electrowinning stage :

NRtAu(CN)2 ---+ Au0 + NRt + 2(CN)" (7)

Although the process chemistry appears to be straightforwardly simple, several investigations and experiments are conducted to develop superior weak-base resins and improved elution procedures for strong-base resins. Weak-base resins are generally selective and have favorable elution features but have low metal-loading capacity. Strong-base resins, on the other hand, have better metal loading capacity but present elution/gold stripping problems (Wan, 1990). The eventual success and application of this technology depends heavily on these developments. 18

2.3.1.3.4 Ion Flotation

Ion flotation is a process of collecting surface-inactive ions from aqueous solutions by the addition of surfactants (collectors) and the introduction of • air bubbles to the solution. The surface-inactive ions (also known as colligend) of one charge are attracted to the surfactant ions of opposite charge, the combination of which are adsorbed in the rising bubbles forming a foam product ( sub late) on the surface of the solution. Ion flotation differs from the more popular froth flotation (which will be discussed later in this thesis) as it collects ions from solutions rather than hydrophobic materials from a suspension of particulate solids (Nicol, 1992). Figure 2.3 is a schematic diagram of a continuous-flow ion flotation unit (Grieves, 1981).

The quantity of surfactant required in ion flotation is directly related to quantity of the colligend and is not influenced by the amount of solution. Theoretically, at least a stoichiometric equivalent is enough (Matis, 1991). The mechanism by which ions physically attain their selectivity is viewed on two assumptions. The first assumption is that the colligend in the bulk undergo an ion exchange reaction with the surfactant until equilibrium is reached and then the surfactant and its associated ions adsorb in the rising bubbles. In the second assumption, ion exchange takes place after the surfactant has adsorbed in the rising bubbles thereby creating a charged interface which must be balanced by the arrival of the counter ions (colligend) (Nicol, 1992). 19

Surfactant Storage Foam Stream

~_.,._ Residual Pump 0 Stream 0 0 0 0 0 0 0

Flowmeter

Moisture Trap

lVIixing Tank

--- Gas Source

Saturator

Figure 2.3 : Schematic Diagram of a Continuous Jon Flotation Unit

(Grieves, 1981) 20

The ion flotation technology, pioneered by Sebba in 1959, had find application in effluent treatment of various metal production and finishing operations, power plants and chemical industries in the 1970s. These industries use large amount of contact process water wherein the spent ' streams generally carry moderate concentrations of metal ions which have to be eliminated. In the past years, ion flotation has shown potential in the recovery of metal ions from dilute solutions (Matis, 1991 ).

The recovery of gold from extremely dilute alkaline cyanide solutions by ion flotation known as Gold Ion Flotation (GIF) was developed in Australia. The development of the GIF method was geared towards introducing a possible alternative to carbon adsorption and stripping stages in conventional gold recovery operation. It is envisioned that the most favorable feed material to GIF is clear leach liquor like heap leach run-off. The process involves the introduction of a cationic surfactant to attract

complex aurocyanide anion (Au(CN)2} Experiments conducted on a clarified heap leach liquor using quaternary ammonium surfactant attained >90% gold recovery (Nicol, 1992).

2.3.2 NON-CYANIDE LIXIVIANTS FOR GOLD DISSOLUTION

The development of other leaching reagents to dissolve gold has been gaining interest for the past years for several reasons. One of these reasons is the growing concern over environmental protection which make cyanidation on the disadvantage due to toxic nature of cyanide. Also, with the depletion of easy-to-mill gold ores, cyanide leaching is becoming less 21

attractive in treating complex ores containing cyanicides (sulfide minerals, copper oxide or silicates) which consume cyanide during processing, not to mention the difficulty of treating refractory gold ores where pre-treatment steps are necessary to enhance gold solubility by cyanide.

The non-cyanide lixiviants that have been given significant attention are thiourea, thiocyanate, thiosulfate and the halides (iodide, chloride and bromide).

2.3.2.1 Thiourea

Thiourea leaching of gold, in contrast to cyanidation, takes place in acidic medium (pH 1). With sufficient solution oxidation potential, a dimer, formamidine disulfide, is formed and act as the extractant. To achieve and maintain the oxidation potential of the solution in the appropriate range for dimer formation, an oxidant is necessary. Oxidants such as oxygen, potassium monopersulfate, potassium persulfate, hydrogen peroxide and ferric sulfate are used, with ferric sulfate as the most common (Eisele, 1989; Huyhua et. al., 1989).

The extraction of gold by thiourea (formamidine disulfide) is represented by the following reaction,

(8)

where gold is dissolved as a cationic complex (Eisele, 1989). 22

The recovery of gold from thiourea leach solution has also attracted a lot of attention in anticipation of the eventual commercialization of the process. Some of the methods studied include cementation by aluminum, lead, iron I or steel wool; carbon adsorption; direct electrowinning using graphite cathodes; acid cationic ion exchangers (resins); neutralization of the solution to pH 6.5 to precipitate gold; electrolysis; and pressure reduction. Though results of these experiments are encouraging they have their inherent disadvantages that need to be resolved (Dupuis, et. al., 1989; Deschenes, et. al., 1989).

Thiourea leaching offers some positive solutions over the use of cyanide. It having a low toxicity and being able to treat mild refractory ores were encouraging enough to gain interest from several researchers. The study by Eisele (1989), however, proved that cyanidation is more preferable over thiourea leaching after comparing the performance of the two processes over 14 different gold ores. It was indicated that if most ores are not leached by cyanide, they will not be leached any better by thiourea. The results of that study somewhat limited the possible applications of thiourea as an alternative lixiviant to cyanide.

2.3.2.2 Thiosulfate

Thiosulfates are compounds containing the group s2O32·, the two most

important salts of which are sodium thiosulfate, Na2S2O3 and ammonium 23

thiosulfate, (NI-Li)2S2O3. Thiosulfate dissolves gold by the following reaction:

The dissolution of gold by thiosulfate which should take place at alkaline medium due to the decomposition of thiosulfate at low pH, is affected by temperature, oxygen partial pressure, agitation and reagent concentration (Davis, 1991 ).

2.3.2.3 Thiocyanate

Thiocyanate ion, SCN-, which has a linear structure with the carbon atom in the center, is capable of forming covalent compounds and complexes. With oxidants such as ferric oxide or oxygen, gold is dissolved in thiocyanate

forming Au(SCN)2- or Au(SCN)4- complexes. Like thiourea, thiocyanate leaching is effected at acidic environment wherein soluble oxidizing agents such as ferric ions, hydrogen peroxide and Caro' s acid can be used. The information on the kinetics of gold dissolution by thiocyanate is not yet exhaustive and thereby needs further investigations to better understand the chemistry of this lixiviant to be applicable in the gold industry (Filho, et. al., 1989). 24

2.3.2.4 Halides

The capability of halides - iodide, bromide and chloride - to form complexes with gold has been long recognized (Davis, 1991). Chlorination, for one, was' used extensively in the 19th century before the cyanidation process was practiced (Eissler, 1896) while bromide was a known lixiviant for gold since 1846 (Davis, 1991).

Chlorination. Gold in the presence of chloride ion and strong oxidant is dissolved when the solution is highly acidic. Gold-chloro complexes such as AuC4- and Aucii- are formed when dissolved chlorine, hypochlorous acid or hypochlorite ion oxidizes gold in chloride media. The recovery of gold from the solution is accomplished by hydrogen sulfide, sulfur dioxide or ferrous sulfide precipitation or by carbon adsorption.

Bromide Leaching. The dissolution reaction of gold by bromide 1s governed by the reaction,

(10)

which is affected by factors such as bromide and gold concentrations and the electrochemical potentials of the anodic and cathodic processes. The reaction can take place at acidic, neutral or slightly alkaline media. The recovery of gold from bromide leach liquor could be attained by solvent extraction, ion exchange, carbon adsorption or cementation (Davis, 1991 ). 25

Iodide Leaching. The chemical reaction involved in the dissolution of gold by iodide ion is given by,

(11)

where the tri-iodide ion, 13-, act as the oxidant.

The gold-iodide complex formed from the reaction is the most stable among the halogens in aqueous environment. It is stable over a wide range of pH (up to pH 12) in contrast with the other two halides. The apparent effectiveness, even at low concentration, in leaching gold and the selectivity in leaching other minerals make iodide leaching a potential alternative to cyanidation (Davis, 1991 ).

2.3.3 GRAVITY CONCENTRATION

Gravity concentration is a method of separating minerals of different specific gravity by their relative movement in response to gravity and one or more forces (i.e., resistance to motion offered by a viscous fluid, such as water or air). For effective separation, it is important that a marked difference in specific gravity exists between the minerals and the gangue. Besides density, the size of the mineral particles also has a great influence in the effectiveness of gravity concentration methods wherein efficiency increases with particle size (Willis, 1979). 26

Gold, having a higher specific gravity over the minerals with which it is commonly associated, has been recovered since the earliest times by gravity methods. The effects however, of gold particle shape (i.e., flattened gold particles resulting from its malleability in alluvial deposits) and the hydrophobicity of fine gold particles significantly reduced its advantage over other gold separating systems developed over the past several years. Although the method is still widely used in recovering gold from alluvial deposits, the advent of cyanidation and flotation has limited its application in milled ores to the removal of coarse free gold (Richards, et. al., 1984).

The most commonly used equipment in the recovery of gold from alluvial and free-milling ores are jigs, riffled tables, shaking tables and sluices. While these equipment have their own advantages (technically or economically) over the other, their common limitation is in the recovery of fine gold particles (<100 µm). In recent years, with increasing energy costs, increasing reagent costs for both flotation and leaching systems, and increasing environmental awareness, there has been resurgence of interest in the use of gravity methods resulting to the development and/or innovations of gravity concentration equipment. This includes cone concentrators, spiral separators ( e.g. waterless spirals) , hydrocyclone systems and centrifugal gravity systems. These technological advancements now offers cost-effective systems for the recovery of gold down to finer sizes (Richards, et. al., 1984).

The establishment of large-scale operations may merit the adoption of the modem equipment described above but the economics of small-scale mining operations in underdeveloped areas may limit their applications. At present, 27

the proliferation of pocket mining in gold-rich countries like the Philippines is recognized where a significant amount of gold is produced from high­ grade vein type deposits. The processing of these free-milling ores involves conventional methods though at times crude by accepted standards. Crushing of ore is usually done manually followed by grinding in low­ capacity ball or rod mills. The ground ore is concentrated in sluice boxes. The sluice concentrates are amalgamated in rod mills and gold is finally recovered by retorting and smelting. Figure 2.4 shows a typical flowsheet in small-scale gold operation in the Philippines.

2.3.4 FROTH FLOTATION

Minerals within the earth's crust usually represent a highly heterogeneous mixture of solidified phases. The separation of these phases is achieved by exploiting their various physical and physiochemical properties such as color, specific gravity, size, shape, electrical charge, magnetic susceptibility, radioactivity and surface properties. Upon liberation to their appropriate sizes through crushing and/or grinding, these minerals could be separated by several methods which include sorting, gravity concentration, electrostatic separation magnetic separation, radiometric sorting, spherical agglomeration, selective flocculation and froth flotation (Hayes, 1993).

Froth flotation, along with spherical agglomeration and selective flocculation, is a process which exploits surface properties of the minerals (i.e., surface energy and surface potential).· When minerals are contacted with two fluid phases, that is, a liquid and gas, the differences of interfacial 28

IDGHGRADE VEIN-TYPE GOLD ORE

CRUSHING/GRINDING

! SLUICING I Taibngs Sluice Concentrate

AMALGAMATION

~ ~ ! PANNING I Tailings Amalgam&Excess Hg

I SQUEEZING! Hg Amalgam

RETORTING Hg Sponge Gold

SMELTING

,~ GOLD BULLION/DORE

Figure 2.4 : Typical Philippine Flowsheet for Small-Scale Gold Recovery Operation 29 free energies lead to a preferential wetting of some minerals by one fluid, namely water. This differentiation enables the non-wetted (hydrophobic) solids to be separated in a froth if the non-wetted solids are allowed into contact with air bubbles (Leja, 1982). Figure 2.5 shows a conventional flotation cell and a Jameson cell used in mineral froth flotation.

The advantage of froth flotation and other separation process exploiting surface properties is that there is a relative ease of adjusting and controlling interfacial and surface charges. While physical properties like magnetic susceptibility and specific gravity can not be changed, surface properties can be modified by the introduction of surfactants such as collectors, depressants, and activators. Collectors, which can either be anionic or cationic, are agents that promote the flotation of certain minerals by modifying their surfaces to be hydrophobic. Depressants are those that induce hydrophilicity of mineral surfaces while activators are those that reinforce the action of a given collector (Leja, 1982). In addition to these surface modifying reagents, another surfactant, a frother, is added to facilitate the attachment between the air bubble and the hydrophobic solid and to stabilize the froth as it arises to surfaces of the pulp. The success of froth flotation depends strongly on the proper and judicious use of specific collectors, modifying agents and frothers.

Froth flotation was first applied to sulfide ores at the end of the 19th century and has progressed into major industrial developments enabling the recovery of valuable minerals at very small particles and in very low grade or complex ores. Besides proper selection of reagents, other variables play significant roles in the efficiency of flotation. These include feed characteristics (particle size distribution for each mineral phase, 30

Compressed air ! Removal of ~~~oJoaded froth 0 O 0 0

0 0

0 0 Slurry in 0 Slurry out

(a)

Air

Washwater

Cell Concentrate

Tails Tailings valve

(b)

Figure 2.5: Flotation Cells - (a) Conventional Cell (b) Jameson Cell 31 mineralogical composition, particle morphology, specific gravity of each mineral phase, etc.), and machine parameters (volume, area, size, aeration rate, etc.) (Barbery, 1985).

Gold in its free state is hydrophobic, thereby, making it possible to be recovered by froth flotation. The use of this process however as the primary method to recover gold is not commonly practiced in favor of the well­ established technologies. Gold from placer deposits is concentrated by gravity methods while free gold from ores is recovered by both gravity and leaching ( cyanidation) methods. Flotation may however be applied to pre­ concentrate low grade ores before cyanidation (Lins, 1993). Since gold is closely associated with sulfides, most of the gold produced by froth flotation comes from the concentration of sulfide minerals. Gold particles may be distributed along the outside of sulfide minerals, at the boundaries between sulfide grains, in fine fractures or even totally enclosed in the sulfide minerals as microscopic or sub-microscopic inclusions. The most common sulfides with which gold is associated are pyrite and arsenopyrite. Other sulfides which may host gold are chalcopyrite, chalcocite-covellite, pyrrhotite, galena, etc. (De Cuyper, et. al., 1985).

In order to maximize recovery of gold in the flotation of sulfides careful evaluation of gold occurrence, liberation size and mineralogy of the host sulfide mineral/s should be taken into consideration. The following are some of the possible measures of maximizing gold recovery by froth flotation (De Cuyper, et. al., 1985; Lins, et. al., 1993; Yang, et. al., 1993): 32

• If the gold remains totally enclosed in the host base-metal sulfide after finest grinding, maximizing the recovery of the host mineral is of primary concern.

• In treating complex sulfide ores, losses of gold values in the less valuable component (i.e. iron sulfide minerals) should be avoided and it should be recovered in the right concentrate.

• As it is known that sodium sulfide, cyanide, lime and iron sulfate suppress gold flotation, their usage should be minimized when gold is present in the ore.

• When the ore contains course free gold particles (>200 µm), recovering them by gravity methods before flotation is advisable.

• The use of stronger collectors or combination of collectors is also recommended especially when the gold surface is coated with an oxide film.

These and other measures are to be considered on a case to case basis to enhance gold recovery by froth flotation.

2.4 AGGLOMERATION OF FINE COAL WITH OIL

'Liquid phase agglomeration' or 'agglomeration in suspension' are general terms used to describe the agglomeration of particles in aqueous or non- 33 aqueous suspension when a second liquid, which is totally immiscible with the first (dispersing/suspending) liquid, preferentially wets the particles and aggregates them forming spherical agglomerates (Sparks, et . al., 1989; Szymocha, et. al., 1989). This process is similar to froth flotation in the sense that the separation of the solid phases is dependent on their differences in surface properties. It differs only in the introduction of another phase which in the case of flotation is gas (i.e. air) while in the present case is another liquid (i.e. oil).

The agglomeration of hydrophobic materials in aqueous suspension with oil or low molecular weight hydrocarbons such as pentane or heptane is commonly referred to as oil agglomeration. It is one of the promising methods of cleaning coal where organic materials tend to be agglomerated in preference to the inorganic minerals (Allen, et. al., 1993). Although froth flotation is commonly practiced in the recovery of coal fines generated during modernized continuous and longwall mining, oil agglomeration offers more flexibility especially in the treatment of very fine (<150 µm) coal where froth flotation does not work effectively.

With adequate amount of oil and sufficient mechanical agitation, oil-coated particles collide with each other and form agglomerates due to the interfacial tension of the oil and the capillary attraction of the oil bridges between particles (Mehrotra, 1983). The amount of bridging liquid in the pore volume of the agglomerates affects to a great extent the tensile strength of the agglomerates, which is associated with capillary pressure. Three states of liquid saturation are identified, namely: pendular, fumicular and capillary. With small amount of bridging liquid, adjacent particles in the 34 agglomerates are held together by pendular bridges of liquid while with large amount where all the pore spaces are completely filled, the particles are held together by capillary forces. The intermediate between the two states of liquid saturation is called fumicular where liquid bridges coexist ' with some pores filled totally with the bonding liquid. It has been established that capillary state is attained when liquid saturation exceeds 80% while at more that 95%, significant changes in strength and plasticity are observed (Szymocha, et. al., 1989; Schubert, 1977). The changes in liquid saturation can be effected by either increasing the amount of the bonding liquid or by compaction as illustrated in Figure 2.6.

The various factors affecting the rate of agglomeration, the size of the agglomerates, the strength of the agglomerates, ash rejection and recovery of coal have been given much attention by researchers for the last several years.

Allen, et al. (1993) reported the effects of pH and ionic strength of the aqueous medium on coal recovery and separation efficiency. They observed that the rate and extent of agglomeration is affected by the changes in pH and ionic strength especially at lower oil dosages and noted that the rate of agglomeration is more sensitive to pH and ionic strength variations than the extent of agglomeration. The rate and extent of agglomeration increased with an increase in ionic strength. 35

Increasing the amount of bonding liquid

Compaction•••• • •• Pendular" Furnicular Capillary Particles suspended in bonding liquid

Figure 2. 6 : The States ofBonding Liquid Saturation

Labuschagne, et. al. (1989) experimented on the influence of oil composition, pH and temperature on selective agglomeration of coal. They reported the effects of double bonds, polar groups and alkyl groups in the bridging liquid on rate of organic recovery and percentage de-ashing. It was also noted that a decrease in slurry-pH (pH 12 - 4) increased the rate of organic recovery (agglomeration) due to the lowering of coal-water interaction or the water-binding energy thereby increasing the degree of hydrophobicity. This change of coal-water interaction as a function of pH is represented as follows,

(12) 36

which illustrates that the degree of coal hydrophobicity as a function of pH depends upon the amount of hydrophilic surface sites. The report further indicated that the moisture content of the coal reflects the tendency of a coal to interact with water such that the lower moisture content, the more hydrophobic' is the coal. The increase of organic recovery with the increase in temperature was also observed in the study. This was due to decrease in coal-oil interfacial tension favoring faster coal-oil contact and the decrease of oil viscosity favoring faster oil dispersion.

