Geometallurgical characterisation of Merensky Reef and UG2 ...

Geometallurgical characterisation of Merensky Reef and UG2 at the Lonmin Marikana mine, Bushveld Complex, South Africa

By

Thomas Dzvinamurungu

Dissertation

Submitted in fulfillment of the requirements for the degree

of

Magister Scientiae

in

Geology

in the

Faculty of Science

at the

University of Johannesburg, South Africa

Supervisor: Prof KS Viljoen

Co-supervisor: M Knoper

December 2012

i

AFFIDAVIT: MASTER’S AND DOCTORAL STUDENTS

TO WHOM IT MAY CONCERN

This serves to confirm that I______________________________________________________ Full Name(s) and Surname

ID Number____________________________________________________________________

Student number__________________________________________________ enrolled for the

Qualification___________________________________________________________________

Faculty of Science______________________________________________________________

Herewith declare that my academic work is in line with the Plagiarism Policy of the University of Johannesburg which I am familiar.

I further declare that the work presented in the ___________________________________(minor dissertation/dissertation/thesis) is authentic and original unless clearly indicated otherwise and in such instances full reference to the source is acknowledged and I do not pretend to receive any credit for such acknowledged quotations, and that there is no copyright infringement in my work. I declare that no unethical research practices were used or material gained through dishonesty. I understand that plagiarism is a serious offence and that should I contravene the Plagiarism Policy notwithstanding signing this affidavit, I may be found guilty of a serious criminal offence (perjury) that would amongst other consequences compel the UJ to inform all other tertiary institutions of the offence and to issue a corresponding certificate of reprehensible academic conduct to whomever request such a certificate from the institution.

Signed at Johannesburg_____________on this _______ of ___________2013

Signature__________________________________ Print name_________________________

STAMP COMMISSIONER OF OATHS

Affidavit certified by a Commissioner of Oaths

THIS AFFIDAVIT CONFORMS WITH THE REQUIREMENTS OF THE JUSTICES OF THE PEACE AND COMMISSIONERS OF OATHS

ACT 16 OF 1963 AND THE APPLICABLE REGULATIONS PUBLISHED IN THE GG GNR 1258 OF 21 JULY 1972; GN 903 OF 10 JULY

1998; GN 109 OF 2 FEBRUARY 2001 AS AMENDED.


47-088130Z-47/ Passport Number: BN649386

201132994

Magister Scientiae

at the University of Johannesburg

MSc Geology

April 11th Day (UJ APK )


ii

TABLE OF CONTENTS TABLE OF CONTENTS .............................................................................................................. i

LIST OF FIGURES ...................................................................................................................... v

LIST OF TABLES ..................................................................................................................... viii

TERMINOLOGY ......................................................................................................................... x

ACKNOWLEDGEMENTS ....................................................................................................... xii

ABSTRACT ................................................................................................................................ xiii

Chapter 1: INTRODUCTION ..................................................................................................... 1

1.0 Introduction .................................................................................................................... 1

1.1 geographical setting, local geology and history ........................................................... 1

1.1.1 Regional Geological setting: introduction ....................................................................................... 4

1.1.2 Regional geological setting .............................................................................................................................. 4

1.1.3 The Merensky Reef ............................................................................................................................................ 7

1.1.4 The UG2 Reef ....................................................................................................................................................... 8

1.1.5 The Platreef ........................................................................................................................................................... 9

CHAPTER 2: AIMS OF THE PRESENT STUDY ................................................................. 11

2.0 Present study ...................................................................................................................... 11

2.1 Previous work and studies ................................................................................................ 11

2.2 Motivation for current study ............................................................................................ 15

2.3 Geometallurgy and geometallurgical assessments ......................................................... 17

CHAPTER 3: SAMPLES COLLECTED, AND SAMPLE MINERALOGY AND GEOCHEMISTRY ..................................................................................................................... 19

3.0 Introduction .................................................................................................................. 19

3.1 samples collected and samples descriptions .................................................................... 19

3.2 Samples mineralogy and mineral modal abundances .................................................... 22

3.2 Mineralogical variation with depth ................................................................................. 23

3.3 Samples geochemistry ....................................................................................................... 29

iii

CHAPTER 4: SAMPLE MILLING, AND ELEMENT DEPORTMENT ............................ 34

4.0 Introduction ....................................................................................................................... 34

4.1 Milling tests ........................................................................................................................ 34

4.2 Grading analysis ................................................................................................................ 38

4.3 Element deportment .......................................................................................................... 39

CHAPTER 5: FLOTATION TESTS ........................................................................................ 44

5.0 Introduction ....................................................................................................................... 44

5.1 Flotation performances ..................................................................................................... 44

5.2 Modal mineralogy of feeds and timed concentrates ....................................................... 46

5.3 Particle and sulphides grain size distribution ................................................................. 49

5.4 Sulphides liberation analyses in feeds ............................................................................. 52

5.5 Comparison of sulphides liberation in ore feeds and concentrates .............................. 54

5.6 Mineral association and locking....................................................................................... 58

5.7 Flotation Recovery Efficiency Analyses .......................................................................... 62

5.8 Flotation Performance analyses ....................................................................................... 66

5.9 Grade and recovery analyses ........................................................................................... 70

CHAPTER 6: DISCUSSIONS ................................................................................................... 72

6.1 Introduction ....................................................................................................................... 72

6.2.1 Mineralogy .......................................................................................................................................................... 72

6.2.2 Geochemistry ..................................................................................................................................................... 72

6.2.3 Milling .................................................................................................................................................................. 73

6.2.4 Grading analysis .............................................................................................................................................. 75

6.2.5 Elemental deportment ................................................................................................................................... 75

6.2.6 Mass pulls and mineralogy .......................................................................................................................... 76

6.2.7 Mineralogy of feeds and concentrates..................................................................................................... 76

iv

6.2.8 Particle size and sulphide grain size distributions ............................................................................. 77

6.2.9 Sulphide liberation in feeds ......................................................................................................................... 77

6.2.10 Comparison of sulphides liberation in feed and concentrates ................................................... 77

6.2.11 Mineral association and locking ............................................................................................................. 78

6.2.12 Flotation recovery efficiency .................................................................................................................... 78

6.2.13 Grade and recovery analyses ................................................................................................................... 79

CHAPTER 7: CONCLUSIONS AND RECOMMENDATIONS .......................................... 80

REFERENCES ............................................................................................................................ 82

APPENDIX 1: METHODS ........................................................................................................ 94

A1.0 Introduction .................................................................................................................... 94

A1.1 Channel Sampling .......................................................................................................... 94

A1.2 Crushing .......................................................................................................................... 95

A1.3 Representative sample splitting .................................................................................... 95

A1.4 Grain mounts preparation ............................................................................................. 95

A1.5 CARBON COATING..................................................................................................... 96

A1.6 Milling.............................................................................................................................. 96

A1.7 Mineral liberation analysis ............................................................................................ 98

A1.8 Flotation procedure ........................................................................................................ 98

A1.9 Chemical analysis ......................................................................................................... 100

APPENDIX 2: MINERALOGICAL DATA ........................................................................... 103

APPENDIX 3: GEOCHEMICAL ANALYSES ..................................................................... 107

APPENDIX 4: MILLING TESTS, ELEMENT DEPORTMENT AND FLOTATION DATA ......................................................................................................................................... 111

APPENDIX 5: DETAILS OF CALCULATIONS PERFORMED FOR DATA REDUCTION ............................................................................................................................ 128

CORRECTIONS BASED ON REVIEWERS COMMENTS. .............................................. 134

v

LIST OF FIGURES

Figure 1.1Geological Map of the Bushveld Complex showing Lonmin Marikana Operations (modified after Von Guenewaldt et al., 1985) ...................................................................................................................................... 1

Figure 1.2 Property Boundaries of Lonmin Platinum Marikana Operations (Adapted from Cawthorn, 1999b; Davey, 1992) ................................................................................................................................................................ 2

Figure 1.3 Regional Geological Map of the Bushveld Complex - Different Lithological Units and Limbs of theBushveld Complex .......................................................................................................................................................... 6

Figure 2.1 Facies Types and PGE Distribution for different Merensky Reef facies. The red bars indicate the abundance and distribution of PGE across the facies (Adapted from Lonmin Group, 2006) ................. 13

Figure 2.2 Vertical distribution of Cu, Ni and PGE in the UG2 Layer (Adapted from Lonmin Group, 2006) ............................................................................................................................................................................................ 16

Figure 3.1 Geological Logs Showing Lithological Variations across the BK (BK), RPM (RPM), WP(WP) facies and UG2 Chromitite reef .......................................................................................................................................... 21

Figure 3.2 Modal abundances (wt.%) distribution of sulphide minerals with depth for 10cm intervals of channel sample of the BK facies of Merensky Reef. Abundant sulphides correlate positively with the position of chromitite stringers. Refer to Figure 3.1 for legend of lithologies shown in this figure ............. 25

Figure 3.3 Modal abundances (wt.%) distribution of sulphide minerals with depth for 10cm intervals of channel sample of the RPM facies of Merensky Reef. Abundant sulphides correlate positively with the position of chromitite stringers. Refer to Figure 3.1 for legend of lithologies shown in this figure ............. 26

Figure 3.4 Modal abundances (wt.%) distribution of sulphide minerals with depth for 10cm intervals of channel sample of the WPfacies of Merensky Reef. Abundant sulphides correlate positively with the position of chromitite stringers. Refer to Figure 3.1 for legend of lithologies shown in this figure ............. 27

Figure 3.5 Modal abundances (wt.%) distribution of sulphide minerals with depth for 10cm intervals of channel sample of the UG2 chromitite facies reef. Refer to Figure 3.1 for legend of lithologies shown in this figure ................................................................................................................................................................................... 28

Figure 3.6 Distribution of Cr, S (wt.%), Pd, Pt and 6PGE (ppm) with depth for individual 10cm intervals of a channel sample of the BK facies of Merensky Reef. Abundant Cr, S, and PGE correlate with the position of chromitite stringers. Refer to Figure 3.1 for legend of lithologies shown in this figure ............. 30

Figure 3.7 Distribution of Cr, S (wt.%), Pd, Pt and 6PGE (ppm) with depth for individual 10cm intervals of a channel sample of the RPM facies of Merensky Reef. Abundant Cr, S, and PGE correlate with the position of chromitite stringers. Refer to Figure 3.1 for legend of lithologies shown in this figure ............. 31

vi

Figure 3.8 Distribution of Cr, S (wt.%), Pd, Pt and 6PGE (ppm) with depth for individual 10cm intervals of a channel sample of the WPfacies of Merensky Reef. Abundant Cr, S, and PGE correlate with the position of chromitite stringers. Refer to Figure 3.1 for legend of lithologies shown in this figure ............. 32

Figure 3.9 Distribution of Cr, S (wt.%), Pd, Pt and 6PGE (ppm) with depth for individual 10cm intervals of a channel sample of the UG2 Chromitite facies Reef. Refer to Figure 3.1 for legend of lithologies shown in this figure ................................................................................................................................................................. 33

Figure 4.1 Milling time against mass % passing 75µm sieve for BK sample ...................................................... 34

Figure 4.2 Milling time against mass % passing 75µm sieve for RPM sample .................................................. 35

Figure 4.3 Milling time against mass % passing 75µm sieve for WPsample ...................................................... 35

Figure 4.4 Milling time against mass % passing 75µm sieve for UG2 sample ................................................... 36

Figure 4.5 Comparison of milling times for the BK, RPM, WPand UG2 facies ................................................. 37

Figure 4.6 Cumulative particle size distributions after crushing and milling for BK, RPM, WPand UG2 38

Figure 4.7 Copper upgrade-downgrade curves for milled ore ................................................................................. 39

Figure 4.8 Nickel upgrade-downgrade curves for milled ore ................................................................................... 40

Figure 4.9 Sulfur upgrade-downgrade curves for milled ore .................................................................................... 41

Figure 4.10 Palladium upgrade-downgrade curves for milled ore. In UG2, -25µm was not analysed due to insufficient sample sizes ................................................................................................................................................... 42

Figure 4.11 Platinum upgrade-downgrade curves for milled ores ......................................................................... 43

Figure 5.1 Summary of time-cumulative mass pull ...................................................................................................... 45

Figure 5.2 Mineral modal abundances (weight%) in BK, RPM, WPand UG2 feed and concentrates ........ 47

Figure 5.3 Comparative mineral modal abundances (weight%) in BK, RPM, WPand UG2 concentrates 48

Figure 5.4 Cumulative particle size distribution for the BK, RPM, WPand UG2 facies milled feeds, -75+38µm fraction (as equivalent circle diameters) in microns ............................................................................... 50

Figure 5.5 Cumulative sulphides grain sizes distribution for the BK, RPM, WPand UG2 facies milled feeds, (as equivalent circle diameters) in microns ........................................................................................................ 51

Figure 5.6 Cumulative liberation yields for sulphides in the milled samples prior to flotation .................... 53

vii

Figure 5.7 Cumulative liberation yields for sulphides in BKFeed, BKConc1, BKConc2, BKConc3 and BKConc4 .................................................................................................................................................................................... 54

Figure 5.8 Cumulative liberation yield for sulphides in RPMFeed, RPMConc1, RPMConc2, RPMConc3 and RPMConc4 ........................................................................................................................................................................ 55

Figure 5.9 Cumulative liberation yield for sulphides in WPFeed, WPConc1, WPConc2, WPConc3 and WPConc4 ................................................................................................................................................................................... 56

Figure 5.10 Cumulative liberation yields for sulphides in UG2Feed, UG2Conc1, UG2Conc2, UG2Conc3 and UG2Conc4......................................................................................................................................................................... 57

Figure 5.11 Comparative mineral recoveries as liberated, binary and ternary composite mineral particles for BKFeed, RPMFeed, WPFeed and UG2Feed ........................................................................................................... 60

Figure 5.12a-d: MLA based SPLGXMAP Chalcopyrite+Pentlandite+Pyrrhotite wt.% locked in binary particles in feed ........................................................................................................................................................................ 61

Figure 5.13a-d: MLA based SPLGXMAP Chalcopyrite+Pentlandite+Pyrrhotite wt.% locked in ternary particles in feed ........................................................................................................................................................................ 62

Figure 5.14a-c: Cumulative flotation recovery of copper, nickel and sulfur as function of flotation time for BK, RPM, WPand UG2 facies ...................................................................................................................................... 67

Figure 5.15a-c: Cumulative grade of Cu, Ni and S in the flotation concentrate as a function of cumulative mass pull percent ..................................................................................................................................................................... 69

Figure 5.16a-c: Cu, Ni, and S grades as function of recovery curves for BK, RPM, WPand UG2 facies . 71

viii

LIST OF TABLES

Table 3.1 Modal mineralogy (area %) of samples ........................................................................................................ 22

Table 4.1 Summary of milling times (minutes) required to attain a grind of 60% mass passing 75µm sieve for BK, RPM, WPand UG2 facies ...................................................................................................................................... 37

Table 5.1 Cumulative mass pull rate tests results (g) .................................................................................................. 44

Table 5.2 Average water recovery (g) per facies type ................................................................................................. 45

Table 5.3 Flotation efficiency percentages (wt.%) of PGE for BK, RPM, WPand UG2 facies based on feed and tails weights and assays ....................................................................................................................................... 64

Table 5.4 Flotation efficiency percentages (wt.%) of base metals and sulfur for BK, RPM, WPand UG2 facies based on feed and concentrate weights and assays ......................................................................................... 65

Table 6. 1 Milling times variation with mineralogy in BK, RPM, WP facies and UG2………….74

Table A1.1 Reagent suite addition and conditioning used in the flotation rate tests ......................................... 99

Table A1.2 Analytical detection limits used for assays in this study .................................................................... 102

Table A2.1 Mineral modal abundance (wt.%) variations of samples of the BK facies of the Merensky reef (-2mm crushed ore sample) (from top to bottom) ...................................................................................................... 103

Table A2.2 Mineral modal abundances (wt.%) variations of samples of the RPM facies of the Merensky reef (-2mm crushed ore sample) (from top to bottom) .............................................................................................. 104

Table A2.3 Mineral modal abundances (wt-%) variations of samples of the WPfacies of the Merensky reef (-2mm crushed ore sample) (from top to bottom) ...................................................................................................... 105

Table A2.4 Mineral modal abundances (wt-%) variations of samples of the UG2 Chromitite facies (-2mm crushed ore sample) (from top to bottom) .................................................................................................................... 106

Table A3.1 Distribution of Cr, (ppm), S (wt.%), and 6PGE (ppb) in the BK facies of Merensky Reef; abundant Cr, S, and PGE correlate with the position of chromitite stringers (-2mm crushed ore sample) (from top to bottom) ............................................................................................................................................................. 107

ix

Table A3.2 Distribution of Cr, (ppm), S (wt.%), and 6PGE (ppb) in the RPM facies of the Merensky Reef (-2mm crushed ore sample) (from top to bottom) ...................................................................................................... 108

Table A3.3 Distribution of Cr (ppm), S (wt.%), and 6PGE (ppb) in the WP facies of Merensky Reef; abundant Cr, S, PGE correlate with the position of chromitite (from top to bottom) .................................... 109

Table A3.4 Distribution of Cr, S (wt.%), Pd, Pt and 6PGE (ppm) in the UG2 Chromitite Reef (from top to bottom) ..................................................................................................................................................................................... 110

Table A4.1 Mass % passing 75µm for BK, RPM, WP, and UG2 facies ore samples ..................................... 111

Table A4.2 Grading Analysis results for BK, RPM, WPand UG2 ore facies .................................................... 112

Table A4.3 Assay results of ore feeds sized fractions ............................................................................................... 113

Table A4.4 Assay of Cu, Ni, S, Pd and Pt in BK, RPM, WPand UG2 ore feed sized fractions ................... 114

Table A4.5 Copper deportment results in BK, RPM, WPand UG2 ore milled feeds ...................................... 115

Table A4.6 Nickel deportment results in BK, RPM, WPand UG2 ore milled feeds ........................................ 116

Table A4.7 Sulfur deportment results in BK, RPM, WPand UG2 ore milled feeds ......................................... 117

Table A4.8 Palladium deportment results in BK, RPM, WPand UG2 ore milled feeds ................................ 118

Table A4.9 Platinum deportment results in BK, RPM, WPand UG2 ore milled feeds ................................... 119

Table A4.10 Mass pull (g) and water recovery (g) variation with time for BK, RPM, WPand UG2 samples (in duplicate) .......................................................................................................................................................................... 120

Table A4.11 Mineral modal abundances of feed and concentrates ...................................................................... 123

Table A4.12 SPLGXMAP Chalc+Pent+Pyrr Wt.% locking in BK, RPM, WPand UG2 feeds .................... 125

Table A4.13 Assay results of ore feeds, concentrates and tailings for BK, RPM, WPand UG2 facies ..... 126

Table A4.14 Flotation performance analyses of BK, RPM, WPand UG2 ore facies ...................................... 127

Table A5.1 Grading analysis of sample of the BK facies type of Merensky Reef ............................................. 128

Table A5.2 Deportment analysis for Pd in sieved mass fractions of samples of the BK facies type of Merensky Reef ....................................................................................................................................................................... 129

Table A5.3 Flotation recovery efficiency values from calculation examples .................................................... 131

Table A5.4 Mass pulls, grades, and recoveries in a sample of the BK facies type of Merensky Reef ....... 132

x

TERMINOLOGY

Reef facies: An ore type or group of ore types having a unique set of textural and compositional

properties from which their metallurgical performance can be predicted.

Platinum group minerals: Naturally occurring chemical compounds of the platinum group

elements, namely platinum, palladium, rhodium, ruthenium, iridium and osmium.

Modal abundance: Percentages of the mineral components of an ore or rock sample.

Grade: Elemental content of an ore, expressed in grams per tonne (g/t), percentage (%), parts per

million (ppm) or parts per billion (ppb) of ore.

Comminution: Reduction of particle sizes of ore to separate the valuable mineral constituents

from the gangue. Comminution involves crushing and milling.

Deportment: Preferential reporting of minerals or elements into a specific grind size fraction of

a milled ore.

Upgrade: Preferential reporting of a mineral or element into a specific grind size fraction of a

milled ore, thereby raising its grade/content in that grind size fraction.

Downgrade: Non-preferential reporting of a mineral or element into a specific grind size

fraction of a milled ore, thereby resulting in low grade of that particular mineral or element in

that grind size fraction.

Mineral liberation: Separation or unlocking of a valuable mineral from gangue, for example by

comminution.

xi

Liberation by free surface: The degree to which a mineral in a particle is exposed at the surface

of that particle.

Froth flotation: Process of separating valuable mineral constituents from gangue and

transferring them into a froth. Flotation is achieved by treating ore slurry with chemical reagents

to make them hydrophobic, followed by bubbling air at a controlled rate.

Mass pull: Amount of each concentrate collected during flotation.

Cumulative mass pull: Concentrate collected over specific time intervals.

Assay: Content of marketable end product in the ore.

Recovery: Percentage of the total element or mineral contained in the ore that is recovered in the

concentrate.

Liberated mineral: Mineral of interest that is considered to be completely unattached (100%

ore mineral) or containing a minor proportion of gangue.

Middlings: Particles composed of at least 50% of mineral interest and the rest gangue.

Locked mineral: Mineral of interest completely or almost completely enclosed in gangue

minerals.

Tailings: Residual material left after recoverying valuable minerals from the milled ore slurry by

flotation or other processes.

Mineral locking: Interpenetration of mineral surfaces into one another, such that comminution is

needed to separate them.

Brakspruit (BK), Rustenburg Platinum Mine (RPM) and Western Platinum Mine (WP) are

facies type names of the Merensky reef at Lonmin Marikana mine.

xii

ACKNOWLEDGEMENTS

I would like to express my sincere gratitude to the following parties for their valuable

contributions to this thesis:

• My supervisor Prof KS Viljoen and co-supervisor, MW Knoper, for motivating,

moulding and guiding me throughout the course of the research by positively criticizing

my work. Thank you for your patience and the freedom you afforded me to explore my

potential. I would also like to thank the Science Faculty of the University of

Johannesburg for allowing me to undertake this study and the Geology Department staff,

the Palaeo-Proterozoic Mineralisation Research Group (PPM) and colleagues for their

important advice and positive criticisms.

• This project could not have been possible without a research grant from the National

Research Foundation and Department of Science and Technology, Geometallurgy Chair

grant to Prof KS Viljoen.

• Many thanks go to Derek Rose from whom I benefited from many critical discussions on

this research and for his assistance during milling and flotation exercises at Dornfontein

Campus, Extractive Metallurgy Department.

• My thanks also go to Dr Reinke at Spectrau for his constant assistance regarding

instrumental analysis work and calibration, and to Lisborn Mangwane and Baldwin

Tshivhiahuvhi for preparing samples for analysis.

• I would also like to express my gratitude to Mrs Elsa Maritz, Mr Hennie and Ms Diana

Khoza for handling all administrative and finance related queries.

• Many thanks also go to Lonmin Marikana Platinum Mine for providing sample material

and data related to this study.

• I would also like to thank many friends and colleagues that I made during my study at the

University of Johannesburg.

• Finally, I would like to thank my family for their support and encouragement. I dedicate

this dissertation to my daughters, Rutendo and Ruvimbo.

xiii

ABSTRACT

The study aims to provide a geometallurgical assessment, including an evaluation of the response of different facies types of the Merensky Reef to mineral processing, and the identification of critical characteristics that determine processing behaviour. This is accomplished by obtaining quantitative mineralogical information, combined with chemical assay, laboratory scale milling and flotation testing. Lonmin Platinum’s Marikana Mine is located on the Western Limb of the Bushveld Complex to the east of Rustenburg. Platinum group elements (PGE) occur in, and are mined from, a variety of facies types of the Merensky Reef, and the UG2. For the purpose of the present study, three facies types of Merensky Reef samples and one sample of UG2 were used. The Merensky facies samples comprise of the BK, RPM, and Western Platinum variants. The mineral assemblages of the various Merensky Reef facies types at this locality comprise varying amounts of orthopyroxene, clinopyroxene, plagioclase, olivine, talc, serpentine, chlorite, chromite, magnetite and sulphides (mainly pyrrhotite, pentlandite and chalcopyrite). Approximately 20 individual 10 cm channel samples were collected from each of the facies variants of the Merensky Reef, and the UG2. These are coarsely crushed, mineral modal abundances determined using the MLA, and then analysed for Co, Cr, Cu, Ni, S and 6 PGE. The samples were then combined per facies type, and each of these composites subjected to laboratory scale milling and flotation testing. Abundant sulphide typically occurs with (is associated with) thin chromitite stringers, as is commonly observed in the Merensky Reef throughout the whole of the Bushveld Complex. Chromitite stringers are characterized by high PGE concentrations. The milling behaviour of the various facies samples, as well as flotation behaviour, was observed to be a function of mineralogy. The influence of ore mineralogy on the various stages of flotation, the mineralogical makeup of the various flotation concentrates, and the level of recovery of the PGE’s during flotation, were also investigated. Ore facies having the most abundant anorthite required the longest milling time to achieve the target grind of 60wt.% passing 75µm; and the ore with the most abundant enstatite produced the largest mass pull on floating. The facies with higher PGE grade, modal abundance of base metal sulphides, higher degree of liberation of base metal sulphides and least enstatite abundance produced the most favourable set of characteristics for efficient PGE recovery.

1

CHAPTER 1: INTRODUCTION 1.0 INTRODUCTION This chapter gives a brief overview of the geographical location of the study area, historical

mining activities, local geology within the broader regional geological setting, and previous

research work done on the study area. It also shows the linkage between the past geological

investigations carried out in this study area and the current research which have been aimed at

providing potential opportunities for more effective and efficient mineral extraction and

processing routes.

1.1 GEOGRAPHICAL SETTING, LOCAL GEOLOGY AND HISTORY Lonmin Platinum’s Marikana Operation is located at 25o 45’ S 27o21’E (Mclaren and De

Villiers, 1982; Von Gruenewaldt et al., 1985 and references therein) on the Rustenburg Layered

Suite of the Western Limb of the Bushveld Igneous Complex about 70km northwest of

Johannesburg, in the North West Province of South Africa, near Marikana town.

Figure 1.1Geological Map of the Bushveld Complex showing Lonmin Marikana Operations (modified after Von

Guenewaldt et al., 1985)

2

Lonmin Platinum Marikana Operation comprises Western Platinum Limited, including Karee

Mine, and Eastern Platinum Limited (Cawthorn, 1999b), all situated in the southern part of the

Western Limb of the Bushveld Complex (Figure 1.1).

The mine’s current lease area (Figure 1.2) covers the farms Zwartkoppies 296JQ , Rooikoppies

297JQ, northern part of Elands drift 467 JQ, Middelkraal 466JQ,Wonderkop 400JQ, Schaapkraal

292 JQ, part of LeeuWPort 402JQ , and Turfontein 462JQ farms (Lonmin Group, 2006; Davey,

1992).

Figure 1.2 Property Boundaries of Lonmin Platinum Marikana Operations (Adapted from Cawthorn, 1999b;

Davey, 1992)

Underground development to exploit the Merensky Reef began in 1970 and milling of the

Merensky Reef ore began in 1971. This was mainly to extract the platinum group minerals,

3

namely: platinum, palladium, rhodium, ruthenium, iridium and osmium. Gold, copper and nickel

are also extracted as by-products (Lonmin Group, 2006).

