Flowsheet development for the Kamoa project – A case study

Minerals Engineering xxx (2013) xxx–xxx

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Flowsheet development for the Kamoa project – A case study N.O. Lotter a,⇑, J.F. Oliveira a, A.L. Hannaford b,1, S.R. Amos b a b

Xstrata Process Support, 6 Edison Road, Falconbridge, Ontario, Canada P0M 1S0 Ivanplats, 82 Maude St., Sandton, South Africa

a r t i c l e

i n f o

Article history: Available online xxxx Keywords: Hypogene Supergene Process mineralogy Sampling Flotation testing

a b s t r a c t Ivanplats Ltd. appointed Xstrata Process Support to perform the flowsheeting development work for their hypogene and supergene geomet units of the Kamoa Copper deposit, located west of Kolwezi in the Democratic Republic of the Congo. Through appropriate use of Gy’s sampling and subsampling models, and systematic flowsheet development using modern Process Mineralogy, an optimised Milestone Flowsheet was developed, delivering a final concentrate grade of 32.8% Cu at 85.4% recovery with hypogene ore, and 45.1% Cu grade at a recovery of 83.4% for supergene ore. These results were obtained from representative samples of drill-core, quantitative mineralogy and high-confidence flotation testing. The significant value of this development is that a single flowsheet will treat both hypogene and supergene ores and produce treatable concentrates. Further work to advance the flowsheet performance beyond this milestone benchmark, and to perform variability testing for this resource, has been identified for investigation in the near future. Ó 2013 Elsevier Ltd. All rights reserved.

1. Introduction

geomet units will be mined and treated in one concentrator, either as blends or as campaigns.

1.1. Location and description of the Kamoa resource The Kamoa copper resource is located some 40 km west of the city of Kolwezi, in the south west of the Katanga Province of the Democratic Republic of the Congo, in the region of Central Africa. Drilling at Kamoa in 2008–2009 had discovered laterally continuous, sediment-hosted, high-grade stratiform copper mineralisation in a newly discovered copper district that forms a previously unrecognised western extension of the famous Central African Copperbelt – which hosts such world-class deposits as Kolwezi, Tenke-Fungurume, Konkola and Nchanga. This discovery showed that the western end of the Copperbelt in the Congo does not terminate at Kolwezi (Friedland and Broughton, 2009). The resource contains hypogene and supergene geomet units with a transitional zone located inbetween the (upper) supergene and the (lower) hypogene. In the hypogene unit, the major copper sulphides are bornite, chalcopyrite and chalcocite, whilst in the supergene, the sequence is led by chalcocite, then bornite, and finally chalcopyrite. Covellite is present in both units as a minor copper sulphide. Both q Originally presented to the MEI Conference ‘‘Process Mineralogy ‘12’’, Cape Town, November 2012, with other parts presented to the Canadian Mineral Processors’ national conference, Ottawa, January 2013. Manuscript now submitted to this journal with permission and revisions. 10th February: manuscript edited in keeping with reviewers’ comments. ⇑ Corresponding author. Tel.: +1 705 693 2761; fax: +1 705 699 3431. E-mail address: [email protected] (N.O. Lotter). 1 Metallurgical Advisor to Ivanplats Ltd.

1.2. Modern process mineralogy The modern best practice of Process Mineralogy was reviewed by Lotter (2011). In this review, the synergy between the separate disciplines of sampling, geology, mineralogy and mineral processing in delivering more advanced flowsheeting practice was discussed. The useful mineralogical information to be obtained from representative samples of drill-core by quantitative and compositional mineralogy by QEMSCAN and EPMA develops clear and useful processing implications that the mineral processor can use in setting up a plan for flotation testing. Rather than the older practice of empirically testing ‘‘samples’’ from drill core using unqualified and questionable singleton flotation tests, the modern practice uses the quantitative and compositional mineralogical information by measuring true samples, combined with High-Confidence Flotation Testing (HCFT), which tightens the associated confidence limits and improves the reproducibility of the metallurgical data (Lotter, 1995a, 1995b; Lotter and Fragomeni, 2010). The approach uses representative ore sampling and correct sample mass reduction in keeping with the minimum sample mass model and safety line of Gy et al. (1979). Thereafter follows a suite of replicate flotation tests with associated quality control to disqualify outliers from a small data set. Subsequently, a composite is prepared from the accepted replicates for the mass and value balance. This generally tightens the metal balance and associated confidence limits of the stated grades and recoveries, and improves

0892-6875/$ - see front matter Ó 2013 Elsevier Ltd. All rights reserved. http://dx.doi.org/10.1016/j.mineng.2013.02.014

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N.O. Lotter et al. / Minerals Engineering xxx (2013) xxx–xxx

the reproducibility of the test data. This is the approach that was taken for this project.

point, however as the project has progressed it is becoming more likely that a direct-to-blister smelter will be included in the project scope, thus the term ‘‘smeltable concentrate’’ will be used.