Various operating parameters/variables in coal agglomeration and recovery were reported by Bhattacharyya, et. al. (1977) and Swanson, et. al. (1977).

2.5 COAL-OIL AGGLOMERATION FOR GOLD RECOVERY

Coal-oil agglomeration technology for the recovery of gold is based on the hydrophobicity/oleophilicity of gold. In the early 1900' s, 'oil concentration' has already been introduced along with 'flotation', but the latter 'made it to the big time' commercially while the former did not quite get there. Clenn'.el in 1915 described 'oil concentration' as follows (Randol, 1992):

"in the Elmore oil concentration process, the mixture of crushed ore and water is fed into a slightly inclined revolving cylinder together with a considerable quantity of crude mineral oil. Owing to some physical action at present imperfectly understood, certain minerals, particularly those 37

having a bright 'metallic' surface, adhere to the oil. On discharging the contents of the cylinder into the tanks containing water, the oil floats to the top carrying with it the adhering minerals, which are then separated from it by means of rapidly revolving centrifugal extractor. By this means gold, galena, copper and iron pyrites and other lustrous minerals may be separated from earthy carbonates, oxides, etc. and from sand and siliceous minerals generally".

Presently, the recovery of gold by oil has been gaining attention by mineral processing engineers/researchers along with the oil agglomeration of carbonaceous materials such as coal. Though it is possible to recover gold by agglomerating them with oil, the amount of gold in the ore is usually so small that there is insufficient gold particles to form agglomerates. Thus, the need to use other hydrophobic materials (e.g. coal) to either form agglomerates together with gold or act as carrier of gold particles.

2.5.1 PATENTS FOR COAL-OIL AGGLOMERATION PROCESS

The use of coal-oil agglomeration technology for the recovery of gold became known to public when a patent, "Recovery of Metal Values from Mineral Ores by Incorporation in Coal-Oil Agglomeration", was granted to BP Australia Ltd in April of 1986. Two further patents were filed by BP group, one in December 1984 by BP Australia Ltd, "Recovery of Metal Values from Mineral Ores as Seeded Hydrocarbon Oil Agglomerates" and the other in July 1987 by the British Petroleum Company, "Separation 38

Process". These patents for the BP processes have become known as the Coal-Gold Agglomeration (CGA), which set out the basic principles of the coal-oil agglomeration for gold, platinum group of elements (PGEs) and base metal minerals recovery. Two other companies have also patented • variations of the coal-oil agglomeration, the Bateman Engineering International Ltd. which was granted a US Patent in July 1986, "Gold Recovery Processes" and the Charlton Mineral Associates Pty. Ltd. which lodged an application in August 1986 for "Mineral Recovery Process" (Bonney, 1992a). The latter is now known as the CARBAD Process.

2.5.1.1 The CGA Process

The CGA Process involves the addition of coal and oil to a slurry of gold ore/minerals followed by the intensive agitation to create the required agglomerate-gold particle collisions. Gold particles which are substantially free become attached to the coal-oil agglomerates when they collide and eventually the gold penetrate into the agglomerates. Parameters such as coal-oil ratio, oil type and agitation intensity determine the size of the agglomerates, which are typically less than 500 µm. They are contacted with the slurry either in counter-current operation or by recycling in co­ current operation. With the co-current operation, wherein majority of the testworks have been done, the agglomerates/slurry mixture passes through a series of contactors having sufficient time and agitation intensity to capture free gold particles onto the agglomerates. The agglomerates are separated from the slurry by flotation and recycled until they reached the desired gold loading. The gold loaded agglomerates are combusted to volatilize the oil 39

and burn the coal followed by smelting or intensive cyanidation and electro­ winning of the resultant ash (Bonney, 1992b).

2.5.1.2 The CARBAD Process

The CARBAD Process on the other hand utilizes pre-formed agglomerates approximately 3mm in diameter which are made into contact with the slurry through counter-current operation in a series of three continuous high shear contact tanks. The agglomerates are contained within each tank by screens until they sufficiently loaded with gold. To maintain hydrocarbon balance in the agglomerates, small amount of oil is continuously fed into the tanks. The loaded agglomerates are recovered by isolating the particular tank and pumping the contents over a screen. They are then agitated with dense solvent (intermediate to the density of the carbon and oil) resulting in the dissolution of the oil bridges and disintegration of the agglomerates to discreet particles. The slurry is pumped through a centrifuge and the resulting concentrate may be directly smelted or may require specific flux before smelting depending on the quantity and type of oleophilic material present. The solvent is distilled for re-use and the media is reformed to provide a continuous recycle of the materials. For small scale operations, where the economics of recycling the media is not justified, burning of the media is recommended (Cadzow, et.al., 1989). The mechanism by which the gold is captured by coal-oil agglomerates is illustrated in Figure 2. 7. 40

Gold Particle

Contact

Carl>ad Media

/ When media collides, capillary forces pwnp the particles within the media

Figure 2. 7: The CARBAD Mechanism

2.5.1.3 The Bateman Process

The operations embodied in the Bateman patent are similar to the CGA and CARBAD processes. The recovery of gold may be carried either mixing simultaneously coal, oil and the gold bearing particulate to form gold loaded agglomerated or using pre-formed agglomerates to collect the gold. The contactors may be stirred tank or a pipe with a degree of internal roughness to promote turbulence and saltation of agglomerates. The operation may be co-current using the pipe or counter-current using the stirred tank wherein the agglomerates are constrained by screens. The residence time is the magnitude of 2 minutes (1-3 meters /second) in the pipe or 6 minutes in 41

stirred tank. The agglomerates are separated from the barren slurry by conventional methods like screening. Pulp density is recommended to be in the range of 20 to 40 % by weight solids and oil to be 15 to 22% by weight of the non-gold particulate bulking material which could be carbon, polystyrene, ground coal, ground lignite, polyethylene, or other hydrophobic materials (Siddal, 1992).

2.5.2 APPLICATION OF COAL-OIL AGGLOMERATION ON GOLD RECOVERY

Bonney ( 1992b) presented laboratory and pilot scale tests conducted on coal-gold agglomeration. The original testwork to prove the merits of the process was carried out using synthetic mixture of high purity silica sand

ground to d80 of 100 µm and gold powder (99.9% purity) at d80 of20 µm.

The coal used which had an ash content of 18.5% was disc milled to d80 400 µm and the hydrophobic liquid used was commercial grade gas oil of 840 gpl density. The testwork was carried with varying slurry densities, agitation intensities, agglomerate concentrations, residence times and gold head grades. Under optimum conditions, over 95% gold recovery was obtained. A 30 kg/hr continuous laboratory test and a 1 tph pilot plant test were also conducted. The pilot plant tests were performed on a gold tailings deposit in Australia. Using diesel fuel with a density of 835 gpl and coal of 8% ash, gold recoveries range from 62% to 75% at different mesh of grind, agitation, intensities, collector (Potassium Amyl Xanthate) addition, residence times, pulp densities and tailings:agglomerate ratios. 42

In a related literature, (House et. al., 1992), laboratory scale tests on four gold bearing materials - synthetic silica/gold mixture, black sand gravity concentrate from Indonesia, gold tailings from France and tailings sample from Australia were discussed. Pilot plant tests on tailings were also ' included. In the laboratory, batch tests were performed on 2 kg. samples in a 5 liter baffled agitator. The slurry is agitated with coal and oil wherein agglomerates are formed very rapidly (1-2 min) and continuously stirred to appropriate ore-agglomerate contact time. Results showed that within 30 minutes, maximum gold recovery is achieved with essentially complete recovery being attained in the first two samples. It was also observed that contact time is independent of gold size and CGA performance is not dependent of head grade. The pilot plant designed for 1 tph capacity is composed of a series of contacter tanks. The ore and the agglomerates are mixed in these tanks where the outflow of the final contacter is pumped to a flotation unit separating the agglomerates/ concentrate from the spent ore. Ore tailings were pumped to 1mm mesh vibro-energy screen to collect any coarse agglomerates. The agglomerates were prepared when needed in an agitated vessel (0.2m3 ) with coal ground to d80 90 µm. The effects of ore particle size and attrition scrubbing were evaluated with other variables such as pulp density and agitation speed kept constant. Results revealed that at further grinding to d80 of 75µm from d80 of 470 µm, the recovery increased to 75% from 65%. Attrition scrubbing gave no significant difference therefore gold recoveries were not limited by the presence of abradable coating on gold particles. A test to demonstrate that agglomerates do not deteriorate in their ability to recover gold during prolonged period of time was also performed. A single batch of agglomerates was retained in the plant for over 200 hours and results showed that during the first 22 43

hours gold loading was up to only 4. lgpt. With the addition of collector however, the loading was constant up to 930 gpt for a 232 hours operation.

Bellamy, et. al. (1989) conducted experiments on a beach sand deposit in New Zealand. Laboratory tests were performed in 0.6 dm3 Pyrex beaker fitted with four vertical baffies using coal ground to d80 of 50 µm and diesel oil of 840 gpl density. As needed potassium amyl zanthate was added as collector. The pulp is conditioned for 5 minutes and then coal was added followed by the oil. After agglomeration for 30 minutes, the agglomerates were separated from the tailings, dried and analyzed by fire assay. The effects of dry and wet grinding and attrition scrubbing were evaluated to assess the degree of the liberation of the gold particles and their response to coal gold agglomeration. Further grinding beyond the optimum grinding conditions resulted to lowering of recoveries which may be caused by the reduction of hydrophobicity of gold by embedding of gangue particles or the deformation of the soft metallic flakes. Attrition scrubbing gave better results to most of the samples and collector additions and higher agitation rates improved recoveries.

Cadzow, et. al., (1989), presented results on the applications of the process as a primary recovery method for hard rock and alluvial ores, for gravity and flotation concentrate upgrade, and for tailings re-treatment. On hard rock application, the fineness of grind is the primary factor on gold recovery. CARBAD process has the capability to recover alluvial gold below the lower limit of gravity concentration (less than 50µm). In cases where there are other hydrophobic minerals present in the ore, pH modification and depressant addition may be necessary. The kinetics of the 44 process is much faster than conventional cyanidation where maxnnum recoveries were attained with residence times of 20 minutes.

Marciano et. al. ( 1994 ), presented results of the laboratory bench scale tests ' aimed at studying the merits of the coal-oil agglomeration as an alternative to amalgamation. The experiments were performed using a non-aerated Denver sub-A flotation cell, with rotation speed control and provided with three axial blade impellers, and Pyrex beakers (2 and 3 liter capacities) as contacting vessels. The agglomerating agents used were diesel oil, kerosene, Soya oil and corn oil whereas the hydrophobic bulking material was a metallurgical grade coal of 16.7% ash and 51.4% fixed carbon ground to 100% passing 0.42 mm and 36% passing 0.037 mm. The three types of gold-containing material used were: an artificial mixture of quartz under 0.21 mm and gold particles with size range of 0.30 to 0.42 mm; a gold concentrate ground to minus 0.21mm with 35% - 0.037 mm, assaying 980 gpt; and a gravity tailing of 13 gpt, a feed to a gold cyanidation plant.

Though both techniques presented satisfactory initial results, the method of using pre-formed agglomerates was used over the simultaneous ore-coal-oil method because recycling of the agglomerates to increase gold loading was included in the laboratory testworks program. Also, coal being cheaper was chosen for the experiments despite synthetic graphite produced similar strong spherical agglomerates. Results of the experiments revealed that with stirring speed of 1000 rpm, emulsification of the oil for 5 minutes prior to 10 minutes agglomeration produced agglomerates similar in size, shape and gold uptake capacity with fifteen minutes agglomeration without prior emulsification. Gold recovery increased with an increase in oil/coal ratio 45

up to 0.3, above which the agglomerates lose strength. Gold uptake was increased by 20% when the agglomerate:ore ratio was increase from 0.25 to 1.1. Though gold recovery increased with longer contact time in one cycle/ pass, it decreased with increasing number of cycles, which was caused by the agglomerates becoming more compact and that penetration of gold was more difficult. Experiments with Soya oil as agglomerant produced better results than with the diesel oil, but the formed agglomerates with Soya oil were more difficult to handle because they have the tendency to aggregate, becoming oily and losing strength while the agglomerates formed using diesel oil had greater strength and behaved as individual agglomerates.

2.6 THE WETTABILITY OF GOLD BY WATER

It has been regarded that gold is 'naturally' hydrophobic due to a large contact angle Q ( 60° - 65°) measured when a sessile drop of water or bubble rests of flat surface of gold. (See Figure 2.8) What renders gold hydrophobic has been an issue among surface chemistry experts and that they are divided whether the hydrophobicity is caused by contaminants or not.

Highly active high-energy surfaces (e.g., metals) have a strong tendency to pick-up contaminants which may either be molecular ( e.g. oxygen on metals or water in glass) or colloidal (e.g., hydrocarbonaceous particles in air contaminating surfaces in seconds). Gold, however, is inert to oxygen in the atmosphere and does not form stable oxide phase. Surface metal oxides may be capable of strong hydrogen bonding interaction (Schrader, 1992). Hydrogen bonding is greatly involved in the hydrophilic behavior of solid 46 surfaces and that its absence 1s correlated strongly with hydrophobic behavior (Good, et. al., 1992).

Sessile Drop

Figure 2.8 : Use of Sessile Drop ofBubble in Contact Angle Determination

Previous experience by Zisman and Bewig ( 1961) with metal surfaces has shown that, in the absence of contaminating organic layer, all metals including gold are hydrophilic. It was reported, however, by White in 1964 that he obtained dropwise condensation of water on gold (Q = 60±5°) under conditions by which the gold surface is free from oxide as well as organic contaminations. His observation was affirmed by Erb in 1965 who concluded that clean "high-energy surface" of metal substantially free of oxygen is not wettable by water and observed contact angles of 55-85° of water on gold. In 1965, Bewig and Zisman however, proved that oxide- and organic contaminants-free gold surfaces are wettable ( contact angle is zero) and indicated that the non-wettability of gold is caused by measurable contamination by hydrophobic organic contamination. White in 1966 in reference to Bewig and Zisman, claimed that the zero contact angle 47 measured by the two scientists was a result of hydrophilic contamination introduced during heating of the sample and imbedding of alumina abrasive during sample polishing. To further prove his hypothesis, Zisman together with Bernett in 1970, conducted experiments designed to eliminate possible hydrophilic contaminations suggested by White and Erb. They measured zero contact angle for water on gold. By using ultrahigh vacuum method of contact angle measurement designed to eliminate all possible contaminants, Schrader (1970) confirmed the findings of Zisman.

These conflicting findings on gold contact angle by water - gold hydrophobicity - are manifested in the two editions of Physical Chemistry of Metals by Adamson (1976 & 1990). In the third edition (1976), gold contact angle by water was listed in Table VII-2 as 66° while in a similar table (Table X-2) in the fifth edition (1990), the contact angle is listed as 0°. The notation of Adamson to the latter table was, to wit: ''As a somewha.t anecdotal aside, there ha.s been an interesting question as to whether gold is or not wet by water, with many publications on either side. This history has been reviewed by Smith. The present consensus seems tha.t gold is water-wet and that reports of non-wetting are a documentation of the ease by which gold surface becomes contaminated "

Following the logic of Adamson's notation, gold particles found in nature should be therefore hydrophobic since they are exposed to conditions (either by oxidation or hydrophobic organic contamin~tion) that could contaminate the surfaces. 48

CHAPTER3 ' SUMMARY OF LITERATURE REVIEW AND RESTATEMENT OF OBJECTIVES

The literature review conducted has manifested the following points:

1. Cyanidation is still the predominant process of recovering gold from its ores. Experiments on other lixiviants proved to be technically viable but economic consideration prevented them from taking off commercially.

2. Gravity concentration coupled with amalgamation is practiced in placer deposits as well as high grade free milling ores especially in small-scale mtmn.g areas.

3. Coal-oil agglomeration process 1s a potential environmental friendly alternative to recover gold.

4. Coal-oil agglomeration like flotation relies on the hydrophobicity of gold to be separated from other minerals.

5. The present consensus among surface chemistry experts is that 'clean' (i.e. free of oxide and organic contaminants) gold is water-wetted (hydrophilic) 49

and that gold hydrophobicity is manifested by its strong tendency to be contaminated by hydrophobic (organic) contaminants.

6. The preferential wetting of oleophilic /hydrophobic particles by oils forms the fundamental basis of separation of particulate materials from aqueous suspensions by oil agglomeration wherein with adequate amount of oil and sufficient mechanical agitation, oil coated particles (e.g. gold and coal) collide with each other and form agglomerates. Oil agglomeration of coal is affected by several variables which include, pH and ionic strength of the aqueous medium, the amount of the bridging liquid (oil), coal particles size, oil composition, temperature, agitation/stirring speed and presence of mineral matters.

7. Although it is possible to agglomerate gold particles, the amount of gold usually contained in ores is insufficient to form agglomerates. Thus the use of hydrophobic bulking material such as coal is necessary to agglomerate together with gold (in the case of BP's patented CGA process) or to act as carrier of gold particles (in the case of the Carbad process).

8. The kinetics of coal-oil agglomeration to recover gold is much faster than cyanidation where maximum recovery are attained with residence times of 20 - 30 minutes.

9. The variables that affect gold recovery are : oil:coal ratio, impeller speed, agglomerate:ore ratio and residence times.

10. On hard-rock applications, the fineness of grind is the primary factor to gold recovery. so

11. Other hydrophobic minerals ( e.g. sulfides) are also recovered by the agglomerates such that pH modification and depressant addition may be necessary.

12. Gold recovery decreases with the number of cycles the agglomerates are charged, which was caused by the agglomerates being more compacted and that gold penetration was more difficult.

13. Agglomerates formed at high stirring speed, with or without pnor emulsification of oil, are similar in size, shape and gold uptake.

By and large, the coal-oil agglomeration technology has been given sufficient attention to be considered as an alternative process to gold recovery. Most of the papers reviewed, however, were more on the application of the technology to specific types of ores and most often than not were targeted to promote the merits of the technologies patented. This study then was undertaken to give a further insight on the technology in an attempt to confirm and explain the causes of variables thus far experimented and try to investigate other areas that may explain further the fundamentals of gold recovery by coal-oil agglomeration.

The specific objectives of this study are:

1. To study the underlying principles/fundamentals by which gold 1s recovered by coal-oil agglomerates. 51

2. To test various types of hydrophobic bulking materials (i.e. coal) and agglomerating liquids (i.e. diesel oil) in relation to agglomeration and gold recovery.