Metallurgical investigations into the PGM recovery from UG2 ore at Western Platinum Limited

were done jointly with the National Institute for Metallurgy (now Mintek) during 1980, and

Western Platinum Limited became the first company to exploit the UG2 Chromitite Layer for its

PGM content on a large scale (Cawthorn, 1999b). Mining of the UG2 at Western Platinum

Limited started in 1982. Currently, both the Merensky Reef and the UG2 are being mined

(Cawthorn, 1999b; Lonmin Group, 2006).

In 1987 the mine began sinking inclined shafts to exploit the UG2 at the Eastern Platinum

Limited, and the milling of ore from Eastern Platinum Limited began in 1989. It is noteworthy

that Eastern Platinum Limited currently produces only UG2 ore (Lonmin Group, 2006).

During 1988, Impala Platinum Limited (Implats) began to sink shafts to exploit the Merensky

Reef and UG2 at Karee before merging with Western Platinum Limited. In January 1990,

Lonmin Platinum Marikana Operation and Implats merged Karee Mine with Western Platinum

Limited. Lonmin Platinum Marikana Operation was given the responsibility of managing the

newly established entity (Lonmin Group, 2006).

The UG2 ore at Lonmin Platinum Marikana Operation is milled separately from the Merensky

Reef, with the chromite from the UG2 being produced as a by-product during the milling

process. Lonmin Marikana’s annual production (2005) was 13.5 million tonnes, consisting of 10

million tonnes of UG2 and 3.5 tonnes of Merensky Reef ores. Mining operations are currently

carried out at various localities along strike of the reefs, using mainly up-dip mining methods.

Mining depths range from 30 metres to 700 metres below the surface (Lonmin Group, 2006).

4

1.1.1 REGIONAL GEOLOGICAL SETTING: INTRODUCTION

This section briefly outlines the regional geological setting of the Bushveld Complex and links it

to the geological setting of Lonmin Platinum Marikana Operation mine lease area to provide a

contextual basis for the study.

1.1.2 REGIONAL GEOLOGICAL SETTING

The Bushveld Complex is the world’s foremost layered intrusion and hosts the largest deposits of

platinum group minerals and economically recoverable amounts of copper, nickel, chromium

and vanadium (Cawthorn, 1999b; Scoon and Mitchell, 2009 and references therein). It is divided

into the Eastern, Western, Far Western and Northern or Potgietersrus Limbs (Figure 1.3)

(Kinloch, 1982; Schouwstra et al., 2000).

The Bushveld Complex consists of ferromagnesian and calcium-aluminium-sodium silicate

rocks. Rock types found in the Merensky Reef range from feldspathic to pegmatoidal

pyroxenites, norites and anorthosites; UG2 has chromitites, pyroxenites and anorthosite; and

Platreef has pyroxenites, serpentinites and cal-silicate rocks. These rocks are repetitively

compositionally fractionated and layered in a cyclic fashion, resulting in a stratigraphic sequence

of tens of kilometres thick of rock layer (Cawthorn, 1999b; Schouwstra et al., 2000; Scoon and

Mitchell, 2009; and references therein). Mineralisation and ore deposit distribution are thought to

be controlled by this layering mechanism.

Chromite, ilmenite, and platinum group minerals are associated with the ferromagnesian (mafic)

component of the cyclic units, and magnetite, cassiterite, zircon and other oxides are associated

with the silicic part. Magmatic and hydrothermal mechanisms are postulated to account for the

variation in stratification and mineralization within the Bushveld Complex reefs (Cawthorn et al.,

1999a; Cawthorn et al., 2002). Common economic minerals found in the Bushveld Complex are

sulphides of iron, nickel, copper and sulphides, arsenides, tellurides and alloys of the platinum

group elements (PGE) forming the platinum group minerals (PGM). The rock layers are

5

generally laterally continuous except in places where they are transgressed by dunite pipes,

young intrusions and potholes (Kinloch, 1982; Naldrett et al., 1986).

The Western Limb of the Bushveld Complex consists of the following stratigraphic units from

bottom to top: the Lower Zone, the Lower Group, the Middle Group, Upper Group, the Critical

Zone, the Main Zone, and the Upper Zone. The Critical Zone encompasses the UG2, the

Merensky Reef pyroxenite and the Bastard pyroxenite/norite (Gruenewalt et al., 1986; Davey,

1992).

The emphasis of this study is on the Merensky Reef and the UG2 Chromitite (Upper Group

Chromitite Number 2) of the Critical Zone. On the Northern Limb is the Platreef, which rests

directly on the Transvaal metasedimentary sequence and Archaean granites, whereas the

Merensky Reef rests on the Bushveld rocks of the Critical Zone, the Middle Group, the Lower

Group and the Lower Zone before the sedimentary rocks of the Transvaal Supergroup

(Cawthorn, 1999b and references therein).

6

Figure 1.3 Regional Geological Map of the Bushveld Complex - Different Lithological Units and Limbs of theBushveld Complex

(Von Gruenewaldt et al., 1985)

7

1.1.3 THE MERENSKY REEF

The Merensky Reef is commonly considered a laterally uniform reef type, but large variations in

reef thickness, reef composition and position of mineralization occur. In its most general sense

the reef consists of a feldspathic pyroxenite underlain and overlain by thin (5mm to 10mm)

chromitite stringers (Brynard et al., 1976; Schouwstra and Kinlock, 2000; and references

therein). The Merensky Reef extends for about 300 kilometres around the whole outcrop of the

Eastern and Western Limbs of the Bushveld Complex, and to depths of 5 kilometres.

Its silicate mineralogy consists predominantly of orthopyroxene (~60wt.%), plagioclase

(~20wt.%), clinopyroxene (~15wt.%), phlogopite (~5wt.%) and minor olivine. Secondary

minerals include talc, serpentine, chlorite and magnetite (Brynard et al., 1976; Schouwstra and

Kinlock, 2000; and references therein)

The major base metal sulphide assemblage includes, of all the sulphides, pyrrhotite (~40wt.%),

pentlandite (~30wt.%) and chalcopyrite (~15wt.%). Millerite (NiS), troilite (FeS), pyrite (FeS2)

and cubanite (Cu5FeS4) also occur as trace minerals (Vermaak and Hendriks, 1976; Schouwstra

et al., 2000). Cooperite (PtS), braggite ((Pt,Pd)NiS)), sperrylite (PtAs2), and platinum group

element alloys are the major platinum group mineral species occurring in the Bushveld Complex.

Laurite (RuS2) can also be abundant, especially in the UG2 Chromitite layer (Kinloch, 1982;

Von Gruenewaldt et al., 1986 and references therein).

PGM distribution, especially Pt-Fe alloys, has been correlated with hot spots, reef disturbances

(potholes) and volatile activity in the Merensky Reef, Platreef and the UG2 chromitite layer

(Farquhar, 1986; Kinloch, 1982; Gain et al., 1982 and references therein). For example, Pt-Fe is

highly associated with dunite pipes and UG2 in close proximity to dunite pipes (Cawthorn,

1999b). Pt-Fe is a significant component of the western Bushveld Complex northeast and

southeast of the Pilanesburg Alkaline Complex and in the dunite pipes of the eastern Bushveld

Complex (Kinloch, 1982; Scoon et al., 2004), and in the UG2 layer in their immediate vicinity.

Sperrylite (PtAs2) is closely associated with potholes, faulting and reef alterations (Kinloch,

1982).

8

In the Merensky Reef the proportions of platinum and palladium average about 55vol.% and

32vol.% respectively, and the other metals constitute about 13vol.% (Cawthorn, 1999b).

The Merensky Reef contains 89vol.% Pt-Pd sulphides types, the pothole reef contains 92vol.%

Pt-Fe alloy, and the contact reef (where the Merensky Reef thins down to only a narrow

chromitite stringer) contains 87vol.% laurite (RuS2) and 11vol.% of the other Pt-Pd sulphide

types. Thus Pt-Fe PGM dominate the pothole reefs (Cawthorn et al., 2002 and references

therein).

The base metal sulphide mineralogy also shows a regional correlation with potholes. In the

Merensky Reef, the sulphide assemblage consists of pentlandite, pyrrhotite and chalcopyrite, but

the pothole reef contains mackinawite and cubanite in addition to the pentlandite, pyrrhotite and

chalcopyrite as the dominant sulphides. The contact reef contains essentially chalcopyrite. The

PGM grain sizes increase towards the potholes (Kinloch, 1982).

1.1.4 THE UG2 REEF

The UG2 reef is a platinum group element bearing chromitite layer, variably situated between

20m to 400m below the Merensky Reef. The UG2 thickness is usually 1m, but it can vary from

0.4m to 2.5m depending on the locality (McLaren and De Villiers, 1982 and references therein)

The major mineral phases of the UG2 are chromite (~60% to 90wt.%), orthopyroxene (~5% to

30%wt.) and plagioclase (~1wt.% to 10wt.%) (Vermaak, 1995). Minor mineral phases are

phlogopite, biotite, clinopyroxene, ilmenite, rutile, magnetite and base metal sulphides. Quartz,

talc and serpentine are present as secondary minerals. Chalcopyrite, pyrrhotite, pyrite,

pentlandite and to a lesser extent millerite constitute the major base metal sulphide assemblage

and occur interstitially within silicates and rarely enclosed in chromite (McLaren and De Villiers,

1982; Penberty et al., 2000; Schouwstra, 2000 and references therein).

The platinum group mineral assemblage of the UG2 chromitite layer ranges from sulphide to

non-sulphide minerals. The sulphides minerals include laurite (RuS2), cooperite (PtS), malanite

9

((Pt,Rh,Ir)2CuSO4)), braggite ((Pt,Pd)NiS)), and vysotskite (PdS) in some cases. Pt-Fe alloys (Pt-

Fe), and (Pt3Fe), tellurides, bismuthinides, bismuthotellurides of Pt and Pd, PGE arsenides and

sulphoarsenides constitute the non-sulphide platinum group minerals. Rustenburgite (Pt3Sn),

isomertierite (Pd11Sb2As2), arsenopalladinite (Pd8(As,Sb)3), plumbopalladinite (Pd3Pb2), potarite

(PdHg) and geversite (PtSb2) also occur. Generally the platinum group minerals in UG2 occur as

fine grains (averaging 12 microns), either associated with base metal sulphides, silicate gangue

and /or chromite grains, especially laurite (Kinloch, 1982; McLaren and De Villiers, 1982;

Gruenewaldt et al., 1986; Penberty et al., 2000; and references therein).

PGE values range from 4g per tonne to 7g per tonne (equivalent to 4-7ppm), depending on the

locality. The chromite content varies from 30wt.% to 35wt.% in the UG2 reef (McLaren and De

Villiers, 1982; Schouwstra et al., 2000). PGE concentrations are distributed throughout the UG2

reef, with larger peak values at the bottom of the seam and a relatively shorter peak at the top of

the seam (McLaren and De Villiers, 1982; Schouwstra et al., 2000).

1.1.5 THE PLATREEF

The Platreef is a mineralization situated in the Northern Limb of the Bushveld Complex, also

known as the Potgietersrus Limb. In the Platreef, the Bushveld Complex rocks rest directly in

contact with the floor rocks of the Archaean granites and the Transvaal metasedimentary

sequence. The Platreef consists of a complex assemblage of pyroxenites, serpentinites, dolomite

xenoliths and calc-silicates, resulting from the reaction caused by heat and material exchange

between the hot Bushveld Complex magma and the lime-rich floor rocks (Gain et al., 1982;

Schouwstra et al., 2000).

The base metal sulphide assemblage consists of pyrrhotite, pentlandite, chalcopyrite, pyrite and

occasional cubanite. The concentrations and distribution of nickel, copper and the PGEs vary

considerably, but the highest values are associated with serpentinites (Gain et al., 1982).

The major platinum group minerals are PGE tellurides, platinum arsenides, platinum sulphides

and platinum group element alloys. Serpentinites are enriched in sperrylite and the upper

10

pyroxenites are enriched with PGE sulphides and PGE alloys. Areas with high potholes, faulting

and reef alteration occurrences are also markedly enriched in sperrylite (Kinloch, 1982).

The PGE alloys in the mineralization are dominant closer to the floor rocks (Schouwstra et al.,

2000). Platinum group minerals often occur enclosed in or on grain boundaries of the base metal

sulphides and in silicate minerals in certain localities (Schouwstra et al., 2000).

11

CHAPTER 2: AIMS OF THE PRESENT STUDY

2.0 PRESENT STUDY

This section briefly describes some work that was already carried out in this study area and

highlights the specific part this current work will cover to add more information to the previous

research findings.

2.1 PREVIOUS WORK AND STUDIES

Exploration drilling, surface mapping, trenching and underground development at Lonmin

Platinum Marikana Operation have revealed remarkable lateral variations within certain

lithological units in the Upper Critical Zone of the Western Bushveld Complex within the mine

lease area (Davey, 1992).

Laterally, the Merensky Reef thickness varies from 2m in the west to 12m in the east of the study

area, with PGE concentrations being higher in the upper chromitite stringer, but higher PGE

concentrations are sometimes found in the lower chromitite stringer (Brynard et al., 1976;

Farquhar, 1986).

The Merensky Reef is that part of the Merensky cyclic unit, which is economically exploitable

for its platinum group elements minerals. In the Marikana mine lease area the Merensky Reef

extends laterally from west to east exhibiting recognizable reef variations in terms of lithology,

reef thickness and platinum group minerals grade distribution.

The Merensky Reef is generally considered a 1m feldspathic pyroxenite sequence which is

underlain by an anorthosite and bounded by upper and lower chromitite stringers or bands

(Brynard et al., 1976; Davey, 1992).

12

Some parts of the Merensky pyroxenite sequence are pegmatoidal in nature. At the top of the

pyroxenite sequence is a 1m orthopyroxenite overlain by an anorthosite (Cawthorn and Boerst,

2006).

There are varying concentrations of platinum group elements within the Merensky Reef, with the

highest concentrations found within the upper and lower chromitite stringers (Brynard et al.,

1976; Davey, 1992; Cawthorn et al., 2002).

Six facies types of the Merensky Reef exist in this area, namely the Brakspruit facies (BK), the

Rustenburg Platinum Mine (RPM) facies, the Thin facies, the Transitional thin facies, the

Marikana facies, the Western Platinum (West Plats or WP) facies and the Eastern Platinum (or

East Plats) facies (Figure 1.4) (Davey, 1992; Lonmin Group, 2006). However, this study focuses

only on the BK, the RPM, the West Plats and the UG2 chromitite layer facies.

The RPM facies has 0.3m to 1m thick Merensky pyroxenite, and the platinum group elements

are concentrated adjacent to the chromitite stringer at the base of the pyroxenite (Figure 2.1).

The West Plats facies consists of pyroxenite and pegmatoidal pyroxenite with thicknesses

varying from 2m to 5m, and PGEs concentrated mainly at the upper chromitite stringer.

Marikana facies together with the Thin and the Transitional Thin facies are transitional facies

between the RPM and the West Plats facies. The Merensky pyroxenite in the Marikana facies

varies from 1m to 2m, with the PGEs almost continuously distributed from the lower chromitite

to the upper chromitite stringers.

The Thin facies consists of a pyroxenite, which varies from 0.5m to 1m, and the PGEs are

concentrated within the basal chromitite at the base of the pyroxenite in the Thin facies-

Transitional thin facies (Davey, 1992).

13

The BK facies, about 2m thick, has two chromitite stringers, one at the base of the Merensky

pegmatoid and another up at the Merensky pyroxenite-pegmatoid contact. The PGEs are

concentrated about these two stringers (Figure 2.1).

The Eastplats facies has an approximately 10m thick Merensky pyroxenite sequence, with the

chromitite stringer only present at the base. The PGEs are predominantly concentrated at the

upper part of the Merensky pyroxenite, and lower PGE concentration occurs about the single

chromitite stringer (Lonmin Group Guide, 2006).

The UG2 chromitite layer is beween 0.8m to 1.5m thick and is underlain variably by a UG2

pegmatoid in some places and a mottled anorthosite in others, and overlain by a UG2 pyroxenite,

followed by one or two thin (0.10mm to 0.15mm) leader chromitite layers laterally in the mine

lease area. Platinum group elements concentrations have peak values near the top and at the base

of the UG2 chromitite layer (Davey, 1992) (Figure 2.2).

Figure 2.1 Facies Types and PGE Distribution for different Merensky Reef facies. The red bars indicate the

abundance and distribution of PGE across the facies (Adapted from Lonmin Group, 2006)

14

The gangue mineralogy of the Merensky Reef at Marikana (Western Platinum Ltd) typically

consists of orthopyroxene (~70wt.%), plagioclase (~20wt.%), clinopyroxene (~4wt.%), biotite

(~2wt.%) and 3wt.% quartz, sulphides and chromite. Secondary talc is common, mainly at the

grain boundaries of sulphides and pyroxene. Hornblende is also present in minor amounts. Major

base metal sulphide mineralogy consists of pentlandite, pyrrhotite, chalcopyrite and pyrite. Trace

base metal sulphides include millerite (NiS), troilite (FeS) and cubanite (Cu5FeS5) (Brynard et

al., 1976; Schouwstra et al., 2000).

Oxides, mainly chromite and rutile are also present in small amounts.

Precious metals, namely platinum, palladium, ruthenium, rhodium, iridium and gold are present

in varying proportions in the Merensky Reef in the Marikana mine lease area (WP, including

Karee Mine). The PGE grade in the Merensky Reef in this area varies between 5g per tonne and

7g per tonne (equivalent to 5-7ppm), consistent with the regional trend in the western limb of the

Bushveld Complex (Schouwstra et al., 2000; Cawthorn et al ., 2002).

These PGEs occur in sulphides, arsenides, tellurides, bismuthotellurides, and alloys as platinum

group minerals. These platinum group minerals consist of mainly PGE sulphides: cooperite

(PtS), braggite (PtPdNiS), and laurite (RuS); telluride: moncheite (PtTe2); bismuthotellurides:

michenerite (PdBiFeTe), kotulskite (Pd(Te,Bi)) and maslovite (PtTeBi); arsenide, sperrylite

(PtAs2); antimonides, stibiopalladinite (PdSb) and geversite (PtSb) and an alloy (PtFe). Gold

exists as an alloy, electrum and as a rare compound, AuBiPdTe (Brynard et al., 1976; Viljoen et

al., 2012). Sperrylite is the most abundant PGM over all the other PGM in this area (Brynard et

al., 1976).

The UG2 chromitite layer in this study area consists of chromite (60wt.% to 90wt.%),

orthopyroxene (5wt.% to 25wt.%), and plagioclase (5wt.% to 15wt.%). Clinopyroxene, base

metal sulphides and other sulphides, platinum group minerals, ilmenite and magnetite occur as

accessory minerals (McLaren and De Villiers., 1982; Penberty et al., 2000). Pentlandite,

pyrrhotite, pyrite and chalcopyrite are the dominant base metal sulphides. Millerite is present in

lesser quantities. The PGM identified in this layer are cooperite, laurite, braggite, Pt-Fe alloy and

15

sperrylite. The relative proportions of precious metals in the UG2 in this study area are

49.5vol.% Pt, 22.5vol.% Pd, 15vol.% Ru, 8.7vol.% Rh, 3.7vol.% Ir and 0.6vol.% Au (Lonmin

Group Guide, 2006).

UG2 thickness varies between 0.7m to 1.3m laterally and becomes generally thicker towards the

eastern part of the lease area. The entire layer is extracted during mining. The UG2 in the

northwestern part of Marikana branches into two and is separated by an internal feldspathic

pyroxenite layer. This variation of the UG2 is referred to as the RPM or NW facies (Davey,

1992). Typically the PGE concentration is highest in the middle and at the base of the UG2

chromitite layer (McLaren and Devilliers, 1982) (Figure 2.2). Nickel and copper also have peak

values at the centre (Figure 2.2), although nickel has a larger proportion than copper (Lonmin

Group Guide, 2006).

2.2 MOTIVATION FOR CURRENT STUDY

Mineralogical investigations have established mineralogical variability of the platinum group

mineral distribution in the Merensky Reef at Lonmin’s Marikana Operation. These investigations

based on drill cores as well as plant feed and concentrate were used to determine the

mineralogical variations of various facies at Lonmin Platinum Marikana Operation and regional

correlations (Brynard et al., 1976; Davey, 1992). More recently an investigation was conducted

on drill core in the Merensky Reef at Marikana Platinum mine to establish the mineralogy of the

PGM within the high grade chromitite stringers using an FEI 600 Mineral Liberation Analyser

(Viljoen et al., 2012).

Data for detailed geometallurgical assessments of the individual reef facies has not been

published in the international scientific literature. The variability of ore and gangue mineralogy

and variations in PGM abundances within the various facies of the Merensky Reef could pose

inherent challenges to PGM liberation behaviour and metallurgical responses in beneficication

processes (Brynard et al., 1976; Becker et al., 2008).

16

Figure 2.2 Vertical distribution of Cu, Ni and PGE in the UG2 Layer (Adapted from Lonmin Group, 2006)

This project seeks to characterize the Merensky Reef facies and UG2 at Lonmin’s Marikana

Operations to establish the influence, if any, of the reef facies variability on comminution and

flotation performance, such as platinum deportment, PGM liberation, and the abundances of

naturally floatable gangue (Xiao and Laplante, 2004; Becker et al., 2008; Becker et al., 2009;

Runge, 2010).

This study also aims to provide a geometallurgical assessment consisting of: (1) an evaluation of

responses of the different facies to mineral processing and (2) the identification of critical

characteristics that determine processing behaviour, by obtaining quantitative mineralogical and

textural information (Coetzee et al., 2011). This study might help refine mineral exploitation

17

(selective reef facies extraction) and processing strategies at Lonmin Platinum Marikana

Operation, and also serves as a basis for further research such as flotation conditions

optimization and grade optimization of the concentrates.

2.3 GEOMETALLURGY AND GEOMETALLURGICAL ASSESSMENTS

Geometallurgy is a multi-disciplinary, applied science which integrates mining, geology and

metallurgy to correctly help assess the processing needs of an orebody (Beniscelli, 2011; Lotter,

2011; Philander and Rozendaal, 2011).

Chemical analysis gives metal grade values in an ore, but does not give the distribution of the

metals in the various minerals within the ore, for example Ni in olivine, pentlandite and

pyrrhotite. Different minerals hosting the same metal type may respond differently to various

processing techniques. Therefore geometallurgy seeks to provide mineralogical information of

an ore to determine causes of different responses of minerals to various metallurgical processes

so that an appropriate metallurgical technique can be employed for effective mineral

beneficiation.

Geometallurgical assessments are carried out using the techniques: geometallurgical unit, ore

domain or ore facies classification, representative sampling, mineralogical measurements,

chemical assays, mill testing, deportment study and mineral separation testing.

Geometallurgical unit definition, ore domain or facies classification divides an orebody into ore

types based on host rock characteristics, alteration, grain sizes, structural geology, grade,

mineralogical variation, and metal ratios with focus on properties which are known to influence

metallurgical performance (Lotter, 2011). Representative sampling is then done for each facies

or geometallurgical unit.

Mineralogical measurements involve quantitative mineralogical characterisation of the coarsely

crushed and milled ore to determine mineral textures, mineral association, mineral modal

18

abundances, mineral grain sizes, grain size distribution, mineral liberation and elemental

deportment by mineral (Fandrich et al., 2007).

Bench top flotation testing is used to determine how different valuable minerals in the milled

feeds respond to mineral separation by froth flotation process. The concentrates and tailings are

all subjected to chemical analysis to determine grade; and to mineralogical analysis, to determine

the mineralogical characteristics causing poor metallurgical performance such as recovery losses

(Lotter et al., 2002, 2003 and 2011; Lastra, 2007; Kormos et al., 2010; Evans et al., 2011).

19

CHAPTER 3: SAMPLES COLLECTED, AND SAMPLE MINERALOGY

AND GEOCHEMISTRY

3.0 INTRODUCTION

Samples for this study were collected from underground workings at Lonmin Marikana Platinum

Mine at the sites, and sampled using the method given in Appendix A1.1.

Visual examination and mineralogical analyses conducted on the three mineralized Merensky

Reef facies channel samples, namely the BK, the Rustengurg Platinum Mine, the Western

Platinum facies, and the UG2 chromitite channel samples are briefly discussed in this chapter.

All the channel samples were visually examined and logged to establish their mineralogical

identities and characteristics (Figure 3.1).

3.1 SAMPLES COLLECTED AND SAMPLES DESCRIPTIONS

The BK facies is characterized, at the top, by medium to coarse grained spotted anorthosite, with

70vol.% plagioclase and 30vol.% pyroxene. This is followed by medium to coarse grained

pyroxenite with up to 90vol.% pyroxene and interstitial plagioclase (10vol.%), disseminated base

metal sulphides and a 5mm to 6mm fine to medium grained chromitite stringer.

Below the chromitite stringer is a coarse grained pegmatoidal pyroxenite, with up to 40vol.%

plagioclase phenocrysts and disseminated base metal sulphides. This is followed by another

chromitite stringer, and a mottled anorthosite with up to 70vol.% plagioclase and 30vol.% large

pyroxene phenocrysts at the bottom.

The RPM facies is characterized by medium to coarse grained pyroxenite, with up to 90vol.%

pyroxene and elongate clinopyroxene phenocrysts (20-30mm) and a chromitite stringer near the

top.

20

A second chromitite stringer (3mm to 5mm) is situated at the base of the pyroxenite, followed by

a coarse grained pegmatoidal pyroxenite, with up to 70vol.% pyroxene phenocrysts and

disseminated trace sulphides, oxides and chromite grains.

A third chromitite stringer (10mm) is located at the contact of the pegmatoidal pyroxenite on top

and a coarse grained mottled anorthosite (70vol.% plagioclase) below.

The WP facies sample consists of a coarse grained pyroxenite, with more than 80vol.%

pyroxene, isolated clinopyroxene phenocrysts (10mm to 20mm), disseminated base metal

sulphide blebs (~0.5vol.%) and a chromitite stringer near the top. A second chromitite stringer

(3mm to 20mm) is situated at the contact of the pyroxenite and a coarse grained mottled

anorthosite (~40vol.% plagioclase) below.

UG2 chromitite reef sample consists of medium to coarse grained pyroxenite at the top, with

disseminated chromite grains throughout, and three thin fine grained chromitite stringers

(≤1mm,3mm and ≤2mm) separated from the main chromitite below by thin pyroxenite bands

(15mm, 1mm and 1mm). The main chromitite consists of very fine grained chromite grains (≥95

vol.%) and very fine grained plagioclase (≤5vol.%). A fine grained anorthosite (≥95vol.%

plagioclase), with little amount of fine grained biotite and pyroxene (≤5vol.%), is situated at the

base.

Detailed quantitative mineralogical studies of each 10cm channel subsample was conducted

using a Scanning Electrom Microscope based Mineral Liberation Analyser (600F MLA) , to

establish accurate mineral assemblages, modal abundances and distribution in each channel

subsample; and across each facies.

21

Figure 3.1 Geological Logs Showing Lithological Variations across the BK (BK), RPM (RPM), WP(WP) facies

and UG2 Chromitite reef

22

3.2 SAMPLES MINERALOGY AND MINERAL MODAL ABUNDANCES

The mounts containing -2mm crushed rock samples for each of the three Merensky Reef facies

and the UG2 were analysed using a Mineral Liberation Analyser (Fandrich et al., 2007; Guy,

2003) (Appendix A1.7) . Their mineral assemblages consist of, in varying amounts, the silicate

minerals plagioclase (anorthite), clinopyroxene (augite), orthopyroxene (enstatite), epidote, K-

feldspar, phlogopite, quartz, amphibole (tremolite), and serpentine. Base metal sulphides

assemblage consists of chalcopyrite, pentlandite, and pyrrhotite. The main oxide is chromite.