1.3. Mixed collectors From the list of different copper sulphide species reported to be present in the Kamoa resource, it was obvious at an early stage that a mixed collector suite would be necessary for successful flotation. The challenge was to formulate a mixture that would successfully float all of the copper sulphides identified in the resource from both liberated and middling classes. These range across primary copper sulphide (chalcopyrite) and secondary copper sulphides (chalcocite, bornite and covellite). These two groups have different electrochemical properties (Allison et al., 1972; Fuerstenau, 1978; Woods, 1994; Guler et al., 2006; Yoon and Basilio, 1993). Mixed collectors, when optimally formulated, deliver improved flotation performance versus single collector systems through a mixed potential mechanism (Bradshaw, 1997). This is especially valuable when a mixture of different metal sulphides with different flotation properties represents the paymetal set, and all have to be recovered by flotation. The typical benefits are increased concentrate grade and recovery, caused by increased froth carrying capacity, more successful flotation across a wider size range, and faster flotation kinetics (Adkins and Pearse, 1992). In particular more successful flotation of middling particles is a key characteristic (Plaskin et al., 1954). The sixty-year history of research and best practice of formulating mixed collectors for synergistic metallurgical performance was reviewed by Lotter and Bradshaw (2010). From this review and other follow-up work, a prototype expert system nicknamed ‘‘Reagent Sudoku’’ was developed, in which candidate collectors for a known mineral set could be selected and ranged before flotation testing. ‘‘Reagent Sudoku’’ selects collectors on the basis of the types and quantities of sulphide minerals present which are required to be recovered to flotation concentrate. It is a proprietary expert system. This theory was successfully validated by certain laboratory and plant scale tests at XPS for and at the Eland Platinum Mine, South Africa (Lotter et al., 2011). For the copper sulphide minerals, the various forms of collector causing hydrophobicity were identified and tabulated (Hangone et al., 2005). It was shown that dithiophosphates could span the electrochemical needs of the primary and secondary copper sulphides, whereas for xanthates, some required free xanthate, and others, dixanthogen. In the case of xanthate as the single collector, therefore, the dilemma was how to produce a mixture of free xanthate and dixanthogen in the same flotation system. Hangone et al. (2005) concluded that, for the O’Kiep copper sulphide ore sample tested, a mixture of 90% xanthate and 10% dithiophosphate was optimal for the successful flotation of this mixture of chalcopryite, bornite and chalcocite. This confirmed earlier operations practice at O’Kiep in the 1980s (Hannaford, personal communication, Kolwezi, 14 April 2011). The distribution and relative occurrence of the copper sulphides in the Kamoa hypogene geomet unit are shown in Table 4, and show a mixture of chalcopyrite, bornite, chalcocite and covellite. The relevant forms of collector known to float the various copper suphides are shown in Table 1. 1.4. Specific objectives The specific objectives of this work were to develop and demonstrate an optimum flowsheet for the hypogene geomet unit, and thereafter to test the supergene geomet unit with this flowsheet to determine what changes may be necessary in order to obtain best metallurgy for the latter. A specific requirement was to attain a final concentrate grade of at least 28% Cu in both concentrates. Initially the project considered saleable concentrate as an end

2. Method 2.1. Sampling The formal fifty-piece experiment of Gy et al. (1979), was adapted by Lotter (2010), to drill-core by considering each drillcore increment as a ‘‘piece’’ in the fifty-piece experiment. It was shown that the model readily adapts to this application, and that different sampling equations result for different geomet units. An empirical estimate of the size range factor g was made using the following equation:

" g¼

#

1 3

Md



n X 3 Mi di

ð1Þ

i¼1

 is where M is the total mass sampled, g (also called the lot mass), d the weighted mean piece diameter, cm, Mi is the mass of the ith piece, g, and di is the diameter of the ith piece, cm. The sampling constant K for the equation between fundamental variance fv and minimum sample mass Ms is shown as Eq. (2). The fundamental variance is expressed as a decimal fraction and is squared, e.g. a value of 8% is written as (0.08)2.

Ms ¼

K fv

ð2Þ

The sampling constant K is found as Eq. (3) in a relationship between the size range factor, the weighted mean grade of paymetal and its associated individual measurements, the lot mass tested and its associated individual masses.

K¼

!!   X n g  v 2 2   ðai  aÞ ðM i Þ =ðv i Þ 2 Ma

ð3Þ

1¼1

 is the weighted mean particle diameter, cm, M the ith fracwhere d i tional mass corresponding to di, vi the volume of the ith ore particle,  the mean particle volume in sample lot, cm3, M the lot mass, cm, v  the weighted mean sample grade of paymetal, expressed as, g, a grammes per tonne, of ore, or percent metal in ore, ai is the grade of metal in the ith size fraction, di the ith particle diameter, cm, fv the fundamental variance, and g is the size range factor. Accordingly the Kamoa hypogene and supergene drill-core logs were used to obtain empirical values of the sampling constant K. After establishing the sampling equations for the two geomet units, the number of flotation tests required for the programme was calculated. The flowsheet development work was to be performed on the hypogene unit, with the supergene undergoing confirmatory tests with adjustments if necessary. It was estimated from the scope and the total number of tests required that some 265 kg of hypogene drill-core would be needed. This core was HQ diameter as quarter core. (See later in Section 3.1, the sampling equation estimated that 66.4 kg of this drill core would be representative at the 8% fundamental variance level. By exceeding this minimum sample mass (265 kg is heavier than 66.4 kg), the bulk sample is representative). A plot of crushing, subsampling and grinding steps was mapped out using Gy’s safety line, which for this project may be represented by Eq. (4). This ensured safe sampling down to the flotation test product level. 3

M P M s ¼ 125000d

ð4Þ

where M is the sample mass, g, Ms the minimum sample mass, g, and d is the topsize of the sample, cm (equivalent to the d95 size).

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N.O. Lotter et al. / Minerals Engineering xxx (2013) xxx–xxx Table 1 Summary of thiolate/dithiolate species responsible for hydrophobicity in major copper sulphides (after Hangone et al., 2005). Mineral

Xanthate

Dithiophosphate

Dithiocarbamate

Species responsible for hydrophobicity

Reference

Species responsible for hydrophobicity

Reference

Species responsible for hydrophobicity

Reference

Bornite Chalcopyrite

CuX X2

1,5,13 1,7,8,9,10,13

Cu(DTP)2 Cu(DTP)2

3 3,13

Chalcocite

CuX⁄, Cu(X)2 CuX

2,5,6,8,9,10,13 4

Cu(DTP)2 Cu(DTP)⁄

3 2,5

Cu(DTC)2 Cu(DTC) Cu(DTC)2 Cu(DTC)

3 11,12 3 3

Covellite

X2, CuX

1,4,8,13

Cu(DTP)2

3

Cu(DTC)2

3

References: 1. Allison et al., 1972, 2. Chander and Fuerstenau (1974), 3. Finkelstein and Gould (1972), 4. Fuerstenau (1978), 5. Fuerstenau (1990), 6. Guy and Trahar (1985), 7. Suoninen and Laajalehto (1993), 8. Valli and Persson (1994), 9. Woods (1984), 10. Yoon and Basilio (1993), 11. Bhaskar Raju and Khangaonkar (1984), 12. Hodgson and Agar, 1989, 13. Guler et al., 2006.