3. To test the applicability of the process to different types of gold bearing materials (i.e. gravity concentrate, amalgamation tails).

4. To assess the effects of gold-associated sulfides on the recovery of gold. 52

CHAPTER4

RESEARCH METHODOLOGY

4.1 MATERIALS

4.1.1 GOLD-BEARING SAMPLES

4.1.1.1 Synthetic Mixtures

4.1.1.1.1 Gold Powder - Cyanidation Residue Synthetic Mixture (Blend A)

A gold powder - cyanidation residue synthetic mixture was prepared for the studies on the effects of various parameters (e.g., oil:coal ratios) on gold recovery including preliminary ( exploratory) experiments. This synthetic mixture (which will be referred to as Blend A in this study) was a blend of pure gold powder (99.99% purity) and residues (from laboratory cyanidation tests) supplied by Metcon Laboratories. The cyanidation residues, based on the mineralogical analysis of flotation concentrate and tailings of same head sample (see Appendix A), contained mostly silica and feldspar (comprising more than 93%) and small amounts of sulfides (<5%), mostly pyrite and of lesser abundance, chalcopyrite and sphalerite. 53

Blend A was prepared by initially mixing a known amount of gold with a small amount of residue in a mortar and pestle. This was done to break some aggregated gold into discreet particles in order for them to be easily and uniformly distributed in the bulk mixture. The resulting highly concentrated blend was mixed with the rest of the residue by repeated coning and quartering and rotary splitting. The bulk mixture weighing approximately 10 kilograms spiked with 0.9392 gram of pure gold was analyzed to have 92 ppm gold. Screen analysis is shown in Appendix B. I. By the use a rotary splitter, samples weighing 100-135 grams were taken, placed in air tight plastic bags and stored, ready for experiments.

4.1.1.1.2 Gold Powder-High Purity Silica Mixture (Blend BJ

Using similar procedure above, a synthetic mixture of gold and pure silica (Blend B) was prepared for the determination of effects of sulfides on gold recovery. Instead however of directly producing a bulk mixture, a concentrated blend was initially prepared. The silica matrix for this mixture is a combination of standard grades 60G and l00G silica at 2:1 ratio, the resulting screen analysis of which is also shown in Appendix B.1. For each test, the concentrated blend, with a gold assay of 875 ppm was mixed with the silica matrix and with desired amount of sulfides to produce 100 grams of sample at 100 ppm Au. For example, a test sample with 2 % sulfides would be a mixture of 2 grams of sulfides, 11.43 grams of concentrated mixture and 86.57 grams of blended silica. 54

4.1.1.2 Real Gold-Bearing Samples

Three real gold bearing samples namely, flotation-sluice concentrate, flotation-Knelson concentrate and amalgamation tailings from Philippines ' were used in the applicability tests. The flotation-sluice concentrate sample came from a copper concentrating plant where final concentrates are made to pass a sack-lined launder (sluice) before going to thickeners. The flotation-Knelson concentrate came from the same plant but was a product of further processing of the flotation-sluice concentrate by a Knelson concentrator. The amalgamation tailings sample came from one of the small-scale mining areas using combination of gravity methods and amalgamation in the recovery of free gold from high grade vein-type deposits. Gold analyses of the three samples were 78.5, 2087 and 9.4 ppm, respectively.

4.1.2 CARBONACEOUS MATERIALS

The carbonaceous minerals used in the study are : bituminous coals from Australia, natural graphite from Sri Lanka and lignitic coal from the Philippines.

4.1.2.1 Bituminous Coals

The first batch of metallurgical coals was taken from Australian Coal Industry Research Laboratories Limited (ACIRL). This consists of eight samples namely: SB024, SB040, SB035, SB008, LHD055, LHD056, 55

LHD040 and LHD045. Initially, each sample was ground to pass 300 µm screen by a high speed disk mill. The milled products were sampled and analyzed for ash contents, which ranged from 9.06% to 18.91%. Based on these ash analysis, three composite samples were formed. The first composite, named low ash coal (LAC) had an ash content of 10.00%; the second one, medium ash coal (MAC) had 12.93% and the third, high ash coal (HAC) had 18.64% ash. The particle size distributions, and proximate and ultimate analyses of these composites are given in Appendix B.2 and Appendix C, respectively.

The second coal from Australia was one of the coal samples from Minerals Engineering Laboratory at UNSW. Though the source was unknown it was found to have good agglomerating properties, thus it was utilized when the three composite coal samples described above were consumed. This sample was also disk milled to pass 300 µm screen. The screen analysis, and proximate and ultimate analyses of this sample (named as Aus Coal) are shown in Appendix B.3 and Appendix C, respectively.

4.1.2.2 Graphite

The natural graphite sample was given by James Cummings and Sons Pty. Ltd. The sample which contained 94% carbon was disk milled to pass 300 µm sieve. Screen analysis is attached in Appendix B.4 56

4.1.2.3 Lignitic Coal

The lignitic coal was taken from Cebu, Philippines. To be effectively used, the samples were to undergo pre-concentration (flotation) to increase carbon content anti lessen the inorganic materials present. The flotation concentrate did not initially formed large agglomerates but rather small and very fragile aggregates. A two-step agglomeration method was devised to form agglomerates suited for gold recovery stage. This will be described later in the study. The analysis of the flotation concentrate is included in Appendix C.

4.1.3 AGGLOMERATING LIQUIDS

The agglomerating liquids in this study were diesel fuel oil with a density of 840 gpl and kerosene with density of 790 gpl.

4.1.4 CHEMICAL REAGENTS

In the analysis of metallurgical products the following chemical reagents were used: • hydrochloric acid ( analytical grade) • nitric acid ( analytical grade) • 25 % (v/v) hydrochloric acid • Au, Fe, Cu, Zn and Pb Standard Solutions (1000 ppm) 57

For pH regulations, the following solutions were prepared using distilled water:

• 10 % (v/v) sulfuric acid solution • 10 % (w/v) sodium hydroxide solution

4.2 EQUIPMENT AND INSTRUMENTATION

4.2.1 EXPERIMENTALAPPARATUS

The experimental set-up used in agglomeration and gold recovery stages is illustrated in Figure 4.1

A F B

E A. Variable speed motor

C B. 4-axial blade impeller

C. Mixing vessel

D D. Water bath

E. Thermometer F. Iron stand

Figure 4.1 : Schematic Diagram of Experimental Set-up 58

The set-up consisted of a four-axial blade stainless steel impeller (0 = 60 mm.) attached to a variable speed motor (speed range : 60 -2000 rpm). The

mixing vessel was a 1.5 liter capacity glass cylinder (0 = 110 mm, H = 170 mm) fitted with four glass baffles. The impeller blade was positioned 8.0 mm from the bottom of the vessel. To maintain constant temperature the mixing vessel was placed in a water bath. In addition to these apparatus, a pH meter (Orion Model 520 A; range : -2.00 to 19.99 pH; resolution of 0.01 and accuracy of± 0.02 pH) was used in experiments involving the effects of variable pH.

4.2.2 SAMPLE PREPARATION EQUIPMENT

The following equipment are utilized for coal and gold bearing samples preparation :

• Jaw crusher • High-speed disk mill • Rotary Splitter • Screen Vibrator (Endecotts, Model E.V.S.I.) • Set of Screens (38 µm to 4. 75 mm)

4.2.3 ANALYTICAL EQUIPMENT

For the analyses of heads and metallurgical products, the following equipment/apparatus were required: 59

• Drying Oven • Hot Plate • Fume Hood • Weighing Balances 1. Mettler - Toledo Model# PB3002 Delta Range, range 0.5 to 3,100 g. Readability of 0.1 g. 2. Mettler Micro-balance Model #AE260 Delta Range, readability of0.0001 g. • Atomic Adsorption Spectrophotometer Varian Model Spectra AA20, Double Beam Operation

4.2.4 SCANNING ELECTRON MICROSCOPE

Microscopic examinations to determine gold attachment/penetration in the agglomerates was done with the aid of JXA-840 Scanning Electron Microscope. Quantitative analysis of gold was determined by SEM-EDS unit.

4.3 EXPERIMENTAL DESIGN

4.3.1 GENERAL PROCEDURES

Unless otherwise stated in some specific tests, the succeeding general procedures for agglomeration and gold recovery stages were followed. 60

4.3.1.1 Agglomeration

Agglomeration was conducted by agitating coal, oil and water in .the mixing vessel in the following sequence : •

1. conditioning 25 grams of coal in 400 ml distilled water for five minutes. 2. introducing desired amount of oil and agglomerating for 30 minutes 3. screening the formed agglomerates by a kitchen strainer 4. washing the agglomerates by a mild spray of wash water to remove un­ agglomerated inorganic particles

4.1.1.2 Gold Recovery Stage

In the same vessel used in agglomeration, the recovery of gold by the agglomerates proceeded as follows :

1. conditioning gold-bearing sample for 5 minutes in distilled water at pulp density of 20-25% solids 2. adding the pre-formed agglomerates and mixing for desired duration. 3. separating the loaded agglomerates from the tailings by simple screemng 4. drying the loaded agglomerates (80°C) and tailings (100°C) 5. weighing of dried products 6. analyzing the products for metal content 61

All tests were conducted at 25°C and at natural pH except when effects of pH were evaluated.

4.3.2 PRELIMINARY EXPERIMENTS

Preliminary test works were conducted to acquire an initial insight on gold recovery by coal-oil agglomerates. A total of nine (9) tests were conducted with varying impeller speed and mixing time (mixing time in this study refers to the agitated mixing of the pre-formed agglomerates and the gold bearing sample, i.e., gold recovery stage) under constant volume of diesel oil and constant weight of low ash coal (LAC). For each test, the impeller speed used for both agglomeration and gold recovery stages was constant. A summary of test conditions used in the preliminary tests are summarized in Table 4.1

Table 4.1: Summary of Test Conditions/or Preliminary Experiments Test# Wt. of Wt. of Vol. of Impeller Mixing Sample, g Coal, g Oil, ml Speed, rpm Time, min 1 131.70 25 11 750 30 2 129.35 25 11 500 30 3 130.52 25 11 300 30 4 132.39 25 11 750 60 5 132.12 25 11 500 60 6 128.68 25 11 300 60 7 131.92 25 11 750 90 8 124.25 25 11 500 90 9 132.18 25 11 300 90 62

4.3.3 EXPERIMENTS TO DETERMINE THE EFFECTS OF VARIOUS PARAMETERS ON GOLD RECOVERY

4.3.3.1 Agglomerate: Ore Ratios

' To determine the effects of different agglomerate to ore ratios, two series of tests were conducted. For the first series, the mixing time was limited to 15 minutes to emphasize the effects of the different ratios while for the second series it lasted for 120 minutes with simultaneous pulp and agglomerate sampling after 15, 30 and 60 minutes. (see section 4.3.5 for sampling procedures). In these tests, the weight of the ore was held 'relatively' constant while the amount of the agglomerates were varied. Agglomeration was done by batch of 25 grams of LAC. The summary of these conditions are tabulated in Table 4.2 below.

Table 4.2 : Experiments to Determine the Effects of Agglomerate:Ore Ratio Variations Series Wt.of Wt. of Vol. of Impeller Mixing #1 Sample, g Coal, g* Oil, ml* Speed, Time, min Test #1 126.37 25 11 750 15 Test #2 130.55 50 22 750 15 Test #3 134.16 75 33 750 15 Test #4 126.89 100 44 750 15 Test #5 133.37 125 55 750 15 Series #2 Test #1 106.68 12.5 5.25 750 120 Test #2 106.35 25 10.5 750 120 Test #3 106.24 50 21 750 120 Test #4 106.36 100 42 750 120 (*) Agglomeration was conducted by batch of 25 grams coal with 11 ml and 10.5 ml of diesel oil for series #1 and series #2, respectively. 63

4.3.3.2 Oil: Coal/Graphite Ratios

Table 4.3 shows the swnmary of test conditions to determine the effects of varying the amount of oil in relation to the amount of carbonaceous material on gold recovery as well as on agglomeration. Different types of carbonaceous materials (LAC, HAC and graphite) and oils ( diesel oil and kerosene) were used in these tests. For the first two series, screen analyses and surface area determinations of the agglomerates were conducted. ( see Appendix D for surface are calculations)

Table 4.3 : Experiments to Determine the Effect of Varying Oil : Coa VG rap, h. ite Ratios . Wt. of Type of Type of Impeller Mixing Series Sample, Carbonaceous OiV Speed, Time, # g MateriaVWt., g vol., ml rpm min

l 123-128 LAC Diesel 750 30 25 8 - 12 2 129-137 LAC Kerosene 750 30 25 9-12 3 126-130 LAC Diesel 1000 30 25 8 - 12 4 110-114 HAC Diesel 750 60 8 - 12 (45)* 5 120-121 Graphite Diesel 750& 90 25 8 - 12 1000 (30/60)* (* sampling time)

4.3.3.3 pH

The evaluation of the effects of pH on gold recovery was conducted using

sulfuric acid (H2SO4) and sodium hydroxide (NaOH) as regulators. 64

Following is the list of conditions on two series using LAC and HAC. {Table 4.4). The effect of pH will also be discussed in the tests conducted on sulfides.

Table 4.4 : Experiments' to Determine the Effects ofpH Wt.of Type of Coal/ Vol. of Impeller Mixing Series Sample, Wt., Oil, pH Speed, Time, g g ml rpm min 1 127-132 LAC/25 11 2-10 750 30 2 119-123 HAC/25 10 2-12 750 60

4.3.3.4 Particle Suspension

The results of the exploratory tests described in section 4.1.3.1 indicated poor gold recovery at lower impeller speed. This maybe due to the large sizes of agglomerates produced at lower speed (lesser surface area) or inadequate mixing which may resulted to poor particle suspension. The following series of tests (Table 4.5) was designed to eliminate the effect of agglomerate size and just center on the effect of gold particle suspension as effected by impeller speed variation. This was achieved by producing agglomerates at same impeller speed, 750 rpm, (i.e. to produce same size of agglomerates) and mixing at different speeds. To further evaluate gold particle suspension specially at low impeller speed (300 rpm), additional tests were conducted wherein the position of the impeller with respect to the bottom of the mixing vessel was varied. From the normal position of the impeller (8.0 mm elevation from the bottom) two tests were conducted, one 65

with a distance of 3 mm and the other, 20 mm. Pulp and agglomerate samples were taken after 30 and 60 minutes.

Table 4. 5 : Experiments to Determine the Effects of Particle Suspension Test# Wt.of Wt. of Vol. of lmpeller Mixing Sample, g Coal, g Oil, ml Speed, rpm Time, min 1 123.84 25 10 750 90 2 121.28 25 10 500 90 3 124.56 25 10 300 90 4 121.59 25 10 3001 90 5 122.95 25 10 300.t 90 1 impeller distance from bottom = 3 mm 2 impeller distance from bottom = 20 mm

4.3.3.5 Coal Particle Size

The attachment of gold particles to the agglomerates depends to a great extent on the strength and stability of the latter. Besides the amount of oil and degree of agitation, the size of the coal particles may affect agglomeration as well as gold recovery. To appraise this effect, three size ranges of coal particles were used in the study. Table 4.6 lists the conditions for the tests undertaken. 66

Table 4. 6 : Experiments to Determine the Effects of Coal Particle Size Wt.of Type of Coal Type of Impeller Mixing Series Sample, Coal/ Particle Oil/ Speed, Time, g Wt., g Size, µm vol., ml rpm min 1 123-127 LAC/ -53 Diesel/ 750 30 ' 25 +53 -75 11 +75 -212 2 129-132 LAC/ -53 Kerosene/ 750 30 25 +53 -75 11 +75 -212 3 123-125 HAC/ -53 Diesel 750 30 25 +53 -75 11 +75 -212

4.3.4 EXPERIMENTS TO DETERMINE THE EFFECTS OF SULFIDES

As discussed in chapter 2, section 2.3.3, gold, being hydrophobic, is recovered with sulfides during froth flotation. The associated occurrence of gold with sulfides may affect gold recovery by oil agglomeration because it has been established that both processes ( oil agglomeration and froth flotation) are influenced by their surface properties. The hydrophobic nature of sulfides though to a lesser degree than gold makes them capable of being captured by coal-oil agglomerates. Up to what extent, then, would the sulfides cause significant effect on the recovery of gold by this process?

To evaluate this point, series of tests using chalcopyrite, galena, sphalerite and pyrite, the more common sulfides with which gold is associated, were conducted. The experiments involved the addition of different amounts of these minerals in a synthetic mixture of gold and silica (Blend B) are summarized in Table 4.7. The effect of pH at constant sulfide content (2%) 67

was also investigated (Table 4.8). Besides gold analysis, the products were also analyzed for copper, iron, zinc and lead to determine the extent of their recovery in the agglomerates.

Table 4. 7 : Experiments to Determine the Effects of Sulfides on Gold R ecovery % Series Wt.of Type of Sulfides Wt. of Vol. of Impeller Mixin~ # Sample, Sulfide in Coal, Oil, Speed, Time, g Sample g. ml rpm min 1 100 Chalco- 0.5-5.0 25 10.25 750 60 pyrite 2 100 Pyrite 0.5-5.0 25 10.25 750 60 3 100 Galena 0.5-2.0 25 10.25 750 60 4 100 Sphale- 0.5-2.0 25 10.25 750 60 rite

Table 4.8: Experiments to Determine the Effects of pH on Gold and Sul/ides Recovery Wt. Series Wt. of Type of of Vol. of Impeller Mixing # Sample, Sulfide pH Coal, Oil, Speed, Time, g g. ml rpm min 1 100 Chalco- 2-12 25 10.25 750 60 pyrite 2 100 Pyrite 2-12 25 10.25 750 60 3 100 Galena 2-12 25 10.25 750 60 4 100 Sphale- 2-12 25 10.25 750 60 rite 68

4.3.5 LABORATORY EXPERIMENTS ·CONDUCTED IN THE PHILIPPINES

To study the applicability of coal-oil agglomeration process on gold bearing ' samples, and coal from the Philippines, actual laboratory test works were conducted at the Philippine Mines and Geosciences Bureau - Metallurgical Research Laboratory.

4.3.5.1 Tests on Real Gold Bearing Samples

The effects of oil:coal ratio, agglomerate:ore ratio, impeller speed and grinding were evaluated on the three real gold bearing samples described in section 4.1.1.2. Tables 4.9 - 4.11 summarize the conditions for these experiments.