Chromite is the major constituent of the UG2 mineral assemblages. Other minerals occur in

minor amounts.

Table 3.1 Modal mineralogy (area %) of samples

Mineral BK RPM WP UG2

Anorthite 31.14 20.59 15.66 25.15

Augite 7.93 5.77 5.84 1.15

Chromite 0.45 1.11 0.75 49.14

Enstatite 53.33 57.68 67.08 15.79

Epidote 0.08 0.31 0.28 0.08

K-Feldspar 0.11 0.18 0.27 0.20

Phlogopite 0.56 1.17 1.28 1.34

Quartz 0.14 0.46 0.76 0.07

Tremolite 0.58 1.38 1.19 0.46

Serpentine 0.33 5.35 1.43 0.27

Chalcopyrite 0.28 0.06 0.15 0.01

Pentlandite 0.49 0.09 0.21 0.02

Pyrrhotite 0.48 0.16 0.29 0.00

Other 4.09 5.68 4.82 6.32

Total 100.0 100.0 100.0 100.0

23

Modal abundances for individual facies channel samples and the UG2 are shown in Appendix 2,

Table A2.1-A2.4 for the BK, RPM, WP facies and UG2 chromitite respectively. Comparative

modal mineralogy of the three facies and the UG2 is presented in Table 3.1.

3.2 MINERALOGICAL VARIATION WITH DEPTH

In the BK (BK) facies sample, chromite is concentrated mostly at the stringers (7.30wt.% and

3.22wt.% at the upper and lower stringers respectively), and also between the two stringers

(1.33wt.%). Sulphides concentrations are generally lower at the chromitite stringers (Figure 3.2),

especially at the upper stringer, but increases away from the stringers (also shown in Appendix 2,

Table A2.1,).

Augite (clinopyroxene) is generally concentrated in the middle of the facies, the highest being at

the upper stringer (21.60wt.%) and the lowest (2.23wt.%) at the lower chromitite stringer.

Enstatite (orthopyroxene) concentrations at the stringers are lower, with 57.72%wt and

48.78%wt enstatite at the upper and lower stringers respectively, compared to about 80%wt on

moving away from the stringers.

In RPM facies chromite concentration increases downwards with depth. The lowermost

chromitite stringer has the highest chromite content (17.94wt.%), followed by the middle stringer

(5.52wt.%), and the uppermost stringer has the least chromite (4.5wt.%) (Appendix 2, Table

A2.2).

The sulphides content is highest at the upper chromitite (3.65wt.%), lowest at the middle stringer

(0.30wt.%), and higher near the bottom stringer (0.86wt.%) and 0.62wt.% at the stringer itself

(Figure 3.3 and Appendix 2, Table A2.2). Serpentine and plagioclase have lower concentrations

at the stringers, and orthopyroxene (enstatite) is distributed evenly across the reef except in the

anorthosite portion.

24

In WP, however, the sulphides are concentrated more around the chromitite stringers (Figure 3.4

and Appendix 2, Table A2.3). Augite and orthopyroxene (enstatite) have lower concentrations at

the chromitite stringers.

The chromite content is higher in the middle of the UG2 chromitite reef and decreases outwards

towards the reef edges (Appendix 2, Table A2.4). Sulphides (Figure 3.5), augite, orthopyroxene

(enstatite), plagioclase, phlogopite, serpentine, tremolite and quartz concentration are generally

higher towards the the edges of the reef (Table A2.4). Exceptions are the lower serpentine and

enstatite values in the anorthosite portion at the bottom of the reef. The highest plagioclase value

occurs in the anorthosite portion at the bottom of the reef, as expected.

25

Figure 3.2 Modal abundances (wt.%) distribution of sulphide minerals with depth for 10cm intervals of channel

sample of the BK facies of Merensky Reef. Abundant sulphides correlate positively with the position of chromitite

stringers. Refer to Figure 3.1 for legend of lithologies shown in this figure

26


sample of the RPM facies of Merensky Reef. Abundant sulphides correlate positively with the position of

chromitite stringers. Refer to Figure 3.1 for legend of lithologies shown in this figure

27


sample of the WPfacies of Merensky Reef. Abundant sulphides correlate positively with the position of chromitite

stringers. Refer to Figure 3.1 for legend of lithologies shown in this figure

28


sample of the UG2 chromitite facies reef. Refer to Figure 3.1 for legend of lithologies shown in this figure

29

3.3 SAMPLES GEOCHEMISTRY

Crushed channel samples were analysed according to Appendix 1, Section A1.9.

The assay values obtained for the PGE, the base metals Cr, Cu, Ni and sulphur are shown in

Appendix 3, Table A3.1-A3.4 for the BK, RPM, WP facies and the UG2 reef respectively. The

assays distribution across the BK, RPM, WP facies and the UG2 reef, from top to bottom, are

also shown in Figure 3.6-3.9 for the BK, RPM, WP and UG2 respectively.

From the assays data it can be observed that the PGE, Cr, Cu, Ni and S are all concentrated at the

chromitite stringers positions in all the Merensky Reef facies (Figure 5.6-3.8).

For BK, the PGE have also another concentration peak between the two chromitite stringers

(Figure 3.6, and Appendix 3, Table A3.1). PGE, Cr, Cu, Ni and S concentration peaks therefore

correlate positively with the chromitite stringer positions in all the Merensky Reef facies.

In the UG2 channel sample, Cr, Cu and Ni are concentrated in the middle of the chromitite reef.

S only is more concentrated at the top, in the UG2 pyroxenite portion (Figure 3.9, and Appendix

3, Table A3.4).

PGE have concentrations peak at the base of the chromitite reef. Ruthenium is the most abundant

of all the other PGE, with the exception of Pd and Pt.

30

Figure 3.6 Distribution of Cr, S (wt.%), Pd, Pt and 6PGE (ppm) with depth for individual 10cm intervals of a

channel sample of the BK facies of Merensky Reef. Abundant Cr, S, and PGE correlate with the position of


31


channel sample of the RPM facies of Merensky Reef. Abundant Cr, S, and PGE correlate with the position of


32


channel sample of the WPfacies of Merensky Reef. Abundant Cr, S, and PGE correlate with the position of


33


channel sample of the UG2 Chromitite facies Reef. Refer to Figure 3.1 for legend of lithologies shown in this

figure

34

CHAPTER 4: SAMPLE MILLING, AND ELEMENT DEPORTMENT

4.0 INTRODUCTION

Milling tests on BK, RPM, WP and UG2 samples were carried out to determine the effect of

mineralogy on milling times required to achieve grinds of 60wt% passing 75µm and liberation

characteristics of the the ore samples. Grading analysis and deportment studies of PGE, Cu, Ni

and S in various size fractions of the Merensky Reef facies and UG2 were done. The milling

tests were carried out using the method outlined in Appendix 1, Section A1.6.

4.1 MILLING TESTS

The milling results for the BK, RPM, WP and UG2 samples are shown in Figures 4.1-4.4, and

also in Appendix 4, Table A4.1.

Figure 4.1 Milling time against mass % passing 75µm sieve for BK sample

y = 1.167x + 12.50

01020304050607080

0 10 20 30 40 50 60

Wei

ght %

pas

sing

75µ

m

Milling times/ minutes

BK

35

Figure 4.2 Milling time against mass % passing 75µm sieve for RPM sample

Figure 4.3 Milling time against mass % passing 75µm sieve for WPsample

The wt.% passing 75µm, of the fines produced by the initial crushing stage, can be obtained by

extrapolation or calculation as intercepts of the graphs at the wt.% passing axes. Thus the wt.%

y = 1.275x + 13.24

0102030405060708090

0 10 20 30 40 50 60

Wei

ght %

pas

sing

75µ

m


RPM

y = 1.455x + 12.16

0102030405060708090

0.00 10.00 20.00 30.00 40.00 50.00

Wei

ght %

pas

sing

75µ

m


WP

36

passing 75µm, prior to milling, are 12.50%, 13.24%, 12.16% and 6.66% for BK, RPM, WP and

UG2 facies respectively (Figures 4.1-4.4).

Figure 4.4 Milling time against mass % passing 75µm sieve for UG2 sample

Figure 4.5 shows a comparison of the milling results for the BK, RPM, WP and UG2.

The milling results for the three Merensky facies and the UG2 samples are also summarized in

Table 4.1.

The milling times required to achieve a grind of 60% mass passing 75µm sieve are 40.67, 36.66,

32.86 and 31.37 minutes for the BK, RPM, WPand UG2 respectively (Figures 4.1-4.5).

y = 1.700x + 6.660

0102030405060708090

100

0 10 20 30 40 50 60

Wei

ght %

pas

sing

75µ


UG2

37

Figure 4.5 Comparison of milling times for the BK, RPM, WPand UG2 facies

Table 4.1 Summary of milling times (minutes) required to attain a grind of 60% mass passing

75µm sieve for BK, RPM, WPand UG2 facies

Reef facies type Mill times (minutes) required to achieve

60% weight passing75µm grind

BK 40.67

RPM 36.66

WP 32.86

UG2 31.37

BK, RPM and WPfacies show linear incremental (of about four minutes from one facies to the

next) milling times trend to achieve a grind of 60% passing 75µm, ranging from 32.86 to 40.67

minutes. The UG2 facies does not conform to the trend displayed by BK, RPM and WPf acies.

0102030405060708090

100

0 10 20 30 40 50 60

wt% passing

Milling time/minutes

WP

BK

RPM

UG2

38

4.2 GRADING ANALYSIS

The sized sample fractions mass percent and cumulative mass percent were calculated as shown

in Appendix 5, Example 1, and the results given in Appendix 4, Table A4.2. The cumulative

mass percentages were then plotted against particle size fractions (Coetzee et al., 2011; Evans et

al., 2011), giving the results shown in Figure 4.6 below.

Grading analysis results show that all the samples BK, RPM, WP and UG2 have finer particles

than the target grind of 60wt.% passing75µm after crushing and milling (Figure 4.6). However,

BK has the finest and WP the coarsest particles, with RPM and UG2 being intermediate and

similar in particle sizes.

Figure 4.6 Cumulative particle size distributions after crushing and milling for BK, RPM, WPand UG2

0

10

20

30

40

50

60

70

80

90

100

-25µm +25µm +53µm +75µm +106µm

wt %

pas

sing

Size Fraction, µm

BK

RPM

WP

UG2

39

4.3 ELEMENT DEPORTMENT

Sample size fraction mass percentages and elemental distribution percentages for each size

fraction were calculated from sized fraction sample assays and sized fraction sample masses

shown in Appendix 4, Table A4.2-A4.4. These results were calculated as shown in Appendix 5,

Example 2.

The deportment results for Cu, Ni, S, Pd and Pt in the BK, RPM, WP and UG2 milled ore

fractions are shown in Figures 4.7-4.11 and also in Appendix 4, Tables A4.5-A4.9.

Figure 4.7 Copper upgrade-downgrade curves for milled ore

-60.00

-40.00

-20.00

0.00

20.00

40.00

60.00

80.00

+106µm +75µm +53µm +25µm -25µm

Cu,

Dow

ngra

de-U

pgra

de

Size fraction

BK

RPM

WP

UG2

40

Results show that copper reports more to the +53µm and +75µm in all the Merensky Reef facies,

i.e, the BK, RPM and W Pfacies. For UG2 copper upgrades into the finer (+25µm) fractions

because the sulphides are smaller (Figure 4.7 and also Appendix 4, Table A4.5).

Figure 4.8 Nickel upgrade-downgrade curves for milled ore

Nickel reports more to finer size fractions (-25µm to +53µm) in BK, WP and UG2 facies and to

+75µm size fraction in RPM facies (Figure 4.8 and also Appendix 4, Table A4.6). The results

generally show that nickel downgrades in the coarser size fractions in all four ore types.

-20.00

-15.00

-10.00

-5.00

0.00

5.00

10.00

15.00

20.00

+106µm +75µm +53µm +25µm -25µm

Ni,

Dow

ngra

de-U

pgra

de

Size fraction

BK

RPM

WP

UG2

41

Figure 4.9 Sulfur upgrade-downgrade curves for milled ore

Sulphur reports preferentially to the +25µm and +53µm size fractions in BK, RPM and WP

facies. In the UG2 facies, copper upgrades preferentially into the -25µm fraction (Figure 4.9 and

also Appendix 4, Table A4.7). Thus UG2 has finer sulfur grain sizes than the Merensky facies.

Sulfur also slightly downgrades into the coarser size fractions, just like copper and nickel, in all

the Merensky Reef facies and the UG2.

-60.00

-40.00

-20.00

0.00

20.00

40.00

60.00

80.00

100.00

+106µm +75µm +53µm +25µm -25µmS, D

owng

rade

-Upg

rade

Size fraction

BK

RPM

WP

UG2

42

Figure 4.10 Palladium upgrade-downgrade curves for milled ore. In UG2, -25µm was not analysed due to

insufficient sample sizes

Palladium upgrades into the +25µm and +53µm, and downgrades in the +75µm fractions in all

the ore facies types (Figure 6.10, and also Appendix 4, Table A4.8). However, a palladium assay

for the -25µm size fraction in UG2 could not be done due to an insufficient sample amount

available for analysis.

-30

-25

-20

-15

-10

-5

0

5

10

15

20

25

+106µm +75µm +53µm +25µm -25µm

Pd, D

owng

rade

-Up

ggra

de

Size fraction

BK

RPM

WP

UG2

43

Figure 4.11 Platinum upgrade-downgrade curves for milled ores

Platinum upgrades into the finer fractions (-25µm to +53µm) in the RPM, WP and UG2, and

downgrades into +106µm fractions in the BK (Figure 4.11, and also Appendix 4, Table A4.9). A

platinum assay could not be performed for the -25µm size fraction in BK and WP facies due to

insufficient sample amounts available for analysis.

-30.00

-20.00

-10.00

0.00

10.00

20.00

30.00

40.00

+106µm +75µm +53µm +25µm -25µm

Pt, D

owng

rade

-Upg

rade

Size fraction

RPM

UG2

BK

WP

44

CHAPTER 5: FLOTATION TESTS

5.0 INTRODUCTION

This chapter briefly describes flotation performances and recovery efficiencies of BK, RPM, WP

and UG2 facies ores. The modal mineralogy of ore feeds and concentrates, and the liberation and

locking characteristics of the flotation products are also discussed.

5.1 FLOTATION PERFORMANCES

Flotation tests were carried out as described in Appendix 1, Section A1.8, using the reagent suite

shown in Appendix 1, Table A1.1. The flotation test results are shown in Figure 5.1, Tables 5.1-

5.2 and Appendix 4, Table A4.10.

Cumulative mass pull data as function of flotation time is summarized below in Table 5.1.

Table 5.1 Cumulative mass pull rate tests results (g)

Time/minutes 2 4 6 8 20

SAMPLE # Conc-1 Conc-2 Conc-3 Conc-4 Totals/g

BK 29.65 18.7 10.5 10.8 69.65

RPM 25.25 13.25 12.85 12.2 63.55

WP 70.25 41.05 24.85 28.15 164.3

UG2 28.95 16.35 8.4 8.15 61.85

The concentrate amounts reported in this chapter are the averages of two corresponding flotation

test runs for each facies as shown in Appendix 4, Table A4.10.

45

Figure 5.1 Summary of time-cumulative mass pull

Total water recoveries to BK, RPM, WP and UG2 concentrates are shown in Table 5.2.

Table 5.2 Average water recovery (g) per facies type

Facies type Average-water recovery/g

BK 405.6

RPM 387.1

WP 430.2

UG2 411.9

WP ore has the fastest flotation rate and highest overall mass pull, followed by BK, RPM and

finally UG2 (Figure 5.1 and Table 5.1).

Mass pull correlates positively with water recovery for each of the Merensky Reef facies and

UG2 ores (Table 5.1-5.2).

0

20

40

60

80

100

120

140

160

180

0 5 10 15 20 25

Cumulative mass pull/g

Flotation time/minutes

BK

RPM

WP

UG2

46

5.2 MODAL MINERALOGY OF FEEDS AND TIMED CONCENTRATES

Mineral modal abundances in milled feeds and timed concentrates were determined using MLA

(Gu, 2003; Fandrich et al., 2007) as described in Appendix 1, Section A1.7.

The mineral modal abundances results for BK, RPM, WP and UG2 feed and concentrates are

shown in Figures 5.2-5.3 and Appendix 4, Table A4.11.

The results show that the most abundant silicate minerals are augite, enstatite and plagioclase;

and the rest are in minor amounts. Chromite and magnetite are the main oxide minerals present

(Figure 5.2, and Appendix 4, Table A4.11).

The amount of augite and enstatite recovered to concentrate generally increases with flotation

time in all the facies. Chromite recovery to concentrate decreases with increasing flotation time

in the UG2 ore (Figure 5.2a-b).

WP has the highest enstatite recovery in all the timed concentrates of all the facies types. Talc

recovery decreases with flotation time in all four facies (Figure 5.3a-d).

Chalcopyrite, pentlandite and pyrrhotie are the major base metal sulphides recovered to

concentrate in all the facies. Minor amounts of galena are also recovered (Figure 5.2a-d).

Mineral modal abundances of chalcopyrite, pentlandite and pyrrhotie decrease with flotation

time in all four facies (Figure 5.3a-d). Most of the chalcopyrite is recovered to Conc1 in all the

facies, and BK has the highest recovery of chalcopyrite, pentlandite, and pyrrhotite to Conc1 in

all the facies (Figure 5.3a-d).

The PGM present in the feeds are RuS, PtFeSnS, PtTeBi, RhPtAsS, PdBi, PdBiTe and PtPdS

(Table A4.11). RuS is only recovered in BK Conc1, and PtTeBi is recovered in BK Conc3 and

UG2 Conc2 only. PtFeSnS and RhPtAsS are recovered to concentrates in all four facies, and

have decreasing mineral modal abundances with flotation time.

47

Figure 5.2 Mineral modal abundances (weight%) in BK, RPM, WPand UG2 feed and concentrates

0.0010.0020.0030.0040.0050.0060.0070.0080.00

BKFeed

BKConc1

BKConc2

BKConc3

BKConc4

0.0010.0020.0030.0040.0050.0060.0070.0080.00

RPMFeed

RPMConc1

RPMConc2

RPMConc3

RPMConc4

0.00

20.00

40.00

60.00

80.00

100.00

WPFeed

WPConc1

WPConc2

WPConc3

WPConc4

0.0010.0020.0030.0040.0050.0060.0070.0080.00

UG2Feed

UG2Conc1

UG2Conc2

UG2Conc3

UG2Conc4

a

b

c

d

48

Figure 5.3 Comparative mineral modal abundances (weight%) in BK, RPM, WPand UG2 concentrates

01020304050607080

CONC1 BK

CONC1 RPM

CONC1 WP

CONC1 UG2

020406080

100

CONC2 BK

CONC2 RPM

CONC2 WP

CONC2 UG2

0

20

40

60

80

100

CONC3 BK

CONC3 RPM

CONC3 WP

CONC3 UG2

020406080

100

CONC4 BK

CONC4 RPM

CONC4 WP

CONC4 UG2

a

b

c

d

49

PdBi has similar modal abundances in all four ore facies feeds, but it is not recovered to

concentrates. PdBiTe is only present in UG2 ore feed but it is not recovered to concentrates.

PtPdS is recovered only in BK Conc3 (Table A4.11).

5.3 PARTICLE AND SULPHIDES GRAIN SIZE DISTRIBUTION

Sample feeds were prepared as described in Appendix 1, Section A1.6 and subjected to MLA

study as as described in Appendix 1, Section A1.7. The results obtained are shown in Figures

5.4-5.5.

The particle size distribution cumulative curves for the -75+38µm BK, RPM, WP and UG2

sieved feeds show that all four facies have almost the same particle sizes, shown by the

superimposition of the four curves on each (Figure 5.4).

The sulphide minerals grain size distributions in BK, RPM, and WP are similar and those in UG2

are finer (Figure 5.5).

50

Figure 5.4 Cumulative particle size distribution for the BK, RPM, WPand UG2 facies milled feeds, -75+38µm

fraction (as equivalent circle diameters) in microns

51

Figure 5.5 Cumulative sulphides grain sizes distribution for the BK, RPM, WPand UG2 facies milled feeds, (as

equivalent circle diameters) in microns

52

5.4 SULPHIDES LIBERATION ANALYSES IN FEEDS

Total sulphides (chalcopyrite+ pentlandite + pyrrhotite) mineral liberation analyses of the BK,

RPM, WP and UG2 ore facies feeds, on the basis of mineral liberation by free surface (i.e,

surface area exposure of sulphide minerals), were done as outlined in Appendix 1, Section A1.7.

The results obtained are shown in Figure 5.6.

The results show that the sulphides liberations in BK, RPM and WP are all similar, with BK

being the most liberated. UG2 is less liberated than the Merensky Reef facies.

The results also show that the degree of sulphides liberation correlates positively with

cumulative mass recovery of liberated sulphides.

53

Figure 5.6 Cumulative liberation yields for sulphides in the milled samples prior to flotation

54

5.5 COMPARISON OF SULPHIDES LIBERATION IN ORE FEEDS AND CONCENTRATES

Comparative sulphides liberation results for BK, RPM, WP and UG2 feed and concentrates are

shown in Figures 5.7-5.10.

Figure 5.7 Cumulative liberation yields for sulphides in BKFeed, BKConc1, BKConc2, BKConc3 and BKConc4

55

Figure 5.8 Cumulative liberation yield for sulphides in RPMFeed, RPMConc1, RPMConc2, RPMConc3 and

RPMConc4

56

Figure 5.9 Cumulative liberation yield for sulphides in WPFeed, WPConc1, WPConc2, WPConc3 and WPConc4

57

Figure 5.10 Cumulative liberation yields for sulphides in UG2Feed, UG2Conc1, UG2Conc2, UG2Conc3 and

UG2Conc4

58

Figures 5.7-5.10 and Table A4.11 show that the flotation speed and recovery of sulphides to

concentrates are a function of liberation, that is, the more liberated the sulphides the faster they

float.

BK, RPM and WP have the highest apparently liberated sulphides in Conc2 (i.e 100% liberated

sulphides), while Conc3 has the highest apparently liberated sulphides in UG2 (Figure 5.7-5.10).

Conc4 has the least apparently liberated sulphides among the concentrates in all the Merensky

Reef facies and the UG2. Merensky Reef facies’ Conc3 and Conc4 have lower apparently

liberated sulphides than their corresponding feeds (Figure 5.7-5.9). All concentrates in UG2 have

higher apparently liberated sulphides than the feed (Figure 5.10).

5.6 MINERAL ASSOCIATION AND LOCKING

Mineral locking refers to an association of a mineral of interest with other minerals in an

unliberated state, either enclosed within or attached to those other minerals. Binary particles are

made up of a mineral of interest and one other mineral, whereas ternary particles are made up of

the mineral of interest in association with two or more other minerals either enclosed or attached

to them (Petruk, 2000).

In this study, the mineral of interest are the grouped sulphides

(chalcopyrite+pentlandite+pyrrhotite). The mineral locking data obtained using

MLA_SPLGXMAP as described in Appendix 1, Section A1.7 are shown in Figure 5.11-5.13 and

Table A4.12.

Figure 5.11 shows that the total liberated sulphides in BK, RPM, WP and UG2 are about

87wt.%, 83wt.%, 82wt.% and 58wt.% respectively, and the balance occurs as locked in or in

association with binary and ternary particles.

59

In the Merensky Reef facies the grouped sulphides form binary particles mainly with enstatite,

plagioclase and RhPtAsS; and in UG2 they form binary particles with chromite, enstatite,

hornblende, plagioclase, PtFeSnS, RuS, PtPdS and RhPtAsS (Figure5.12a-d).

Ternary particles form in association with augite, enstatite, hornblende, plagioclase and

magnetite in the Merensky Reef facies, whereas in UG2 grouped sulphides form ternary particles

in association with augite, chromite, enstatite, hornblende, forsterite, orthoclase, biotite,

plagioclase, PtS, quartz, PtPdS and magnetite (Figure 5.13a-d).

60

Figure 5.11 Comparative mineral recoveries as liberated, binary and ternary composite mineral particles for

BKFeed, RPMFeed, WPFeed and UG2Feed

61

Figure 5.12a-d: MLA based SPLGXMAP Chalcopyrite+Pentlandite+Pyrrhotite wt.% locked in binary particles in

feed

Figures 5.11-5.12 indicate that a larger fraction of the locked sulphides is hosted in binary

particles than in ternary particles.

0.001.002.003.004.005.00

BK_Mineral Locking for Chalc+Pent+Pyrr - Binary Particle (%)

0.001.002.003.004.00

RPM_Mineral Locking for Chalc+Pent+Pyrr - Binary Particle (%)

0.002.004.006.008.00

Aug

iteC

alci

teC

hrom

iteE

nsta

tite

Hor

nble

nde

Bio

tite

Plag

iocl

ase

PtFe

SnS

Qua

rtz

Tal

cT

rem

olite

RhP

tAsS

Mag

netit

e

WP_Mineral Locking for Chalc+Pent+Pyrr - Binary Particle (%)

0.002.004.006.008.00

Aug

iteC

alci

teC

hlor

iteC

hrom

iteE

nsta

tite

Hor

nble

nde

Ort

hocl

ase

Bio

tite

Plag

iocl

ase

PtFe

SnS

RuS

PtPd

SR

hPtA

sSM

agne

tite

UG2_Mineral Locking for Chalc+Pent+Pyrr- Binary Particle (%)

62

Figure 5.13a-d: MLA based SPLGXMAP Chalcopyrite+Pentlandite+Pyrrhotite wt.% locked in ternary particles

in feed

5.7 FLOTATION RECOVERY EFFICIENCY ANALYSES

PGE flotation recovery efficiencies could not be determined directly from recovered

concentrates due to low concentrate sample recoveries, but were calculated based on weights of

feeds and tails and assay values (Table A4.13) obtained as described in Appendix 1, Section

A1.9 and Table A1.2.

The PGE and gold recovery efficiencies were calculated as shown in Appendix 5, Example 3.

Flotation recovery efficiency results for Au, PGE, base metals and sulfur are shown in Tables

5.3-5.4.

0.000.200.400.600.80

Aug

iteC

alci

teC

hlor

iteE

nsta

tite

Hor

nble

nde

Fors

teri

teO

rtho

clas

eB

iotit

ePl

agio

clas

eQ

uart

zT

alc

Tre

mol

iteR

hPtA

sSM

agne

tite

BK_Mineral Locking for Chalc+Pent+Pyrr - Ternary+ Particle (%)

0.00

0.50

1.00

1.50

Aug

iteC

alci

teC

hrom

iteE

nsta

tite

Hor

nble

nde

Fors

teri

teB

iotit

ePl

agio

clas

ePt

FeSn

SQ

uart

zT

alc

RhP

tAsS

Mag

netit

e

RPM_Mineral Locking for Chalc+Pent+Pyrr - Ternary+ Particle (%)

0.000.200.400.600.801.00

Aug

iteC

alci

teC

hrom

iteE

nsta

tite

Gal

ena

Hor

nble

nde

Fors

teri

teO

rtho

clas

ePl

agio

clas

ePt

FeSn

SQ

uart

zSe

rpen

tine

Tal

cT

rem

olite

PtPd

SR

hPtA

sSM

agne

tite

WP_Mineral Locking for Chalc+Pent+Pyrr - Ternary+ Particle (%)

0.000.501.001.502.002.503.003.50

UG2_Mineral Locking for Chalc+Pent+Pyrr - Ternary+ Particle (%)

a b

c d

63

Gold recovery in the Merensky Reef facies ranges from 78.5% to 84.4wt.%, while the gold

recovery in UG2 is ~ 62.8wt.%. BK facies has the highest PGE recovery, followed by WP, all

with recoveries of above 90%, except Ru in WP. RPM and UG2 have similar PGE recoveries

(Table 5.3).