2.2. Mineralogy Milled test charges of 2.0 kg dry solids ground to a p80 size of 75 lm were prepared from the hypogene and supergene geomet units. After filtration and drying, these were separately subsampled using a spinning riffler to a subsample of approximately 150 g for sizing. This practice ensures that the 150 g subsample is representative of the lot (see Fig. 4). After being passed through a 53 lm wet test sieve (with the oversize retained for further separation on coarser test sieves), the undersize was pre-cycloned at a cutsize of 3 lm. The oversize was filtered and saved for later cyclosizing into CS1–7 size classes. The undersize was retained as sink size 3 + 0 lm. The total mass at each size class was filtered, dried and labelled, then subsampled using the miniature spinning riffler into 2 g subsamples for polished section preparation. At all stages of filtration, alcohol was added to the slurry during filtration to reduce agglomeration. After polished section preparation, the size classes were presented to QEMSCAN and the CAMECA microprobe for quantitative and compositional mineral analysis. Appropriate quality controls were used to assure a high level of accountability in the measurements. The liberation pattern of the copper sulphides present in the hypogene geomet unit is shown in Table 5. 2.3. Flotation testing 2.3.1. High confidence flotation testing – a summary of the approach The flotation testing method known as ‘‘High-Confidence Flotation Testing’’ was developed at Rustenburg Platinum Mines Limited, South Africa (now known as Amplats), by Lotter and Munro (1994), for the specific problem of true sampling and reproducible flotation testing of PGE and PGM-bearing ores. This system was further advanced and validated in a postgraduate study at the University of Cape Town (Lotter, 1995a, 1995b). The platform provided a means of producing flotation test data sets at the 95% level of confidence. Further refinements were made at Falconbridge Ltd and Xstrata Process Support in Canada to focus this model on base metal applications (Lotter et al., 2002; Lotter and Fragomeni, 2010). There are two parts to the model: the correct sampling and subsampling of the ore to the flotation test charge level; and the replicate flotation test system using the powerful averaging effects of the Central Limit Theorem (Box et al., 1978), with an outlier rejection system written and proposed for small data sets by Grubbs (1969). Ore samples that are crushed and milled to smaller size distributions before assaying present easier platforms for the accurate measurement of ore grade (or, as Gy refers to this measurement, as ‘‘content’’). Provided that the primary sample satisfies the sampling equation with a known and acceptable fundamental variance, and that the crushing and milling are performed in keeping with

the provisions of the Safety Line, the subsample integrity is maintained (Gy et al., 1979). This high-confidence flotation system system recognises and manages this challenge by modelling the correct sampling equation for that size distribution and paymetal variance from Gy’s fifty-piece experiment, and by using the Safety Line to appropriately reduce the primary sample after crushing and blending to replicate test charges with smaller size distributions and standard deviations. The ‘‘assay heads’’ are thus obtained from milled test charges at the rougher float feed size distribution (commonly with a topsize of 106 lm). Thereafter, the system uses sets of replicate flotation tests with application of an outlier rejection model for small data sets as proposed by Grubbs (1969). First concentrate replicate mass data must satisfy the criterion of demonstrating a relative standard deviation of less than 5%. Finally, composite formulation and replicate assaying lead to metal balancing of built-up head grades to the mean assay head grade to satisfy a criterion of less than ±3.3% error. Use of this approach enables the comparison of flotation test conditions with smaller differences in grade and recovery, and reducing the scale-up risk. 2.3.2. The mill-float-mill-float (MF2) circuit For ores which carry a wide range of sulphide grain sizes, including very fine-grained species, such as at Kamoa, or the South African Merensky reef, a two-stage mill-float-mill-float (MF2) circuit is preferable. This approach was developed by several South African platinum operations, and delivers a higher paymetal recovery to the cleaner circuit feed and final concentrate than would be the case for a single-stage grind (Nel et al., 2004; Deeplaul and Bryson, 2004 for example). The flowsheet development work was based on the sample of hypogene ore provided from a site visit to Kamoa in April 2011. The baseline platform for rougher and scavenger flotation followed an MF2 arrangement, with primary and secondary grinds set at p80 sizes of 75 and 38 lm respectively, and using the High-Confidence Flotation Testing approach. These initial grinds were selected from a desktop study of prior mineralogical work, which had estimated the grain size distributions of the copper sulphides from earlier samples of hypogene drill-core. Flotation times were 20 and 15 min for rougher and scavenger stages at natural pH (8.5–9.0). The total collector dose was 125 g/t milled, and was split between the flotation stages. This arrangement is shown in Fig. 1. These treatments were assembled from a study of the prior mineralogical work on drill core that revealed the very fine-grained nature of the copper sulphides, ranging from 25 to less than 3 lm. Obviously it would not be economical to grind to a completely liberated state in the primary circuit. Rather, in keeping with typical South American practice where similar challenges are found, the treatments in Fig. 1 will partially liberate the copper sulphides,

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N.O. Lotter et al. / Minerals Engineering xxx (2013) xxx–xxx

leaving the flotation to recover a mixture of liberated and middling particles (Bulatovic et al., 1998). Subsequent niche regrinding of the rougher and scavenger concentrates before cleaner flotation would be economical, since by then the mass treated will be substantially less than that of ore milled. The scope of the mixed collector suite to be developed will be capable of this task; the more successful flotation of middlings by a tuned mixed collector suite was reported by Plaskin (1954). Other work by Bristette and Roman (2012), has demonstrated that, for this type of hypogene and supergene copper operations such as at the Kemess operations, the finishing size of the regrind mill treating the rougher concentrate has a significant effect on the grade and recovery of saleable concentrate, and needs to be optimised for different mining areas.