Table 4.9 : Effects of Oil : Coal Ratios on Flotation-Knelson Concentrates Test Wt. of Wt. of Vol. of Impeller Mixing # Sample, g Coal, g Oil, ml Speed, rpm Time, min (Aus. Coal)

1 114.7 25 8 750 60

2 114.2 25 10 750 60

3 114.4 25 12 750 60 69

Table 4.10: Effect of Grinding and Impeller Speed on Amalgamation Tails Test Wt. of Wt. of Vol. of Impeller Mixin # Sample, g Coal, g Grinding Oil, ml Speed, g rpm Time,

1 93.5 25 none 9.5 750 60

2 95.5 25 none 9.5 1000 60

3 96.3 25 with 9.5 750 60

4 95.8 25 with 9.5 1000 60

Table 4.11 : Effect ofAgglomerate:Ore and Oil:Coal Ratios on Flotation- Gravity Concentrate Test Wt. of Wt. of Vol. of Impeller Mixing # Sample, Coal, Oil, Speed, Time, g g* ml* rpm min

1 49.45 12.5 4 750 60 2 49.45 25 8 750 60 3 49.58 37.5 12 750 60 4 49.75 12.5 4.75 750 60 5 49.82 25 9.5 750 60 6 49.85 37.5 14.25 750 60 7 49.32 12.5 5.5 750 60 8 49.41 25 11 750 60 9 49.86 37.5 16.5 750 60 (*) agglomeration was conducted by batch of 25 grams of coal with 8 ml oil (tests 1-3), 9.5 ml oil (tests 4-6) and 11ml oil (tests 7-9). 70

4.3.5.2 Tests on Philippine Lignitic Coal

The effectiveness of a Philippine lignitic coal in gold recovery in comparison with bituminous coals from Australia was also investigated ' (Table 4.12). The agglomerating property of the local coal was inferior to the bituminous coals due to the inherent property of lignite of being less hydrophobic. To able to use the local coal in the gold recovery stage, a two-stage agglomeration was undertaken. The first stage involved agitating a certain amount of oil with coal-water mixture and separating the initially formed small agglomerates (by screening) from the un-agglomerated inorganic minerals present in the coal. In the second step, the initial agglomerates were again agitated in water with additional amount of oil where the actual growth of the agglomerates took place. During the first step, the growth of the agglomerates was hindered by the presence of inorganic (hydrophilic) particles (Szymocha, et. al., 1989).

Table 4.12 :Experiments on the Efficiency of Philippine Coal on Svnthetic Mixture Test Wt.of Wt. of Vol. of Impeller Agglomeration Mixing # Sample, Coal, Oil, Speed, Time, Time, g g ml rpm min min 1 119.30 35 5.01/7.52 750 151/302 60

2 122.40 35 5.0/9.0 750 15/30 60

3 119.70 35 5.0/7.5 1000 15/30 60

4 120.40 35 5.0/6.0 1000 15/30 60

5 115.00 25 10/2.0 750 60/30 60

6 116.70 25 10/2.0 750 60/30 60 1 and 2 denotes the first and second stages of agglomeration, respectively 71

4.3.6 DETERMINATION OF THE EXTENT OF GOLD ATTACHMENT/ PENETRATION IN THE AGGLOMERATE

The CARBAD mechanism in Figure 2. 7 illustrates how the gold is captured by the agglomerates through the capillary forces of the bridging liquid. The extent of gold penetration, however, was not fully explained to define where the gold is captured and/or attached. Are the gold particles captured just at the surface or have they penetrated within the agglomerates? If the gold particle is made an integral part of the agglomerate, how many particle-deep is it situated? To answer these vague issues, an in-depth examination of the gold-agglomerate structure through the aid of a powerful electron scanning microscope was undertaken.

To prepare gold loaded agglomerates for microscopic examination, same procedures described in section 4.1.1 were followed. In this case however, the synthetic mixture prepared was of very high gold content to produce highly loaded agglomerates. Several attempts to scan loaded agglomerates from the previous tests did not produce positive results since looking for gold particles in these agglomerates with gold contents of 300 - 400 ppm, was very difficult and time consuming. Hence, 100 gram synthetic mixture (gold powder and pure silica) with a gold content of 2,000 ppm, was mixed with approximately 10 grams of agglomerates to produce agglomerates with gold content of about 20,000 ppm. The loading lasted for 120 minutes with agglomerate samples taken after 15, 30, 45 and 60 minutes.

To investigate the actual depth of attachment or penetration, both the surface and the cross-section of the agglomerates were scanned. The cross- 72

section investigation was made possible by carefully cutting the agglomerates with a razor blade and mounting them upside down. To successfully make a near perfect cleave, several attempts were necessary due to the very fragile nature of the agglomerates. The cut agglomerates as ' well as whole agglomerate samples were mounted in aluminum stubs ( l "0) using silver paste. A pair of surgical tweezers was used to pick-up and transfer the agglomerate samples. Figure 4.2 illustrates the procedure in agglomerate sample preparation.

In the actual microscopic examination, the cross-section is viewed in areas not affected by cutting (denoted by check marks in Fig. 4.2) due to the possibility that gold particles may have been moved to the inner part of the agglomerate by the cutting edge of the razor blade. In this case,. any gold particle viewed in the affected area ( denoted by X marks) may not be the representative gold particle which actually penetrated/moved through the agglomerate. In similar case, mounting the specimens in resins was not possible because by polishing action, gold may be moved from one place to another since the agglomerates were not rigidly compact solids.

4.4 ANALYTICAL PROCEDURES

4.4.1 ANALYSIS OF HEADS AND METALLURGICAL PRODUCTS

The determination of metal contents of head samples and metallurgical products was undertaken by Atomic Adsorption Spectrophotometry method. The solid samples were digested in aqua regia (3:1, hydrochloric:nitric acid 73

I Razor --./ Blade //

+-- Loaded Agglomerate

a. Cutting of the agglo1nerate

b. Cleaved agglomerate shmving cutting-affected areas.

Silver AltUlll.lllUll ---- Paste Stub -----.

c. Mounted samples ready for carbon coating

Figure 4. 2 : Schematic Diagram of Agglomerate Preparation for Scanning Electron Microscopy 74

solution) and diluted in 25% (v/v) hydrochloric solution. Au, Cu, Fe, Zn and Pb standard solutions were prepared with 25% (v/v) hydrochloric solution matrix. Prior to digestion, the samples were roasted at 815°C to eliminate carbon and sulfides which may otherwise adsorb gold from the ' solution. The detailed procedure is discussed in Appendix E. I.

4.4.2 ASH DETERMINATION

The ash contents of the coal samples were determined by following the ASTM Standard Method for Ash Determination attached in Appendix E.2.

4.5 SIMULTANEOUS PULP AND AGGLOMERATE SAMPLING PROCEDURE

The use of a pipette has been the general practice in solution or pulp sampling in agitated systems. To get a representative sample from the system, sampling is done while the solution/pulp is in motion. In the case of this study however, simultaneous sampling of pulp and agglomerate during specific time intervals was not possible using the ordinary pipette. A sampling apparatus was therefore devised to accommodate large agglomerates together with the pulp. The device consisted with a round flask fabricated with two glass tube openings at opposite ends, a 8-mm 0 rubber hose fitted at one end and a rubber pipette filler at the other end. (Figure 4.3) Operating it like an ordinary pipette gave satisfactory results. 75

round flask A la, // ~~ ~ _____} f rubber pipette ~ ) filler rubber hose ~

Figure 4.3 : Simultaneous Pulp and Agglomerate Sampler

4.6 METAL RECOVERY CALCULATIONS

The evaluations of test results in this study were based on the amount of metal recovered in the agglomerates. In the case of the experiments related to the studies on the effects of different variables, only gold recoveries were calculated while those testworks related to the effects of sulfides, metals of copper, iron, zinc and lead were included. Percent metal recovery was calculated as follows :

Aa % Metal Recovery = X 100 Aa + Tt 76

where A = weight of the agglomerate T = weight of the tailings a = metal analysis of agglomerate in ppm ' t = metal analysis of tailings in ppm 77

CHAPTERS

RESULTS AND DISCUSSION

5.1 RESULTS OF PRELIMINARY EXPERIMENTS

Exploratory experiments were conducted to determine how coal-oil agglomerates are formed, how agglomerates behave when they are mixed with a pulp and how much gold is recovered in the agglomerates. The initial results gave good indications on the effects of mixing time and impeller speed on agglomeration and gold recovery.

Agglomerates formed at higher stirring speed were found to be smaller than those formed at lower speed due to higher destructive forces effected by high shear energy. (Swanson, 1977). With two balancing forces affecting agglomeration - adhesion force (mainly caused by capillary force exerted by bridging liquid, e.g., oil) and destructive force (caused by the shearing effect of the impeller) - agglomeration was observed to be completed after 30 minutes. Within this period and with the amount of oil added, the equilibrium state may have been reached wherein the agglomerates were strong, rigid and did not posses sufficient plasticity to deform and form bonds on collision. (Szymocha, et.al., 1989) 78

It was observed during the gold recovery stage that some agglomerate particles were chipped-off from the main agglomerates especially at higher impeller speeds, thereby requiring the use of screen with finer opening ( other than, the ordinary kitchen strainer) during agglomerate-pulp separation. In order to minimize possible abrading effect of the pulp, the coarser portions of the synthetic mixture ( + 212 µm) were screened-out in the succeeding experiments.

Gold recoveries at different mixing time and impeller speed from these preliminary experiments are tabulated and graphically illustrated in Table 5.1, and Figure 5.1 and 5.2.

Table 5.1 : Results of Preliminary Experiments IMPELLER· GOLD RECOVERY SPEED (%) rpm MIXING TIME , min. 30 60 90 750 93.1 98.4 98.9 500 66.9 87.6 88.1 300 43.2 62.8 64.3 Agglomerate:Ore Ratio (AOR) = 0.23; Oil:Coal Ratio, w/w (OCR)= 0.37 Pulp Density = 25% solids 79

Fi ure 5.1: E ecto Mixin Time on Gold Recover

100 -<> 0 0--- 80 -~ -t- ~ 60 ~ u0 ~ 40 =: -<>-750 rpm "C - --0-500rpm ~ 20 -A-300rpm

0 0 20 40 60 80 100 Mixing Time (min)

AOR = 0.23, OCR= 0.37, 25% solids

Figure 5.2 : Effect of Impeller Speed on Gold Recovery (Variable A lomerate Size 100

- 80 -~ t 60 u~ ~ 40 "C -<>-30 min ~ 20 --0-60min -A-90min

0+----i------+----+-----+----+------1 200 300 400 500 600 700 800 Impeller Speed (rpm) AOR = 0.23, OCR= 0.37, 25% solids

It can be deduced from these figures that gold recovery increased with time and impeller speed. The dependence of gold recovery on mixing time however was evident only up to 60 minutes, after which, increase in gold recovery was minimal if not insignificant. 80

The increase on gold recovery with stirring speed could be attributed to the combined effects of inter-particle collisions, size of agglomerates and gold particle suspension. Higher speeds gave better chances of agglomerate­ gold collisiofis that even only after 30 minutes, gold recovered at 750 rpm was already more that 93% while at 300 rpm only 43% of gold was collected by the agglomerates. With same amount of coal and oil, larger agglomerates formed at lower speed were lesser in number and had lower surface area than those produced at higher speed. As coal-oil agglomeration relies on the attachment of gold particles in the agglomerates, the larger surface area available and greater number of particles, the better chances of gold capture. (Bellamy, et. al., 1989) The low gold uptake at lower impeller speed even at longer mixing time could have been caused by limited probability of gold-agglomerate collisions due to poor gold particle suspension in the pulp. Some of the gold particles could have remained in the bottom of the vessel due to high specific gravity.

Further evaluation of these initial findings especially on the effects of agglomerate number and surface area, inter-particle contact and particle suspension are presented in the succeeding discussions.

5.2 EFFECTS OF AGGLOMERATE:ORE RATIOS

To evaluate the effects of agglomerate number and available surface area on gold uptake without the influence of impeller speed and agglomerate size, varying amounts of agglomerates, formed at constant speed and oil:coal ratio (OCR), were mixed with ore at fairly constant weight (i.e., different 81

agglomerate:ore ratios - AOR). The results of these tests are shown in · Tables 5.2 and 5.3, and Figures 5.3 and 5.4

Time

Agglomerate:Ore Ratio 0.24 0.48 0.73 1.02 1.36 Gold Recove % 91.3 94.3 98.0 98.9 99.0 OCR= 0.37, 750 rpm, 25% solids

Table 5.3 : Effects ofAeelomerate:Ore Ratio at Extended Mixing Time Mixing Gold Recovery Time % min A1!1domerate:Ore Ratio 0.13 0.28 0.54 1.03 15 91.6 94.2 96.5 98.7 30 93.4 96.4 98.5 98.9 60 96.3 99.1 99.2 99.3 120 99.3 99.5 99.7 99.8 OCR= 0.35, 20% solids, 750 rpm

100 99

-C~ 98 -t' 97 ~ ~ 96 u 95 ~ "Cl 94 -~ 93 92 91 0 0.2 0.4 0.6 0.8 1.2 1.4

Agglomerate:Ore Ratio

OCR= 0.37, 25% solids, 750 rpm 82

lomerate:Ore Ratio on Gold Recove 120 min

100 99

';{e. 98 ' -0 - 97 G,)t- > 96 C MixingTiJne u 95 ~ ~15min: "C 94 ---

OCR= o.,35, 20% solids, 750 rpm

Fi ure 5.5: E 100 99 -. 98 "I- -t- 97 ~ 96 8 Agglomerate:Ore ~ 95 ~0.13 ::S 94 ri ---Cr0.28 \sl 93 -A-0.54 -0--1.03 92 91 -+----+----+----t------.1------t-----t 0 20 40 60 80 100 120 Mixing Time, min OCR= 0.35, 20% solids, 750 rpm 83

The results of varying the agglomerate:ore ratio (Figure 5.3) revealed that gold recovery was increased by about 8% when the ratio was increased from 0.24 to 1.38 (agglomerate number and surface area was increased by a factor of 5.75). Above AOR of 1, there was no appreciable increase in gold recovery which means that at high agglomerate to ore ratios, kinetics of gold extraction is fast such that maximum recovery is already achieved even only after 15 minutes of residence time. Maximum attainable recovery (or final recovery) will be referred to in this paper as the point where recovery is not significantly increasing even at extended mixing time.

The results of the second series of tests (Figure 5.4), while confirming the outcome of first series, showed that at lower AOR (0.13), maximum recovery was not yet attained even at 60 minutes contact time as compared to those attained at AOR of 0.28 and above. This indicates that at lower AOR, more time is needed (or increased impeller speed as the case may be) for successful inter-particle collisions between the agglomerates and gold particles. Maximum gold recoveries were attained for all AORs after 120 minutes (Figure 5.5) with gold loadings of 88-750 ppm, way below the suggested normal operating loading of 20,000 ppm (Cadzow, et. al., 1989). This observation suggests that while available surface area and number of agglomerates affect the rate of gold recovery, it does not have much influence on maximum attainable gold recovery.

5.3 EFFECTS OF OIL:COAL RATIOS

In the agglomeration of coal by oil, the amount of oil and the degree of agitation control the size and strength of agglomerates. At constant stirring 84

speed, the size is mainly dependent on the amount of oil, where below the saturation point (over-critical oil addition), the diameter of the agglomerate increases with oil addition. At lower oil dosage, initial agglomerates do not have enough oil in their surfaces to bond with each other thereby restricting

$ them to have secondary or tertiary growth. Second growth of agglomerates takes place when, after the consolidation period, primary agglomerates achieved wet surfaces ( enough oil to bond with other primary agglomerates). Tertiary growth is possible after the compacted secondary agglomerates still attained oily surfaces (Szymocha, 1989). With this in mind, it is possible therefore to vary the number and the surface area of agglomerates without changing the weight of coal thereby having constant agglomerate to ore ratio. In gold recovery by coal-oil agglomerates, however, gold particles are attracted by oil due to their hydrophobic/oleophilic nature, hence, the question arises whether the lowering of oil in the system would also lower gold recovery or vice versa. The results given in Tables 5.4 to 5.6, and Figures 5.6 to 5.10, explain the significance of different oil:coal ratios on agglomeration and gold recovery.

5.3.1 EFFECT OF OIL:COAL RATIOS ON AGGLOMERATE SIZE, SURFACE AREA AND NUMBER

From Table 5.4 and Figure 5.6 below, increasing the amount of oil or kerosene at constant weight of coal (LAC) increases agglomerate size while it decreases surface area and number of agglomerates formed. ( see Appendix D. I and D.2 for surface area determination) At higher oil dosages, secondary or tertiary growth were possible such that bigger agglomerates were formed. At ratios of 0.403 and 0.384 for diesel and 85 kerosene, respectively, however, the agglomerates formed were observed to be mushy and unstable such that during the gold recovery stage small particles were chipped-off and reported to the tailings. At these ratios, the critical oil dosages may have been reached wherein agglomerates tend to lose strength .

It can be seen in Figure 5.6 that at high amount of oil (above OCR= 0.35), agglomerates formed with kerosene were larger than those formed with diesel. Actually, some of the bigger agglomerates formed with kerosene (+3.35 mm) were aggregates of smaller agglomerates starting to form bigger agglomerates.

Table 5.4 : Effects of Variable Diesel Oil/Kerosene:Coal Ratios on Agglomerate Size and Surface Area, No. of Agglomerates an dG0 ldRecovery Oil:Coal Agglomerates Available Surface No. of Gold 2 Ratios Size, d50 mm Area, mm Agglomerates Recovery (%) Diesel Oil 0.269 0.623 208,993 181,864 99.1 0.302 1.322 103,242 22,102 96.3 0.336 2.359 55,380 3,396 96.0 0.370 2.974 45,545 2,086 94.7 0.403 3.215 37,766 1,796 91.4 Kerosene 0.288 0.900 158,445 70,598 97.7 0.320 1.836 77,436 8,098 92.9 0.352 2.825 49,618 2,576 92.2 0.384 3.252 35,355 1,284 90.4 AOR (Diesel)= 0.24, AOR (kerosene)= 0.20; 750 rpm, agglomeration time= 30 minutes, mixing time = 3 0 minutes 86

Figure 5.6: Effects ofOiVKerosene Ratios on Agglomerate Surface Area and Size 2.SE+0S 4.5

N 4 ~ 2.0E+oS 3.5 ~ { 3 -=i l.5E+o5 GS' 8 2.5 .!::i 00 .e ~diesel .., ; 2 $ 00 l.0E+oS -x-kerosene =... $ ~ 1.5 El f e a 1 'i .S 5.0E+o4 -< ., 0.5

O.OE+oo ------+ 0 0.25 0.27 0.29 0.31 0.33 0.35 0.37 0.39 0.41 Oil:Coal Ratio (w/w) AOR (Diesel) = 0.24, AOR (kerosene) = 0.20; weight of coal = 25 grams, 750 rpm impeller speed, agglomeration time = 30 minutes, mixing time = 30 minutes

5.3.2 EFFECTS OF AGGLOMERATE SURFACE AREA AND SIZE ON GOLD RECOVERY

Table 5.4 above and Figures 5.7 and 5.8 show that gold recovery depends on the amount of oil available at the swface of the agglomerates as dictated by the size and number of agglomerates, not on the total amount of oil in the system. With smaller agglomerates and corresponding greater in number, available swface area of agglomerates formed at lower oil dosage was greater, hence higher gold uptake. 87

99

~ 97 0 -t, 95 ~... uQ 93 ~ 'C 91 J- odiesel 89 a kerosene

87 0 0.5 1 1.5 2 2.5 3 3.5 Agglomerates Size, d50 mm

AOR (Diesel) = 0.24, AOR (kerosene)= 0.20; 750 rpm, agglomeration time= 30 minutes, mixing time = 30 minutes

Figure 5.8: Effect ofAvailable Surface Area on Gold Recovery

100 99 98 ~ -0 97 -t- 96 <> ~... uQ 95 ~ 94 'C 93 -Q o diesel c., 92 a kerosene 91 90 0 5 10 15 20 25 Available Surface Area, (x 10,000) mm1 AOR (Diesel) = 0.24, AOR (kerosene) = 0.20; 750 rpm, agglomeration time= 30 minutes, mixing time = 30 minutes

From the above figures, gold recoveries for diesel-formed agglomerates were greater ·than those made with kerosene even at same surface area 88

available. This could be attributed to the more stable agglomerates formed with diesel. With kerosene, some agglomerate particles were abraded during mixing with the pulp and by screening, these particles which maybe loaded with gold went with the tailings. As mentioned by Labuschange (1989), lighte'r oils form weaker agglomerates and kerosene is lighter than diesel.