The gold and PGE recovery in the Merensky Reef facies positively correlates with sulphide

modal abundances (Table 3.1) and sulphide liberation (Figure 5.6). This suggests that these PGE

are mostly hosted in sulphides as PGM since they were recovered by the bulk sulphide flotation

process (Wiese et al., 2005, 2006, and 2007). Lower recovery of iridium, osmium and ruthenium

in UG2 and RPM suggests that these PGE are not entirely hosted in base metal sulphides.

The Au recovery in both the Merensky Reef and UG2 could imply that most of the gold could

have been hosted in sulphides or in solid solution in sulphide hosted PGM or as part of the PGM

mineral chemistry association or alloys. Or it exists as free gold.

Flotation recovery efficiencies for the base metals and sulphur were calculated as shown in

Appendix 5, Example 4. Overall copper flotation recovery efficiencies range from 96% to

100wt.% in all the facies (Table 5.4). Co, Cu and Ni generally correlate positively with sulphide

modal abundances (Table 3.1). Ni flotation recovery efficiency is generally lowest and the most

variable.

The low nickel flotation recovery efficiencies could imply that some of the nickel not recovered

by bulk sulphide flotation is probably hosted in non-sulphide mineral phases such as pyroxenes,

olivine or serpentine (Wiese et al., 2005).

64

Table 5.3 Flotation efficiency percentages (wt.%) of PGE for BK, RPM, WPand UG2 facies

based on feed and tails weights and assays

Facies Type

Element type

Flotation efficiency (%)

Based on Feed and Tails

Facies Type Element type Flotation efficiency (%)

Based on Feed and Tails

BK Au 80.80 WP Au 84.37

Ir 92.26 Ir 91.10

Os 91.39 Os 90.92

Pt 96.03 Pt 93.58

Pd 96.95 Pd 94.27

Rh 97.29 Rh 91.69

Ru 90.17 Ru 87.62

RPM Au 78.46 UG2 Au 62.77

Ir 84.62 Ir 89.28

Os 82.77 Os 82.89

Pt 94.74 Pt 93.77

Pd 91.14 Pd 94.04

Rh 92.34 Rh 93.23

Ru 80.69 Ru 83.35

65

Table 5.4 Flotation efficiency percentages (wt.%) of base metals and sulfur for BK, RPM,

WPand UG2 facies based on feed and concentrate weights and assays

Facies Type

Element type


Based on Feed and Conc

Facies Type

Element type


Based on Feed and Conc

BK Co 83.39 WP Co 32.95

Cu 98.01 Cu 96.11

Cr 5.91 Cr 11.39

Ni 87.87 Ni 63.39

S 100 S 100

RPM Co 14.63 UG2 Co 4.42

Cu 100 Cu 100

Cr 4.05 Cr 1.38

Ni 28.28 Ni 12.46

S 100 S 100

66

5.8 FLOTATION PERFORMANCE ANALYSES

Flotation performances of the four reef types were evaluated in terms of mass pull, grade and

recovery, based only on copper, nickel and sulphur. The flotation performance results for the

Merensky Reef facies and UG2 ore are shown in Figures 5.14-5.15 and also Appendix 4, Table

A4.14. These results were calculated as shown in Appendix 5, Example 4.

Cumulative percentage recoveries of copper, nickel and sulphur as functions of flotation times

are presented in Figure 5.14a-c.

Th results in Figure 5.14a indicate that BK facies ore has the fastest initial copper recovery rate

in the first 10 minutes, and then slows down, and finally reaching an overall copper recovery of

about 98%. The RPM, WPand UG2 ores have similar initial recovery rates. RPM and UG2 have

overall copper recoveries of 100wt.%, while BK and WP have slightly lower than 100wt.%

recoveries.

Ni flotation recovery rates are variable, decreasing from BK, WP, RPM to UG2 (Figure 5.14b).

The BK ore had the fastest nickel flotation recovery rate, with about 80wt.% nickel recovery in

the first 10 minutes and an overall recovery of about 88wt.% (Figure 5.14b, and also Table 5.4).

The WP ore had the second fastest nickel flotation recovery rate, with about 60% recovery in the

first 10 minutes and an overall nickel recovery of 64wt.%.

The nickel recovery was slightly faster in the RPM ore than UG2 ore. The RPM and UG2 ores

had overall flotation recoveries of about 28.3% and 12.5wt.% respectively.

This trend correlates positively with sulphide modal abundances (Table 3.1) and sulphides

liberation extent (Figure 5.6).

Sulphur flotation recovery rates are identical for the BK, RPM, WPand UG facies; however the

initial recovery rates for BK and UG2 are slightly higher than those for RPM and WP. All four

facies have overall sulfur flotation recoveries of 100wt.% (Figure 5.14c).

67

Figure 5.14a-c: Cumulative flotation recovery of copper, nickel and sulphur as function of flotation time for BK,

RPM, WPand UG2 facies

0.0020.0040.0060.0080.00

100.00120.00

0 5 10 15 20 25

Cu

reco

very

,%

Flotation time/min

BKRPMWPUG2

0.00

20.00

40.00

60.00

80.00

100.00

0 5 10 15 20 25

Ni r

ecov

ery,

%

Flotation time/min

BKRPMWPUG2

0

20

40

60

80

100

120

0 5 10 15 20 25

Cum

ulat

ive

S re

cove

ry,%

Flotation time, min

BKRPMWPUG2

a

b

c

68

Results of cumulative grades as function of mass pull percentages for BK, RPM, WP and UG2

ores are shown in Figure 5.15a-c.

BK has the highest overall copper and Ni grades and UG2 has the lowest. RPM and WP have

similar overall grades, but slightly higher than the UG2 (Figure 5.15a-b). BK also has the highest

overall sulfur grade as function of cumulative mass pull percentage, followed by WP (Figure

5.15c). RPM and UG2 have the lowest overall sulfur grade. These results correlate positively

with sulphide modal abundances and liberation in BK facies, that is, BK has the highest sulphide

abundance and the most liberated sulphides (Table 3.1, Figure 5. 6).

The overall lower Cu, Ni and S grades in RPM, WP and UG2 correlate positively to the

abundant orthopyroxene in RPM and WP and the chromite in UG2 (Table 3.1, and also

Appendix 4, Table A4.11). BK has the highest Ni, Cu and S grades and the least orthopyroxene

modal abundance.

69

Figure 5.15a-c: Cumulative grade of Cu, Ni and S in the flotation concentrate as a function of cumulative mass

pull percent

0

0.5

1

1.5

2

2.5

3

0 5 10 15 20

Cum

ulat

ive

Cu

gra

de %

Cumulative mass pull %

BKRPMWPUG2

0

1

2

3

4

5

6

7

0 5 10 15 20

Cum

ulat

ive

Ni g

rade

%

Cumulative mass pull%

BKRPMWPUG2

0

2

4

6

8

10

12

0 5 10 15 20

Cum

ulat

ive

S gr

ade

%

Cumulative mass pull %

BKRPMWPUG2

a

b

c

70

5.9 GRADE AND RECOVERY ANALYSES

Results of Cu, Ni and S grades as function of cumulative recovery percentages for the Merensky

Reef facies and UG2 ores are shown in Figure 5.16a-c and also in Appendix 4, Table A4.14.

BK, RPM and UG2 have 100% Cu recovery. WP has slightly less than 100% Cu recovery. BK

has the highest and UG2 the lowest Cu grades. RPM and WP have similar Cu grades (Figure

5.16a).

BK also has the highest Ni grade and recovery and UG2 has the lowest grade and recovery

(Figure 5.16b). Sulfur recovery is 100% for all the facies (Figure 5.16c). However, BK has the

highest grade followed by WP. RPM and UG2 have the lowest grades. Consistently higher

grades and recoveries for BK correlate positively with sulphides modal abundances (Table 3.1)

and the degree of sulphides liberation (Figure 5.6).

The 100% S recovery in all the facies may imply that all base metals, gold and PGM associated

with or hosted in sulphides are recovered to concentrates. The low Ni recoveries could also

imply that Ni in the facies is not hosted in pentlandite and pyrrhotite only. It could be hosted in

other mineral phases which are not amenable to bulk sulphide flotation such as orthopyroxene,

serpentine and olivine (Appendix 4, Table A4.11).

71

Figure 5.16a-c: Cu, Ni, and S grades as function of recovery curves for BK, RPM, WPand UG2 facies

0

0.5

1

1.5

2

2.5

3

0 20 40 60 80 100 120

Cum

ulat

ive

Cu

grad

e, %

Cumulative Cu recovery,%

BKRPMWPUG2

0

1

2

3

4

5

6

7

0.00 20.00 40.00 60.00 80.00 100.00

Cum

ulat

ive

Ni g

rade

,%

Cumulative Ni recovery,%

BKRPMWPUG2

0

2

4

6

8

10

12

0 20 40 60 80 100 120

Cum

ulat

ive

S gr

ade,

%

Cumulative S recovery, %

BKRPMWPUG2

a

b

c

72

CHAPTER 6: DISCUSSIONS 6.1 INTRODUCTION

This chapter discusses the results from the preceding chapters and relates them to some of the

findings of other researchers in the field of geometallurgy.

6.2.1 MINERALOGY

This study has shown that BK, RPM, WP and UG2 reef facies at Lonmin’s Marikana Mine have

variable silicate and base metal sulphide modal mineralogy (Table 3.1), which is in agreement

with previous researchers (Kinloch, 1982; Gruenewaldt et al., 1986; Penberty et al., 2000 and

references therein). The base metal sulphides are concentrated at and /or around the chromitite

stringers within the Merensky Reef facies (Figure 3.2-3.4, Appendix 2, and Table A2.1-A2.3).

The UG2 reef facies has the least sulphide modal abundance (Table 3.1, and Table A2.4), and the

sulphides are distributed more in the pyroxenite portion, that is, towards the edge of the

chromitite reef (Figure 3.5, and Table A2.4).

6.2.2 GEOCHEMISTRY

PGE, Cr, Cu, Ni, and S are concentrated at and/ or about the chromitite stringers positions in the

BK, RPM, and WP facies (Figure 3.6-3.8), as observed by previous researchers (Davey., 1992;

Lonmin Guide., 2006; and Viljoen et al., 2012).

The assay values obtained for the PGE, the base metals and sulphur for the BK, RPM and WP

facies respectively, are shown in Appendix 3, Table A3.1-A3.3.

In the UG2 reef, Cr, Cu and Ni are concentrated in the middle of the reef. S only occurs more

abundantly in the pyroxenite portion of the reef, though in trace amounts (Figure 3.9 and Table

A3.4).

73

The PGE are highly concentrated at the base of the chromitite reef with values of 11.57ppm,

8.82ppm, 2.11ppm and 2.27ppm for Pd, Pt, Rh and Ru respectively, which is in agreement with

previous observations (McLaren and De Villiers, 1982; Davey, 1992; Schouwstra et al., 2000

and references therein). Ruthenium is the most abundant of all the other PGE, with the exception

of Pd and Pt, which is in agreement with some previous researchers (Kinloch, 1982; McLaren

and De Villiers, 1982; von Gruenewaldt et al., 1986; Schouwstra., 2000 and references therein),

who observed that laurite (RuS2) is more associated with chromite grains, whereas the other PGE

are more associated with the base metal sulphides.

6.2.3 MILLING

The milling times required to achieve a grind of 60wt.% passing 75µm sieve were 40.7, 36.7,

32.9 and 31.4 minutes for the BK, RPM, WP and UG2 respectively (Figures 4.1-4.5 and Table

4.1).

The milling times decrease from BK to WP facies. This trend correlates positively with the

decreases in modal abundances of anorthite (Table 3.1). This can be explained in terms of the

mineralogical hardness of anorthite as compared to enstatite, augite and chromite. Anorthite has

a hardness of 6-6.5 on Mohs scale, compared to enstatite with 5.5, augite 5-6 and chromite 5.5

respectively (Deer et al., 1992).

Therefore the BK facies (with 31.1wt.% anorthite) needs the longest milling time (40.7 minutes),

in contrast to RPM (with 20.6wt.% anorthite, 36.7 minutes) and WP (with 15.7wt.% anorthite,

32.9 minutes) (Table 3.1 and Table 6.1). Quartz, though occurring in minor amounts, probably

acts as a grinding medium due to its hardness (Craig and Vaughan, 1994). That is, as quartz

content increases the effective grinding media (steel rods plus quartz) also increase leading to

reduced times required to achieve the grind of 60% passing 75µm, and the less the quartz content

the longer the time required to achieve the required grind (0.14% quartz required 40.7 minutes,

compared to 0.46wt.% quartz requiring 36.7 minutes and 0.76wt.% quartz requiring 32.9 minutes

74

to achieve the same grind of 60wt.% passing 75µm in the BK, RPM, WP and UG2 facies

respectively (Table 3.1)).

Table 6. 2 Milling times variation with mineralogy in BK, RPM, WP facies and UG2

Facies

type

Anorthite

(wt.%)

Augite

(wt.%)

Enstatite

(wt.%)

Chromite

(wt.%)

Mill times (minutes),

60% passing 75µm

BK 31.14 7.93 53.33 0.45 40.67

RPM 20.59 5.77 57.68 1.11 36.66

WP 15.66 5.84 67.08 0.75 32.86

UG2 25.15 1.15 15.79 49.14 31.37

The milling times also decrease with increasing modal abundances of K-feldspars (though in

minor amounts) from BK, RPM, and WP (Table 3.1) (Becker et al., 2008). The effect of the

higher value of the combined alteration minerals for RPM (8.36wt.%), which has the effect of

making the milling time shorter than that of WP is reduced by the higher RPM anorthite content

(20.6wt.%), thereby resulting in a longer milling time for the RPM ore than the WP facies ore,

which has a much higher orthopyroxene content (67.1wt.%). Thus a mineralogy-milling time

correlation is also observed, since the degree of mineral alteration has an effect on milling times

(Becker et al., 2008).

The UG2 facies shows a different milling behaviour due to mineralogical differences from the

Merensky Reef facies types. The UG2 facies has a much higher chromite modal abundance

(49.14wt.% chromite) (Table 6.1) than the BK facies (0.45wt.% chromite), RPM facies

(1.11wt.% chromite) and WPfacies (0.75wt.% chromite). The chromite abundance, and the fact

that it is least hard (5.5 on Mohs scale), makes the UG2 facies milling time (31.4 minutes) the

shortest of them all. The higher content of the harder anorthite content (25.15wt.%) could also

act as co-grinding media (Craig and Vaughan, 1994) with the steel rods in the mill. However,

75

UG2 ore has a much lower combined modal abundance of alteration minerals (2.14wt.%) than

both the RPM and WPfacies ores (Table 3.1).

Mineralogical characteristics such as mineral type, texture, modal abundance, mineral alteration

and hardness of the Merenky Reef facies and the UG2 also account for the amount of initial fines

produced by the initial crushing stage, with BK facies having 12.5wt.% of particles passing

75µm, RPM (13.24wt.% passing 75µm), WP (12.2wt% passing 75µm) and UG2 (6.7wt.%

passing 75µm) (Figures 4.1-4.4 (intercepts)) respectively. The higher modal abundances of the

harder anorthite and alteration minerals tend to produce more fines than the softer minerals such

as chromite (in UG2) during the crushing stage to produce a -2mm sample (Becker et al., 2008).

However, the milling results obtained in this study contrast with the results of Brough et al

(2010), who found that the anorthite rich NP2 reef of the Merensky Reef at Northam Platinum

Mine (with ~63wt.% plagioclase) has the shortest milling time compared to the Normal and the

P2 reef (with <17wt.% plagioclase).

6.2.4 GRADING ANALYSIS

Grading analysis (Figure 4.6) indicates that BK, RPM, WP and UG2 milled ore feeds all have

particles finer than the target grind of 60wt.% passing 75µm. However, WP particles are the

coarsest and BK the finest. This result correlates positively with the observation made in Section

4.1, which shows that the ore with the most abundant anorthite generally produces the largest

amount of fine particles among the Merensky Reef facies.

6.2.5 ELEMENTAL DEPORTMENT

Deportment analyses (Section 4.3) show that copper, nickel, sulphur, platinum and palladium are

distributed in various fractions of the milled ore (Grammatikopoulos et al., 2004; Olubambi et

al., 2006; Dai et al., 2008; Goodall, 2008; Kapsios et al., 2010 and Coetzee et al., 2011),

but more preferentially reporting to finer size fractions (≤75µm fraction). This is more dominant

76

especially in the -25µm and +25µm size fractions, probably due to the small grain size nature of

sulphides (Figure 5.5) and PGM in the Merensky Reef and UG2 chromitite reef (Kinloch, 1982;

McLaren and De Villiers, 1982; Gruenewaldt et al., 1986; Penberty et al., 2000 and references

therein). This might also imply that the base metal sulphides and PGM are better liberated in the

finer fractions (Ford et al., 2011). The downgrading of copper, nickel, sulfur, palladium and

platinum in the coarser size fractions may imply incomplete liberation of the base metal

sulphides which host these elements.

6.2.6 MASS PULLS AND MINERALOGY

Cumulative mass pulls for the facies decrease in the order WP, BK, RPM and UG2 (Figure 5.1).

This correlates positively with the combined modal abundances of naturally floatable gangue

(pyroxenes) (Table 3.1) and talc-induced, copper ions and nickel ions activated floatable gangue

minerals (augite, enstatite, plagioclase, chromite, and talc) (Senior et al., 1995; Malysiak et al.,

2004; Jasieniak and Smart, 2009 and 2010) (Section 5.2).

6.2.7 MINERALOGY OF FEEDS AND CONCENTRATES

The increasing modal abundances of augite and enstatite in concentrates with flotation time

(Figure 5.2) show that these minerals are naturally floatable (Lotter et al., 2008; Becker et al.,

2009), and they are the main diluents of metal grades observed in Section 5.9.

Almost all the chalcopyrite is recovered in Conc1; and pentlandite is recovered in Conc1-Conc2

in decreasing amounts. Pyrrhotite is recovered in Conc1-Conc4 also in decreasing amounts in the

Merensky Reef facies, but in increasing amounts in the UG2 (Figure 5.2). This trend is in

agreement with observations that chalcopyrite is the fastest floating sulphide, followed by

pentlandite, and pyrrhotite being the slowest floating sulphide (Miller et al., 2005; Wiese et al.,

2006; Becker et al., 2008).

77

PtFeSnS, RhPtAsS and PtTeBi are recovered in Conc1-Conc4 in almost all the facies, but RuS is

recovered only in Conc1. These results are in agreement with Penberty et al (2000), who

observed that these PGM are slow floaters, and RuS is a fast floating mineral.

6.2.8 PARTICLE SIZE AND SULPHIDE GRAIN SIZE DISTRIBUTIONS

The particle size distributions in the -75+38µm sized fractions are all similar for the Merensky

Reef facies types and finer for the UG2 (Figure 5.4).

Sulphides grain size distribution variations in all three Merensky Reef ore types are also

generally similar, but finer in UG2 ore (Figure 5.5). Therefore a finer grind for UG2 ore is

usually required to completely liberate the base metal sulphides and PGM, such as 70-80wt.%

passing 75µm (Hay and Roy, 2010).

6.2.9 SULPHIDE LIBERATION IN FEEDS

Cumulative liberation yields for sulphides in the milled samples prior to flotation decrease in the

order BK, RPM, WP and UG2 (Figure 5.6). This shows a positive correlation between sulphides

liberation, flotation rate, sulphides recovery (Section 5.2), base metal and sulphur recovery

(Section 5.7 and Figure 5.14a-c) (Lastra, 2007; Ford et al., 2011; Bushell, 2012).

6.2.10 COMPARISON OF SULPHIDES LIBERATION IN FEED AND CONCENTRATES

Conc1 and Conc2 in BK and WP have more liberated sulphides than their corresponding feeds,

and Conc3 and Conc4 have less liberated sulphides since most of the liberated sulphides are

recovered in Conc1-Conc2 (Figure 5.7 and Figure 5.9). The remaining sulphides are present

mostly as composite particles in binary and ternary particles (Fig 5.11-5.13 and Appendix 4,

Table A4.11-A4.12).

78

In RPM Conc1 has more liberated sulphides than the feed and this means most of the liberated

sulphides report to Conc1 (Figure 5.8). Little liberated sulphides remain and report to Conc2-

Conc4 and the balance is locked also in binary and ternary composite particles (Fig 5.11-5.13

and Appendix 4, Table A4.11-A4.12).

In UG2, Conc1-Conc4 all have more liberated sulphides than the feed, and Conc3 has the highest

liberated sulphides. This means that most of the liberated sulphides in UG2 are slow floating and

therefore report to Conc3 (Penberthy et al., 2000 and references therein). The little liberated fast

floating sulphides report to Conc1 and Conc2. Only very few liberated slow floating sulphides

remain after recovery of Conc3 and report to Conc4 (Figure 5.10 and Appendix 4, Table A4.11-

A4.12).

6.2.11 MINERAL ASSOCIATION AND LOCKING

UG2 has highest amount of sulphides in binary and ternary particles compared to the Merensky

facies (Figure 5.11-5.13), as it has finer sulphides grain sizes (Figure 5.5) and thus the least

amount of liberated sulphides of the four ore types (Figure 5.6). This is in agreement with Hay

and Roy (2010; and references therein), who observed that finer grained minerals are less easily

liberated and need finer grinding than the coarser grained minerals.

6.2.12 FLOTATION RECOVERY EFFICIENCY

PGE recoveries are generally above 90wt.% in all four facies, except Ir, Os and Ru in RPM and

UG2 (Table 5.3). Au recovery is generally lower than 90wt.% in all the facies. The lower

recoveries of Ir, Os, Ru and Au may suggest that these elements are not only associated with

base metal sulphides, but could also be in other forms which cannot be recovered by bulk

sulphide flotation such as alloys and free gold, or enclosed in silicate minerals (Vermaark and

Hendriks, 1976).

79

Flotation efficiency results (Table 5.4) also indicate that copper flotation recovery efficiencies

range from 96% to 100wt.% for BK, RPM and WP ores probably due to the presence of liberated

and highly floatable chalcopyrite contents (Ekmekci et al., 2005) in these ore. Nickel flotation

recovery efficiency values decrease in the order BK, WP, RPM and UG2. This shows a positive

correlation with decreasing pentlandite modal abundances in BK, WP, RPM and UG2 ores

(Table 3.1).

However, nickel recoveries for BK, RPM, WP and UG2 ores do not correlate with the degree of

sulphides liberation (Figure 5.6), especially for the RPM ore. Thus the low nickel recoveries in

RPM and UG2 might imply that some of the nickel is not hosted in sulphides (such as

pentlandite and pyrrhotite) but in other mineral phases probably pyroxenes, olivines and

serpentine which are depressed by Sendep30D addition during flotation (Wiese et al., 2005,

2007), especially after achieving a 100wt.% sulfur recovery in all four facies.

6.2.13 GRADE AND RECOVERY ANALYSES

Grade and recovery data indicate that grades are generally low and recoveries are high,

especially for Cu and S, for all the facies type (Figure 5.16a-c). The low grades are likely due to

dilution (Neethling et al., 2008) from increasingly high recoveries to concentrates of augite,

enstatite, plagioclase and talc (Figure 5.2a-b; Figure 5.15a and Figure 5.15c; and Appendix 4,

Tables A4.11) probably due to copper ions and talc activation (Gasparrini, 1981; Malysiak et al.,

2002 and 2004; Wiese et al., 2006 and 2011; Lotter et al., 2008; Becker et al., 2009; Jasieniak et

al., 2009).

80

CHAPTER 7: CONCLUSIONS AND RECOMMENDATIONS

Geometallurgical evaluations of orebodies from exploration up to the end of the life of a mine or

that of part of a mine, have become an integral part of many of mining operations (Lotter et al.,

2003, 2011a, 2011b). This involves continually assessing the variability in the metallurgical

responses of the ore, and thus highlighting the need, if any, to adjust the ore processing flowsheet

in response to mineralogical variations. The following conclusions are drawn from this

geometallurgical study.

1) The present study shows that the amount of plagioclase and orthopyroxene in the ores of the

Merensky Reef has a direct influence on the amount of energy required to produce a grind of

60% passing 75 micron, with longer milling times required for plagioclase-rich and

orthopyroxene-poor ores. If ores are blended (Adams, 2007; Van Tonder et al., 2010) prior to

milling, a reduction in milling time (and energy consumption) could be achieved by limiting the

plagioclase content of the blend.

2) The amount of orthopyroxene in the ores of the Merensky Reef has a direct influence on mass

pull, as orthopyroxene is naturally floating in character (Becker et al., 2009). High

orthopyroxene content is therefore detrimental to the efficient recovery of the PGE, due to

concentrate dilution during flotation. If ores are blended prior to milling to reduce the amount of

plagioclase as suggested in (1), then care should be taken not to increase the orthopyroxene

content, or a high mass pull will result, and thus consequently a dilution of the PGE grade.

3) Of the three facies types of Merensky Reef examined, the overall characteristics of the BK

facies type, that is, a high PGE grade, low abundance of enstatite, a high modal abundance of

base metal sulphides, and a higher degree of liberation of the base metal sulphides on milling of

the ore, represent the most favourable set of characteristics for the efficient recovery of PGE

(Brough et al., 2010). It is therefore the best quality ore of the three samples of Merensky Reef

examined, as is confirmed by the flotation testing, with the highest recovery of Pt and Pd

observed.

81

4) The finer grain size of base metal sulphides in the UG2 relative to the Merensky Reef requires

a finer grind in order to achieve maximum sulphides liberation and recovery of the PGE (Hay

and Roy, 2010). This does not necessarily imply a longer milling time than in the case of the

Merensky Reef, as mill testing indicate that UG2 mills slightly faster than Merensky Reef.

However, even at a grind of 60% passing 75 micron in the present study, the recovery of PGE in

UG2 is acceptable, and only exceeded by the sample of the BK facies of Merensky Reef.

The following recommendations are worth considering:

1) Milling and flotation tests could be done on the composites of the Merensky Reef facies and

the UG2 ores (Lee et al., 2008), since practically no single facies ore would be enough to meet a

processing plant’s daily target throughput. Thus a study to determine the processing performance

behaviour of the composite samples of all the four ore types could be essential.

2) Further work on flotation conditions optimization studies could be carried out in order to

improve grades and recoveries in the ore facies (Shackleton et al., 2007a, 2007b), possibly

through flotation reagents optimizations or use of reagents combinations (Senior et al., 1995;

Pietrobon et al., 1997; Wiese et al., 2005, 2006 and 2007; Becker et al., 2009; Corin et al., 2010;

Lotter et al., 2011).

3) Grade optimization tests could also be carried out through cleaner and re-recleaner kinetics

studies on the rougher concentrates to reduce the amount of floatable gangue reporting to final

concentrates, and thus improve the final metal grades in concentrates (Hay, 2010).

82

REFERENCES

Adams, M.D., 2007. Towards a virtual metallurgical plant 2: Application of mineralogical data.

Minerals Engineering 20, 472-479.