3. Results 3.1. Sampling The minimum sample mass equations that resulted from a study of the drill-core logs yielded the following equations: 3.1.1. Hypogene ore

Ms ¼

424:95 fv

ð5Þ

For a fundamental variance of 8%, the minimum sample mass required is thus 66.40 kg. Since this was the geomet unit upon which the flowsheet development would be performed, a total of 265.62 kg of quarter HQ core was configured as a representative sample, since this mass exceeds the minimum sample mass, and despatched to Xstrata Process Support.

2.3.3. Cleaner circuit The development of the cleaner circuit would start with an initial prototype, as shown in Fig. 2, and be further led by the microprobe and QEMSCAN data for type, quantity and grain size/ association data of the copper sulphides, together with data on the silicates, as well as the mixed collector programme to tune the mixture so as to recover both primary and secondary copper sulphides. It was recognised at an early stage of the project that the rougher and scavenger concentrates would require separate cleaner circuits so as to disconnect any potential competition between their relative flotation speeds and electrochemistry for the primary and secondary copper sulphides. Furthermore, whatever secondary copper sulphides had floated into the rougher concentrate may show only partial recovery to rougher recleaner concentrate, because of their slower flotation kinetics. An empirical – or prototyping – approach was thus formulated, and using the High-Confidence Flotation Testing approach.

3.1.2. Supergene ore

Ms ¼

218:1 fv

ð6Þ

For a fundamental variance of 8%, the minimum sample mass required is thus 34.08 kg. Since less development work was anticipated on this geomet unit, a total of 193.21 kg of representative sample was configured from quarter HQ core to match the calculated number of tests, and despatched to Xstrata Process Support. The sample material presented to the laboratory scale testwork is thus representative in terms of the definitions of minimum sample mass according to Gy et al. (1979).

2.3.4. Assaying It was critically important to provide a highly accurate assaying measurement system. By discussion with the chemistry staff at various laboratories, it was agreed that a sodium peroxide fusion method followed by acid dissolve and an ICP-OES finish would provide the best measurements. Samples were assayed in duplicate.

3.2. Sample preparation In the case of the hypogene sample, the 265.62 kg of quarter HQ drill core located itself at A on the Safety Line in Fig. 3 (the quarter core has a diameter of 19 mm equivalent). It was then crushed to a

SIBX (40) 3477 (4) DF (38) SIBX (20) 3477 (2) DF (12.5)

Ore

t = 20’ 80% -75 microns

Rougher Concentrate 80% -38 microns

SIBX (52.5) 3477 (5.5) DF (25)

t = 15’

Scavenger Tailings Scavenger Concentrate Fig. 1. MF2 baseline flowsheet (‘‘DF’’ denotes Dowfroth; figures in parentheses denote dosage in g/t ore milled).

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N.O. Lotter et al. / Minerals Engineering xxx (2013) xxx–xxx

topsize of 1.7 mm, relocating itself to B. Sampling is now safe, since B is on the left of the Safety Line. The whole sample was then blended using a spinning riffler, then subsampled on the same spinning riffler to replicate test charges of 2.0 kg each at C (Lotter, 1995a; Lotter and Fragomeni, 2010). The 2.0 kg replicate test charges are thus true subsamples of the primary sample, and may be used for flotation tests. A similar process was followed for the supergene sample. It should be pointed out at this stage that the milling of the 2-kg test charges for flotation further advances the test charge location with respect to the Safety Line to point D, allowing very small samples to be safely taken for mineralogical and chemical analysis. Extension of a vertical line downwards from point D shows that a true sample as small as 1– 100 g may be taken from the milled sample at D to E. The assay head subsamples and mineralogical subsamples were extracted in this manner. This arrangement is shown in Fig. 3. The supergene sample underwent equivalent treatment.

3.3. Assay head grade The hypogene ore sample yielded a mean copper head grade of 3.30% Cu with a relative standard deviation of 2.6% from ten replicate subsamples each assayed in duplicate. The low standard deviation reflects the correct use of the Safety Line, and the impact of

5

the ball milling of the 2.0 kg replicate test charges to flotation feed topsize before releasing the assay head subsamples.

3.4. Rougher float feed mineralogy – hypogene The average compositional mineralogy of the main copper minerals, as measured by the Cameca SX100 microprobe, in the hypogene rougher float feed is shown in Table 2. These measured compositions are typical of these minerals. The bulk modal assembly (BMA) of the hypogene rougher float feed, measured by QEMSCAN, is shown in Table 3. Within the group of copper minerals, the distribution of total copper between copper mineral types is shown in Table 4, with the associated copper sulphide liberation data in Table 5. In Fig. 4, showing the grain size range of the copper sulphides on the y-axis against overall particle size on the x-axis, it is clear that the milled ore presented at a p80 size of 75 lm as Rougher Float Feed provides clues as to the secondary milling requirements. The liberated copper sulphide grain sizes change gradient at cyclosizer product CS1–2, which has a nominal mean particle size of nearly 38 lm (QEMSCAN measured average 32 lm). To a lesser extent a similar observation may be made for the middlings. The foregoing mineralogical information leads to the following process implications:

Fig. 2. Prototype cleaner circuit.

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N.O. Lotter et al. / Minerals Engineering xxx (2013) xxx–xxx

Topsize, mm 0.1

1

10

100 1000

A

B

100

10

C

Mass, Kg

D

1

0.1

E 0.01

0.001 Fig. 3. Pathway with respect to the safety line of the hypogene bulk ore from primary sample (A) to replicate 2.0 kg test charges (C) and finally replicate assay head subsamples (E).