5.3.3 SUMMARY OF EFFECTS OF DIFFERENT OIL:COAL RATIOS ON GOLD RECOVERY

Table 5.5 and Figure 5.9 show the results of the tests conducted on oil:coal ratios using other type of coal, high ash coal (HAC), extended mixing time and higher impeller speed. At higher agitation rate and longer mixing time, the influence of varying oil:coal ratios is lessen due to the amplification of agglomerate-gold particle collisions. This observation is similar to what was perceived in the agglomerate:ore ratio tests, where available surface area and number of agglomerates do affect the rate of gold recovery but not the maximum attainable gold recovery. Final gold recoveries are dictated by mixing time and impeller speed.

Table 5.5: Effect of Oil:Coal Ratios on Gold Recovery at Extended Time an dlltht~i er Lmpe ll er S'.ipee d Oil:Coal Gold Recovery (%) Ratio LAC, 750 rpm LAC, 1000 rpm HAC, 750 rpm J0min J0min 4Smin 60min 0.269 99.0 99.5 98.5 99.0 0.302 96.3 97.5 98.6 0.336 96.0 98.6 96.2 98.1 0.370 94.7 95.8 97.4 0.403 91.4 96.9 94.3 96.4 AOR = 0.24, 25% solids, LAC: low ash coal, HAC: high ash coal 89

100 99

'-;;R- 98 ~ 97 t' 96 u~ 95 -lr-LAC:Diesel, 1000 rpm, 30 minutes ~ 94 -x-HAC:Diesel, 750 rpm, 60 minutes "C 93 l JC HAC:Diesel, 750 rpm,45 minutes '-' 92 o LAC:Diesel, 750 rpm, 30 minutes 91 90 -+------f---+----+------f---+----+------l----+-- 0.25 0.27 0.29 0.31 0.33 0.35 0.37 0.39 0.41 Oil:Coal Ratio (w/w) AOR = 0.24, 25% solids

5.3.4 EFFECT OF OIL:GRAPHITE RATIOS ON GOLD RECOVERY

Another series of tests was conducted to verify the effect of the amount oil on gold recovery. This was conducted using graphite instead of coal. The results shown in Table 5.6 and Figure 5.10 were similar to what was obtained using coal, where gold recovery increases with decreasing amount of oil, increasing agitation rate and longer mixing time 90

Table 5.6: Effect of Oil:Graphite Ratios at Different Impeller Speeds and Mt.,x,ng Tt,meson GldR0 ecovery Impeller Speed, rpm Oil: 750 1000 Graphite Mixing Time, min Ratio (w/w} 30 60 90 30 60 90 0.302 90.2 95.1 96.8 94.3 98.5 99.3 0.336 89.0 93.2 95.7 93.7 97.9 99.2 0.370 82.0 92.1 95.7 93.2 97.8 99.0 0.403 78.1 90.4 94.4 93.1 96.8 98.9 AOR = 0.23, 23%sohds

Figure 5.10: Effect of Oil:Graphite Ratios at Different Impeller Speeds and Mixin Times on Gold Recove 100 0 0 2 I: i l!Jr--- ~ 6 6 95 ~ -6 ":J:. =ts= :)( -0 -=l!:: - 90 --a t-G,) u> -0- I 000 rpm, 90 min Q 85 ~ -x:- 1000 rpm, 60 min "CS 80 -x- 1000 rpm, 30 min -~ ---lr-750 rpm, 90 min 75 -0-750 rpm, 60 min ~ 750 rpm, 30 min 70 0.290 0.310 0.330 0.350 0.370 0.390 0.410 Oil:Graphite Ratio (w/w)

AOR = 0.23, 23%solids

5.4 EFFECTS OF pH ON GOLD RECOVERY

The effects of pH variations are shown in Table 5. 7 and Figure 5.11. The results showed that while lower recovery was obtained at pH 2, gold 91 extraction was not affected significantly by the variation of pH from 4 to 12. The low recovery obtained at pH 2 with 30 minutes mixing time may be due to the removal of organic and/or oxide contamination on gold surface by excess H+ ions. As indicated in the findings of Zisman (1970) and Schrader (1970) that clean gold surfaces free of contaminants are hydrophilic, then it is likely that at pH 2, there was an initial 'cleaning' effect that rendered them to be less hydrophobic. At longer exposure to the pulp however, the initially 'cleaned' gold surfaces may have been re-contaminated such that after 60 minutes gold recovery was increased (see 60 minute line of Figure 5.11). The findings of Zisman and Schrader that oxide or organic contaminants are the factors that make gold hydrophobic were in agreement with the fairly constant high gold recovery at high pH where superficial oxide contamination is possible. As reported further by Adamson (1990), the non-wetting (hydrophobicity) of gold is caused by the ease by which gold surface becomes contaminated.

Table 5. 7 : Et 'l"ect ofpH on Gold Recovery Gold Recovery (%) pH LAC, 30min HAC, 60 min 2 79.8 94.8 3 98.1 4 97.8 99.1 5 98.6 6 97.2 98.2 8 97.4 10 96.9 98.2 12 98.0 AOR (LAC) = 0.23, AOR (HAC) = 0.25; 24% solids (LAC), 22% solids (HAC); 750 rpm 92

100

95 -C~ -t- ' ~ 90 u~ ~ 85 "C - ~LAC,30min ~ 80 -0-HAC, 60 min

75 0 2 4 6 8 10 12 14 pH AOR (LAC)= 0.23, AOR (RAC)= 0.25; 24% solids (LAC), 22% solids (RAC); 750 rpm

5.5 EFFECT OF PARTICLE SUSPENSION

As observed in the preliminary experiments, gold recovery was influenced by the combined effects of number of agglomerates, available surface area and stirring rate. From the previous two sections the dependence of gold recovery on the first two factors as dictated by the amount and size of agglomerate was evaluated. This section focuses on the effects of gold particles suspension by varying the agitation rate and position of impeller at constant agglomerate size and number. The results are shown in Tables 5.8 and 5.9; and Figure 5.12 and 5.13. 93

Table 5. 8 : Effects or Impeller Speed on Gold Recoverv Mixing Time Gold Recovery (%) min Impeller Speed, rpm 300 500 750 30 48.9 69.5 91.9 60 62.5 86.7 97.8 90 64.4 90.5 98.5 AOR = 0.21, OCR= 0.336, 23% solids, impeller elevation (normal)= 8.0mm, impeller speed (agglomeration)= 750 rpm

Table 5.9: Effect ofImpeller Elevation on Gold Recovery Mixing Time Gold Recovery (%) min Impeller Elevation (mm) 3 8 20 30 61.5 48.9 34.1 60 78.9 62.5 40.2 90 82.2 64.4 43.7 AOR = 0.21, OCR= 0.336, 23% solids, impeller speed (gold recovery)= 300 rpm, impeller speed (agglomeration)= 750 rpm

100

90

~ -e.., 80 t t 70 c.J ~ 60 -+-30 min = -0-60min < -tr-90min 50

40 +------t----+----+------t----+------1 200 300 400 500 600 700 800 Impeller Speed (rpm) AOR = 0.21, OCR= 0.336, 23% solids, impeller elevation (normal)= 8.0mm, impeller speed (agglomeration)= 750 rpm 94

90

80 -~ !!,... 70

Q,jt' 8 60 Q,j i=i:: 50 = --0-30min < -0-60min 40 ---6-90 min

30 +------+------+------+------1 0 5 10 15 20 Impeller Elevation (mm)

AOR = 0.21, OCR= 0.336, 23% solids, impeller speed= 300 rpm

Figure 5.12 manifested that even without the influence of amount and sizes of agglomerates, gold recovery is greatly affected by the rate of agitation. This clearly illustrates that an increase in inter-particle collisions increases the chances of successful adhesion of gold particles to the agglomerates. The low gold recoveries obtained at slower stirring rates even at prolonged mixing time indicates that gold particles may not be uniformly agitated (i.e., just stay at the bottom due to high density) within the pulp such that some of the gold particles were not able to collide with the agglomerates. This observation was validated by the subsequent experiments wherein the impeller position was adjusted relative to the bottom of the mixing vessel. In Figure 5.13, gold recovery at 300 rpm was increased by almost 18% when the impeller elevation was lowered from 8 mm to 3 mm. This resulting increase explains that gold particles were not adequately agitated and suspended at low rpm and high impeller elevation. This was further confirmed by the analysis of the samples taken after 30 and 60 minutes of 95

lDlXlllg. Compared to the final tailings analysis, the 30- and 60- minute samples gave lower gold content. This implies that samples taken at the mid-section of the pulp were not true (representative) samples because some of the gold particles were not suspended. In theses cases, computation of gold recoveries were based on the agglomerates and adjusted tailings analysis calculated from the head grade. Gold recovery was even more minimized when the impeller was positioned 20 mm from the bottom.

5.6 EFFECTS OF COAL PARTICLE SIZE ON AGGLOMERATION AND GOLD RECOVERY

The outcome of the test works evaluating the importance of coal particle size (Table 5.10) revealed that using agglomerates formed with fine coal sizes gave higher gold recoveries. This observation could be ascribed to the size and strength of the agglomerates where smaller and stronger agglomerates were produced using finer coal particles (Swanson, 1977). According to Hogg (1989), the overall (tensile) strength of an agglomerate could be estimated from the equation:

(1- Ea) F er=

where Ea is the agglomerate porosity, x is the size of individual particles and F is the force between them. From the equation, agglomerate strength is inversely proportional to particle size. Weaker agglomerates produced from +75 - 212 µm coal fractions gave poorer gold recovery due the chipping-off of small agglomerate particles during agitation which reported to the tailings carrying with them attached gold particles. 96

Table 5.10: Effect of Coal Particle Size on Gold Recovery Coal Particle Size LAC/ HAC/ MAC/ Range, µm Kerosene Diesel Diesel -53 95.3 95.4 97.2 + 53 , 75 92.5 92.9 95.6 + 75 -212 87.4 89.5 91.6 OCR= 0.37; AOR (LAC)= 0.189, AOR (HAC) = 0.23, AOR (MAC)= 0.23; 24% solids mixing time= 30 minutes

5.7 EFFECTS OF SULFIDES ON GOLD RECOVERY AND RECOVERY OF SULFIDES BY COAL-OIL AGGLOMERATES

5. 7.1 EFFECTS OF VARYING AMOUNTS OF SULFIDES ON GOLD RECOVERY

The results shown in Table 5.11 and Figure 5.14 indicates that gold recovery is not affected by the presence of commonly associated sulfides. This could be attributed to the fact that gold is more hydrophobic than the sulfides and the amount of sulfides in the feed ( up to 5% in the case of chalcopyrite and pyrite) were not able to significantly interfere in gold attachment to the agglomerates. Although the graph shows a decreasing trend of gold recovery especially at 5% chalcopyrite, the decrease was just a fraction of a percent. This decreasing trend in gold recovery with increasing amounts of sulfides however, may indicate that at much higher sulfides content gold recovery could be significantly affected. 97

Table 5.11: Effects of Varying Amounts ofSulfides on Gold Recovery % Sulfides Gold Recovery (%) in Feed Chalcopyrite Pyrite Galena Sphalerite

0 99.9 99.9 99.9 99.9 0.5 99.9 99.9 99.9 99.9 1.0 99.8 99.9 99.8 99.9 2.0 99.8 99.9 99.8 99.8 5.0 99.2 99.8 AOR = 0.34, OCR= 0.344, 20% solids, 750 rpm

Figure 5.14: Effects of Varying Amounts ofSulfides on Gold Recovery

100

'I-- -t- ~.... uQ 99 ~ -O-CuFeS2 "'O -O---FeS2 Q -t.!) ---lr-PbS --M-ZnS

98 0 1 2 3 4 5 % Sulfides in Feed

AOR = 0.34, OCR= 0.344, 20% solids, 750 rpm

5.7.2 SULFIDES RECOVERY BY COAL-OIL AGGLOMERATES

The base metals analysis (Cu for chalcopyrite, Fe for pyrite, Pb for galena and Zn for sphalerite) of the metallurgical products obtained from the same experiments as above showed that sulfides are also recoverable by the agglomerates (Randol, 1992). It can be seen from Table 5.12 and Figure 98

5.15 that chalcopyrite, galena and pyrite exhibited good response to oil agglomeration while sphalerite did not. The poor recovery of sphalerite could be attributed to its being oxidized since the sample used was a ground stock material while the other three were freshly milled. '

Table 5.12: Recovery ofSulfides by Coal-Oil Agglomerates % Sulfides Sulfides Recovery (%) in Feed CuFeS2 FeS2 PbS ZnS 0.5 99.1 94.6 98.9 29.4 1.0 99.2 94.5 98.2 9.0 2.0 99.2 92.9 98.3 4.1 5.0 99.2 86.4 AOR = 0.34, OCR= 0.344, 20% solids, 750 rpm

Figure 5.15: Recovery ofSulfides by Coal-Oil Agglomerates

100 S---a ~ 90 : 80 -C~ 70 -t' ~ 60 > u0 50 --<>- CuFeS2 ~ 40 ~FeS2 -....o:s 30 ~ ---tr-PbS ~ 20 """*""ZnS 10 0 0 I 2 3 4 5 % Sulfides in Feed AOR = 0.34, OCR= 0.344, 20% solids, 750 rpm 99

5. 7.3 EFFECT OF pH ON GOLD AND SULFIDES RECOVERY

The significance of pH variation without the presence of sulfides evaluated in section 5.5, wherein above pH 4, gold recovery was not affected. The ' results shown below (Tables 5.13 and 5.14 and Figures 5.16 and 5.17 ) gave similar results on gold recovery. The recovery of sulfides however, varied with pH especially with pyrite and sphalerite. Pyrite recovery was significantly decreased at pH range 8 to 12 by the addition of sodium hydroxide. This trend is similar to flotation where pyrite is depressed at high pH by lime (CaO). The slight increase in sphalerite recovery at pH 2 suggests that high amount of sulfuric acid addition may have sulfidized the mineral but not enough to maximize recovery. Longer conditioning time or

addition of an activator ( e.g. CuSO4) similar to flotation may be necessary to enhance sphalerite recovery by coal-oil agglomerates. Figures 5.18 to 5 .21 show the actual pictures of agglomerates loaded with sulfides at natural pH.

Table 5.13: Effects ofpH on Gold Recovery at 2% Sulfide Matrix % Au Recovery pH no sulfide CuFeS2 FeS2 PbS ZnS 2 99.6 99.6 99.6 99.6 99.6 4 99.8 99.8 99.8 99.7 99.7 6 99.9 99.8 99.9 99.8 99.8 8 99.9 99.9 99.9 99.8 99.8 10 99.9 99.9 99.9 99.8 99.7 12 99.8 99.9 99.8 99.8 99.8 AOR = 0.313, OCR= 0.344, 20% solids, 750 rpm, 2% sulfides in feed, mixing time = 60 minutes 100

Figure 5.16: Effects ofpH on Gold Recovery at 2% Sulfide Matrix

100.0

99.8 t- ~ e> u 99.6 ~ ~no sulfide <= 99.4 -a- CuFeS2 -6- FeS2 0~ --M- PbS 99.2 -ZnS

99.0 0 2 4 6 8 10 12 pH AOR = 0.313, OCR= 0.344, 20% solids, 750 rpm, 2% sulfides in feed, mixing time = 60 minutes

able 5.14: EJ '"ects of pH on Sul(ides Recovery Sulfides Recovery pH CuFeS2 FeS2 PbS ZnS 99.0 90.1 98.4 25.5 4 98.9 69.1 97.9 3.1 6 99.2 92.8 98.3 4.2 8 98.9 14.8 98.6 1.8 10 98.9 3.7 98.7 1.7 12 97.2 0.0 96.7 1.6 AOR = 0.313, OCR= 0.344, 20% solids, 750 rpm, 2% sulfides in feed, mixing time = 60 minutes 101

100 ti ti 0 0 ~ 90

;;.-.. 80 i. ...~ 70 0 ~ 60 ~ i::i::: 50 "1 ~ CuFeS2 ~ "'O 40 i;:: -0- FeS2 30 - -f:r- PbS rJ)= 20 ---*"- ZnS 10 0 0 2 4 6 8 10 12 pH AOR = 0.313, OCR= 0.344, 20% solids, 750 rpm, 2% sulfides in feed, mixing time = 60 minutes

lomerates with Cha/co rite

5% Chalcopyrite in Feed 1mm 102

5 % Pyrite in Feed 1mm

lomerates with Galena

2% Galena in Feed 1mm 103

2% Sphalerite in Feed 1mm

5.8 RESULTS OF LABORATORY EXPERIMENTS CONDUCTED IN THE PHILIPPINES

5.8. l TESTS ON REAL GOLD BEARING SAMPLES

The data obtained from the experiments conducted on three gold bearing samples from the Philippines showed that gold mineralogy and its association with the rest of the minerals in the ore are important factors to be considered in the extraction of gold by coal-oil agglomerates.

The low gold recovery from a highly loaded copper concentrate (2,087 ppm Au) suggests that the presence of large amount of sulfides may affect gold attachment to the agglomerates (Table 5.15). To consider that the sample was a gravity concentrate from a final copper concentrate means that the amount of chalcopyrite could be around 65% (computed from assumed 104 marketable copper concentrate grade of 22% Cu). From the previous experiments on sulfides using synthetic mixtures, there was a decreasing trend of gold recovery as sulfide percentage was increase. Though the decrease was not significant up to 5%, a 65% sulfide matrix would likely affect gold recovery' not to mention the enhanced hydrophobicity of the sulfides from residual collectors. The excessive amounts of sulfides may have competed with the gold for available sites in the agglomerate surfaces such that even at low oil:coal ratio (higher surface area), gold recovery was not significantly increased.