Allison, S.A., and O’Connor, C.T., 2011. An investigation into the flotation behaviour of

pyrrhotite. International Journal of Mineral Processing 98, 202-207

Andersen, J.C.O., Rollison, G.K., Snook, B., Herrington, R., and Fairhurst, R., 2009. Use of

QEMSCAN for the characterization of Ni-rich and Ni-poor goethite in laterite ores. Minerals

Engineering 22, 1119-1129.

Becker, M., Brough, C., Reid, D., Smith, D., and Bradshaw, D., 2008. Geometallurgical

characterization of the Merensky Reef at Northam Platinum Mine–Comparison of Normal,

Pothole and Transitional Reef types. Proceedings Ninth International Congress of Applied

Mineralogy, Brisbane, 391-399.

Becker, M., Harris, P.J., Wiese, J.G., and Bradshaw, D.J., 2009. Mineralogical characterization

of naturally floatable gangue in Merensky Reef ore flotation. International Journal of Mineral

Processing 93, 246-255.

Becker, M., de Villiers, J., and Bradshaw, D., 2010. The flotation of magnetic and non-magnetic

pyrrhotite from selected nickel ore deposits. Minerals Engineering 23, 1045-1052.

Beniscelli, J., 2011. Geometallurgy-Fifteen Years of Developments in Codelco: Pedro Carrasco

Contributions. In: The First AusIMM International Geometallurgy Conference 2011, 5-7

September 2011, Brisbane, Australia, pp3-7.

83

Brough, C.P., Bradshaw, D.J., and Becker, M., 2010. A comparison of the flotation behaviour

and effect of copper activation on three reef type from the Merensky reef at Northam. Minerals

Engineering 23, 846-854

Brynard, H.J., De Villiers, J.P.R., and Villjoen, E.A., 1976. A mineralogical investigation of the

Merensky Reef at the WP, near Marikana, South Africa. Economic Geology 71, 1299-1307.

Cawthorn, R.G., 1999a. Seventy-fifth Anniversary of the Discovery of the Platiniferous

Merensky Reef. The largest platinum deposits in the world. Platinum Metals Reviews 43, (4),

146-148.

Cawthorn, R.G., 1999b. The platinum and palladium resources of the Bushveld Complex. South

African Journal of Science 95, 481-489.

Cawthorn, R.G., Lee, C.A., Schouwstra, R.P., and Mellowship, P., 2002. Relationship between

PGE and PGM in the Bushveld Complex. The Canadian Mineralogist 40, 311-328.

Chapman, N.A., Shackleton, N.J., Malysiak, V., and O’Connor, C.T., 2011. The effect of using

different comminution procedures on flotation of platinum-group minerals. Minerals Engineering

24, 731-736.

Chetty, D., Gryffenberb, L., Lekgetho, T.B., and Molebale, I.L., 2009. Automated SEM study of

PGM distribution across a UG2 flotation concentrate bank: implications for understanding PGM

floatability. The Journal of the South African Institute of Mining and Metallurgy 109, 587-593.

Chryssoulis, S.L., 2001.Using Mineralogy to Optimize Gold Recovery by Flotation. Journal of

mineralogy, 48-50.

Coetzee, L.L., Theron, S.L., Martin, G.J., Van der Merwe, J.D., and Stanek, T.A., 2011. Modern

gold deportments and its application to industry. Minerals Engineering 24, 565-575.

84

Corin, K.C., and Harris, P.J. 2010. Investigation into the flotation response of a sulphide ore to

depressant mixtures. Minerals Engineering 23, 915-920.

Craig, J.R., and Vaughan, D.J., 1994. Ore microscopy and Ore petrography. Wiley Interscience

publication, John Wiley & Sons, Inc, 2nd edition, 331-332.

Czamanske, G.K., Haffty, J., and Nabbs, S.W., 1981. Pt, Pd and Rh Analyses and Beneficiation

of Mineralized Mafic Rocks from the La Perouse layered Gabbro, Alaska. Economic Geology

76, 2001-2011.

Dai, Z., Bos, J., Lee, A., and Wells, P., 2008. Mass balance and mineralogical analysis of

flotation plant survey samples to improve plant metallurgy. Minerals Engineering 21, 826-831.

Davey, S.R., 1992. Lateral variations within the Upper Critical Zone of the Bushveld Complex

on the farm Rooikoppies 297 JQ, Marikana, South Africa. South African Journal of Geology 95,

141-149.

Ekmekci, Z., Can, M., Ergiin, S.L., Gulsoy, O.Y., Benzer, H., and Celik, I.B., 2005. Performance

Evaluation of CBI Flotation Plant Using Mineralogical Analysis. The 19th International Mining

Congress and Fair of Turkey, IMCET2005, Izmir, Turkey, June 09-12, 233-240.

Evans, C.L., Wightman, E.M., Manlapig, E.V., and Coulter, B.L., 2011. Application of process

mineralogy as a tool in sustainable processing. Minerals Engineering 24, 1242-1248.

Fandrich, R., Gu, Y., Burrows, D. and Moeller, K., 2007. Modern SEM-based mineral liberation

analysis. International Journal of Mineral Processing 84, 310-320.

Farquhar, J., 1986. The WP. In: Annhaeusser, C.R., and Maske, S (Ed.). Mineral deposits of

southern Africa. Geological Society of South Africa, Johannesburg, 1135-1142.

85

Ford, F.D., Wercholaz, C.R. and Lee, A., 2011. Predicting process outcomes for Sudbury

platinum group minerals using grade-recovery modelling from mineral liberation analyzer

(MLA) data. The Canadian Mineralogist 49, 1627-1642. DOI: 10.3749/canmin.49.6.1627.

Gain, S.B and Mostert, A.B., 1982. The geological setting of the platinoid and base metal

sulphide mineralization in the Platreef of the Bushveld Complex in Drenthe, North of

Portgietersrus. Economic Geology 77, 1395-1404.

Gasparrini, C., 1981. The significance of applied mineralogy to metal extraction. In: Hausen,

D.M and Park, W.C (editors), Process Mineralogy, Extractive Metallurgy, Mineral Exploration,

Energy Resources: Proceedings of Symposium, February 22-26, 33-49.

Goodall, W.R., Scales, P.J. and Butcher, A.R., 2005. The use of QEMSCAN and diagnostic

leaching in the characterisation of visible gold in complex ores. Minerals Engineering 18, 877-

886.

Goodall, W.R., and Scales, P.J., 2007. An overview of the advantages and disadvantages of the

determination of gold mineralogy by automated mineralogy. Minerals Engineering, 20, 506-517

Goodall, W.R., 2008. Characterisation of mineralogy and gold deportment for complex tailings

deposits using QEMSCAN. Minerals Engineering 21, 518-523.

Gottlieb, P., Willie, G., Sutherland, D., Ho-Tun, E., Suthers, S., Perera, K., Jenkins, B., Spencer,

S. and Butcher, A., 2000. Using Quantitative Electron Microscopy for Process Mineralogy

Applications. JOM, 24-25.

Grammatikopoulos, T.A., Porritt, L., Petersen, J.S. and Christensen, K., 2004. Mineralogical

characterisation and process mineralogy of gold-bearing rocks from the Nalunaq gold deposit,

Greenland. Applied Earth Science: (Transactions of the Institution of Mining and Metallurgy,

Section B), September 2004, 113, B197-B203.

86

Grammatikopoulos, T.A., Kapsios, A., Zaccarini, F. and Tsikouras, B., 2007. Investigation of

platinum group minerals (PGM) from Pindos chromitites (Greece) using hydroseparation

concentrates. Minerals Engineering 20, 1170-1178.

Gu, Y., 2003. Automated Scanning electron Microscope based Mineral liberation Analysis. An

introduction to JKMRC/FEI Mineral Liberation Analyser. Journal of Minerals & Materials

Characterization & Engineering 2, 1, 33-41.

Gulsoy, O.Y., 2005. A simple model for the calculation of Entrainment in Flotation. Korean

Journal of Chemical Engineering 22, 4, 628-634.

Hagni, A.M., 2008. Phase Identification, Phase Quantification, and Phase Association

Determinations Utilizing Automated Mineralogy Technology. Journal of Mineralogy, 33-37.

Hay, M.P., 2010. A case study of optimisng UG2 flotation performance part 2: Modelling

improved PGM recovery and C2O3 rejection at Northam’s UG2 concentrator. Minerals


Hay, M.P and Roy, R., 2010. A case study of optimizing UG2 flotation performance. Part 1:

Bench, pilot and plant factors which influence Cr2O3 entrainment in UG2 flotation. Minerals


Hunt, J., Berry, R. and Bradshaw, D., 2011. Characterising chalcopyrite liberation and flotation

potential: Examples from an IOCG deposit. Minerals Enginerring 24, 1271-1276.

Jasieniak, M., and Smart, R.S.C., 2009. Collectorless flotation of pyroxene in Merensky ores:

Residual layer identification using ToF-SIMS analysis. International Journal of Mineral


87

Jasieniak, M., and Smart, R.S.C., 2010. Surface chemical mechanisms of inadvertent recovery of

chromite in UG2 ore flotation: Residual layer identification using statistical ToF-SIMS analysis.

International Journal of Mineral Processing 94, 72-82.

Kapsiotis, A., Grammatikopoulos, T.A., Zaccarini, F., Tsikouras, B., Garuti, G. and

Hatzipanagiotou, K., 2006. Platinum-group mineral characterisation in concentrates from low-

grade chromitites from the Vourinos ophiolite Complex, northern Greece. Applied Earth

Science: Transactions of the Institution of Mining and Metallurgy, Section B, 115, 2, 49-57.

Kapsiotis, A., Grammatikopoulos, T.A., Tsikouras, B. and Hatzipangiotou, K., 2009. Platinum-

Group Mineral Characterization in Concentrates from High-Grade PGE Al-rich Chromitites of

Korydallos area in the Pindos ophiolite Complex (NW Greece). Resource Geology 60, 2, 178-

191 (doi:10.1111/j.1751-3928.2010.00124.x)

Kerrick, D.M., Eminhizer, L.B. and Villaume, J.F., 1973. The role of carbon thickness in

electron microprobe analysis. The American Mineralogist 58, 920-925.

Kinloch, E.D., 1982. Regional trends in the platinum-group mineralogy of the Critical Zone of

the Bushveld Complex, South Africa. Economic Geology 77, 1328-1347.

Kormos, L., Oliveira, J., Fragomeni, D., Whiteman, E. and Carrion De la Cruz, J., 2010. Process

Diagnosis using Quantitative Mineralogy. 42nd Annual CMP Conference January 19-21, 1-39

Lastra, R., 2007. Seven practical application cases of liberation analysis. International Journal of

Mineral Processing 84, 337-347.

Lee, K., Archibald, D., McLean, J. and Reuter, M.A., 2008. Flotation of mixed copper oxide and

sulphide minerals with xanthate and hydroxamate collectors. Minerals Engineering (2008),

doi:10.1016/j.mineng.2008.11.005.

88

Lenahan, W.C., and Re de Murray-Smith, 1986. Assay and Analytical Practice in the South

African Mining Industry. The South African Institute of Mining and metallurgy. Monograph

series M6, pp 41-594.

Lonmin Group, 2006. A guide to the geology at Lonmin Platinum, Marikana operations.

Unpublished Company Report, pp1-20 (Unpublished company report).

Lotter, N.O., Whittaker, L., Kormos, L., Stickling, J.S. and Wilkie, G.J., 2002. The development

of process mineralogy at Falconbridge Limited, and application to the Raglan Mill. CIM Bulletin

95, 1066, 85-92.

Lotter, N.O., Kowal, D.L., Tuzun, M.A., Whittaker, P.J. and Kormos, L., 2003. Sampling and

flotation of Sudbury Basin drill core for process mineralogy modelling. Minerals Engineering 16,

857-864.

Lotter, N.O., 2007. Distribution modelling of the nickel assay grades in the final tailings at

Raglan, Quebec. Minerals Engineering 20, 1067-1074.

Lotter, N.O., Bradshaw, D.J., Becker, M., Parolis, L.A.S. and Kormos, L.J., 2008. A discussion

of the occurrence and undesirable flotation behaviour of orthopyroxene and talc in the processing

of mafic deposits. Minerals Engineering 21, 905-912.

Lotter, N.O., Di Feo, A., Kormos, L., Fragomeni, D. and Comeau, G., 2010. Design and

measurement of small recovery gains: A case study at Raglan concentratror. Minerals


Lotter, N.O., Kormos, L.J., Oliveira, J., Fragomeni, D. and Whiteman, E., 2011. Modern process

Mineralogy: Two case studies. Minerals Engineering 24, 638 – 650.

89

Lotter, N.O., 2011. Modern Process Mineralogy: An integrated multi-discipline approach to

flowsheeting. Minerals Engineering 24, 1229-1237.

Malysiak, V., Shackleton, N.J. and O’Connor, C.T., 2004. An investigation into the floatability

of a pentlandite-pyroxene system. International Journal of Mineral processing 74, 251-262.

McLaren, C.H., and De Villiers, J.P.R., 1982. The Platinum-Group Chemistry and Mineralogy of

the UG-2 Chromitite Layer of the Bushveld Complex. Economic Geology 77, 1348-1366.

Miller, J.D., Li, J., Davidtz, J.C and Vos, F., 2005. A review of pyrrhotite flotation chemistry in

the processing of PGM ores. Minerals Engineering 18, 855-865.

Naldrett, A.J., Gasparrini, E.C., Barnes, S.J., Von Gruenewaldt, G; and Sharpe, M.R., 1986. The

Upper Critical Zone of the Bushveld Complex and the Origins of Merensky-type Ores.

Economic Geology 81, 1105-1117.

Neethling, S.J., and Cilliers, J.J., 2008. Predicting and correcting grade-recovery curves:

Theoretical aspects. International Journal of Mineral Processing 89, 17-22.

Newell, A.J.H., Bradshaw, D.J. and Harris, P.J., 2006. The effect of heavy oxidation upon

flotation and potential remedies for Merensky type sulphides. Minerals Engineering 19, 675-686.

Olubambi, P.A., Ndlovu, S., Potgieter, J.H., and Borode, J.O., 2006. Influence of applied

mineralogy in developing an optimal hydrometallurgical processing route for complex sulphide

ores. Mineral Processing and extractive Metallurgy Review 27, 143-158.

Olubambi, P.A., Ndlovu, S., Potgieter, J.H., and Borode, J.O., 2008. Mineralogical

characterization of Ishiagu (Nigeria) complex sulphide ore. International Journal of Mineral


90

Pearse, M.J., 2005. An overview of the use of chemical reagents in mineral processing. Minerals


Penberthy, C.J., Oosthuyzen, E.J and Merkle, R.K.W., 2000. The recovery of platinum group

elements from the UG2 Chromitite, Bushveld Complex – a mineralogical perspective.

Mineralogy and Petrology 68, 213-222.

Petruk, W., 2000. Applied Mineralogy in the Mining Industry. Elsevier Science B.V,

Amsterdam, The Netherlands, First edition, pp24-25, 48-49.

Philander, C and Rozendaal, A., 2011. The contributions of geometallurgy to recovery of

lithified heavy mineral resources at the Namakwa Sands mine, West Coast of South Africa.

Minerals Engineering 24, 1357-1364.

Pietrobon, M.C., Grano, S.R., Sobieraj, S and Ralston, J., 1997. Recovery mechanisms for

pentlandite and MgO-bearing minerals in nickel ores from Western Australia. Minerals

Engineering 10, 8, 775-7786.

Runge, K., 2010. Laboratory flotation testing - An essential tool for characterization. In:

Flotation plant optimization: A metallurgical guide to identifying and solving problems in

flotation plants. Australasian Institute of Mining and Metallurgy, Spectrum Series (Carlton, Vic)

16, 155-173.

Scoon, R.N and Mitchell, A.A., 2004. The platiniferous dunite pipes in the eastern limb of the

Bushveld Complex: Review and comparison with unmineralized discordant ultramafic bodies.

South African Journal of Geology 107, 505-520

Scoon, R.N and Mitchell, A.A, 2009. Discovery and Geology of the Platinum Group Element

Deposits of the Bushveld Complex, South Africa. SEG NEWSLETTER 78, 13-19

91

Senior, G.D., Trahar, W.J., and Guy, P.J., 1995. The selective flotation of pentlandite from a

nickel ore. International Journal of mineral Processing 43, 209-234.

Shackleton, N.J., Malysiak, V., and O’Connor, C.T., 2007a. Surface characteristics and flotation

behaviour of platinum and palladium tellurides. Minerals Engineering 20, 1232-1245.

Shackleton, N.J., Malysiak, V., and O’Connor, C.T., 2007b. Surface characteristics and flotation

behaviour of platinum and palladium arsenides. International Journal of Mineral Processing 85,

25-40.

Schouwstra, R.P., and Kinloch, E.D., 2000. A short geological Review of the Bushveld

Complex. Platinum Metals Rev., 44, 1, 33 – 39.

Schouwstra, R., de Vaux, D., Hey, p., Malysiak, V., Shackleton, N., and Bramdeo, S., 2010.

Understanding Gamsberg – A geometallurgical study of a large stratiform zinc deposit. Minerals


Sizgoric, N.B and Alcock, R.A, 1981. Quantitative Mineralogy as an aid to Beneficiation of

Nickel Sulphide ore. In: Hausen, D.M and Park, W.C (editors), Process Mineralogy: Extractive

Metallurgy, Exploration, Energy Resources. Proceedings of Symposium, February 22-26, pp

419-428.

Smith, L.K., Senior, G.D., Bruckard, W.J., and Davey, K.J., 2011.The flotation of millerite –A

single mineral study. International Journal of Mineral Processing 99, 27 – 31.

Sutherland, D., 2007. Estimation of mineral grain size using automated mineralogy. Minerals


Van Tonder, E., Deglon, D.A and Napier-Munn, T.J., 2010. The effect of ore blends on mineral

processing of platinum ores. Minerals Engineering 23, 621-626.

92

Vermaak, C.F., and Hendriks, L.P., 1976. A review of the Mineralogy of the Merensky Reef,

with Specific Reference to New Data on the Precious Metal Mineralogy. Economic Geology 71,

1244-1269

Vermaak, C.F., 1995. The Platinum Group Metals: A global Perspective. Mintek, Randburg,

247pp

Viljoen, F., Knoper, M., Rajesh, H., Rose, D and Greef, T., 2012. Application of a field emission

mineral liberation analyser to the in-situ study of platinum group element mineralization in the

Merensky reef of the Bushveld Complex, South Africa. Proceedings of the 10th International

Congress for Applied Mineralogy (ICAM), 757-764, Springer Berlin Heidelber.

Von Gruenewaldt, G., Martin, R.S; and Hatton, C.J., 1985. The Bushveld Complex: Introduction

and Review. Society of Economic Geologists, Inc. Economic Geology 80, 803-812.

Von Gruenewaldt, G; Hatton, C.J., Merkle, R.K.W., and Gain, S.B., 1986. Platinum Group

Element-Chromitite Associations in the Bushveld Complex. Economic Geology 81, 1067-1079.

Wiese, J., Harris, P., and Bradshaw, D., 2005. The influence of the reagent suite on the flotation

of ores from the Merensky reef. Minerals Engineering 18, 189-198.

Wiese, J., Harris, P and Bradshaw, D., 2005. Investigation of the role and interactions of a

dithiophosphate collector in the flotation of sulphides from Merensky reef. Minerals Engineering

18, 791-800.

Wiese, J., Harris, P and Bradshaw, D., 2006. The role of the reagent suite in optimising

pentlandite recoveries from the Merensky reef. Minerals Engineering 19, 1290-1300.

93

Wiese, J.G., Becker, M., Bradshaw, D.J., and Harris, P.J., 2007. Interpreting the role of reagents

in the flotation of platinum-bearing Merensky ores.The Journal of the South African Institute of

Mining and Metallurgy 107, 29-36.

Wiese, J., Harris, P and Bradshaw, D., 2011. The effect of the reagent suite on froth stability in

laboratory scale batch flotation tests. Minerals Engineering 24, 995-1003.

Wills, B.A., 1997. Mineral Processing Technology: An introduction to the practical aspects of

ore treatment and mineral recovery. 6th edition, pp110, 258-341. Butterworth-heinemann

Xiao, Z., and Laplante, A.R., 2004. Characterization and recovering the platinum group minerals

– a review. Minerals Engineering 17, 961-979.

Xiao, Z., Laplante, A.R., and Finch, J.A., 2009. Quantifying the content of gravity recoverable

platinum group minerals in ore samples. Minerals Engineering 22, 304-310

94

APPENDIX 1: METHODS

A1.0 INTRODUCTION

This chapter briefly describes the various techniques employed and experimental procedures

carried out to perform a particular activity. These include sampling, crushing, representative

sample splitting, grain mounts preparation, and grain mounts carbon coating, milling, flotation,

screening and chemical assaying.

A1.1 CHANNEL SAMPLING

The following are the source locations of channel samples of the reef facies used in this study,

from Lonmin Marikana Platinum Mine:

BK facies: 15 MW 6E1 RSE

RPM facies: 15 MW 22 RSE

WP facies: K3 # 13 ME 69 RSE

UG2: 4B UG2 UNDERGROUND

The reef ores were channel sampled using the panel sampling method through the reef, in which

10cm long (and 5cm wide) interval subsamples were collected from each of the accessible facies

types at Marikana, that is: RPM facies, WPfacies and BK facies. One set of samples per facies

type was collected and one sample of the UG2 chromitite layer was also sampled in the same

manner.

Each of the reef facies subsamples was weighed and geologically logged (Figure 3.1). Various

techniques were used to investigate the mineralogical and textural characteristics of the ores.

These included chemical analysis (fire assay), scanning electron microscope based mineral

liberation analysis (Gu, 2003 and Fandrich et al., 2007), milling and bench-top flotation. The

mineral liberation analysis technique was used to quantify modal mineralogy, particle and grain

size distribution, mineral associations and mineral locking, and mineral liberation by surface area

and by composition.

95

All sample preparations and analyses were done at the University of Johannesburg’s central

analytical facility, Spectrau. Chemical analysis was done at Genalysis Laboratories in

Johannesburg; milling and bench-top flotation were carried out at the University of

Johannesburg’s Doornfontein Campus, Extractive Metallurgy Department.

A1.2 CRUSHING

Jaw-crushing of the samples was done to achieve 100% passing 2mm. Each subsample was

stage-crushed and sieved (using a 2mm sieve) to avoid excessive fine grains or powders.

Intensive cleaning of the jaw crusher and sieve was done using water and a hot air dryer. Final

rinsing was done using acetone and wiping with paper towel to eliminate cross-contamination

between samples.

A1.3 REPRESENTATIVE SAMPLE SPLITTING

Crushed representative samples were split using riffler splitters until three fractions of about 6g

each were obtained. These were made into blocks for modal abundances determination.

A1.4 GRAIN MOUNTS PREPARATION

The 6g subsample aliquots were used to make 30mm flat rock mounts. The process of making

grain mounts involved initially lubricating the inside of the 30mm mounting cups using Vaseline

jelly for easy removal of the mounts once dried. A hardener-resin mixture was then made from 1

part hardener and 7 parts resin, thoroughly mixed and placed in an oven at 50 degrees Celsius for

2 to 5 minutes to expel air bubbles.

The air-free hardener-resin mixture from the oven was poured into the 30mm mounting cups,

followed by the crushed sample aliquots and then stirred to avoid air spaces or bubbles

forming/setting at the bottom with the sample grains. Sample labels were then placed at the top

facing upwards but covered with the hardener-resin mixture, making sure sample names were

96

clearly visible. The labelled samples were placed in a pressure vessel machine for 24 hours to

remove air bubbles and drying the samples.

After drying, the sample mounts were ground at the base to expose the mineral grains, followed

by stage polishing using a 6micron, 3micron, 1micron and 0.25micron diamond paste (DP-stick

P) to create final polished rock mounts blocks for mineralogical analysis.

A1.5 CARBON COATING

The polished grain mounts were then coated with a 25 nm thick layer of carbon (Kerrick et al.,

1973 and Andersen et al., 2009; Jasieniak and Smart., 2010) to create a conductive surface to

prevent excess charge build up in the Mineral Liberation analyser and Electron Microprobe

during analysis.

Coating was done using a 230V Agar Turbo Carbon Coater, model 208C, at a setting of 4.5V

and operating on a manual mode. A brass disk standard was employed, which produced blue

interference colours when a good enough carbon layer was achieved. A total of 240 grain mounts

were carbon coated for onward analysis using an FEI 600F Mineral Liberation Analyser (MLA)

(Appendix 1), housed at Spectrau.

A1.6 MILLING

A 210mm mill was used, in conjunction with 6x25mm, 9x20mm and 6x16mm mill rods (Wiese

et al., 2005), to mill 1kg aliquots of representatively riffled splits of the BK, RPM, WPand UG2

facies (respectively labelled as BK_A, RPM_A, WP_A and UG2_A) in order to produce milling

curves for the respective facies for determining the exact times required to mill a 1kg sample

aliquot to 60% mass passing 75µm sieve. To clean the mill, the above configuration of rods was

placed in the mill, followed by adding 1kg of pure quartz pebbles (Senior et al., 1995) and then

520ml of water was added into the mill contents to produce 66% solid slurry by weight.

97

The mill was then closed securely, placed on rollers set at a rotation speed of 61 RPM and then

milled for 10 minutes. The mill was then opened and its contents discarded, followed by

thorough washing with water. Then the above procedure was repeated using the 1kg sample

aliquot for each facies under study for the times shown in Appendix 4, Table A4.1, Table A4.2,

Table A4.3 and Table A4.4 for the BK, RPM, WP and UG2 ore facies respectively.

After milling the mill was taken and placed onto a stand in an upright position, opened, and the

slurry was washed off from the lid using a wash bottle with minimal water into the mill. The rods

were taken from the mill and the slurry washed back into the mill with minimal water. The rod

mill contents were then tipped into a clean bucket labelled with a sample number (such as

BK_A), and all the remaining slurry from the mill was washed into the bucket with minimal

water using a wash bottle.

A clean bucket labelled -75µm was placed under a 75µm sieve to receive particle slurry

consisting of -75µm sizes. The slurry was vigorously stirred and poured through the 75µm sieve

and also using minimal water to help the slurry through the sieve. The +75µm sludge was

transferred into another clean bucket labelled +75µm.

The -75µm and +75µm fractions were filtered separately and separately dried in an oven at low

temperature (170oC). The dried +75µm fraction lumps were broken down using steel rollers and

then sieved for 20 minutes through a sieve stack consisting of 300µm, 150µm and 75µm aperture

sieves and a collecting pan at the bottom. The -75µm fraction was added to the other -75µm

portion separately filtered and dried earlier. The +300µm, +150µm and +75µm fractions were

also combined. Both the combined -75µm and +75µm portions were separately weighed and the

masses recorded, including the % - passing 75µm as shown in the Table A4.1, Appendix 4 for

the BK_A sample.

The same procedure was repeated for the varied times shown in the table, using the same rod

mill, same rods configuration, same amount of water and mill rotation speed after combining the

weighed -75µm and +75µm portions of the sample.

98

A1.7 MINERAL LIBERATION ANALYSIS

A Mineral Liberation Analyser (MLA) (model 600F), an automated mineral analysis platform,

was used to perform mineralogical analyses on 30mm polished mount blocks of crushed ore

feeds (-2mm), milled ore feeds (60wt% passing 75µm) and concentrate (-75+38µm size fraction)

samples of the BK, RPM, WP and UG2 facies.