Table 2 Compositional mineralogy: hypogene. Mineral

Bornite Chalcocite Chalcopyrite Covellite Pyrite Azurite

Table 4 Distribution of total copper by mineral type: hypogene.

Composition (%) S

Fe

Cu

Zn

As

Ag

Total

26.03 21.05 35.21 32.76 52.73 –

11.19 0.47 30.13 0.44 45.88 –

62.40 77.78 34.31 65.62 0.10 55.51

0.06 0.08 0.05 0.07 0.01 –

0.00 0.00 0.02 0.01 0.09 –

0.01 0.00 0.01 0.01 0.01 –

99.70 99.41 99.74 98.92 98.81

Table 3 Modal mineralogy: hypogene.

Mineral

%

Bornite Chalcocite Chalcopyrite Covellite Azurite Chlorite Fe (Ti) oxides Other

50.89 17.11 26.33 3.66 1.59 0.21 0.22 0.01

Table 5 Liberation data: hypogene rougher float feed.

Mineral

%

Mineral

%

Mineral

Size class

Locked

Middling

Liberated

Bornite Chalcocite Chalcopyrite Covellite Pyrite Azurite Orthopyroxene Chlorite Quartz

2.72 1.62 3.19 0.19 0.66 0.10 0.29 14.01 29.76

Muscovite Biotite Kaolinite Orthoclase Plagioclase Fe–Ti oxides Carbonates Apatite Other

11.70 2.48 1.13 26.10 0.69 2.66 2.05 0.16 0.49

Overall copper sulphides

+106 106 + 75 75 + 53 CS1–2 CS3 CS4–5 CS6 CS7 Total

5.17 5.01 3.91 3.85 1.27 0.91 0.42 0.38 20.92

1.44 1.94 2.68 3.48 1.30 1.77 2.11 1.88 16.60

0.76 3.23 5.93 15.23 6.50 11.08 10.42 9.33 62.49

1. The compositional data on the major and minor copper sulphides are typical. 2. Chalcopyrite, bornite and chalcocite are the major copper sulphide minerals present, therefore the milling and flotation strategy should be focussed on these minerals, their relative abundances, and grain sizes/textures. 3. In terms of total copper distribution, bornite carries 50.89%, with chalcopyrite and chalcocite a lesser 26.33% 17.11%, respectively. Thus, bornite flotation has to be a key consideration towards higher overall copper recoveries. 4. At the primary grind of p80 = 75 lm, which was formulated from an initial desktop study of prior work, is rather fine for a primary grind. Only 62.49% of total copper sulphides are liberated, with 16.6% as middlings, and 20.92% as locks. This significantly implies that successful flotation of middling particles will be a key strategy in attaining a high primary circuit recovery of copper before cleaner circuit flotation, and that niche regrinding of the

rougher and scavenger concentrates will be necessary before cleaner flotation in order to attain saleable concentrate grade in the order of >28% Cu grade. 5. At the primary grind size, the sample mean copper sulphide grain sizes are: liberated: 27 lm; middling: 15 lm; and locks: 7 lm. In all, a very fine-grained size range across which to successfully float chalcopyrite, bornite and chalcocite. 3.5. MF2 flotation testing The flowsheet development was performed on the hypogene geomet unit. The prototype MF2 rougher/scavenger platform was tested as described in Fig. 1. Note that the key features were: natural pH (which was 8.5); two mainstream grinds to p80 sizes of 75 and 38 lm respectively; and that a mixed collector system based on the O’Kiep work by Hannaford, and later, by Hangone et al. (2005), at 90:10 by mass of SIBX:3477, was used to start with.

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N.O. Lotter et al. / Minerals Engineering xxx (2013) xxx–xxx

Table 6 Grade and recovery data for prototype MF2 platform. Cum. stream

RC1 RC1. . .RC2 RC1. . .RC3 RC1. . .SC1 RC1. . .SC2 BUH

Mass pull%

7.3 12.6 14.3 16.1 18.1 100.0

Fig. 4. CuS grain size by size class in rougher float feed at d80 – 75 lm.

30

Cum. Grade % Cu

The results, culminating in a primary recovery of 87.2% Cu at an overall grade of 16.01% Cu, are shown in Table 6 and Fig. 5. The change in gradient in the grade-recovery curve (Fig. 5) at the point where the secondary grind and scavenger flotation commence indicates further liberation of copper sulphides in the rougher tailings. Another 7.0% Cu recovery is delivered from this process, leaving 12.9% of mill head Cu as losses to Scavenger Tailings. Reference to the liberation data suggests that the prototype MF2 platform has successfully floated most of the middling copper sulphides into rougher concentrate. The cumulative rougher recovery to this point was 80.1% Cu. Only 62.5% of the copper sulphides in the Rougher Float Feed were liberated. If all of these were indeed floated, this implies that (80.5  62.5) = 17.5% were floated as nonliberated species, possibly as middlings. Several different treatment conditions were formulated and tested on the MF2 platform. These were considered from other equivalent operations experience, and included a ‘‘zero’’ baseline (at which the pH was natural – typically 8.5–8.6); addition of lime to pH 10.2; addition of sodium silicate; addition of sodium hydrosulphide; and use of high chrome grinding media. The purpose was to scope a range of known treatments to determine which offered the most promising result. Additionally, the extra cost, safety and associated hazop procedures implied by these reagents had to be considered together with any gain in performance. The zero baseline, with natural pH, was selected. The mineralogy of the hypogene Scavenger Tailings produced by this baseline MF2 float was investigated by QEMSCAN in a series of size classes. The results showed that 59.0% of the total copper sulphides were locked, 25.0% were middlings, and 16.0% were liberated. The distribution of the copper losses by mineral type showed that these losses were led by bornite (36.1%), followed by chalcocite (25.8%), with minor chalcopyrite (9.7%) and covellite (3.7%). If all these liberated copper losses could be recovered to concentrate, there would be a potential recovery gain of 2.1% Cu, offering a higher potential MF2 recovery of 89.3% to rougher and scavenger concentrate. A module of replicated sighter flotation tests using the MF2 rougher-scavenger platform was arranged so as to scope out a few likely reagent combinations based on the expert mixed collector system Reagent Sudoku. These flotation tests were conducted in replicate using the High Confidence Flotation Testing approach. Based on the known mineralogy measured at the Rougher Float Feed stage, in addition to the baseline SIBX:3477 90:10 formulation which had thusfar been used in the testwork, five new combinations were tested. All tests were conducted at a constant collector dosage of 125 g/t milled. These are shown in Table 7 with the results. Note that, due to the fine-grained nature of the primary and secondary copper sulphides, and due to the known incomplete liberation of these across the MF2 platform, it was the expectation that a successful new reagent formulation would increase the copper recovery and drop the concentrate grade due to successful middling particle flotation. From these results it is clear that the