Table 5.15: Recovery of Gold from Flotation-Knelson Concentrate at Different Oil:Coal Ratios Oil :Coal Gold Recovery (%) Ratio (w/w) 0.269 50.8 0.336 49.6 0.403 42.5 AOR = 0.27, 22%solids, 750 rpm, 60 minutes mixing time

The combined effects of agglomerate:ore and oil:coal ratios on gold recovery on a flotation-sluice concentrate (78 ppm Au) are shown in Table 5.16 and Figure 5.22. It can be seen from the results that only up to 83.3% of the gold was recovered from the concentrate as compared to more than 99% extraction obtained from the synthetic mixture (Blend A) at similar operating parameters. This relatively low recovery may be attributed to the interference of too much sulfides present in the system, similar to what was discussed above. 105

Table 5.16: Recovery of Gold from Flotation-Sluice Concentrate at Diffii eren t A ee,omerl ati e: Orean d Oi: .l C oa l R at· JOS. Gold Recovery (%) Agglomerate: Oil:Coal Ratio Ore Ratio 0.256 0.314 0.352 0.299 59.5 44.8 44.5 0.613 77.6 72.9 64.1 0.918 83.3 79.0 67.8 10.5% solids, 60 minutes mixing time

Figure 5.22: Recovery of Gold from Flotation-Sluice Concentrate at Di erent A lomerate:Ore and Oil:Coal Ratios. 90

80 ';;;R -0 -t- 70 Go> ;i.. =CJ 60 ~ Oil:Coal Ratio "Cl 50 ~0.256 ~- ~0.314 40 ---a\-0.352 30 0.2 0.3 0.4 0.5 0.6 0.7 0.8 0.9 l Agglomerate:Ore Ratio (w/w) 10.5% solids, 60 minutes mixing time

The significance of the degree of particle liberation on gold recovery is shown in Table 5.17. By disk milling the amalgamation tails (9.4 ppm Au) from d80 of 258 µm to 136µm, gold extraction was increased three-fold. Higher impeller speed also increased recovery. 106

Table 5.17: Effect of Grinding and Impeller Speed on Gold Recovery firom A margal mat· ion 1lai "ls • Impeller Ore Particle Size Gold Speed dsoµm Recovery (%) 750 258 16.87 1000 258 19.21 750 136 61.56 1000 136 77.94 AOR = 0.29, OCR= 0.32, 190/o solids, 60 minutes mixing time

5.8.2 TESTS ON PHILIPPINE LIGNITIC COAL

The results of the tests (Table 5.18) conducted using a local lignitic coal on synthetic mixture (Blend A) revealed that although gold extraction were over 90%, these were still inferior to the recoveries using bituminous coals. This may be credited to weaker and less stable agglomerates formed by the less hydrophobic coal. Even if it was possible to agglomerate the lignitic coal sample by two-stage agglomeration, the resulting maximum gold recovery was only 96.35 as compared to more than 99% using the bituminous coal.

Table 5.18: Agglomeration and Gold Recovery Using Philippine Lignitic Coal Wt. of Initial Additional Weight of Impeller Gold Coal, Oil, Oil, Agglomerate, Speed Recovery 2 ml ml 2 rpm (%) 25 5 7.5 18.1 750 91.8 25 5 9 19.8 750 90.5 35 5 7.5 25.6 1000 94.0 35 5 6 25.8 1000 96.4 107

5.8.3 SUMMARY ON THE RESULTS OF TESTS CONDUCTED IN THE PHILIPPINES

The outcome of the applicability tests conducted in the Philippines indicated that beside knowing the importance of operating variables such as oil:coal ratio, agglomerate:ore ratio, impeller speed and pH, the mineralogy of gold including liberation size and its association with other minerals should be taken into account for optimum recovery by coal-oil agglomerates. Furthermore, the strength and stability of the agglomerates, as affected by the type of coal, to withstand abrasion brought about by the solid particlulates in the pulp and degree of agitation should always be considered to minimize gold losses in the tailings. 108

CHAPTER6

DETERMINATION OF THE EXTENT OF GOLD ATTACHMENT/PENETRATION IN THE AGGLOMERATES

It has been established in the preceding chapter that the stability of agglomerate surfaces has a notable consequence on gold recovery in the sense that gold particles situated in the surfaces may be lost when fine agglomerate particles are shaved-off from the main agglomerates due to shear regime infused by the solid particulates in the pulp. In addition to the losses due to agitation on weaker agglomerates, gold particles may also be detached in the final separation of the agglomerates from the pulp ( screening) if they are just attached at the surface. Also, in actual application, agglomerates are re-loaded for several times to increase the amount of gold in the agglomerates before final recovery of gold either by direct smelting or physical separation ( e.g. gravity concentration) of gold from coal after dis-integrating the agglomerates into discreet particles. (Cadzow, 1989) It is therefore appropriate to study the mechanism by which gold is captured by the agglomerates. This chapter explores the depth of penetration of gold particles in the agglomerates as effected by the length of mixing time with impeller speed at 750 rpm. 109

6.1 PHOTOGRAPHS OF AGGLOMERATE SURFACES

The following photographs reveal that, as mixing time progresses, gold particles are drawn within the agglomerates. Within 30 minutes mixing, (Figures 6.1 and 6.2), a lot of noticeable gold particles are situated in the surface while after 45 minutes (Figure 6.3), lesser gold particles are seen. In the photographs of agglomerates at 60 and 120 minutes (Figures 6.4 and 6.5) no gold particles are recognized at the surfaces. Actually, the gold particles observed in the first three images are not individual/discreet gold grains but rather clusters of them as would be further revealed in the SEM images. After 60 minutes, these clusters of gold powders may have been embedded in agglomerate surfaces as individual grains such that at the given resolution of the photographs with 60 and 120 minutes, these were not noticeable. 110

Imm 111 112

6.2 SEM IMAGES OF AGGLOMERATE SURFACES

To explore further the attachment of gold in the agglomerate surfaces, scanning electron microscopy was conducted where qualitative analysis of gold particles were determined by SEM-EDS unit. Figure 6.6 is a typical EDS pattern confirming the presence of gold in the agglomerates.

Figure 6. 6. Typical EDS Pattern for Gold Analysis

X-RAY: 0 - 20 keU L i v e: 10 Os Pre set: 10 0 s R e:ma. i n i n g: 0 s Rea. l: 113s 12% De:a.d

< •7 10 • 920 k e:U FS= 2K eh 556= 44 cts MEM1:SAMPLE GOLD 113

The following SEM images illustrate further the observations in the previous photographs that gold particles are made as integral part of the agglomerates. The clusters of gold grains (viewed like individual particles in corresponding photograph) however in Figure 6. 7 may not be observed in actual operation or in the previous experiments conducted because the feed used in the present study was a highly concentrated mixture (2,000 ppm) producing agglomerates with almost 20,000 ppm Au. Although it was suggested by Cadzow (1989) that agglomerates could be loaded up to 20,000 ppm for normal operations, the mechanism by which the gold grains are captured can be different in such a way that several loadings are necessary since common gold bearing ores are way below 2,000 ppm Au. It could be viewed in Figure 6.8 that even at only 15 minutes mixing time some of the gold particles are already embedded in the surface. The images after 60 and 120 minutes (Figures 6.9 and 6.10) revealed that gold grains are already trapped between coal particles. 11 4

Figure 6.8: SEM Image ofAgglomerate Surface at 15 minutes (Higher Magnification) 11 5

6.3 SEM IMAGES OF AGGLOMERATE CROSS-SECTIONS

To finally determine the extent of gold penetration, cross-sectional areas of the agglomerates were extensively scanned and qualitatively analyzed for the presence of gold particles within the agglomerates. Examination of several agglomerates at 15 minutes mixing time gave negative results, where no gold grain was observed to have penetrated inside. On agglomerates loaded for 60 and 120 minutes however, gold particles were found inside the agglomerates within several microns away from the surface. Figure 6.11 shows two gold grains (-10 µm) present about 10 to 30 µm from the surface. Figure 6.12 represents another 60 minute-loaded

agglomerate where three gold particles (- 2 to 5 µm) are observed about 40 to 50 µm from the surface. Figure 6.13 exhibits a view of an agglomerate after 120 minutes contact time. Two gold particles were scanned, one having a size of about 10 µm embedded just at the surface and the other of 11 6 about 6 µm situated more or less 50 µm from the surface. In the last figure ( 6.14 ), three gold grains are found at different locations relative to the surface. Two small particles (- 2-3 µm) detected at distances 45 and 30 µm from the surface and one bigger grain (- 10 µm) at 40 µm from the surface.

It would be of interest to note that some of the gold grains found inside the agglomerates (2 grains in Figure 6.11 , the smaller grain in Figure 6.13 and the three particles in Figure 6.14) are situated in a round pit. This space around the particles may represent the space of oil that enveloped them as they were driven inside the agglomerates. It could be suggested that the driving forces that caused the gold particles moved within the agglomerates were: (1) capillary force exerted by the bridging fluid ( oil); (2) external forces brought about by the inter-particle (i.e. agglomerate-gold and/or agglomerate-agglomerate) collisions; and (3) the additional shearing forces contributed by the baffles and the impellers.

Figure 6.11: SEM Image of Agglomerate Cross-section at 60 minutes (#])~--~~~=~~-~-~~!'!!!"'l"'-~~- 117

Figure 6.12: SEM Image of Agglomerate Cross-section at 60 minutes (#2)

Figure 6.13: SEM Image of Agglomerate Cross-section at 120 minutes (#]) 118

Figure 6.14: SEM Image of Agglomerate Cross-section atl20 minutes (#2) - . agglomerate surface ' 119

CHAPTER 7

CONCLUSIONS AND RECOMMENDATIONS

7.1 CONCLUSIONS

The fundamental studies and applicability test works conducted on the merits of coal-oil agglomeration for the recovery of gold draw the following conclusions :

1. Coal-Oil Agglomeration is a technically viable process of recovering liberated gold from gold-bearing ores and minerals. Experience with a synthetic mixture revealed that almost 100 % of the gold is recoverable including very fine gold powders down to 2 µm in size.

2. The recovery of gold depends heavily on the successful collisions of agglomerates and gold particles brought about by adequate agitation and enough mixing time. Stirring rates below 750 rpm were not sufficient to maximize gold particles suspension (i.e., gold particles remained at the bottom due to high density) thus, preventing gold­ agglomerate contact resulting to poorer final gold recoveries. Maximum recoveries for 300, 500 and 750 rpm were 64%, 90% and 99% respectively. Improving stirring mechanism, however, by 120

lowering the impeller from 8 mm to 3 mm above the bottom of the vessel, increased gold particle distribution in the pulp and gold recovery by 20% at 300 rpm agitation rate.

3. Inter-particle collisions are also governed by amount of agglomerate in the system - available agglomerate surface area and number. While the surface area and number of agglomerates dictate the rate of gold recovery, mixing time dictates maximum gold recovery. Stirring rate on the other hand, affects both rate and final recovery.

4. The agglomerate surface area and number of agglomerates, which control the rate of gold recovery, depend upon the amount of agglomerates relative to the weight of the ore - agglomerate:ore ratio (AOR). With a 1: 1 agglomerate:ore ratio, 99% gold recovery was obtained with only 15 minutes of mixing time while with 0.13:1 ratio, same percentage gold recovery was attained only after 120 minutes.

5. The available surface area and number of agglomerates can also be altered, without changing the agglomerate:ore ratio, by varying the amount of oil relative to the amount of coal - oil:coal ratio (OCR). Lesser oil dosage formed smaller agglomerates, resulting to greater number of agglomerates and available surface area. At 0.24: 1 AOR and 30 minutes mixing time, decreasing OCR from 0.40 to 0.27 increased surface area, number of agglomerates and gold recovery

from 38,000 mm2 to 209,000 mm2, from 1,800 to 182,000 and from 91 to 99%, respectively.

6. Experiments with pH variations indicate that gold recovery is not significantly affected from pH 4 to 12. At pH 2 however, excess H+ 121

ions could have initially 'cleaned' the gold surfaces of organic and/or oxide contamination thus rendering them less hydrophobic resulting to a drop in gold recovery. Longer exposure to the pulp (from 30 to 60 minutes) however, had likely re-contaminated the surfaces that gold uptake ~as increased considerably from 80% to 95%.

7. Gold recovery is to a great extent dependent on the strength and stability of the agglomerates. Using less stable agglomerates may result to gold loss due to the shaving off of fine agglomerate particles. If these fine particles, which may be loaded with gold, are not collected (i.e., during screening) with the main agglomerate products but rather are discarded with the tailings then gold recovery is adversely affected. Comparatively weaker agglomerates are formed with kerosene than with diesel. Also, coarser coal particles gave less stable agglomerates and so with excessive amount of oil ( over-critical oil addition). Less hydrophobic coals like the lignitic coal used in the applicability tests likewise produced weaker agglomerates.

9. At up to 5% sulfides in the feed (synthetic mixture) gold recovery was not significantly affected and was consistently above 99%. There was a decreasing trend however, which may indicate that excessive amount of sulfides may have an adverse effect on gold recovery.

10. Sulfides, being hydrophobic are also recoverable by coal-oil agglomerates. Chalcopyrite, pyrite and galena gave good response to the process with over 98% recovery for chalcopyrite and galena, and 94% for pyrite at natural pH. 122

11. The response of sulfides to different pH suggests that the chemistry of coal-oil agglomeration is similar to that of flotation. At high pH range, pyrite was depressed (rendered hydrophilic) by sodium hydroxide very similar to the action of lime (CaO) in flotation. At pH 12 virtually no pyrite was recovered by the agglomerates. Also, sphalerite was likely activated by sulfuric acid that recovery was increased from 4 to 25% when the pH was changed from pH 6 to pH 2. Chalcopyrite and galena recoveries were lowered by 2% from pH 2 to pH 12.

12. Applicability test works conducted on real gold bearing samples showed that, similar to other gold recovery processes, the mineralogy of gold occurrence (i.e., liberation size and association with other minerals) is of prime importance in coal-oil agglomeration. Results of experiments concerning gold in a copper concentrate manifested that excessive sulfides in the ore could affect gold recovery significantly. While the tests with synthetic mixtures with 5% sulfides did not alter gold uptake ( consistently above 99%), 65% chalcopyrite in the feed lowered gold recovery considerably (only up to 83%). In the tests on an amalgamation tails, grinding the sample (i.e., increase gold liberation) increased gold recovery.

13. Photographs and SEM images of agglomerate surfaces and cross­ sections revealed that gold particles are pushed within several microns inside the agglomerates. Although the actual mechanism of gold particle transfer inside the agglomerates is not totally understood, it could be suggested that the forces influencing such transfer are the capillary force exerted by the bridging liquid (e.g., oil), the external forces inputted by inter-particle collisions ( agglomerate-agglomerate 123

and/or agglomerate-gold) and the impact forces contributed by the baffles and the impellers.

7.2 RECOMMENDATIONS FOR FUTURE WORK

Coal-oil agglomeration for gold recovery was patented over a decade ago but until now it has not yet found its ground in commercial scale in the mining industry. It is indeed a viable process that is at par with the conventional and more popular recovery methods but it is still in a stage where more fundamental studies are to be undertaken to make it accepted and practiced. In relation to the findings of the present study the following recommendations for future studies are presented.

1. It has been observed that agglomerate strength and stability do influence gold recovery in the sense that gold losses are possible due to separation of small agglomerate particles during mixing with the pulp. While there were several studies concerning the tensile strength of the agglomerates in the recovery of fine coal, a more detailed study is recommended to look into their stability and strength as they are agitated with solid particulates in the pulp. The abrading effects of the solids may differ according to their sizes and shapes where coarser particles and higher degree of angularity may cause more severe abrasion. This could also be correlated with the type and amount of oil, and the hydrophobicity of the carbonaceous bulking material. Furthermore, a study of surfactants that will improve agglomerate 124

stability but will not, in any case, affect gold capture could also be included.

2. It has been established that sulfides are also recovered by oil­ agglomerates. With the close association of gold with common sulfides, a study on their selectivity by coal-oil agglomeration would not only to improve recoveries but also would give a better understanding of the process chemistry. An improved knowledge on sulfides recovery and selectivity will increase gold recovery in the sense that sulfides could either be activated and be captured by the agglomerates (i.e., gold is enclosed in the host sulfide) or depressed and just remain with the tailings (i.e., gold is liberated from the sulfide matrix). 125

REFERENCES

Adamson, A. W., Physical Chemistry ofSurfaces, 3rd Edition, Wiley, New York, Chapter 7,' 1976

Adamson, A. W., Physical Chemistry ofSurfaces, 5th Edition, Wiley, New York, Chapter 10, 1990

Allen, R. W. and Wheelock, T. D., "Effects of pH and Ionic Strength on Kinetics of Oil Agglomeration of Fine Coal", Minerals Engineering, Vol. 6, No. 1, 1993, pp. 87-97

Barbery, G., "Recent Progress in the Flotation and Hydrometallurgy of Sulfi.de Minerals", Developments in Mineral Processing 6, 1985, pp. 1-38

Bellamy, S. R., House, CJ., Veal, C. J., "Recovery of Fine Gold from a Placer Ore by Coal Gold Agglomeration", World Gold '89, Chapter 41, 1989

Bewig, K. W. and Zisman, W. A., "The Wetting of Gold and Platinum by Water", The Journal of Physical Chemistry, Vol. 69, No. 12, December 1965,pp.4238-4242

Bhappu, R. B., "Hydrometallurgical Advances in Precious Metal Extraction" Advances in Mineral Processing, Proceedings of A Symposium Honoring Nathaniel Arbiter, New Orleans, Louisiana, Chapter 27, March 1986, pp. 463-479 126

Bhappu, R. B., "Hydrometallurgical Processing of Precious Metal Ores", Mineral Processing and Extractive Metallurgy Review, Vol. 6, 1990, pp. 67-80

Bhattacharyya, R. N., Moza, A. K. and Sarkar, G. G., "Role of Operating Variables in Oil Agglomeration of Coal", Proc. r' International Symposium onAgglomeration, Chapter 55, 1977, pp. 931-938

Bonney, C. F., "The BP-CGA and CARBAD Processes: Technical and Cost Differences", Innovations in Gold and Silver Recovery, Randol, Chapter 16, 1992a, pp. 2320-2322

Bonney, C. F .. , "Coal-Gold Agglomeration - A Novel Approach to Gold Recovery", Innovations in Gold and Silver Recovery, Randol, Chapter 16, 1992b

Cadzow, M. and Lamb, R., "CARBAD Gold Recovery", World Gold '89, Edited by R. B. Bhappu and R.J. Harden, AIME, Littleton, Colorado, USA, Chapter 45, pp. 375-379

Davis, A., "Kinetic Studies on Gold Extraction in Iodide Systems", Ph.D. Thesis, University ofNew South Wales, April 1991