The MLA was operated under the following conditions during analyses:

Operating voltage………………………25kV

Beam spot………………………………X6

Magnification…………………………...450X, for -75+38µm sieved fraction (SPL-GXMAP)

and the -2mm crushed (XMOD) sample analyses.

Two operating modes, XMOD and GXMAP procedures, were used (Gu, 2003; Fandrich et al.,

2007.

XMOD was used to determine bulk modal mineralogy data (mineral modal abundance in wt.%)

of the various minerals in the samples, resulting in the mineral data presented in Table 3.1,

Figures 3.2-3.5 and Appendix 2, Table A2.1-A2.4).

SPL-GXMAP measurement mode was used to determine sulphides particles in the -75+38µm

sieved size fraction of the BK, RPM, WPand UG2 concentrates, in order to get more statistically

representative particles in each concentrate block for further analysis and also for particle size

distribution study (Sutherland, 2007).

A1.8 FLOTATION PROCEDURE

1kg representative riffled split aliquot of each of the BK, RPM, WP and UG2 facies was placed

in the exact same rod mill as was used for the mill testing (210mm inside diameter), with the

same rods configuration (6x25mm rods, 9x20mm rods and 6x16mm rods) (Wiese at al., 2005)

99

and the same milling speed (61 revolutions per minute). 500 ml of water was added into the mill

to give pulp solids content of 66%. The BK, RPM, WPand UG2 facies were milled for 41, 37, 33

and 31 minutes respectively according to the previous mill test results (Chapter 4) to obtain

grinds of 60%-75µm passing.

After the milling was completed the slurry from the rods and mill was washed carefully into a

basin and then transferred into a 2.5 litre flotation cell ( also known as a 1kg cell) on the Denver

flotation machine. A total water volume of 1.86 litres to 1kg sample was used, resulting in solids

content of 35%, giving a froth height of 2cm and scrapping depth of 0.5cm above the slurry level

after the impeller and armature of the Denver flotation machine were lowered into the cell. The

flotation machine was started at a speed of 1200 revolutions per minute (Senior et al., 1995;

Wiese et al., 2006), with air closed, and a 160ml volume of sample slurry was scooped out,

filtered and dried for mineral liberation study. Water was then added to the cell to make up for

the volume scooped out and to maintain the 2.5cm froth height in the cell.

Table A1.1 Reagent suite addition and conditioning used in the flotation rate tests

Reagent Solution

Strength/%

Conc Active

Ingredient/%

Required

Dosage(g/t)

Volume to

add to cell

(1kg ore)/ml

Conditioning

Time/minutes

Activator:CuSO2 1 100 40 4 5

Collectors:SIBX

Senkol 5

1

1

90

50

37.5

37.5

4.17

7.50

2

Add with

SIBX

Depressant:

Sendep

1 100 100 10 2

Frother:

Senfroth XP 200

1 100 40 4 1

100

The flotation reagents suite used in these testworks were supplied by Senmin and their additions

were based on Wiese et al (2005), adjusted for Senmin reagents suite, as summarized in the table

above.

Still with the flotation machine running, air closed, 4ml CuSO2 (Becker et al., 2010; Miller et al.,

2005; Schouwstra et al., 2010; Senior et al., 1995 and references therein) was added to the

swirling pulp and conditioned for 5 minutes, followed by adding simultaneously 4.17 ml of

SIBX and Senkol 5 phosphate collectors (Becker et al., 2005; Becker et al., 2006; Becker et al.,

2010; Corin et al., 2010; Miller et al., 2005; Wiese et al., 2005, 2006 and 2011; Pearse., 2005

and references therein). Agitation was continued for further 2 minutes and then 10 ml of Sendep

30D (depressant) added and conditioned for 2 minutes (Allison et al., 2011; Senior et al., 1995

and references therein). 1ml of frother Dow 200 was added and then conditioned for 1 minute.

Then air flow was opened and controlled manually to keep the froth height constant at 2cm and

scrapping depth at 0.5cm; and immediately a timer was also started. Concentrate collection was

done by scrapping the froth into a collecting pan every 15 seconds. Concentrate 1 was collected

every 15 seconds for the first 2 minutes. Concentrate 2 was collected for every 15 seconds for the

next 4 minutes. Concentrate 3 was collected every 15 seconds for a further 6 minutes.

The airflow was then stopped and 2ml of frother Dow 200 was added. Then the airflow was

opened again and concentrate 4 was collected every 15 seconds for 8 minutes. The flotation

process was then stopped. Duplicate flotation test rates were done for each facies.

A1.9 CHEMICAL ANALYSIS

At least 75g representative aliquots of each crushed (-2mm) channel sample block (unit),

composite milled ore, milled screened ore feed (+106µm, +75µm, +25µm) and tailings, and at

least 8g of flotation concentrates from each timed flotation test run from the BK, the RPM, the

WP facies and the UG2 reef were submitted to Genalysis Laboratories Ltd (South Africa) for

assays.

101

All representative samples from the BK, RPM, WP and UG2 ore facies were analysed for base

metals ( Co, Cr, Cu, Ni), sulfur and the platinum group elements (Ir, Os, Pd, Pt, Rh, and Ru) and

Au.

Co, Cr, Cu, Ni and S were recovered by the sodium peroxide Fusion Zirconium Crucible Fusion

process and hydrochloric acid dissolution of the melt from 75g crushed (-2mm) and milled

(60wt% passing 75µm) ore feeds and tailings, and 8g concentrates. The base metals and sulfur

were then quantitatively analysed by the Inductively Couple Plasma Optical Emission

Spectrometry (ICP-OES) Finish Technique (Lenahan et al., 1986).

Platinum group elements (Pt, Pd, Rh, Ru, Ir and Os) and Au were quantitatively recovered by the

Nickel Sulphide Collection Fire Assay procedure from the 25g crushed sample (-2mm) and

milled (60wt% passing 75µm) ore feeds aliquots and milled screened ore feed (+106µm, +75µm,

+25µm) and tailings. The recovered PGE were quantitatively analysed by the Inductively

Coupled Plasma Mass Spectrometry (ICP-MS) technique (Lenahan et al., 1986).

Only Co, Cr, Cu, Ni and S were analysed in the concentrates due to low sample volume.

PGE+Au were not analysed in the concentrates also due to inadequate sample amounts which

were available for assays.

The detection limits that were used for assays are given in the table below.

102

Table A1.2 Analytical detection limits used for assays in this study

Element Detection limits

Cr 50ppm

Cu, Ni, Co 20ppm

S 0.05%

Au 5ppb

Ir, Os, Pd, Pt, Ru 2ppb

Rh 1ppb

The assay results of the ore feeds, concentrates and tailings, including sized fractions, for the

four facies are shown in Appendix 3, Tables A3.1-A3.4 and Appendix 4, Tables A4.3-A4.4 and

Table A4.13. Assays of each timed flotation concentrate and tailings were subsequently

reconciled with the head grades of the respective reef facies (Ekmekci et al., 2005).

103

APPENDIX 2: MINERALOGICAL DATA

Appendix 2 contains tables of mineral modal abundances for BK, RPM, WP and UG2 facies ores

referred to in various parts of this study.

Table A2.1 Mineral modal abundance (wt.%) variations of samples of the BK facies of the

Merensky reef (-2mm crushed ore sample) (from top to bottom)

Sample

ID

Wt%

Anor

Wt%

Aug

Wt%

Chr

Wt%

Enst

Wt%

Epid

Wt%

Kfsp

Wt%

Phlg

Wt%

Qtz

Wt%

trem

Wt%

Serp

Wt%

Ccp

Wt%

Pn

Wt%

Po

Wt%

Other

BK21-top 75.55 4.50 0.01 16.65 0.11 0.19 0.78 0.23 0.12 0.02 0.09 0.08 0.09 1.58

BK20 70.67 5.19 0.01 20.63 0.08 0.27 0.69 0.14 0.17 0.06 0.15 0.14 0.22 1.57

BK19 46.79 7.18 0.02 43.05 0.03 0.08 0.43 0.04 0.33 0.16 0.18 0.20 0.21 1.30

BK18 25.18 8.39 0.03 62.91 0.02 0.03 0.25 0.02 0.61 0.29 0.22 0.26 0.28 1.51

BK17 19.27 6.74 0.05 68.52 0.01 0.01 0.35 0.01 0.59 0.33 0.58 0.84 1.34 1.36

BK16 10.48 7.43 0.02 78.08 0.00 0.01 0.21 0.01 0.65 0.39 0.42 0.29 0.54 1.48

BK15 12.06 5.12 0.06 79.59 0.00 0.01 0.19 0.00 0.49 0.26 0.23 0.40 0.37 1.22

BK14 7.00 21.60 7.30 57.72 0.00 0.01 0.54 0.00 0.60 0.20 0.68 1.22 1.29 1.85

BK13 24.43 11.31 0.44 57.11 0.03 0.14 0.60 0.03 0.42 0.16 0.81 1.33 1.52 1.66

BK12 15.34 12.64 0.13 63.50 0.13 0.22 1.09 0.29 0.53 0.10 1.19 2.22 1.32 1.30

BK11 10.93 10.14 0.36 74.72 0.06 0.08 0.63 0.16 0.52 0.19 0.24 0.65 0.20 1.11

BK10 5.40 15.56 1.33 73.00 0.08 0.11 0.94 0.23 0.66 0.37 0.18 0.34 0.14 1.66

BK9 5.85 9.61 0.06 77.86 0.08 0.06 0.84 0.30 0.97 0.35 0.45 1.18 1.33 1.06

BK8 14.74 19.74 0.13 59.89 0.09 0.08 0.54 0.15 0.63 0.20 0.34 0.78 1.23 1.45

BK7 14.40 8.24 0.03 73.16 0.33 0.07 0.41 0.08 0.63 0.25 0.31 0.40 0.46 1.23

BK6 21.09 7.12 0.21 67.06 0.24 0.07 0.63 0.36 1.05 0.40 0.10 0.49 0.11 1.08

BK5 7.56 5.06 0.19 82.42 0.20 0.04 1.34 0.24 1.17 0.44 0.12 0.20 0.03 0.99

BK4 22.00 4.10 0.16 68.49 0.10 0.07 0.49 0.11 0.86 0.33 0.90 0.93 0.38 1.09

BK3 34.10 2.23 3.22 48.78 0.01 0.05 0.29 0.02 0.44 0.64 0.59 4.53 3.89 1.19

BK2 94.60 3.08 0.00 0.68 0.08 0.21 0.11 0.02 0.02 0.00 0.00 0.00 0.01 1.19

BK1(btm) 92.65 2.80 0.00 2.50 0.06 0.15 0.12 0.04 0.09 0.01 0.01 0.01 0.00 1.56

Note that the rows in bold indicate chromite stringer positions across the reef facies.

104

Table A2.2 Mineral modal abundances (wt.%) variations of samples of the RPM facies of the


Sample

ID

Wt%

Anor

Wt%

Aug

Wt%

Chr

Wt%

Enst

Wt%

Epid

Wt%

Kfsp

Wt%

Phlg

Wt%

Qtz

Wt%

trem

Wt%

Serp

Wt%

Ccp

Wt%

Pn

Wt%

Po

Wt%

Other

RPM25 11.04 6.03 4.45 67.97 0.13 0.06 1.00 0.11 2.00 1.99 0.64 1.17 1.84 1.57

RPM24 14.94 4.76 0.33 73.63 0.05 0.03 0.42 0.01 2.01 2.01 0.24 0.2 0.21 1.16

RPM23 3.82 12.40 0.97 68.04 0.13 0.01 2.48 0.83 4.66 3.60 0.41 0.62 0.46 1.56

RPM22 9.93 9.59 0.47 66.77 1.05 0.09 2.21 1.32 3.41 3.37 0.06 0.09 0.11 1.54

RPM21 15.90 3.85 0.38 65.03 1.87 0.24 3.56 1.66 2.58 3.07 0.03 0.03 0.02 1.79

RPM20 17.76 3.71 0.46 69.29 0.43 0.27 1.06 0.47 1.70 3.25 0.04 0.03 0.02 1.50

RPM19 9.30 7.73 0.48 74.49 0.60 0.41 1.79 0.72 1.10 1.66 0.03 0.04 0.02 1.63

RPM18 13.92 6.50 0.04 74.90 0.19 0.06 0.48 0.22 0.67 1.74 0.04 0.06 0.00 1.18

RPM17 10.12 10.81 0.05 73.49 0.38 0.13 0.82 0.31 0.70 1.61 0.02 0.04 0.02 1.51

RPM16 14.52 9.00 0.03 72.13 0.11 0.03 0.32 0.11 0.68 1.59 0.03 0.04 0.02 1.40

RPM15 9.44 8.77 0.05 77.70 0.11 0.03 0.51 0.09 0.68 1.11 0.03 0.03 0.02 1.43

RPM14 10.81 10.07 0.04 74.94 0.05 0.01 0.53 0.05 0.72 1.31 0.03 0.03 0.02 1.39

RPM13 8.74 4.09 5.52 75.42 0.26 0.21 1.32 0.39 1.09 0.87 0.08 0.11 0.11 1.79

RPM12 4.19 3.16 0.32 80.99 0.98 0.50 2.64 1.70 2.44 1.14 0.01 0.04 0.01 1.87

RPM11 23.91 0.85 1.19 67.16 0.19 0.17 0.84 0.28 0.57 2.61 0.01 0.16 0.05 2.01

RPM10 15.76 5.26 0.89 65.24 0.77 0.58 1.79 0.68 1.72 4.22 0.01 0.08 0.12 2.87

RPM9 23.17 0.92 0.55 58.52 0.19 0.19 0.84 0.07 0.71 10.57 0.02 0.20 0.56 3.50

RPM8 11.09 6.64 1.00 69.46 0.13 0.07 0.91 0.22 0.86 7.17 0.01 0.06 0.23 2.17

RPM7 8.15 17.36 0.76 65.21 0.22 0.16 1.19 0.19 1.19 3.76 0.02 0.05 0.16 1.59

RPM6 10.26 8.31 2.25 65.08 0.17 0.09 1.01 0.28 1.05 8.82 0.03 0.14 0.43 2.07

RPM5 19.29 3.46 1.21 54.58 0.04 0.12 1.24 0.00 0.49 15.09 0.02 0.16 0.68 3.60

RPM4 21.55 2.20 1.78 51.56 0.02 0.05 0.67 0.00 0.40 17.79 0.01 0.07 0.69 3.20

RPM3 20.63 3.85 17.94 46.10 0.09 0.03 0.83 0.13 1.17 6.08 0.09 0.19 0.34 2.51

RPM2 92.48 3.35 0.00 1.75 0.16 0.22 0.13 0.03 0.05 0.02 0.01 0.01 0.01 1.77

RPM1 94.14 2.15 0.04 1.51 0.18 0.06 0.09 0.05 0.07 0.03 0.01 0.00 0.00 1.43


105

Table A2.3 Mineral modal abundances (wt-%) variations of samples of the WPfacies of the


sample

ID

Wt%

Anor

Wt%

Aug

Wt%

Chr

Wt%

Enst

Wt%

Epid

Wt%

Kfsp

Wt%

Phlg

Wt%

Qtz

Wt%

trem

Wt%

Serp

Wt%

Ccp

Wt%

Pn

Wt%

Po

Wt%

Other

WP25 25.43 6.48 0.10 62.43 0.07 0.10 0.47 0.17 0.66 0.50 0.32 0.78 0.89 1.60

WP24 14.13 6.65 0.03 71.19 0.13 0.19 0.51 0.29 0.75 0.58 0.62 1.05 1.78 2.10

WP23 13.29 6.81 0.04 70.70 0.10 0.14 0.98 0.18 1.24 1.31 0.51 1.15 1.79 1.75

WP22 15.60 4.29 0.09 72.07 0.16 0.11 0.49 0.12 1.26 1.87 0.56 0.71 0.82 1.86

WP21 12.83 4.06 1.48 67.88 0.17 0.02 2.04 1.39 1.53 2.19 0.56 1.43 1.88 2.37

WP20 10.26 6.42 1.97 70.34 0.13 0.02 2.08 0.71 1.40 1.25 0.87 0.93 1.54 2.09

WP19 8.14 6.42 7.95 66.99 0.06 0.02 2.24 0.41 1.30 0.88 0.72 1.61 1.82 1.43

WP18 11.71 7.19 0.32 66.62 1.43 1.75 3.03 2.86 1.12 1.29 0.15 0.55 0.52 1.46

WP17 9.40 11.24 0.28 71.18 0.53 0.24 1.72 0.88 1.18 0.58 0.25 0.34 0.48 1.71

RPM16 4.39 10.02 0.23 69.15 1.31 2.19 4.59 3.36 0.96 0.88 0.05 0.05 0.08 2.75

WP15 16.99 3.10 0.31 76.27 0.07 0.03 0.29 0.08 0.89 0.55 0.04 0.02 0.00 1.37

WP14 15.94 3.30 0.30 76.39 0.02 0.01 0.42 0.01 1.33 0.70 0.10 0.11 0.01 1.36

WP13 14.14 4.00 0.39 77.28 0.13 0.02 0.71 0.19 1.26 0.82 0.02 0.02 0.01 1.03

WP12 11.72 5.46 0.40 74.18 0.78 0.15 2.04 1.56 1.06 0.99 0.23 0.08 0.09 1.25

WP11 18.34 6.62 0.10 71.29 0.07 0.07 0.42 0.29 0.93 0.84 0.03 0.01 0.00 0.98

WP10 13.99 5.30 0.19 75.92 0.15 0.02 0.77 0.35 0.97 1.01 0.03 0.03 0.00 1.28

WP9 9.60 11.35 0.12 73.72 0.46 0.19 0.95 0.75 1.15 0.68 0.01 0.01 0.01 1.00

WP8 13.24 3.54 0.28 77.37 0.31 0.13 0.77 0.48 1.18 1.30 0.01 0.02 0.01 1.37

WP7 15.85 4.93 0.15 73.73 0.09 0.01 0.61 0.05 1.53 1.56 0.02 0.03 0.00 1.44

WP6 12.47 9.62 0.20 71.71 0.10 0.01 1.20 0.17 1.44 1.26 0.01 0.05 0.01 1.74

WP5 15.66 5.12 0.59 72.79 0.29 0.13 0.90 0.43 0.80 1.09 0.02 0.03 0.01 2.14

WP4 9.91 9.78 0.34 72.53 0.66 0.12 1.78 1.01 1.26 0.84 0.02 0.04 0.01 1.70

WP3 11.63 9.56 0.44 71.55 0.38 0.11 1.57 0.58 1.21 1.48 0.02 0.03 0.02 1.44

WP2 15.86 4.75 0.47 72.45 0.21 0.23 0.91 0.34 1.28 1.56 0.02 0.06 0.06 1.79

WP1 30.71 1.95 10.50 53.35 0.05 0.08 0.35 0.06 0.76 0.76 0.07 0.12 0.04 1.20


106

Table A2.4 Mineral modal abundances (wt-%) variations of samples of the UG2 Chromitite

facies (-2mm crushed ore sample) (from top to bottom)

sample ID

Wt%

Anor

Wt%

Aug

Wt%

Chr

Wt%

Enst

Wt%

Epid

Wt%

Kfsp

Wt%

Phlg

Wt%

Qtz

Wt%

trem

Wt%

Serp

Wt%

Ccp

Wt%

Pn

Wt%

Po

Wt%

Other

UG286 9.05 3.05 51.49 30.72 0.08 0.01 2.13 0.03 1.68 0.13 0.02 0.03 0.01 1.57

UG285 10.76 4.56 9.73 68.61 0.25 0.73 2.34 0.42 0.73 0.23 0.01 0.03 0.01 1.59

UG284 2.65 0.07 87.86 6.65 0.02 0.15 0.80 0.03 0.07 0.12 0.01 0.02 0.00 1.55

UG283 4.78 0.04 89.52 2.13 0.12 0.08 1.26 0.02 0.07 0.10 0.01 0.02 0.00 1.85

UG282 7.22 0.06 85.21 5.11 0.01 0.12 0.56 0.01 0.07 0.06 0.01 0.02 0.00 1.54

UG281 17.91 0.05 71.46 8.41 0.01 0.06 0.35 0.00 0.08 0.14 0.01 0.03 0.00 1.51

UG280 18.53 0.43 69.46 6.82 0.10 0.02 1.07 0.01 0.43 0.51 0.01 0.02 0.00 2.58

UG279 28.97 0.36 60.08 6.50 0.09 0.09 1.37 0.01 0.27 0.21 0.01 0.05 0.00 2.00

UG278 91.92 1.87 0.71 3.41 0.07 0.15 0.28 0.04 0.11 0.03 0.01 0.00 0.00 1.42

Full names of minerals abbreviations used in Tables A2.1-A2.4 are given below:

Anor…..anorthite, Aug…..augite, Chr….chromite, Enst….enstatite, Epid…epidote,

Kfsp…K-feldspar, Phlg….phlogopite, Qtz….quartz, Trem….tremolite, Serp….serpentine,

Ccp….chalcopyrite, Pn….pentlandite, Po….pyrrhotite,

107

APPENDIX 3: GEOCHEMICAL ANALYSES Table A3.1 Distribution of Cr, (ppm), S (wt.%), and 6PGE (ppb) in the BK facies of Merensky

Reef; abundant Cr, S, and PGE correlate with the position of chromitite stringers (-2mm

crushed ore sample) (from top to bottom) Element Cr Cu Ni S Au Ir Os Pd Pt Rh Ru

Units ppm ppm ppm Wt.% ppb ppb ppb ppb ppb ppb ppb

BK21 407 203 465 0.11 28 <2 <2 11 11 <1 <2

BK20 485 277 597 0.09 40 <2 <2 12 28 <1 <2

BK19 964 396 995 0.21 59 <2 <2 19 32 <1 5

BK18 1607 469 1242 0.16 74 <2 <2 27 37 2 6

BK17 1703 973 2361 0.45 163 4 <2 94 208 11 33

BK16 2112 613 1567 0.3 119 9 <2 203 385 26 65

BK15 2103 513 1562 0.28 148 15 <2 261 820 39 104

BK14 18194 1486 2971 0.57 240 202 175 1473 5711 564 1264

BK13 2664 1167 2286 0.44 425 118 83 1635 2845 351 772

BK12 2880 1756 3141 0.62 484 132 91 2789 7374 421 748

BK11 2490 460 1608 0.3 124 50 15 842 1973 145 333

BK10 3200 292 1180 0.17 86 28 <2 598 1489 78 194

BK9 2325 595 2304 0.38 293 161 126 3680 6683 506 1046

BK8 2101 471 1523 0.19 146 88 69 2299 9250 348 603

BK7 2000 771 2648 0.43 253 685 591 12851 11730 1890 3815

BK6 1940 142 761 0.1 35 43 4 1115 1296 94 280

BK5 2130 111 967 0.07 32 2 <2 131 132 9 26

BK4 1713 726 2428 0.44 1232 89 70 7157 17959 290 529

BK3 7292 1177 9602 1.43 202 460 440 14438 7354 1296 2677

BK2 87 20 72 <0.05 <5 <2 <2 4 8 3 6

BK1 99 22 104 <0.05 12 <2 <2 <2 <2 <1 2


108

Table A3.2 Distribution of Cr, (ppm), S (wt.%), and 6PGE (ppb) in the RPM facies of the

Merensky Reef (-2mm crushed ore sample) (from top to bottom) Element Cr Cu Ni S Au Ir Os Pd Pt Rh Ru


RPM25 10457 1430 3796 0.79 846 175 118 4377 9752 514 869

RPM24 2688 615 1389 0.32 253 26 <2 744 2600 77 157

RPM23 4051 846 2275 0.44 430 39 <2 1740 3619 116 200

RPM22 2791 73 723 0.05 56 2 <2 66 160 15 37

RPM21 2988 81 666 <0.05 13 2 <2 15 123 11 36

RPM20 3345 96 624 <0.05 5 7 <2 40 266 22 56

RPM19 3187 81 730 <0.05 <5 14 <2 67 607 56 81

RPM18 2111 76 576 0.08 <5 <2 <2 <2 14 2 <2

RPM17 2153 44 568 <0.05 <5 <2 <2 <2 29 3 4

RPM16 2099 65 567 <0.05 <5 <2 <2 <2 132 10 3

RPM15 2262 62 611 0.05 <5 <2 <2 4 94 12 12

RPM14 2287 61 656 <0.05 <5 <2 <2 11 69 3 3

RPM13 10837 210 1192 0.15 33 250 175 869 7757 807 1435

RPM12 3234 28 930 0.08 7 <2 <2 72 52 3 5

RPM11 2227 30 1290 <0.05 <5 <2 <2 63 28 2 5

RPM10 2085 42 1143 <0.05 <5 <2 <2 25 6 <1 3

RPM9 1042 20 1283 0.08 <5 <2 <2 32 <2 <1 <2

RPM8 2396 29 1247 <0.05 <5 <2 <2 44 33 3 6

RPM7 2621 41 1218 0.06 10 <2 <2 36 28 2 5

RPM6 3199 66 1544 0.06 18 13 <2 116 320 38 121

RPM5 1716 38 1415 <0.05 13 <2 <2 54 48 5 8

RPM4 1650 24 1138 <0.05 9 2 <2 29 112 15 12

RPM3 24858 173 1216 0.09 55 117 71 891 2812 368 705

RPM2 132 39 99 <0.05 8 <2 <2 27 14 5 11

RPM1 201 22 64 <0.05 <5 <2 <2 9 20 2 3

Note that the rows in bold indicate chromite stringer positions across the reef facies

109

Table A3.3 Distribution of Cr (ppm), S (wt.%), and 6PGE (ppb) in the WP facies of Merensky

Reef; abundant Cr, S, PGE correlate with the position of chromitite (from top to bottom) Element Cr Cu Ni S Au Ir Os Pd Pt Rh Ru


WP25 1739 549 1652 0.31 106 <2 <2 52 135 5 14

WP24 2144 1071 2586 0.49 177 3 <2 90 205 8 27

WP23 2184 1152 2739 0.55 215 6 7 150 283 17 52

WP22 2257 994 2793 0.54 301 28 27 734 1443 79 202

WP21 6141 2436 5210 1.16 761 64 64 1583 4122 177 409

WP20 4864 1662 3339 0.72 509 67 59 2013 6214 191 335

WP19 15450 2212 4678 1 552 214 186 3325 10946 628 1377

WP18 2314 551 1662 0.3 263 29 <2 1028 1629 78 163

WP17 2559 507 1765 0.22 265 25 <2 906 1394 66 153

WP16 2837 140 1038 <0.05 38 3 <2 153 385 15 31

WP15 3466 144 809 <0.05 61 <2 <2 62 165 11 24

WP14 2904 145 857 0.11 81 6 <2 232 355 15 46

WP13 2998 68 628 <0.05 10 <2 <2 3 19 2 11

WP12 3062 331 1009 0.15 181 2 <2 423 385 15 27

WP11 2830 76 637 0.06 <5 <2 <2 <2 18 1 8

WP10 2388 58 607 0.1 6 <2 <2 15 44 4 20

WP9 2450 42 587 <0.05 7 <2 <2 4 31 3 11

WP8 3076 61 745 0.11 6 <2 <2 <2 37 2 12

WP7 2496 48 567 <0.05 6 2 <2 27 63 9 28

WP6 2579 48 617 0.07 12 <2 <2 10 35 3 17

WP5 3076 47 611 <0.05 <5 <2 <2 2 34 5 17

WP4 3632 64 0 <0.05 11 <2 <2 14 67 6 15

WP3 2848 57 640 <0.05 <5 3 <2 22 88 11 26

WP2 2817 100 793 0.1 41 4 <2 183 332 17 36

WP1 16645 134 919 0.1 <5 3 <2 16 132 15 17

Note that the rows in bold indicate chromite stringer positions across the reef facies

110

Table A3.4 Distribution of Cr, S (wt.%), Pd, Pt and 6PGE (ppm) in the UG2 Chromitite Reef

(from top to bottom)

Element Cr Cu Ni S Au Ir Os Pd Pt Rh Ru


UG286 152231 74 1200 0.08 8 86 15 514 1120 200 430

UG285 39549 38 791 <0.05 <5 21 <2 90 237 48 117

UG284 264184 48 1246 <0.05 7 175 85 326 2288 429 856

UG283 283937 49 1263 <0.05 <5 245 123 186 2501 376 1061

UG282 236697 61 1181 <0.05 17 235 146 528 3636 558 1235

UG281 222244 57 1192 <0.05 <5 155 63 362 1665 335 668

UG280 208765 50 1198 <0.05 13 297 192 3074 4224 863 1312

UG279 148396 45 914 <0.05 57 685 381 11574 8815 2113 2270

UG278 1789 40 167 <0.05 <5 <2 <2 10 10 4 13

111

APPENDIX 4: MILLING TESTS, ELEMENT DEPORTMENT AND FLOTATION DATA

The following tables of data show milling times for 60% mass passing 75µm sieve, grading

analysis, elemental deportment, and flotation performance data for BK, RPM, WPand UG2

facies ore samples.