25 20 15 10 5 0 60

70

80

90

100

Cum. Recovery % Cu Fig. 5. Grade-recovery curve for the MF2 platform described in Fig. 1, using 90:10 SIBX:3477 as collectors.

Table 7 Mixed collector sighter tests. Reagent formulation

Baseline SIBX:3477 90:10 SIBX:SEX:Senkol 65 Mix 1 SIBX:SEX:Senkol 65 Mix 2 SIBX:3477 64:36 (Note 1) SIBX:3477:5100 SIBX:3477:3894:5100

Mass pull%

17.6 16.5 16.2 22.1 19.5 16.9

Copper Grade % Cu

Recovery % Cu

16.01 15.64 15.02 12.38 13.35 13.70

87.17 87.17 87.77 89.30 88.91 88.30

Note 1: Formulated by the XPS Expert System ‘‘Reagent Sudoku’’.

formulation SIBX:3477 64:36 by mass (a formulation by the XPS Expert System Reagent Sudoku), produces the highest copper recovery at the lowest concentrate grade, an indication that this combination successfully improves the flotation of middling particles. This formulation was accepted for further work on the flowsheet, and a full circuit test performed at an appropriate stage of cleaner circuit development – i.e. when regrinding strategy in the cleaner circuit had been advanced; to determine whether this suite would deliver the higher recovery to saleable concentrate, and whether the regrinding of the rougher and scavenger concentrates would effectively deal with the extra middling particles. 3.6. Cleaner circuit

COPPER Grade % Cu

Recovery % Cu

28.2 20.4 18.6 17.5 16.0 3.32

61.8 77.4 80.1 84.7 87.1 100.0

3.6.1. Prototype cleaner circuit The cleaner circuit was developed from an initial prototype (see Fig. 2), based on the premise that the rougher and scavenger concentrates would require separate cleaner sections due to their different flotation kinetics. The rougher concentrate was cleaned in two stages to release Rougher Recleaner Concentrate, discarding

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45

B

Cum. Grade % Cu

40 35 30 25 20 15 10

A

5 0 20

30

40

50

60

70

80

90

100

Cum. Recovery % Cu Fig. 6. Grade-recovery curve for the prototype cleaner circuit.

the Rougher Cleaner Tailings and presenting the Rougher Recleaner Tailings to the Scavenger Cleaner Section for retreatment. The latter treated the Scavenger Concentrate plus the Rougher Recleaner Tailings in a single-stage cleaner with a discard Scavenger Cleaner Tailing. Additionally, it was highly desirable to attain a position whereby circulating loads could be minimised and the cleaner tailings would be discard grade. The first prototype cleaner circuit took these criteria into its layout, and added the regrinding of the rougher concentrate prior to its treatment in the rougher cleaner circuit with an Isamill, recognising this technology as appropriate for this type of regrinding (Pease et al., 2006). Initial sizing of

this stream before regrinding estimated an f80 size of 55 lm. Whereas 62.5% of all copper sulphides in the rougher float feed were liberated, these liberated sulphides averaged 27 lm in size. By contrast the middlings amounted to 16.6% abundance with a mean grain size of 15 lm. An initial regrind p80 product size of 15 lm was set using an Isamill. The grade-recovery curve is shown in Fig. 6, delivering a smeltable concentrate grading 31.95% Cu at a recovery of only 65.6% at point ‘‘A’’. A significant recovery loss occurs in the Rougher Cleaner Tailings, amounting to 19.43% of mill head copper (point ‘‘B’’ grade: 20.01% Cu, recovery 85.03% Cu as compared to point ‘‘A’’, viz. (85.03  65.6) = 19.43% copper recovery loss). This cleaner tailing weighed 7.1% mass at a copper grade of 6.2% Cu, and as such could not be regarded as a discard cleaner tailing. Rather, it would have to undergo a retreatment process before discard. Modelling of the unit recoveries around the Prototype Cleaner Circuit showed that this tailing would be best retreated in the Scavenger Cleaner unit. 3.6.2. Breakthrough cleaner circuit From the information produced by the Prototype Cleaner Circuit, the Breakthrough Cleaner Circuit (Fig. 7) was arranged to retreat the Rougher Cleaner Tailings in the Scavenger Cleaner, and niche copper collectors Cytec 3894 and 5100 added to the Rougher Concentrate regrind stage. The results for this circuit change are shown in Fig. 8, showing a direct comparison between the Prototype and Breakthrough circuits. In Fig. 8, point ‘‘A’’ shows the original release point of smeltable concentrate at 31.95% Cu grade with 65.58% recovery in the Prototype Circuit, with point

Rougher Concentrate SIBX (4.5) 3477 (2) DF (A.N.) Rougher Cleaner t=10’

IsaMill to 15 microns

Rougher Cleaner Tailing

Rougher Recleanert=10’

Scavenger Concentrate SIBX (4.5) 3477 (2) DF (A.N.) Scavenger Cleaner t=10’

4 Rougher Recleaner Concentrates

Discard Scavenger Cleaner Tailing 3 Scavenger Cleaner Concentrates Fig. 7. Breakthrough cleaner circuit.