De Cuyper, J. and Oudenne, P. D., "Experience in the Treatment of Sulfide Ores Containing Precious Metals", Developments in Mineral Processing 6, 1985, pp. 239-254 127

Deschenes, D., Ritcey, G. M. and Ghali, E., "Reduction of Gold and Silver by Hydrogen from Sulfuric Thiourea Solutions", Precious Metals '89, edited by Jha, M. C., Hill, S. D. and El Guindy, M., a publication of the Minerals, Metals & Materials Society, pp. 341-357

Dupuis, C. and Ghali, E., "Gold Recovery from Thiourea Solutions by Direct Electrowinning", Precious Metals '89, edited by Jha, M. C., Hill, S. D. and El Guindy, M., a publication of the Minerals, Metals & Materials Society, pp. 359-373

Eisele, J. A., "Comparative Leaching of Gold and Silver Ores with Cyanide and Thiourea Solutions", Precious Metals '89, edited by Jha, M. C., Hill, S. D. and El Guindy, M., a publication of the Minerals, Metals & Materials Society, pp. 253-259

Eissler, M., The Metallurgy of Gold : A Practical Treatise on the Metallurgical Treatment ofGold Bearing Ores, 4th Edition, Lockwood, London, 1896

Erb, R. A., "Wettability of Metals under Continuous Condensing Conditions", Journal of Physical Chemistry, Vol. 69, No. 4, April 1965, pp.1306- 1309

Filho, 0. B. and Monhemius, A. J., "Thermochemistry of Thiocyanate Systems for Leaching Gold and Silver Ores", Precious Metals '89, edited by Jha, M. C., Hill, S. D. and El Guindy, M., a publication of the Minerals, Metals & Materials Society, pp. 307-339 128

Gasparrini, C., "The Mineralogy of Gold and its Significance in Metal Extraction", CIM Bulletin, Vol. 76, No. 851, March 1983, pp. 144-153

Good, R. J. and Van Oss, C. J., "The Modem Theory of Contact Angles and the Hydrogen Bond Components of Surface Energies", Modem Approaches to Wettability - Theory and Applications, edited by Schrader, M. E. and Loeb, G. I., Plenuin Press, New York, 1992

Grieves, R. B., "Surfactant-Anion Interactions in Foam Fractionation", Interfacial Phenomenon in Mineral Processing, Proceedings of the Engineering Foundation Conference, Franklin College, Ridge, New Hamsphire, August 1981, pp. 357-374

Habashi, F., "The Theory of Cyanidation", Transactions of the Society ofMining Engineers, Vol. 235, September 1966, pp. 236-239

Habashi, F., "Historical Metallurgy Notes - One Hundred Years of Cyanidation", CIM Bulletin, Vol. 80, No. 905, September 1987, pp. 108-114

Hayes, P. C., Process Principles in Minerals and Materials Production, 2nd Edition, Hayes Publishing, Sherwood, Qld., 1993

House, C.I., Townsend, LG. and Veal, C. J., "Coal-Gold Agglomeration - Pilot Retreatment of Tailings", Innovations in Gold and Silver Recovery, Randol, Chapter 16, 1992, pp. 2301-2308 129

Hogg, R., "Role of Colloid and Interface Science in Agglomeration, 5th International Symposium on Agglomeration, 1989, pp. 484-494

' Huyhua, J. C., Zegarra, C. R. and Gundiler, I. H., "A Comparative Study of Oxidants of Gold and Silver Dissolution in Acidic Thiourea Solutions, Precious Metals '89, edited by Jha, M. C., Hill, S. D. and El Guindy, M., a publication of the Minerals, Metals & MateriaJ.s Society, pp. 287-303

Jha, M. C., "Recovery of Gold and Silver from Cyanide Solutions: A Comparative Study of Various Processes", Precious Metals Recovery, Proc. 1st International Symposium, Nevada, Chapter 21, June, 1984

Labuschagne, B. C. J., Rex, V. and Prinsloo, J. J., "Influence of Oil Composition, pH and Temperature on the Selective Agglomeration of Coal", Proc. 5th International Symposium on Agglomeration, 1989, pp. 505-514

Leja, J., Surface Chemistry ofFroth Flotation, Plenum Press, New York, 1982

Lins, F. F. and Adamian, R., "Some Chemical Aspects of Gold Particles Flotation", Proc. 1B'h International Mineral Processing Congress, Sydney, May 1993, pp. 1119-1122 130

Marciano, A., Costa, L.S.N. and Lins, F.F., "Utilization of Coal-Oil Agglomerates to Recover Gold Particles", Minerals Engineering, Vol. 7, No. 11, 1994, pp. 1401-1409

Marsden, J. 0. and Fuerstanau, M. C., "Comparison Between Merrill-Crowe Precipitation and Carbon Adsoprtion for Precious Metals Recovery", Ufh International Mineral Processing Congress, May 1993, pp. 1189- 1194

Matis, K. A. and Mavros, P., "Recovery of Metals by Ion Flotation from Dilute Aqueous Solutions", Separation and Purification Methods, Vol. 20, No. 1, 1991, pp. 1-48

Mehrotra, V. P., Sastry, K.V.S. and Morey, B.W., "Review of Agglomeration Techniques for Processing of Fine Coals", International Journal of Mineral Processing, November 1983, pp. 175-201

Nicol, S. K., Galvin, K. P. and Engel, M. D., "Ion Flotation - Potential Applications to Mineral Processing", Mineral Processing, Vol. 5, Nos. 10-12, 1992, pp.1259-1275

Paterson, C. J., "Ore Deposits of Gold and Silver", Mineral Processing and Extractive Metallurgy Review, Vol. 6, 1990, pp. 43-46

Randol, "Novel Gold Pre-Concentration Systems", Innovations in Gold and Silver Recovery, Chapter 16, 1992 131

Richards, R G. and Bangester, P. J., "Gravity Concentrating Systems in Gold Recovery", Proceedings of the Aus.I.MM Perth and Kalgoorlie Branches Regional Conference on Gold Mining, Metallurgy and Geology, October 1984, pp. 1-12 '

Schrader, M. E., "High- and Medium-Energy Surfaces : Ultrahigh Vacuum Approach", Modem Approaches to Wettability - Theory and Applications, edited by Schrader, M. E. and Loeb, G. I., Plenum Press, New York, Chapter 3, 1992, pp.53-71

Schrader, M. E., "Ultrahigh-Vacuum Techniques in the Measurement of Contact Angles - Water on Gold", The Journal of Physical Chemistry, Vol. 74, 1970,p. 2313-2317

Schubert, H., "Tensile Strength and Capillary Pressure of Moist Agglomerates", Proc. 2"" International Symposium onAgglomeration, pp. 144-155

Siddall, G. B., "Gold Recovery Process", (United States Patent No. 4,597,791, July 1, 1986), Innovations in Gold and Silver Recovery, Randol, Chapter 16, 1992

Sparks, B. D., Meadus, F. W., and Kumar, A., "Liquid Phase Agglomeration as an Aid to Particle Separation from Hydrocarbon Suspensions", Proc. 5th International Symposium onAgglomeration, 1989, pp.495-504

Stanley, G. G., "Gold Extraction Plant Practice in South Africa", Mineral Processing and Extractive Metallurgy Review, Vol. 6, 1990, pp. 191-216 132

Swanson, A. R., Bensley, C. N. and. Nicol, S. K., "Some Fundamental Aspects of the Selective Agglomeration of Fine Coal", Proc. 2"" International Symposium on Agglomeration, 1977, pp. 939-951

Szymocha, K., Pawlak, W. and Ignaisak, B., "Kinetics of Coal At Extended Agitation Times", Proc. 5th International Symposium on Agglomeration, 1989, pp.525-536

Wan, R. Y. and Miller, "Research and Development Activities for the Recovery of Gold from Alkaline Cyanide Solutions", Mineral Processing and Extractive Metallurgy Review, Vol. 6, 1990, pp. 143-190

White, M. L. and Drobe~ J., "The Effect of Residual Abrasives on the Wettability of Polished Gold Surfaces", The Journal of Physical Chemistry, Vol. 70, No. 11, November 1966, p. 3432-3436

White, M. L., "The Wetting of Gold Surfaces by Water", The Journal ofPhysical Chemistry, Vol. 68, No. 10, October, 1964, p. 3083-3085

Willis, B. A, Mineral Processing Technology, An Introduction to the Practical Aspects of Ore Treatment and Mineral Recovery, Pergamon Press, Oxford; New Yor~ 1979

Yang, F. and Gao, H., "Discussion of Some Technological Problems in Increasing the Recovery of Gold and Silver Associated Sulfides", J8"' International Mineral Processing Congress, Sydney, May 1993 133

Zisman, W. A., and Bernett, M. K., "Confirmation of Spontaneous Spreading by Water on Pure Gold", The Journal ofPhysical Chemistry, Vol. 74, 1970, p. 2309-2312 '

Zisman, W. A., and Bewig, K. W., " Low Energy Reference Electrodes for Investigating Adsoprtion by Contact Potential Measurements", Advances in Chemistry, Series No. 33, American Chemical Society, Washington D. C., 1961, pp. 100-113 134

APPENDICES 135

'

Appendix A Mineralogy of Cyanide Leaching Residues (Blend A Matrix) (Flotation Products of the Same Sample Used in Leaching Tests) 136

LAKE COWAL FLOAT TEST DATA SHEET

GRIND mill 2kg, s/s ROUGHER FLOAT CLEANER FLOAT medium site water cell type Denver cell type grams solids 1984 litres volume 5 litres volume % solids 65 rpm 1500 rpm minutes 33

7.8 rougher 1 0.37 8.2 10 6 5 7 .5 rougher 2 5 3 0 .5 2 7 .5 finished b 2 minutes rougher 3 -500 5 7.9 5 2 5 rougher 4 -500 2 7.9 5 2 3 git

rougher 1 & 2 93 4.6 32.9 152.5 20.8 96.4 87.7 94.7 rougher 3 & 4 15 0.8 4.56 3.4 3.4 2.6 2 .0 2.5 tailings 1890 94.6 0.19 18 .0 0 .03 2.8 10.3 2.8 total 1998 100.0 1 .74 173.9 100.0 100.0 head 2.35 1.6 repl icate assay tail 019, 0.18g/tAu

425 0.0 100.0 100 0.0 100.0 212 0.0 100.0 150 9.8 3.6 86.4 106 7. 0 9 .7 76.6 75 9.4 13.1 63.5 63 4.6 6.4 57.1 53 3.8 5 .3 5 1.8 45 2.8 3 .9 47.9 38 2.4 3.3 44.6 -38 44.6 total 71.8 100.0 137

SUMMARY COMMENTS (including reference to your notes)

The individual detailed description indicate the following generalised compositions: rougher cone. pyrite particles dominant , rock fragments without pyrite mmor rock fragments carrying fine pyrite mmor chalcopyrite, sphalerite, magnetite, hematite accessory visible gold ( composite with pyrite) rare trace

CPS cone. rock fragments without sulfides dominant rock fragments carrying pyrite subordinate liberated pyrite mmor chalcopyrite, sphalerite accessory visible gold ( enclosed in carbonate) rare trace tail rock fragments without pyrite highly dominant rock fragments carrying pyrite accessory free grains of pyrite rare trace

The sparseness of visible gold in the rougher cone. (one grain in pyrite) suggests that 30 git assay value given for this sample must surely include 'invisible' gold. As solid solution in the abundant pyrite.

There is no evidence that the pyrite in any of those samples is 'partly oxidised' ( as suggested in your notes) even though it is agreed that some of the pyrite examined under binocular microscope appears to be tarnished. Note that minor grains of sphalerite Fe + Ti oxide are dark coloured under binocular microscope.

However the amount of fine pyrite, unliberated and enclosed within gangue particles is probably 'significant' (particularly if this pyrite carries gold) and these composite particles are proportionately much higher in the CPS cone. than in the rougher cone. 138

DPAF18 Rougher Cone.

Grain size range (deslimed), as seen in the polished thin section grain mount is 0.01mm to (rarely) 0.45mm maximum dimension.

The greater proportion of this sample consists of basically independent, liberated pyrite grains albeit commonly with minute silicate-mineral inclusions. Rock fragments occur in relatively minor numbers of individual grains, but they are generally larger than the pyrite grains. These appear to consist of basaltic or andesitic colcanics, but pervasively altered to various proportions of extremely fine quartz, secondary feldspar, carbonate, clay­ sericite, and minor chlorite. Importantly, some of these rock fragments carry crystals of pyrite. Minor sphalerite, chalcopyrite>magnetite grains are also present, mostly as individuals, rarely in rock fragments.

Trace gold was seen (see below). Visually estimated volume percent abundances of these components are as follows: Estimated Abundances (semi quantitative to relative • Monomineralic, single crystal fragments of pyrite, with 10% nil or negligible inclusions • Grains of pyrite, with minute inclusions, forming variably 5% to 15% of a given pyrite grain, mostly silicate mineral, rarely of iron-oxide, chalcopyrite, sphalerite. None of the pyrite grains are 'oxidised'. -65% • Fragments of rock, carrying sparse small grains of iron and/or titaniferous oxide, but no sulfide 7-10% • Fragments of rock carrying accessory small (to 0.05mm) grains of pyrite and of trace chalcopyrite, ± F e/fi oxide [% of pyrite of whole sample with this inclusion mode of occurrence could be as high as 7%] 12-15% • Grains of chalcopyrite commonly with sparse minute inclusions of pyrite, some composite with fine hematite 3% • Grains of sphalerite, with minor minute inclusions of chalcopyrite, very rarely composite with pyrite 1-2% • Grains of magnetite, and hematite <1% 139

DPAF18 Tail

This ( deslimed) tails sample is seen in polished thin section to consist almost entirely of rock fragments, O:'lmm to 0.5mm in size. These probably had an original basaltic to andesitic composition, but they have been very extensively, prevasively altered, to extremely fine secondary quartz, feldspar, carbonate, clay-sericite and chlorite. About 20% of these incorporate accessory, very fine secondary (leucoxenitic) Fe-Ti oxide grams.

Ther are minor (2-3%) small separate fragments oflimonite.

Sulphide forms far less than 1% of the whole sample, with pyrite marginally more abundant than chalcopyrite. Most of this pyrite and chalcopyrite occurs as sparse, very small inclusions, mostly <25 µm size within a small number of rock chips. Rare grains of these sulphides inclusions measure up o 100 µm size, rare small splinters of steel occur independently. 140

AppendixB SCREEN ANALYSIS OF GOLD-BEARING AND COAL SAMPLES

Appendix B.1 : Screen Analysis of Synthetic Mixtures - Blends A and B

Blend A

Screen Weight %Wt. Cum.% Size Retained Retained Passing Screen Analysis of Blend A L1m g. 425 0.18 0.08 99.92 ~ 100 ~- ....,,= 212 10.78 5.00 94.91 ; 80 150 26.02 12.08 82.83 j:l,,c 'I- 80 106 32.81 15.23 67.60 ...~ 40 75 63.58 29.52 38.09 -= 63 68.92 31.99 6.09 -I 20 ..0 53 11.57 5.37 0.72 u= 0 38 1.05 0.49 0.23 0 100 200 300 400 500 Pan 0.50 0.23 0.00 Particle Size µm Totals 215.41 100.00

BlendB

Screen Weight* 0/o Wt. Cum.% Size Retained Retained Passing Screen Analysis of Blend B µm g. 250 5 1.69 98.31 100 .. - 150 11 3.73 94.58 ~ .e.,, 80 75 40 13.56 81.02 ; j:l,,c 53 34 11.53 69.50 'I- 60 45 40 13.56 55.94 -~ 40 30 40 13.56 42.38 --= 20 33 11.19 31.19 e= 20 10 53 17.97 13.23 u= 5 13 4.41 8.82 0 3 13 4.41 4.41 0 100 200 300 Pan 13 4.41 0.00 Particle Size, µm Totals 295 100.00 * Computed from product specifications supplied by manufacturer, Commercial Minerals Limited 141

Appendix B.2 : Screen Analysis of Coal Composite Samples

LAC - Low Ash Coal

Screen Weight %Wt. Cum.% Screen Analysis of LAC Size Retained Retained Passing 100 2 .. - u.m 90 300 0 0.00 100.00 Cl) 1::1 80 98.68 1111 212 0.6 1.32 ·- 70 ~ 150 2.2 4.82 93.86 = 60 2.7 5.92 87.94 ?te. 106 ~ 50 75 13.5 29.61 58.33 ~• 40 63 16.4 35.96 22.37 -i 30 53 7.9 17.32 5.04 d 20 45 0.9 1.97 3.07 10 j 38 0.9 1.97 1.10 0 Pan 0.50 1.10 0.00 0 100 200 300 Totals 45.6 100.00 Particle Size µm

MAC - Medium Ash Coal

Screen Weight %Wt. Cum.% Screen Analysis of MAC Size Retained Retained Passing 100 - u.m e:. 90 300 0 0.00 100 r so 212 0.1 0.43 99.57 ·1 70 ~ 150 0.8 3.42 96.15 ?te, 60 106 1.4 5.98 90.17 ~ 50 75 2.8 11.97 78.21 i 40 63 1.1 4.70 73.50 'i 30 53 7.3 31.20 42.31 !20 45 6.9 29.49 12.82 10 ~ 38 1.4 5.98 6.84 0 Pan 1.6 6.84 0 100 200 300 Totals 23.4 100.00 Particle Size, mm 142

HAC - High Ash Coal

Screen Weight %Wt. Cum.% Screen Analysis of HAC Size Retained Retained Passing

100 ~ g. u.m 90 0 0.00 100 Cl) 300 80 97.48 ....,,= 212 0.6 2.52 70 150 I.I 4.62 92.86 ~= 60 ~ 106 2.8 11.76 81.09 0 50 7.1 29.83 51.26 ...~ 40 75 ~ 63 5.8 24.37 26.89 -I 30 53 2.5 10.50 16.39 20 a 10 45 1.3 5.46 10.92 0 38 1.4 5.88 5.04 0 100 200 300 1.2 5.04 0.00 Pan Particle Size, mm Totals 23.8 100.00 143

Appendix B.3 : Screen Analysis of Aus Coal

Screen Weight %Wt. Cum.% Screen Analysis of Aus Coal Size Retained Retained Passing 100 u.m 2:." 90 300 0 0.00 100.00 DI) t::I 80 2.12 97.88 ....,, 212 1.06 ; 70 ~ 150 2.53 5.05 92.83 60 '::?.0 106 3.46 6.91 85.92 50 ~ 75 15.26 30.48 55.43 --=CIIS 40 63 17.56 35.08 20.36 -El= 30 53 7.5 14.98 5.37 u= 20 45 0.6 1.20 4.17 10 ) 38 1.2 2.40 1.78 0 Pan 0.89 1.78 0.00 0 100 200 300 Totals 50.06 100.00 Particle Size µm