Table A4.1 Mass % passing 75µm for BK, RPM, WP, and UG2 facies ore samples

Facies type Milling

times/minutes

Cumulative

mill

times/minutes

+75µ

mass/g

-75µ

mass/g

% passing

75µ

sieve

BK 10 10 765.6 225.4 22.74

15 25 548.9 431.4 44.01

25 50 292.9 684.3 70.03

RPM 10 10 746.0 245.5 24.76

15 25 522.0 465.0 47.11

25 50 229.9 739.5 76.28

WP 10 10 737.1 261.1 26.16

15 25 501.3 492.1 49.54

20 45 224.4 761.8 77.25

BK 10 10 761.5 236.7 23.71

15 25 504.9 487.1 49.10

25 50 82.3 910.0 91.71

112

Table A4.2 Grading Analysis results for BK, RPM, WPand UG2 ore facies

Facies Type Size fraction Mass-Discrete Discrete mass Cumulative mass

µm g % %

BK -25µm 12.24 1.23 1.23

+25µm 302.89 30.47 31.7

+53µm 307.24 30.9 62.6

+75µm 296.71 29.85 92.45

+106µm 75.08 7.55 100

Total 994.16

RPM -25µm 7.18 0.37 0.37

+25µm 426.24 21.92 22.29

+53µm 636.24 32.73 55.02

+75µm 736.24 37.87 92.89

+106µm 138.2 7.11 100

Total 1944.1

WP -25µm 16.02 0.86 0.86

+25µm 320.39 17.25 18.11

+53µm 616.24 33.18 51.29

+75µm 428.08 23.05 74.34

+106µm 476.65 25.66 100.00

Total 1857.38

UG2 -25µm 16.53 1.66 1.66

+25µm 284.6 28.55 30.21

+53µm 249.12 24.99 55.20

+75µm 354.51 35.56 90.76

+106µm 92.09 9.24 100.00

Total 996.85

The values in the above table were calculated as shown in Appendix 5, Example1.

113

Table A4.3 Assay results of ore feeds sized fractions

Element Co Cr Cu Ni S Au Ir Os Pd Pt Rh Ru

Units(µm) ppm ppm ppm ppm Wt.% ppb ppb ppb ppb ppb ppb ppb

BK+106 71 3421 56 733 0.07 1683 16 12 747 21212 24 122

BK+75 61 2855 201 895 0.15 432 43 29 811 2181 113 304

BK+53 78 2918 743 1857 0.4 179 80 53 2403 2076 327 565

BK+25 113 3556 1278 3079 0.6 277 124 76 3593 2593 495 785

BK-25 277 6349 2833 1.09 2.06 … … … … … … …

RPM+106 89 3669 25 880 <0.05 214 5 2 56 246 12 30

RPM+75 158 7641 89 1435 0.07 25 12 7 127 503 32 80

RPM+53 82 4933 222 1180 0.1 55 37 22 421 1030 115 243

RPM+25 91 4375 346 1296 0.11 106 51 29 615 1515 171 309

RPM-25 134 5547 572 2000 0.25 … … … … … … …

WP+106 81 4261 124 911 0.09 234 10 9 140 434 19 53

WP+75 99 4764 261 1136 0.18 94 14 9 241 608 30 73

WP+53 103 4297 771 1833 0.35 208 33 19 692 1356 90 186

WP+25 93 4204 1085 2175 0.44 181 46 27 940 2104 124 281

WP-25 153 5857 1989 4432 0.88 … … … … … … …

UG2+106 178 18.89 <20 947 <0.05 9 33 17 337 465 92 191

UG2+75 171 19.90 <20 935 <0.05 13 48 25 474 625 133 269

UG2+53 157 17.34 <20 920 <0.05 16 138 61 786 1856 351 648

UG2+25 164 16.29 56 946 <0.05 27 353 151 1609 4502 820 1566

UG2-25 202 19.55 145 1387 0.09 … … … … … …

NB: Underlined figures were in percentage. Figures in BOLD were calculated and the rest were

measured.

… represent figures that could not be calculated.

All the data are as analysed and reported by a commercial laboratory

114

Table A4.4 Assay of Cu, Ni, S, Pd and Pt in BK, RPM, WPand UG2 ore feed sized fractions

ELEMENT Cu Ni S Pd Pt

Units(µm) ppm ppm Wt.% ppb ppb

BK+106µm 56 733 0.07 747 21212

BK+75µm 201 895 0.15 811 2181

BK+53µm 743 1857 0.4 2403 2076

BK+25µm 1278 3079 0.6 3593 2593

BK-25µm 2833 1.09% 2.06 18946 …

RPM+106µm 25 880 <0.05 56 246

RPM+75µm 89 1435 0.07 127 503

RPM+53µm 222 1180 0.1 421 1030

RPM+25µm 346 1296 0.11 615 1515

RPM-25µm 572 2000 0.25 738 112577

WP+106µm 124 911 0.09 140 434

WP+75µm 261 1136 0.18 241 608

WP+53µm 771 1833 0.35 692 1356

WP+25µm 1085 2175 0.44 940 2104

WP-25µm 1989 4432 0.88 1704 …

UG2+106µm <20 947 <0.05 337 465

UG2+75µm <20 935 <0.05 474 625

UG2+53µm <20 920 <0.05 786 1856

UG2+25µm 56 946 <0.05 1609 4502

UG2-25µm 145 1387 0.09 … 10532

NB: Figures in BOLD were calculated and the rest were measured.

… represent figures that could not be calculated.

All the data are as analysed and reported by a commercial laboratory

115

Table A4.5 Copper deportment results in BK, RPM, WPand UG2 ore milled feeds

Facies type Size fraction Fraction Mass Fraction Mass Cu mass Difference

µm g % % %

BK +106µm 75.08 7.55 2.01 -5.54

+75µm 296.71 29.85 31.45 1.60

+53µm 307.24 30.9 33.72 2.82

+25µm 302.89 30.47 32.77 2.30

-25µm 12.24 1.23 0.05 -1.18

Total 994.18 100 100

RPM +106µm 138.2 7.11 1.66 -5.45

+75µm 736.24 37.87 47.23 9.36

+53µm 636.24 32.73 35.27 2.54

+25µm 426.24 21.92 15.83 -6.09

-25µm 7.18 0.37 0.00 -0.37

Total 1944.1 100 99.99

WP +106µm 476.65 25.66 25.44 -0.22

+75µm 428.08 23.05 20.52 -2.53

+53µm 616.24 33.18 42.52 9.34

+25µm 320.39 17.25 11.49 -5.76

-25µm 16.02 0.86 0.03 -0.83

1857.38 100.00 100.00

UG2 +106µm 92.09 9.24 0.00 -9.24

+75µm 354.51 35.56 0.00 -35.56

+53µm 249.12 24.99 0.00 -24.99

+25µm 284.6 28.55 99.66 71.11

-25µm 16.53 1.66 0.34 -1.32

Total 996.85 100 100.00

116

Table A4.6 Nickel deportment results in BK, RPM, WPand UG2 ore milled feeds

Facies

Size

fraction

Fraction

Mass

Fraction

Mass Ni mass %difference

µm g % % %

BK +106µm 75.08 7.55 2.81 -4.74

+75µm 296.71 29.85 13.57 -16.28

+53µm 307.24 30.9 29.15 -1.75

+25µm 302.89 30.47 47.65 17.18

-25µm 12.24 1.23 6.82 5.59

Total 994.18 100 100.00

RPM +106µm 138.2 7.11 4.87 -2.24

+75µm 736.24 37.87 42.33 4.46

+53µm 636.24 32.73 30.08 -2.65

+25µm 426.24 21.92 22.13 0.21

-25µm 7.18 0.37 0.58 0.21

Total 1944.1 100 100.00

WP +106µm 476.65 25.66 15.41 -10.25

+75µm 428.08 23.05 17.26 -5.79

+53µm 616.24 33.18 40.08 6.90

+25µm 320.39 17.25 24.73 7.48

-25µm 16.02 0.86 2.52 1.66

Total 1857.38 100.00 100.00

UG2 +106µm 92.09 9.24 9.28 0.04

+75µm 354.51 35.56 35.26 -0.30

+53µm 249.12 24.99 24.38 -0.61

+25µm 284.6 28.55 28.64 0.09

-25µm 16.53 1.66 2.44 0.78

Total 996.85 100 100.00

117

Table A4.7 Sulfur deportment results in BK, RPM, WPand UG2 ore milled feeds

Facies

Size

fraction

Fraction

Mass

Fraction

Mass S mass %difference

µm g % % %

BK +106µm 75.08 7.55 1.38 -6.17

+75µm 296.71 29.85 11.72 -18.13

+53µm 307.24 30.9 32.37 1.47

+25µm 302.89 30.47 47.87 17.40

-25µm 12.24 1.23 6.64 5.41

Total 994.18 100 100.00 0.00

RPM +106µm 138.2 7.11 0.00 -7.11

+75µm 736.24 37.87 31.46 -6.41

+53µm 636.24 32.73 38.83 6.10

+25µm 426.24 21.92 28.62 6.70

-25µm 7.18 0.37 1.10 0.73

Total 1944.1 100 100.00 0.00

WP +106µm 476.65 25.66 8.74 -16.92

+75µm 428.08 23.05 15.70 -7.35

+53µm 616.24 33.18 43.95 10.77

+25µm 320.39 17.25 28.73 11.48

-25µm 16.02 0.86 2.87 2.01

Total 1857.38 100.00 100.00 0.00

UG2 +106µm 92.09 9.24 0.00 -9.24

+75µm 354.51 35.56 0.00 -35.56

+53µm 249.12 24.99 0.00 -24.99

+25µm 284.6 28.55 0.00 -28.55

-25µm 16.53 1.66 100.00 98.34

Total 996.85 100 100.00 0.00

118

Table A4.8 Palladium deportment results in BK, RPM, WPand UG2 ore milled feeds

Facies

Size

fraction

Fraction

Mass

Fraction

Mass Pd mass Difference

µm g % % %

BK +106µm 75.08 7.55 2.38 -5.17

+75µm 296.71 29.85 10.22 -19.63

+53µm 307.24 30.9 31.35 0.45

+25µm 302.89 30.47 46.21 15.74

-25µm 12.24 1.23 9.85 8.62

Total 994.16 100 100

RPM +106µm 138.2 7.11 1.22 -5.89

+75µm 736.24 37.87 14.69 -23.18

+53µm 636.24 32.73 42.08 9.35

+25µm 426.24 21.92 41.18 19.26

-25µm 7.18 0.37 0.83 0.46

Total 1944.1 100 100

WP +106µm 476.65 25.66 7.22 -18.44

+75µm 428.08 23.05 11.16 -11.89

+53µm 616.24 33.18 46.11 12.93

+25µm 320.39 17.25 32.57 15.32

-25µm 16.02 0.86 2.95 2.09

1857.38 100.00 100

UG2 +106µm 92.09 9.24 10.51 1.27

+75µm 354.51 35.56 14.78 -20.78

+53µm 249.12 24.99 24.52 -0.47

+25µm 284.6 28.55 50.19 21.64

-25µm 16.53 1.66 xxx xxx

Total 996.85 100

119

Table A4.9 Platinum deportment results in BK, RPM, WPand UG2 ore milled feeds

Facies

Size

fraction

Fraction

Mass

Fraction

Mass Pt mass Difference

µm g % % %

BK +106µm 75.08 7.55 43.48 35.93

+75µm 296.71 29.85 17.67 -12.18

+53µm 307.24 30.9 17.41 -13.49

+25µm 302.89 30.47 21.44 -9.03

-25µm 12.24 1.23 xxx xxx

Total 994.16 100

RPM +106µm 138.2 7.11 1.35 -5.76

+75µm 736.24 37.87 14.73 -23.14

+53µm 636.24 32.73 26.07 -6.66

+25µm 426.24 21.92 25.69 3.77

-25µm 7.18 0.37 32.16 31.79

Total 1944.1 100 100

WP +106µm 476.65 25.66 10.46 -15.20

+75µm 428.08 23.05 13.17 -9.88

+53µm 616.24 33.18 42.27 9.09

+25µm 320.39 17.25 34.1 16.85

-25µm 16.02 0.86 xxx xxx

1857.38 100.00

UG2 +106µm 92.09 9.24 1.96 -7.28

+75µm 354.51 35.56 10.15 -25.41

+53µm 249.12 24.99 21.19 -3.80

+25µm 284.6 28.55 58.72 30.17

-25µm 16.53 1.66 7.98 6.32

Total 996.85 100 100

NB: XXX represented valued that could not be calculated.

120

Table A4.10 Mass pull (g) and water recovery (g) variation with time for BK, RPM, WPand

UG2 samples (in duplicate)

Conc Name Flotation

Time/minutes

Wet conc

mass/g

Dry conc

mass/g

Water

recovery

mass/g

BK_B_Conc1 2 93.2 34 59.2

BK_B_Conc2 4 61.6 16.7 89.1

BK_B_Conc3 6 134 11.1 129

BK_B_Conc4 8 168.3 8.8 132.9

BK_C_Conc1 2 65.2 25.3 39.9

BK_C_Conc2 4 101.5 20.7 80.8

BK_C_Conc3 6 101.6 9.9 91.7

BK_C_Conc4 8 201.4 12.8 188.6

RPM_D_Conc1 2 71 23.9 47

RPM_D_Conc2 4 67.4 14.9 52

RPM_D_Conc3 6 115.8 12.4 103

RPM_D_Conc4 8 200.3 13.9 186

RPM_E_Conc1 2 82.8 26.6 56.2

RPM_E_Conc2 4 61.6 11.6 50

RPM_E_Conc3 6 134 13.3 120.7

RPM_E_Conc4 8 168.3 10.5 157.8

WP_E_Conc1 2 139.9 64.4 75.5

WP_E_Conc2 4 78.3 32.8 45.5

WP_E_Conc3 6 104.5 25.8 78.7

WP_E_Conc4 8 237.3 31.6 205.7

121

Table A4.10 (continued) Mass pull (g) and water recovery (g) variation with time for BK, RPM, WPand UG2 samples (in duplicate)

Conc Name Flotation

Time/minutes

Wet conc

mass/g

Dry conc

mass/g

Water

recovery

mass/g

WP_F_Conc1 2 163.7 76.1 87.6

WP_F_Conc2 4 134.9 49.3 85.6

WP_F_Conc3 6 123.3 23.9 99.4

WP_F_Conc4 8 207.1 24.7 182.4

UG2_B_Conc1 2 116.2 28 88.2

UG2_B_Conc2 4 89.7 17.2 72.5

UG2_B_Conc3 6 133.4 9.4 0.124

UG2_B_Conc4 8 0.1362 0.0079 128.3

UG2_C_Conc1 2 116.3 29.9 86.4

UG2_C_Conc2 4 92.1 15.5 76.6

UG2_C_Conc3 6 101 7.4 93.6

UG2_C_Conc4 8 162.6 8.4 154.2

Concentrate and tailings masses were reconciled with the original sample mass which was

initially subjected to flotation to determinethe sample material lost, if any, during milling,

sampling from cell, flotation, drying, and weighing, using the following calculations:

BK_B: Original sample mass= 993.7g

122

BK_B_FEED = 61 g (dry) – sampled from flotation cell before reagents additions (for assay).

BK_B_Tailings mass (dry) =850g.

Mass balance = Feed mass ( sampled)+Tailings mass+ sum of all concs = 981.6g

Mass loss = 12.1g

The average flotation test value of the duplicate sample was obtained by adding corresponding

timed concentrate masses of the same facies and dividing the sum by two. For example, the

average mass pull for BK_Conc1 is given by: (mass of BK_B_Conc1+ mass of

BK_C_Conc1)/2= (34g+25.3g)/2=29.65g

123

Table A4.11 Mineral modal abundances of feed and concentrates

Mineral FEED FEED FEED FEED CONC1 CONC1 CONC1 CONC1 BK RPM WP UG2 BK RPM WP UG2

Apatite 0.01 0.03 0.16 0.02 0 0 0.15 0

Augite 11.28 10.21 10.03 1.07 5.03 10.27 9.47 5.07

Calcite 0 0.01 0.02 0.01 0.05 0.32 0.18 0.24

Chalcopyrite 0.26 0.03 0.64 0 7.23 1.9 2.56 0.37

Chlorite 0.06 0.17 0.07 0.07 0.02 0.08 0.03 1.09

Chromite 1.09 3.1 1.56 72.96 0.09 0.27 0.23 17.41

Dolomite 0 0 0 0 0 0.01 0 0

Enstatite 57.63 63.54 68.36 9.5 39.19 68.49 66.74 56.11

Galena 0 0 0 0 0.01 0.01 0.02 0

Hornblende 0.68 1.04 0.63 0.21 1.18 1.24 0.95 2.13

Ilmenite 0.02 0.07 0 0 0 0 0.04 0

Forsterite 0.08 2.08 0.06 0.43 0.19 1.01 0.46 3.82

Orthoclase 0.1 0.15 0.17 0.12 0.04 0.05 0.02 0.25

Pentlandite 0.91 0.07 0.66 0.03 23.29 4.39 5.9 1.21

Biotite 0.26 0.59 0.61 0.57 0.35 1.85 1.09 1.58

Plagioclase 23.44 14.31 12.77 12.22 1.66 1.82 1.11 5.84

PtFeSnS 0.02 0.02 0.01 0.01 0.03 0.04 0.03 0.01

PtTeBi 0.81 0.92 0.78 0.9 0.03 0 0 0

Pyrrhotite 1.2 0.08 0.79 0.03 18.93 3.2 6.27 0.58

Quartz 0.15 0.61 0.56 0.05 0.02 0.25 0.45 0.95

RuS 0 0 0 0 0.15 0 0 0

Rutile 0 0 0 0.02 0 0 0 0.12

Serpentine 0.02 0.04 0.05 0.02 0.05 0.16 0.16 0.32

Talc 0.77 0.84 0.83 0.06 1.72 2.57 2.56 0.87

Tremolite 0.15 0.17 0.24 0.09 0.42 1.3 1.02 1.81

Wollastonite 0.01 0.02 0.02 0 0 0.01 0.02 0.11

PtPdS 0 0 0 0 0 0 0 0

PdBiTe 0 0 0 0.04 0 0 0 0

ThPO4 0.02 0.03 0.02 0.6 0 0 0 0.05

RhPtAsS 0.38 0.49 0.3 0.77 0.18 0.05 0.09 0.02

ThSiO3 0.29 0.34 0.38 0.05 0 0 0 0

Magnetite 0.29 0.97 0.2 0.06 0.13 0.7 0.46 0.01

PdBi 0.04 0.06 0.04 0.1 0 0 0 0

Total 100 100 100 100 100 100 100 100

124

Table A4.11 (continued) Mineral modal abundances of feed and concentrates

Mineral CONC

2 CONC

2 CONC

2 CONC

2 CONC

3 CONC

3 CONC

3 CONC

3 CONC

4 CONC

4 CONC

4 CONC

4

BK RPM WP UG2 BK RPM WP UG2 BK RPM WP UG2 Apatite 0 0.02 0.07 0 0.02 0 0 0 0 0 0.06 0.16

Augite 9.91 10.39 9.53 5 9.4 11.71 11.02 13.3 10.94 11.18 10.09 3.78

Calcite 0.01 0.09 0.09 0.09 0.11 0.07 0.03 0.01 0.02 0.03 0.05 0.03

Chalcopyrite

1.67 0.51 0.62 0.02 0.4 0.22 0.57 0.83 0.15 0.12 0.31 0.01

Chlorite 0.04 0.1 0.04 0.84 0.32 0.13 0.04 0.03 0.05 0.13 0.04 0.64

Chromite 0.18 0.34 0.69 16.82 5.68 0.17 0.2 0.7 0.01 0.37 0.25 21.3

Enstatite 66.29 75.46 77.84 61.02 67.78 75.03 76.8 72.34 75.59 73.93 79.16 55.9

Galena 0.02 0.01 0.01 0 0.01 0 0.01 0 0 0 0 0

Hornblende 1.21 0.92 0.64 1.77 1.94 1.91 0.93 1.75 2.32 2.64 1.04 2.55

Ilmenite 0 0.01 0 0 0 0.01 0 0 0 0 0 0

Forsterite 0.14 0.83 0.19 2.55 1.85 1.18 0.17 0.11 0.16 1.23 0.15 4.25

Orthoclase 0.05 0.04 0.05 0.27 0.17 0.08 0.13 0.08 0.04 0.14 0.11 0.14

Pentlandite 4.87 1.51 2.13 0.04 0.86 0.21 0.51 0.44 0.44 0.19 0.21 0.04

Biotite 0.51 2.05 1.31 1.71 0.88 2.18 2.06 0.63 0.83 2.11 1.19 1.65

Plagioclase 3.84 1.42 1.45 6.71 4.75 1.3 1.71 4.01 4.24 2.04 1.99 6.45

PtFeSnS 0.02 0.01 0.01 0.01 0.01 0.01 0.02 0.02 0.04 0 0 0

PtTeBi 0 0 0 0.01 0.01 0 0 0 0 0 0 0.03

Pyrrhotite 8.77 2.87 1.99 0.01 2.25 0.72 1.03 2.07 0.43 1.47 0.87 0.05

Quartz 0.28 0.26 0.64 0.66 0.31 0.51 0.79 0.18 0.23 0.37 0.8 0.82

Rutile 0.01 0 0 0.1 0 0 0.02 0 0 0 0.1 0.01

Serpentine 0.06 0.16 0.1 0.21 0.09 0.12 0.06 0.06 0.06 0.06 0.08 0.09

Talc 1.52 1.51 1.49 0.44 2.12 3.12 2.64 2.46 3.22 2.67 2.52 0.9

Tremolite 0.54 1.16 0.77 1.53 0.84 1.14 0.89 0.87 1.03 1.07 0.83 0.92

Wollastonite

0.01 0.01 0.03 0.1 0.04 0.04 0.04 0.02 0.02 0.03 0.04 0.05

PtPdS 0 0 0 0 0.01 0 0 0 0 0 0 0

ThPO4 0 0 0 0.08 0.06 0 0 0 0 0 0 0.16

RhPtAsS 0.03 0.02 0.03 0 0.02 0.02 0.01 0 0.01 0.01 0.01 0

ThSiO3 0 0 0 0 0.01 0 0 0 0 0 0 0

Magnetite 0.03 0.31 0.28 0.01 0.07 0.12 0.33 0.08 0.18 0.19 0.1 0.04

Total 100 100 100 100 100 100 100 100 100 100 100 100

125

Table A4.12 SPLGXMAP Chalc+Pent+Pyrr Wt.% locking in BK, RPM, WPand UG2 feeds

Total Sulphides locked in…

BK-Binary Particle (%)

RPM-Binary Particle (%)

UG2-Binary Particle (%)

WP-Binary Particle (%)

BK-Ternary particle (%)

RPM-Ternary Particle (%)

UG2-Ternary Particle (%)

WP-Ternary Particle (%)

Calcite 0.56 0.00 0.33 0.11 0.05 0.14 0.01 0.09 Chlorite 0.00 0.00 0.62 0.00 0.01 0.00 0.04 0.00 Chromite 0.08 0.06 5.15 0.06 0.00 0.01 1.31 0.01 Dolomite 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 Enstatite 2.63 3.41 2.35 5.66 0.64 1.16 2.25 0.78 Epidote 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 Galena 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.02 Hornblende 0.50 1.06 2.65 0.19 0.18 0.31 0.78 0.28 Ilmenite 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 Forsterite 0.03 0.00 0.00 0.00 0.02 0.11 0.88 0.01 Orthoclase 0.00 0.00 0.06 0.00 0.01 0.00 0.58 0.01 Biotite 0.00 0.45 0.31 0.25 0.01 0.01 1.71 0.00 Plagioclase 1.64 1.66 3.56 0.90 0.15 0.26 3.33 0.26 PtAs 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 PtFe 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 PtFeSnS 0.49 0.23 1.04 0.40 0.00 0.05 0.00 0.08 PtS 0.00 0.00 0.00 0.00 0.00 0.00 1.36 0.00 PtTeBi 0.06 0.00 0.00 0.00 0.00 0.00 0.06 0.00 Quartz 0.00 0.00 0.00 0.08 0.03 0.02 0.77 0.07 RuS 0.00 0.00 1.32 0.00 0.00 0.00 0.00 0.00 Rutile 0.00 0.00 0.00 0.00 0.00 0.00 0.05 0.00 Serpentine 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.01 Talc 0.16 0.00 0.00 0.02 0.05 0.11 0.03 0.07 Tremolite 0.00 0.00 0.00 0.02 0.01 0.00 0.10 0.01 PtPdS 0.00 0.00 1.39 0.00 0.00 0.00 0.66 0.01 RhPtAsS 4.75 3.67 6.85 5.77 0.01 0.24 0.30 0.18 ThSiO3 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 Magnetite 0.19 1.51 0.52 0.92 0.12 0.84 0.67 0.51 Total 11.90 12.60 26.35 15.73 1.53 3.99 15.70 2.66

126

Table A4.13 Assay results of ore feeds, concentrates and tailings for BK, RPM, WPand UG2

facies

ELEMENT Au Co Cr Cu Ir Ni Os Pd Pt Rh Ru S

Sample ID ppb ppm ppm ppm ppb ppm ppb ppb ppb ppb ppb %

BK_Feed 139 77 2925 764 89 1987 50 2369 1780 286 579 0.4

BK_Tails 31 71 3232 34 8 600 5 84 82 9 66 <0.05

BK_Conc1 … 776 2191 1.84% … 4.05% … … … … … 10.02

BK_Conc2 … 247 2174 3293 … 9650 … … … … … 2.96

BK_Conc3 … 168 2455 1715 … 4622 … … … … … 1.37

BK_Conc4 … 106 2656 877 … 2143 … … … … … 0.62

RPM_Feed 64 94 4082 143 28 938 15 321 1293 90 174 0.07

RPM_Tails 16 94 4598 <0.05% 5 789 3 33 8 39 <0.05

RPM_Conc1 … 345 2069 5106 … 7246 … … … … … 1.75

RPM_Conc2 … 175 2317 761 … 3250 … … … … … 0.72

RPM_Conc3 … 148 2886 392 … 2216 … … … … … 0.53

RPM_Conc4 … 104 2798 243 … 1283 … … … … … 0.24

WP_Feed 153 74 4045 554 26 1413 17 498 938 65 162 0.3

WP_Tails 31 79 4803 28 3 664 2 37 78 7 26 <0.05

WP_Conc1 … 266 2361 6277 … 9047 … … … … … 2.79

WP_Conc2 … 142 2710 1506 … 3710 … … … … … 0.92

WP_Conc3 … 121 2780 892 … 2473 … … … … … 0.62

WP_Conc4 … 91 3073 386 … 1486 … … … … … 0.33

UG2_Feed 21 182 18.63% <0.05% 162 958 66 714 2189 385 746 <0.05

UG2_Tails 9 201 19.39% <0.05% 20 930 13 49 157 30 143

<0.05

UG2_Conc1 … 122 3.61% 1066 … 2748 … … … … … 0.4

UG2_Conc2 … 82 3.64% 246 … 1179 … … … … … 0.13

UG2_Conc3 … 82 4.18% 174 … 974 … … … … … 0.07

UG2_Conc4 … 93 4.86% 177 … 930 … … … … … <0.05

… represent values that could not be measured. All data are given as analysed and reported by a

commercial laboratory.