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N.O. Lotter et al. / Minerals Engineering xxx (2013) xxx–xxx

concentrate: 62.7%; rougher recleaner tailings: 67.2%), dominated by bornite (scavenger concentrate: 58% and rougher recleaner tailings: 75%), suggesting that regrinding before cleaner flotation/ retreatment would improve performance. Discussions in the review group led to a reformulation of the Breakthrough Cleaner Circuit with expectations of even better performance. The circuit was named the ‘‘Sandton Cleaner Circuit’’, and underwent several iterations and revisions before its final formulation as the ‘‘Milestone Flowsheet’’, shown in Figs. 9a and 9b, with results in Table 8 and Fig. 10. The salient features of the changes are:

Cumulative Copper Grade %

45 40 35 30 25 20 15 10 Prototype 5

Breakthrough

0 20

30

40

50

60

70

80

90

100

Cumulative Copper Recovery % Fig. 8. Copper grade-recovery curves for the prototype and breakthrough cleaner circuits.

‘‘B’’ indicating the substantial copper loss to the cleaner tailings. The Breakthrough Cleaner Circuit delivered a smeltable concentrate at point ‘‘C’’, or 28.17% Cu grade with 83.43% recovery, an advance of 17.85% over the Prototype, and still delivered on-spec grade in excess of 28% Cu. It was concluded that the recoverable copper minerals in the Rougher Cleaner Tailing were probably slow-floating and that they responded well to their retreatment in the Scavenger Cleaner, where less fast-floating copper species would be found. This was a key interpretation. 3.6.3. Milestone circuit Various improvements to the Breakthrough Cleaner Circuit were considered by the project team during several review meetings. These concepts had to be strengthened by quantitative mineralogy performed on samples of the laboratory scale concentrates and cleaner tailings in the Breakthrough Cleaner Circuit. Key information was provided by QEMSCAN measurement of the Scavenger Concentrate and Rougher Recleaner Tailings from the Breakthrough Cleaner Circuit. This information showed that both of these streams carried poorly liberated fine-grained copper sulphides (scavenger

1. The Rougher Cleaner Tailings are scavenged on their own in a new flotation bank, then discarded. The arising concentrate was presented to the Rougher Recleaner together with the Rougher Cleaner Concentrate. 2. The Rougher Recleaner Tailings and Scavenger Concentrate were reground in an IsaMill to a p80 size of 10 lm before presentation to the Scavenger Cleaner. Niche copper collectors Cytec 3894 and 5100 were added at this point in the IsaMill. The similarities in degree of copper sulphide liberation and in copper speciation, with bornite dominating the copper sulphides present, led to the conclusion that these two streams would best be processed together. 3. Lengthening of the flotation time in the Scavenger Cleaner bank from 10 to 15 min. 4. Addition of a Scavenger Recleaner bank to reduce the entrainment of silicates in the Scavenger Cleaner Concentrate.

3.6.4. Analysis of scavenger tailings losses from the milestone circuit The Scavenger Tailings losses from the Milestone Circuit were measured by QEMSCAN to inform the project of further processing opportunities. These tailings graded 0.50% Cu and carried 11.71% of the mill head copper recovery. A size-by-size mass and value balance, shown in Table 9, was first constructed by separating a weighed sample across a series of test sieves and the cyclosizer. This analysis quickly showed that the two topsize classes 75 + 53 lm and cyclosizer sizes CS 1–2 (between 38 and 53 lm) carried a concentration of copper that is disproportionate

SIBX (42.6) 3477 (24) DF (38) SIBX (14.2 3477 (8) DF (12.5)

Ore

t = 20’ 80% - 75 microns

Rougher Concentrate 80% - 38 microns

SIBX (23.2) 3477 (13.1) DF (25)

t = 15’

Scavenger Tailings Scavenger Concentrate Fig. 9a. Milestone MF2 circuit.

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N.O. Lotter et al. / Minerals Engineering xxx (2013) xxx–xxx

Rougher Concentrate SIBX (4.5) 3477 (2) DF (A.N.) Rougher Cleaner t=10’ IsaMill to 15 microns

Rougher Cleaner Scavenger t=10’

Rougher Recleaner t=10’

Discard Rougher Cleaner Scavenger Tailing

Scavenger Concentrate SIBX (4.5) 3477 (2) DF (A.N.) Scavenger Cleaner t=15’ IsaMill to 10 microns

4 Rougher Recleaner Concentrates

Scavenger Recleaner t = 10’

Scavenger Cleaner t=10’

3 Scavenger Recleaner Concentrates

Discard Scavenger Cleaner Tailing

Discard Scavenger Recleaner Tailing

Fig. 9b. Milestone cleaner circuit.

locked, 38.1% occur in these topsizes. Further work to address this opportunity was identified.

Table 8 Grade and recovery data for the Milestone Flowsheet (Fig. 9). Cum. stream

RRC1 RRC1. . .RRC2 RRC1. . .RRC3 RRC1. . .RRC4 RRC1. . .SRC1 RRC1. . .SRC2 RRC1. . .SRC3 RRC1. . .RCT RRC1. . .RCST RRC1. . .SCT RRC1. . .ST

Mass pull%

2.5 4.7 5.9 6.7 7.3 7.9 8.6 13.6 21.8 22.9 100.0

Copper Grade % Cu

Recovery % Cu

39.3 40.15 39.49 37.39 36.56 35.12 32.78 20.94 13.26 12.66 3.29

29.8 57.77 70.31 75.66 81.64 84.33 85.37 86.36 87.96 88.29 100.00

to the total masses in these size classes. It was probable that these sizes also carried these copper sulphides as locks. QEMSCAN measurement that followed showed the liberation pattern shown in Table 10. This analysis shows that, for total copper sulphides lost to Scavenger Tailings, 59.0% are locked, 25.0% are middlings, and 16.0% are liberated. The copper sulphide liberation pattern suggests that the locking is concentrated in the topsizes 75 + 53 lm and CS 1–2. Overall, of the 59.0% of all copper sulphides present which are

3.7. Supergene geomet unit The Milestone Flowsheet thus delivered a promising result for the hypogene geomet unit. At this stage of the project, the supergene geomet unit was introduced to the testing programme, and mineralogically characterised, then treated through the Hypogene Milestone Flowsheet with no variations in the process. The results of this work follow.