Appendix B.4 : Screen Analysis of Graphite

Screen Weight %Wt. Cum.% Screen Analysis of Graphite Size Retained Retained Passing 100 1.1,m g. 90 300 0 0.00 100.00 DI) t::I 80 34 11.21 88.79 ....,, 212 ; 70 150 36.6 12.07 76.71 ~ '::?. 60 106 27.6 9.10 67.61 0 ~ 50 75 34.4 11.35 56.27 ~• 40 63 23.7 7.82 48.45 - 30 § •t 53 12.2 4.02 44.42 u= 20 45 56.3 18.57 25.86 10 38 77.5 25.56 0.29 0 J Pan 0.89 0.29 0.00 0 100 200 300 Totals 303.19 100.00 Particle Size µm 144

AppendixC COAL ANALYSES

LAC MAC HAC Aus Coal Lig. Coal Proximate Analysis Flot. Cone. Moisture (M), % 1.38 1.62 1.44 3.51 4.18 Volatile Combustible Matter (VCM). % 20.40 15.92 20.45 27.8 38.78 Ash (A),% 10.00 12.93 18.64 14.54 14.49 Fixed Carbon (FC), % 68.22 69.53 59.47 54.15 42.55

Ultimate Analysis Carbon (C), % 78.11 75.95 69.4 66.64 63.66 Hydrogen (H), % 4.08 3.52 3.6 4.07 4.73 Nitrogen (N), % 0.96 0.83 0.83 0.7 0.88 Sulfur (S), % 0.46 0.59 0.38 0.35 1.7

Note: The analyses were conducted by the Coal Section, Energy Research Laboratory Department of Energy, Philippines AppendixD SCREEN ANALYSES AND SURFACE AREA CALCULATIONS OF AGGLOMERATES AT DIFFERENT OIL/KEROSENE:COAL RATIOS Surface Area Calculations: I. Weigh a representative sample from each screen fraction. (w 1) " 2. Count the number of agglomerates of each sample. (n 1)

3. Calculate the average weight of the agglomerates. (w 1 In 1 = w 0 ) 4. Compute for the total number of agglomerates for each screen fraction. (total weight of agglomerates/w a = n t) 5. Assuming the agglomerates are perfect spheres, calculate the average surface area of agglomerates within each screen fraction. The diameter of agglomerates between two screens is assumed to be the average of the two screen sizes (e.g. -0.85 mm +0.60 mm.: diameter= (0.86 + 0.60)/2 while the diameter

of the top most agglomerates is assumed to be the size of the largest screen. (sa 0 ) 6. Calculate the total surface area for each screen fraction by multiplying the average surface area by the total number of agglomerates. (sa a *n ,)

Appendix D.1 : OIL - COAL AGGLOMERATES A. Oil:Coal Ratio - 8ml:25g [0.27(w/w)] Screen A.naly.ris Surface Area Calculations Screen Size Weight %Wt. Cum.% Weight of Number Ave. Wt Total No. Ave.Agg S.Area/ Total mm Retained, g Retained Passing Sample, R of Agg of Agg, 2 of A22 Dia., mm agg,mm2 S. Area, mm2 0.850 0.61 2.02 97.98 0.0410 80 0.00051 1,190 0.85 2.27 2,701.62 0.600 15.99 52.86 45.12 0.0342 170 0.00020 79,482 0.73 1.65 131,249.35 0.425 12.50 41.32 3.80 0.0550 355 0.00015 80,682 0.51 0.83 66,575.32 0.300 1.15 3.80 0.00 0.0157 280 0.00006 20,510 0.36 0.41 8,466.85 Pan 0.00 0.00 ~ Total 30.25 100.00 181,864 208,993.15 -VI .,I:>. - °'

2

2

34.74

64.90

53.26

147.25

254.87

1,627.55

1,395.31 1,292.55

34,015.09 27,115.89 26,282.22

66,338.46 SS,380.16

Total

Total

103,241.93

Area,mm

Area,mm

S.

S.

2

2

1.65 1.65

3.24

0.83 0.41 6.51 3.24

9.08 6.51

12.95

35.26

25.61

Area/

Area/

S.

S.

agg,mm

agg,mm

1.44 1.02

1.02

1.70 1.44

3.35 0.73 2.03 0.51 0.36

2.86 0.73

mm

mm

Agg

Agg

Ave.

Ave.

Dia.,

Dia.,

7

11

39

65

142

845

357

250

No.

No.

1,059

2,030

3,396

A22

Agg

10,183 10,510

22,102

of

of

Total

Total

2

g

Wt

Wt

0.0009 0.0005

Agg,

A22,

0.00015 0.00006 0.00093 0.02767 0.00051 0.01396 0.00702 0.00702 0.00200 0.00300

Ave.

Ave.

of

of

3

92

36 45

132 103

137

122

Agg

Agg

of

of

Number Number

Calculations

Calculations

2

g

of

of

Area Area

1.9122

0.0857 0.0830 0.0229 0.1291 0.9272 0.2438 0.3092

Sample,

Weight

Sample, Weight

Surface Surface

1.46

0.07 0.09 2.60 0.06 0.10

96.84 50.10

32.44 99.33

Cum.%

Cum.%

Passing

Passing

[0.34(w/w)]

[0.30(w/w)]

1.36

0.03 0.00 3.16 0.06 0.00 0.00 0.67 0.03

2.50

30.98

64.40

47.50

49.23

100.00

100.00

%Wt. %Wt.

Retained

Retained

g

10ml:25g

9ml:25g

1.00

0.00 0.75 0.20 0.01 9.79 0.02 0.07 0.00 0.43 0.01 0.02 0.00

-

14.26

14.78

-

30.02

20.35

31.60

Weight

Weight

Retained,g

Retained,

Ratio

1.180 1.700 1.700

1.180

0.425 0.850 0.850 3.350 2.360 0.600 0.600

0.300

Size Size

Analysis Analysis

mm

mm

Oil:Coal

Oil:Coal Ratio

C.

B.

Screen Pan Screen Pan Screen Total Screen Total .i:,. -..J

-

2

2

61.74

mm

104.23

198.32

227.14 315.36

1,206.54

8,965.95 5,504.12

5,449.44

10,577.74

30,428.56

45,545.37 20,272.09

37,765.88

Total

Total

Area,mm

Area,

S.

S.

...

2

2

1.65

1.65

3.24

6.51

6.51 3.24

12.95

12.95

35.26

35.26

25.61

25.61

Area/

Area/

s. S.

agg,mm

agg,mm

1.44 1.02

1.44 1.02

3.35

2.86 3.35 2.03

0.73 2.86

2.03

0.73

mm

mm

Agg

Agg

Ave.

Ave.

Dia.,

Dia.,

32

48

61 37

138

185

254 300

425

792

421

No.

No.

1,188

2,086

1,796

A22

A22

of

of

Total

Total

g

g

Wt

Wt

0.0077 0.0021 0.0009 0.0005

0.0324 0.0167

0.0441 0.0169 0.0070 0.0023 0.0009 0.0005

Agg,

A22,

Ave.

Ave.

of

of

65

88 53

62

150 105

152 143

Agg

A22

of

of

Number

Number

Calculations

Calculations

g

g

of

of

Area

Area

1.7173 1.0880 1.1537

1.0051

0.2169

2.7326 2.5736

0.2034

Sample,

Weight

Sample,

Weight

Surface

Surface

weight)

1.68

0.63

0.00

0.00 0.00

0.25

10.98

11.51

73.92

56.05

40(by

Cum.%

Cum.%

Passing

Passing

..

[0.37(w/w)]

- 0

1.42

0.09 0.22 0.00 0.22

0.00

9.83

10.35

26.08

62.94

43.95

44.55

100.00

100.00

%Wt.-

%Wt.

Retained

Retained

g

g

llml:25g

12ml:25g

3.27 8.24

0.10 0.32 0.32 0.03 0.07

0.00

0.02 0.06 0.00

0.43

0.06 0.19 0.06

2.96

-

19.89

-

13.22 13.40

30.09

31.60

Weight

Weight

Retained,

Retained,

Ratio

Ratio

1.700

1.180

1.700 1.180

3.350

0.850

3.350 2.360

0.600 2.360

0.850 0.600

Size

Size

Analysis

Analysis

mm

mm

Oil:Coal

Oil:Coal

D.

Screen

E. Screen

Pan Total Screen

Pan

Screen

Total ~

00

......

2

2

363.45

7,518.61

2,087.59

8,857.97

77,436.21

44,283.68

23,546.33

46,174.23

26,451.00

29,589.92 47,008.43

Total

Total

158,445.00

Area,mm

S.

S.Area,mm

2

2

Area/

Area/

s.

s.

a22,mm

agg,mm

17.4974144 12.9461892

1.34782179

1.91134497

7.54767635

5.22792433

3.73252623

9.07920277 2.68802521 6.51440653

1.7

1.44 1.09

0.78

mm

mm

0.655

0.925

Agg

Agg

1.2900

1.5500

2.3600 2.0300

Ave.

Ave.

Dia.,

Dia.,

40

119

No.

No.

1,438

1,360

3,421 3,120

7,928 8,098

Agg

Agg

19,625

17,488

70,598

24,158

of

of

Total

Total

g

AGGLOMERATES

Wt

Wt

A22,g

Ae:e:,

0.00497

0.00016

0.000303

0.004213

Ave.

Ave.

of

of

0.0103818

0.0029656 0.0013924

0.0013425 0.0007333 0.0004584

COAL

-

11

20

32

23

73

57

105

366

215

245

A22

Agg

of

of

Number

Number

Calculations

Calculations

g

g

of

of

Area

Area

0.098

0.1142

0.0949 0.1462

0.1109 0.0344

0.0994 0.0969

0.1123

KEROSENE

0.0418

Sample,

Sample,

Weight

Weight

:

Su,face

Su,face

D.2

0.00

6.19

0.00

o/o

o/o

11.95

38.16

39.80

95.80

99.36 92.41 70.29

[0.288(w/w)]

[0.32(w/w)]

Cum.

Cum.

Passine:

Passing

6.79 0.00

4.20

0.64

6.95

Appendix

11.95

57.64 31.37

22.12

27.85 30.50

100.00

100.00

%Wt.

%Wt.

- 9 ml:25g

-10ml:25g

Retained

Retained

I

I

1.83

1.24

2.00

0.00

8.02

9.25

5.81

7.32 3.14 0.00 0.00

0.17

Ratio

Ratio

17.00

29.49

26.28

Weight

Weight

Retained,

Retained,

1.180

1.700

1.700 1.400

1.000

1.180

2.360

0.850 0.710 0.600

Size

Size

Analysis

Analysis

mm

mm

Kerosene:Coal

Kerosene:Coal

Pan B.

Screen

Total A.

Scrun Pan

Screen

Total Screen ~

-

"°

2 2

112.35

1,127.06

8,391.31 2,032.21

14,087.01 14,015.86

20,099.97 25,107.32 49,617.85

35,355.24

Total Total

Area,mm

Area,mm

S. S.

..

2

2

Area/

Area/

S. S.

agg,mm

agg,mm

12.9461892

12.9461892

7.54767635 35.2565236

35.2565236 25.6072003 25.6072003

7.54767635

1.55

1.55

3.35

2.86 2.03

3.35 2.86 2.03

mm mm

Agg

Agg

Ave.

Ave.

Dia.,

Dia.,

15

87

980 269 238 785

398

No.

No.

1,088

2,576

1,284

Ae

Ae

of

of

Total

Total

g

g

Wt

Wt

Agg,

Agg,

0.01386 0.00509

0.00222

Ave.

Ave.

of

of

0.0224067

0.0023146

0.0281481 0.0168804 0.0059783

5

15 15

20 41

51

27

23

Agg

Agg

of

of

Number

Number

Calculations

Calculations

g

of

of

Area

Area

0.76

0.3361 0.2079 0.1018 0.0949

0.8609 0.1375 0.0111

weight)

Sample,g

Weight

Sample, Weight

Surface

Surface

0.00

2.48

2.21

0.13 0.00

0.384(by

78.74 24.56

55.23

[0.352(w/w)]

Cum.%

Passing Cum.%

Passine:

0.00 0.00

2.48

0.13

2.08

21.26

54.18 22.08

53.01 44.77

100.00

100.00

llml:25g

12ml:25g-

%Wt.

%Wt.

Retained

- Retained

-

2

g

5.33 5.54 0.62 0.00

0.00 0.00

0.03

0.52

13.59

11.19 13.25

25.08

24.99

Ratio

Ratio

Weight

Weight

Retained,

Retained,

1.4

1.7

3.35 2.36

1.700 1.400

3.350

2.360

Size

Size

Analysis

Analysis

mm

mm

Kerosene:Coal

Kerosene:Coal

C.

Screen D. Screen

Pan

Screen Total Screen Pan

Total 150

Appendix E ANALYTICAL PROCEDURES

Appendix E.1: Determination of Au, Cu, Fe, Zn and Pb in Heads and Metallurgical Products

Reagents:

Concentrated Hydrochloric Acid Concentrated Nitric Acid 3M Hydrochloric Acid

Apparatus:

Furnace Hotplate Atomic Adsorption Spectrophotometer Vacuum filter, ceramic funnel, Erlenmeyer flask Porcelain crucibles 400 ml beakers, watch glasses 100 ml volumetric flasks

Procedures:

For Coal-Oil Agglomerates

1. Weigh 3-5 grams of chy agglomerates in a porcelain crucible.

2. Roast the sample at 650°C for one hour, then raise the temperature to 815°C for another hour. This is to convert coal into ash and/or sulfides to avoid gold losses due to adsorption on carbon or elemental sulfur during decomposition.

3. Transfer the roasted sample in 400 ml beaker.

4. Add 24 ml of aqua regia (3 HCl: 1 HN03) and leave the samples until all reactions cease.

5. Boil the samples on a hot plate for 15 minutes, then remove from hot plate and dilute slightly with water.

6. Vacuum filter the digested mixture washing thoroughly the residue with 3M HCI. 151

7. Transfer the solution to 100 ml volumetric flask and bring to volume with 3M HCl.

8. Analyze the sample through AAS using standard solutions with 3M HCl matrix.

For Heads and Tailings l. Weigh 15-20 grams of dry tailingss in a porcelain crucible.

2. Roast the sample at 650°C for one hour, then raise the temperature to 815°C for another hour..

3. Transfer the roasted sample in 400 ml beaker.

4. Add 40 ml of aqua regia (3 HCl: 1 HN03) and leave the samples till all reaction cease.

5. Boil the sample covered with a watch glass on a hot plate for 30 minutes, then remove the watch glass and evaporate to a honey-like consistency. If sample evaporate to dryness, add aqua regia and boil again.

6. Add 40 ml 3M HCl, boil briefly and take off from the hot plate.

7. Vacuum filter the digested mixture washing thoroughly the residue with 3M HCl.

8. Transfer the solution to 100 ml volumetric flask and bring to volume with 3M HCl.

9. Analyze the sample through AAS using standard solutions with· 3M HCl matrix. 152

Appendix E.2 : Determination of Ash (Australian Standard - Methods for the Analysis and Testing ofCoal and Coke, AS 1038. 3 - 1989)

Introduction. During the incineration of coal, the associated mineral matter undergoes chemical changes and the ash remaining is generally of less mass than the mineral matter originally present in the coal. These changes include the loss of water of hydration from mineral matter, the loss of carbon dioxide from the carbonates, the oxidation of pyrites to iron oxide and a partial fixation of oxides to sulfur. The determination of ash is empirical because the conditions of incineration control the extent to which these reactions occur, and it is essential, therefore, to adhere strictly to the procedure specified. A moisture determination is carried out concurrently.

Scope of Section. This section sets out a method for the determination of ash in higher rank coal. A single furnace or two furnaces can be used.

Principle. A known mass of sample is heated in air to 500°C in 30 minutes, maintained at this temperature for 30 minutes and heated at 815°C until constant mass of the residue remaining after incineration.

Note : Due to fixation of oxides of sulfur, the presence in the furnace of coals of high sulfur may cause variably high results in coals producing a highly basic ash.

Apparatus.

Muffle Furnace - either of the following muffle furnace combinations is required: a. A muffle furnace capable of -

1. achieving an adequate zone at a uniform temperature of 500±10°C in 30 minutes from room temperature 153

11. being maintained at this temperature, andof then being raised to 815±10°C; and

iii. maintaining an adequate zone at the latter temperature.

The ventilation shall be such as to give at least 10 atmosphere changes per minute at 5009C. b. The following two furnaces:

1. A muffle furnace capable of achieving an adequate zone at a uniform temperature of 500±10°C in 30 minutes from room temperature, and of being maintained at this temperature. The ventilation shall be such as to give at least 10 to 12 atmosphere changes per minute at 500°C.

11. A muffle furnace capable of maintaining a temperature of 815±10°C. The ventilation shall be such as to give at least 4 atmosphere changes per minute at 815°C.

Dish - a silica dish, 10 mm to 15 mm dee~, with lid, such a size that the loading of the coal layer does not exceed 0.15g/cm .

Note : Silica dishes deteriorate with continued use, and should be discarded when they show significant roughening or any tendency to flake.

Procedure. The procedure shall be as follows: a. Weigh a clean, dry, empty dish with its lid to the nearest .0lmg (m 1). Spread into the dish an even layer of the coal sample, of approximately 1 g, and replace the lid. Determine the mass of the sample taken, reweigh to the nearest 0.1mg (m2). b. Place the lid in a desiccator and insert the uncovered dish into the muffle furnace. Raise the furnace temperature to 500°C over a period in any 5-min interval, and maintain at this temperature for 30 min. c. Continue heating to 815°C in the same furnace. Alternatively, transfer the dish to a second muffle furnace, previously heated to 815°C. (See Note 1) d. When incineration is complete (See Note 2), remove the dish from the muffle furnace, replace the lid (See Note 1), allow to cool on a thick metal plate for 10 min, and transfer to a desiccator until it attains room temperature. 154

e. Weigh the covered dish to the nearest 0.01mg (m3). f. Brush out the ash completely and reweigh the empty dish and lid to the nearest 0.01mg (m4). If the change in mass is greater than 1mg, repeat the determination and check if dish should be discarded.

Notes: 1. Ifthe ash is light and fluffy, place the lid on the dish before transferring. Remove the lid from the dish after transferring. 2. Heating at 815°C to 60 min is sufficient in many cases. If incomplete combustion is suspected, e.g. with high ash coals, do not brush out the ash after the first weighing, but re-ignite at 815°C for further periods of 30 min until constant in mass (±0.03 mg); then brush out and re-weigh the empty dish. The brushing out and re-weighing procedure eliminates errors due to the hygroscopicity of the silica dish.

Calculation of Results. The percentage of ash remaining after incineration of the coal sample, determined as described in the procedure, shall be calculated from the following equation.

where

Asd = ash in the air-dry sample, in percent

m3 = mass of dish plus lid plus ash, in grams

m4 = mass of dish plus lid after ash has been brushed out, in grams

m2 = mass of dish plus lid plus sample, in grams

m 1 = mass of dish plus lid, in grams