127

Table A4.14 Flotation performance analyses of BK, RPM, WPand UG2 ore facies

Facies

Type

Sample

ID Time/min

Cum

Mass

pull%

Cum

Cu

grade,%

Cum

Cu

Rec,%

Cum Ni

grade,%

Cum

Ni

Rec,%

Cum S

grade,%

Cum S

Rec,%

BK Conc1 2 3.20 1.84 37.23 4.05 28.35 10.02 79.54

Conc2 6 5.22 2.40 79.26 6.22 70.95 7.29 94.36

Conc3 12 6.35 2.28 91.55 5.93 82.40 6.23 98.21

Conc4 20 7.52 2.06 98.01 5.34 87.87 5.36 100

RPM Conc1 2 2.71 0.51 87.70 0.72 19.15 1.75 69.63

Conc2 6 4.13 0.36 94.56 0.59 23.66 1.40 84.66

Conc3 12 5.51 0.28 97.98 0.50 26.64 1.18 95.36

Conc4 20 6.82 0.23 100.00 0.43 28.28 0.07 100

WP Conc1 2 7.48 0.63 79.09 0.90 45.21 2.79 75.83

Conc2 6 11.85 0.45 90.18 0.71 56.04 2.1 90.44

Conc3 12 14.49 0.39 94.16 0.62 60.41 1.83 96.41

Conc4 20 17.49 0.33 96.11 0.54 63.39 1.57 100

UG2 Conc1 2 3.12 0.11 81.67 0.27 8.65 0.4 81.02

Conc2 6 4.88 0.08 92.31 0.22 10.75 0.3 95.89

Conc3 12 5.79 0.07 96.18 0.20 11.64 0.27 100

Conc4 20 6.67 0.06 100.00 0.19 12.46 0.23

128

APPENDIX 5: DETAILS OF CALCULATIONS PERFORMED FOR DATA REDUCTION

EXAMPLE 1: Grading analysis calculations, using BK facies sample mass

Table A5.1 Grading analysis of sample of the BK facies type of Merensky Reef

Size fraction

Sample fraction Mass

Sample fraction Mass

Cumulative Mass

µm g % %

-25µm 12.24 1.23 1.23

+25µm 302.89 30.47 31.7

+53µm 307.24 30.9 62.6

+75µm 296.71 29.85 92.45

+106µm 75.08 7.55 100

Total 994.16

Sample fraction mass %

To get sample fraction mass %, the sample fraction mass for a given size fraction is divided by

the total sample mass, and then multiplied by 100. For example, +25µm fraction mass:

Sample fraction mass% = (302.89g/994.16g)x100=30.47%.

The rest are calculated in the same way.

Cumulative mass %

Cumulative mass % is calculated by adding a given fraction mass % to the next, for example:

Cumulative mass % for -25µm fraction =1.23%

Cumulative mass % for +25µm fraction =1.23%+30.47%=31.7%

Cumulative mass % for +53µm fraction =31.7%+30.95=62.6%, and so on.

The same calculation procedure was applied to all the other facies.

129

EXAMPLE 2: Deportment study calculations, using BK facies sample mass fractions and Pd

assays

Table A5.2 Deportment analysis for Pd in sieved mass fractions of samples of the BK facies

type of Merensky Reef

Size fraction

Sample fraction Mass, g

Sample fraction Mass,%

Pd assay

in fraction,%

Pd mass in

fraction,

g

Pd mass % in

fraction

Pd upgrade/or downgrade

+106µm 75.08 7.55 0.0000747 0.0000561 2.38 -5.17

+75µm 296.71 29.85 0.0000811 0.000241 10.22 -19.63

+53µm 307.24 30.9 0.00024 0.000738 31.35 0.45

+25µm 302.89 30.47 0.000359 0.001088 46.21 15.74

-25µm 12.24 1.23 0.00189 0.000232 9.85 8.62

Total 994.16 100 0.002355 100.00

Sample fraction mass%

This is calculated as shown in Appendix 5, Example 1.

Pd mass in size fraction

Using assay values from Appendix 4, Table A4.6, Pd masses in the size fractions calculated as

follows:

Pd mass in +106µm fraction=0.0000747% x75.08g=0.0000561g.

Pd mass in +75µm fraction=0.0000811% x296.71g=0.000241g



Pd mass in -25µm fraction=0.00189% x12.24g=0.000232g

130

Total Pd mass in all fractions (sum of all the above Pd masses in all fractions) = 0.002355g

Pd mass % in fractions

To get Pd mass % for a given fraction, the Pd mass in that fraction is divided by the total Pd

mass in all the fractions, and then multiplied by 100 as shown below:

Pd mass % in +106µm fraction= (0.0000561g/0.002355g) x100=2.38%


Pd mass % in +53µm fraction = (0.000738g/0.002355g) x100=31.35%


Pd mass % in -25µm fraction= (0.000232g/0.002355g) x100=9.85%

Pd upgrade/or downgrade values

Upgrade or downgrade values for a given size fraction is given by the formula below:

Upgrade or downgrade value=Pd mass% in fraction-Sample fraction mass%.

For example, upgrade/or downgrade value for +106µm fraction=2.38% -7.55%=-5.17%

For +75µm fraction= 10.22% -29.85%= -19.63%

For +53µm fraction=31.35% -30.9%= 0.45%

For +25µm fraction=46.21% -30.47%= 15.74%

For -25µm fraction=9.85% -1.23%= 8.62%

The same calculation procedure is applied to all facies sized sample fractions for Cu, Ni, S and

Pt. The upgrade or downgrade values are then plotted as function of size fraction to determine

the deportment pattern of Cu, Ni, S, Pd and Pt in the various sized sample fractions.

131

EXAMPLE 3: Flotation recovery efficiency calculations

Table A5.3 Flotation recovery efficiency values from calculation examples

Facies type Ore feed mass/g Head grade/ppb

Pt

Tails mass/g Tails grade/ppb

Pt

BK 1990.4 1780 1713.3 82

RPM 2016.3 1293 1737 79

WP 2009.4 938 1550.2 78

UG2 1993 2189 1731.3 157

The values in Appendix 4, Table 5.3 were obtained as shown by the following calculations:

1ppb =1/10 000 000%.

Working:

Total Pt in ore feed = 1990.4x1780/10 000 000 = 0.003542912g

Total Pt in tails = 1713.3x82/10 000 000 = 0.000140491g

Total Pt recovered to concentrate = 0.00340242g

Percentage of Pt recovered to concentrate = (0.00340242g/0.003542912g)x100%

= 96.03% as shown in Table 5.3.

All values for the other PGE and gold recovered to concentrate were calculated in the same way.

132

EXAMPLE 4: Cumulative mass pull, grade and recovery performance calculations.

Table A5.4 Mass pulls, grades, and recoveries in a sample of the BK facies type of Merensky Reef

Sample

ID

Time min

Mass pull,

g

Cum mass pull,

g

Cum Mass pull,

%

Cu Assay,

%

Cu Mass,

g

Cum Cu Mass,

g

Cum Cu

grade,

%

Cum Cu

Rec, %

C1 2 59.30 59.30 3.20 1.84 1.09 1.09 1.84 37.23

C2 4 37.40 96.70 5.22 3.29 1.23 2.32 2.40 79.26

C3 6 21.00 117.70 6.35 1.72 0.36 2.68 2.28 91.55

C4 8 21.60 139.30 7.52 0.88 0.19 2.87 2.06 98.01

Tails 1713.30 0.0034 0.06 2.93 100.00

Total 1852.60

The values in the above table are obtained as shown by the following calculations:

C1 to C4 are Concentrate1 to Concentrate4 (mass pulls) collected at 2, 6, 12 and 20 minutes of

flotation times respectively.

Cumulative mass pulls are successive sums of mass pulls, for eaxmple:

Cumulative mass pull corresponding to C1=59.30g

Cumulative mass pull corresponding to C2 (59.3g+37.4g) =96.70g

Cumulative mass pull corresponding to C3 (96.7g+21g) =117.70g

Cumulative mass pull corresponding to C4 (117.7g+21.6g) =139.30g

Cumulative mass pull% is obtained by dividing each cumulative mass pull value by total mass of

concentrates and tails, and then multiplying by 100, for example:

Cumulative mass pull% corresponding to C1= (59.30g/1852.60g) x100=3.20%

Assay values are taken from Appendix 4, Table A4.16. For example, for C1, %Cu=1.84%

133

Mass of Cu, corresponding to C1=(1.84x59.30g)/100=1.09g. All values corresponding to C2-C4

are obtained in the same way.

Cumulative Cu masses are obtained by adding the next Cu mass pull to the previous consecutive

value, for example:

Cumulative Cu mass pull corresponding to C2=(1.09g+1.23g) =2.32g

Cumulative Cu mass pull corresponding to C3=(2.32g+0.36g) =2.68g, and so on.

Cumulative Cu grade % is obtained by dividing cumulative Cu mass by corresponding

cumulative mass pull, and then multiplying by 100%, for example:

Comulative Cu grade % corresponding to C2=(2.32g/96.7g)x100 =2.40%, and so on.

Cumulative Cu recovery % is obtained by dividing each cumulative Cu mass by the total Cu

mass recovered, and then multiplying by 100%, for example:

Cumulative Cu recovery corresponding to C3=(2.68g/2.93g)x100%=91.5%

The same calculation procedure is done for Co, Cr, Ni, and S for all the RPM, WPand UG2

concentrates, and all the results are shown in Table 5.4.

134

CORRECTIONS BASED ON REVIEWERS COMMENTS.

REVIEWER 1

Evaluation of MSc dissertation entitled: Geometallurgical characterization of Merensky reef and

UG2 at the Lonmin Marikana mine, Bushveld Complex, South Africa.

MSc candidate: Mr Thomas Dzvinamurungu (Department of Geology).

Supervisors: Prof. KS Viljoen and Mr. M Knoper

The dissertation describes and applies a protocol for a geometallurgical assessment for the

different facies types in the Bushveld complex. His applied research is relevant for the platinum

mining industry as the result can be used to improve their ore processing procedure. Further, the

thesis lays out a methodology that can be used by the mining industry to investigate the

geometallurgical characteristics of ore material. As such, this research is relevant and beneficial

to the mining industry.

The dissertation shows that the candidate has the ability to define a problem statement and that

he can successfully generate results to address the defined problem. I have indicated below

detailed comments that can help the candidate to improve his dissertation. The corrections can be

done under the guidance of the supervisors.

General comments:

• The main issue that must be addressed by the candidate is the addition of a chapter that

explains the geometallurgical assessment method used in this study. The candidate

clearly states in the study aim and abstract that the purpose of the study is to develop a

geometallurgical assessment that can aid in mineral processing. One would therefore

expect a chapter that outlines the assessment method and also briefly motivates/explains

the different techniques used in the assessment methods. The absence of this chapter

makes the reader continuously guess what the purpose is of the different assessment

techniques described in chapters 3, 4, and 5.

The following section was added:

2.3 Geometallurgy and geometallurgical assessments

135

• Format of the thesis: the author uses many one-sentence paragraphs (typical examples are

on p.12 and p.14), which in many cases can be grouped together. I would recommend

avoiding one-sentence paragraphs in the text.

Some one-sentence paragraphs were grouped together.

• Use of abbreviations: For a non-Bushveld specialist like me, the use of abbreviations like

BK, RPM, WP are confusing; rather write these names in full throughout the text.

BK facies type of Merensky Reef, RPM facies type of Merensky Reef and WP facies type of

Merensky Reef were used, and abbreviations (BK, RPM and WP) added to the terminology

section.

• Weight % vs wt.%: Both are used in the text. Rather use either weight % or wt.% but not

both. The candidate should be consistent.

Wt.% is now used in the text

• The candidate is not consistent in using a space between number and unit.

Spaces between numbers and units were deleted

• g per tonne and ppm are both used in the text. The candidate should be consistent.

ppm is now used in the text as analysed and reported by a commercial labopratory

Figures with a % scale on the y-axis: some of these diagrams (e.g Figs.4.6, 5.4) have a

value up to 120% which is obviously impossible.

% scales on y-axis were corrected to have values up to 100%

136

• At numerous places in the text, a space between words is left out. The author must

carefully check the final document on this before final submission.

Spaces between words were deleted

• Where more than one reference is used in the text, the order of these references appears

random. The standard practice is to refer to these references in chronological order.

References were chronologically re-ordered in the text (e.g.): Kapsios et al., 2006; Kapsios

et al., 2009

• The author must carefully check the number of significant numbers used in the text and

tables. This appears to be random, which does not look very scientific.

Numbers are expressed to two decimal places

• The format of the tables: (1) they are not consistent in terms of aligning and centring. (2)

Mineral names 9e.g. table A4.11, p123) are incomplete or do not fit. (3) The size of the

tables can be significantly reduced by decreasing the font size. (4) When reporting

numbers in tables, it is easier to read if the numbers are right-aligned than centred.

Font sizes were reduced for mineral names to fit in the tables. All figures in the tables

were right-aligned for easier reading.

Detailed comments per page:

p.xii, Abstract: The abstract ends with: “The influence of ore mineralogy….also investigated.”

How was this done and what were the results of these investigation? To end an abstract like this

raises questions.

137

The following was added to the text: ‘Ore facies having the most abundant anorthite required the

longest milling time to achieve the target grind of 60wt.% passing 75µm; and the ore with the

most abundant enstatite (orthopyroxene) produced the largest mass pull on floating. The facies

with higher PGE grade, modal abundance of base metal sulphides, higher degree of liberation of

base metal sulphides and least enstatite abundance produced the most favourable set of

characteristics for efficient PGE recovery’.

p.x, grade: “percentage (%) should be weight percentage (wt.%).”

Wt.% is now used in the text

p.4. Regional setting: what are the rock types found in the Bushveld Complex?

The following was added to the text:

‘The Bushveld Complex consists of ferromagnesian and calcium-aluminium-sodium silicate

rocks. Rock types found in the Merensky Reef range from feldspathic to pegmatoidal pyroxenites,

norites and anorthosites; UG2 has chromitites, pyroxenites and anorthosite; and Platreef has

pyroxenites, serpentinites and cal-silicate rocks.’

p.7.3rd paragraph: “Vermaark” should be “Vermaak”, “Gruenewaldt et al” should be

“Gruenewaldt et al.,”.

Vermaak replaced Vermaark; and Von Gruenewaldt et al., replaced Gruenewaldt et al.,

p.8, first sentence (and other places in the text): Specify the % here. Also 55+32+15 does not add

up to 100%.

Quantities are now in wt.%, and 15% is replaced by 13wt.%

p.8, 2nd paragraph: % volume should be vol.%.

Vol.% has now replaced % volume

138

p.8, 2nd paragraph: Cawthoorn should be Cawthorn.

Cawthorn replaced Cawthoorn in the text

p.9, 2nd paragraph (and other places in the text): 30% and 35%; these percentages must be

specified.

30wt.% and 35wt.% have now replaced 30% and 35% respectively

p.13, Fig.2.1: What do the red bars indicate in the diagram?

The following text was added :

‘The red bars indicate the abundance and distribution of PGE across the facies (Adapted from

Lonmin Group, 2006)’

p.14, 1st paragraph: The mineral percentages must be specified (i.e., modal percentages?).

mineral modal percentages are now expressed in wt.%

p.17, last sentence: “… further research.” Such as…?

The following text was added:

…. further research such as flotation conditions optimization and grade optimization of the

concentrates.

p.21, Table 3.1: Is it really possible to report the modal % in 4 significant numbers?

The figures are reported to two decimal places as generated by the MLA software.

p.22, 2nd paragraph (and other places): %wt should be wt.%.

139

%wt was replaced by wt.%

p.33, Fig.4.1 (and subsequent Figs.): To indicate a value for R2 is a useless exercise if there are

only 3 data points. Remove this from the table.

R2 was deleted from Figs.4.1-4.4

p.33, Fig. 4.1: The values of the % on the y-axis are given in 4 significant numbers whereas for

all the other figures it is just two.

% values on the y-axis were changed to two significant figures (Figs.4.1 and 4.3)

p.36, Table 4.1: Indicate that the mill time is in minutes.

The milling times are now expressed in minutes

p.72, section 6.2.3, Milling: The milling time appears to be controlled by the mineralogy. I

would suggest that the candidate summarises his results in a table and/ or figure to show this.

The way it is written now requires the reader to read it several times before it is clear how the

mineralogy controls the milling time.

Table 6. 1 was added

p.106, Table A3.1 (and other tables where applicable): The candidate reports values of 0.00. I

assume that these concentrations are below the detection limit. If so, they must be reported as

such.

Values of 0.00 are now expressed as below detection limits in tables

140

p.112, Table A4.3: The table shows three dots (…) at numerous places in the table. What is the

meaning of this? Also, some values are given in wt.% and not ppm. Rather use one consistent

unit for the entire table.

Three dots (…) represent figures that could not be calculated.

Wt.% and ppm units are used as analysed and reported by a commercial laboratory

p.123, Table A4.11: Do the values 0 really mean 0 or does it indicate below detection?

0 values were reported as given by MLA instrument

p.125, Table A4.13: What is the meaning of 3 (…), 4(…), or 5(…) dots in the table?

Three dots (…) represent values that could not be measured.

p.127, 128, Tables A5.1 and A5.2: What is the point of specifying the facies type in the first

column if there is only one reported on anyway? Rather use the Table caption to indicate that

grading analysis was done for the specific facies type.

Table captions were amended

p.128, Table A5.2: The candidate uses the scientific notation randomly in this table. The same is

also done in the text (p.128, last four lines).

Scientific notations were replaced by decimal fractions

141

REVIEWER 2

Supervisors: Prof. KS. Viljoen, Mr. M. Knoper

General: In this MSc thesis an MLA and flotation study of core material, feeds and concentrates

of various Merensky reef facies and UG2 samples at the Lonmin Marikana mine is carried for a

chemical and mineralogical deportment investigation.

The aim of the study is to characterize the variability of ore and gangue mineralogy and variation

in PGE abundances within the various MR facies by automated mineralogical techniques. This

characterization would then allow a geometallurgical assessment such as evaluation of responses

of the different facies to mineral processing and identification of critical characteristics

determining processing behavior, by obtaining quantitative mineralogical and textural

information.

The candidate Thomas Dzvinamurungu thereby shows that he is able to undertake scientific

research and to report it. The candidate masters the chosen topic and is able to operate

competently in the broader subject discipline. He also generates original results, which is highly

rated. The standard and level of language and technical presentation are appropriate.

The thesis states clearly the problem and purpose of the investigation, and its motivation and

objectives. It contains a review of the regional and local geological setting of the Merensky, UG2

and Platreefs and the source area for samples investigated. A description of previous work and

studies in this field at Lonmin in the western Bushveld includes stratigraphy, rock and facies

types and mineralization and grades. An appendix chapter A.1 deals with the methods employed

such as sampling, crushing, splitting, grain mounts preparation, MLA and flotation procedures.

Chapters on sample mineralogy and geochemistry as found by MLA, sample milling and element

deportment, as well as flotation tests form the data body of the thesis. On the literature side,

sufficient relevant authoritative literature is cited, which allows the evaluation and interpretation

in the context of the defined topic.

The research programme is appropriate to address the defined topic and successfully generates

new quantitative mineralogical data of selected merensky Reef (MR) and the UG2 facies at

142

Marikana allowing a geometallurgical assessment of previously existing chemical assay data.

The database created is more quantifiable and representative than previous datasets, largely

confirming existing data but creating a data set which is more useful for metallurgists. It allows

an evaluation of the responses of the different MR facies to mineral processing.

The thesis marks are the logical structure, detailed description and the appropriate

style/terminology of presentation of the findings, corresponding with the scientific conventions

of the discipline. The sources are correctly quoted and a reference list is given in the format

appropriate to our discipline.

Obviously some investigations to characterize the Merensky ore by MLA and flotation have

been done before (i.e., Becker et al., 2008, 2009; Brough et al., 2010; Wiese et al., 2005a+b,

2006, 2007; Viljoen et al., 2012). Did those studies not characterize similar facies quantitatively?

It has the impression that this is the first MLA study of MR in the area; the others have been

performed in the Northam area. The scientific question is less a dispute about genetic processes

or a formation model but more of a practical nature in assisting refinement of mineral

exploitation and processing strategies of the ore at Lonmin’s Marikana mine.

The technical editing is mostly good and the standard and level of language and style are

appropriate.

Subject, investigation: The candidate is familiar with the nature and purpose of his research.

Research: Thomas Dzvinamurungu obviously has mastered the techniques that are relevant to

his research.

Literature: He has adequate knowledge of relevant literature, although I am not convinced that

he interpretes the literature of his field of study entirely.

Scientific methodology: Thomas has a good grounding in the theory and application of

scientific methodology.

143

Research report structure: Thomas is able to structure the research report in a scientifically

justifiable manner. The arrangement of the material of the thesis is systematic and well

documented. The arrangement of the material is done in a logical way from theory as stated by

literature over description and results of experiments, where the kinetics of mineral reactions is

the most important aspects.

Detailed comments on content:

The Abstract is relatively general, one might expect slightly more detail on the results such as

how samples of the various MR facies types did mill and float in terms of mineralogy.

The terminology section offers very useful selection of technical terms for the understanding.

The introduction, chapter 1, is a brief and concise section aiming at the objectives-efficient and

effective mineral extraction and processing routes. A short geological setting and a brief

elaboration on the mineralogy of MR, UG2 Reef and the Platreef is given.

The aims of the present study, chapter 2, gives a description of facies in terms of thickness and

rock type, gangue and ore minerals, as well as PGE content and distribution and the motivation

for current study which is:

-A quantitative description of the variability of ore and gangue mineralogy and variation in PGE

abundances within the various MR facies which could pose inherent challenges to PGM

liberation behavior and metallurgical responses in beneficiation processes.

-An assessment of the influence of reef facies variability on comminution and flotation

performance such as Pt deportment.

-A development of a geometallurgical assessment such as an evaluation of responses of the

different facies to mineral processing and the identification of critical characteristics determing

processing behavior, by obtaining quantitative mineralogical and textural information.

-Lastly the study may help to refine mineral exploitation (selective reef facies extraction) and

processing strategies as a basis for further research.

144

Chapter 3 ‘Samples collected, and sample mineralogy and geochemistry’, gives a brief

description, good figures/illustrations of the results of the investigation.

The quality of the figures is high; those are well readable, right size; except figure 5.12a-d and

5.13a-d, where the font of the header is too large.

Font sizes were reduced (from 12 to 10) for numbers, headers and footers in Figs.5.12a-d and

5.13a-d

Chapter 4, the ‘Sample milling, and element deportment’ shows mostly grinding and Cu, Ni, S,

Pt and Pd downgrade/upgrade behavior which are later interpreted. The BK facies represents the

most favourable set of characteristics for the efficient recovery of PGEs; finer grain size of PGM

requires finer grinding not necessarily longer grinding times a grind of 60% passing 75 microns.

Chapter 5 ‘Flotation tests’ gives data on flotation performance, modal mineralogy of feeds and

timed concentrates with particle and sulphides grain size distribution; also sulphide liberation

analyses in the feeds are given and compared to sulphide liberation in ore feeds and concentrates.

Furthermore mineral association and locking, liberated, binary and ternary composite mineral

particles, flotation recovery efficiency and flotation performance analyses as well as grade and

recovery analyses are displayed. The amount of opx in MR has direct influence on mass pull, as

orthopyroxene is naturally floatable in character. Care is taken by not blending too much opx in

high mass pull, but dilution of the PGE grade.

In the discussion most of the results of the investigation are discussed and related to the relevant

findings of other researchers. The discussion is controversial in that it emphasizes where there

are differences to other results from other operations and which consequences it has on

downstream processes. The conclusions are a brief geometallurgical assessment, answering some

of the aims of the study from chapter 2.2, such as the influence of various proportions of gangue

minerals on grinding and flotation behavior of the various rock facies. Blending solutions are

offered to overcome problematic behaviour. The BK facies type ore is identified as offering the

145

most favourable set of characteristics for efficient recovery of PGEs; this facies seems to exist

also at Northam platinum (Brough et al., 2010).

The final paragraph gives very useful recommendations of where to progress with future test

work to clarify the effects of processing composites of MR facies and the UG2 ores such as 1.

Composites of MR and UG2; 2.Further work on flotation conditions optimization; and 3. Grade

optimization tests could be carried out.

The reference chapter is given in a proper manner and the format and layout of the bibliography

is correct and including the most important and recent sources offers a high amount of citation

(94) on 10 pages proving that discussion and conclusions are valid and presented and evaluated

in the context of authoritative published literature.

There is an instructive methods appendix A1.1-A1.9, containing clear mineralogical data

appendices 2 and clear chemical data appendices 3 with milling tests, element deportment and

flotation data making the data more digestible to the reader.

With the help of literature the candidate has identified the geometallurgical analyses of the

various MR and UG2 facies at Lonmin Marikana as feasible topic, thereby formulating the

hypothesis that different amounts of gangue minerals and amounts and sizes of BM sulphides

and PGM in various ore facies have a different effect on grinding, liberation and flotation

behaviour, calling for measures to increase effective PGE recovery. The strength of the study are

the rigorous research approach, a possible weakness could be the non-genetic approach of

mineralogy to purely assist technical solutions.

Recommendation: Thomas Dzvinamurungu has largely realized his research aims. He knows

and has applied the necessary techniques, demonstrates a sound training in research

methodologies and an understanding of the research process. The statement of the problem, and

the application of MLA investigation for geometallurgical analyses of the various MR and UG2

facies at Lonmin Marikana Mine are the essence of the thesis, the text is well structured and

readable. The statement of research question, concepts and methodology are in a good balance

146

results presentation and discussion. Formally, with 92 pages of text plus further 40 pages of

appendix, the author manages to adapt his Master thesis to a reasonable length. Taking the above

criteria into consideration, I would mark this research thesis with 75%. I trust that comments will

be addressed in the bound version of the thesis under guidance of the supervisors.

Geometallurgical characterisation of Merensky Reef and UG2 ...

Documents

Transcript of Geometallurgical characterisation of Merensky Reef and UG2 ...