3.7.1. Supergene mineralogy The modal mineralogy of the supergene ore sample is shown in Table 11, with the copper speciation, in Table 12. The copper mineralogy is dominated by secondary copper sulphides, especially chalcocite. Table 13 shows the size-by-size liberation data for the copper sulphides in the supergene ore milled to rougher Float Feed size (p80 = 75 lm). These data confirm the hypogene pattern of incomplete liberation at the Rougher Float Feed stage, and emphasise the importance of successful middlings flotation.

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N.O. Lotter et al. / Minerals Engineering xxx (2013) xxx–xxx Table 12 Distribution of total copper by mineral type: supergene.

45

Cum. Grade % Cu

40 35 30 25 20 15

Mineral

%

Bornite Chalcocite Chalcopyrite Covellite Azurite Other

2.2 78.9 0.2 7.7 9.3 1.7

Prototype

10

Breakthrough

5

Milestone

0 20

30

40

50

60

70

80

90

100

Cum. Recovery % Cu Fig. 10. Comparative grade-recovery curves for the prototype, breakthrough and milestone cleaner circuits.

Table 13 Liberation data: copper sulphides in supergene rougher float feed. Mineral

Size Class

Locked

Middling

Liberated

Overall copper sulphides

+106 106 + 75 75 + 53 CS1–2 CS3 CS4–5 CS6 CS7 Total

6.00 5.51 4.87 4.76 1.94 1.14 0.29 0.20 24.71

1.06 2.04 2.90 5.94 3.07 2.76 2.60 1.78 22.15

0.22 1.19 2.58 14.65 8.68 14.68 6.63 4.53 53.14

Table 9 Size-by-size copper losses in hypogene scavenger tailings. Size class (lm)

Mass (%)

Grade Cu %

Dist’n % Cu

75 + 53 CS1–2 CS3 CS4–5 CS6 CS7

16.89 7.82 10.70 14.93 15.47 34.20

1.04 0.96 0.64 0.45 0.30 0.33

32.19 13.76 12.55 12.31 8.51 20.68

Table 14 Grade and recovery data for the Milestone Flowsheet. Cum. stream

Table 10 Size-by-size copper sulphide liberation in hypogene scavenger tailings. Size class (lm)

Locks

Middlings

Liberated

75 + 53 CS1–2 CS3 CS4–5 CS6 CS7 Total

26.69 11.41 8.61 7.19 1.57 3.55 59.02

2.70 2.95 2.74 4.49 3.72 8.43 25.02

0.69 0.49 1.04 1.15 3.85 8.73 15.96

Table 11 Modal mineralogy: supergene. Mineral

%

Mineral

%

Bornite Chalcocite Chalcopyrite Covellite Pyrite Azurite Pyroxenes/amphiboles Chlorite Quartz

0.2 0.1 4.8 0.4 0.0 0.6 0.1 14.8 37.3

Muscovite Biotite Kaolinite Orthoclase Plagioclase Fe–Ti oxides Carbonates Apatite Other

10.3 0.6 2.4 22.3 0.8 4.9 0.0 0.0 0.3

3.7.2. Supergene flotation behaviour in the Milestone Flowsheet The metallurgical performance of copper in the Milestone Flowsheet is shown in Table 14 as cumulative grade-recovery data. Successful smeltable concentrate was produced at a grade of 45.11% Cu with a recovery of 83.20% using identical flowsheeting conditions to those developed for the hypogene geomet unit. The two grade-recovery curves are shown in Fig. 11. 4. Discussion The key principle of modern Process Mineralogy is to develop a single flowsheet that captures the range of processing needs of the

RRC1 RRC1. . .RRC2 RRC1. . .RRC3 RRC1. . .RRC4 RRC1. . .SRC1 RRC1. . .SRC2 RRC1. . .SRC3 RRC1. . .RCT RRC1. . .RCST RRC1. . .SCT RRC1. . .ST

Mass pull%

1.9 3.2 3.9 4.4 5.2 5.9 6.9 11.2 19.0 20.5 100.0

Copper Grade % Cu

Recovery % Cu

69.80 68.41 64.35 59.96 56.16 52.09 45.11 27.96 16.68 15.56 3.73

34.93 58.02 67.35 70.91 78.76 82.06 83.20 83.95 85.02 85.29 100.00

main geomet units in the resource, and to deliver therefrom saleable, or treatable, final concentrates on a sustainable basis. For the samples tested to date, this has been achieved for the Kamoa project. The fact that the Milestone Flowsheet was developed and validated within ten months of the first site visit to view drill-core and formulate first sample material reflects the value delivery of Modern Process Mineralogy. As interesting is the fact that exactly the same collector suite succeeded in floating two very different copper mineral mixtures. The 64:36 w/w mixture of SIBX and 3477 seems to cover the mixed potential needs of both the primary chalcopyrite and the secondary chalcocite, bornite and covellite in both hypogene and supergene. Further work to investigate this feature has been identified. Further recovery opportunity has been identified in the size distribution characteristics of the Scavenger Tailings, where the two topsize classes in the distribution show enrichment of copper grades as locked particles. Should the grinding arrangements at the secondary mill be improved to reduce the mass reporting to these sizes before scavenger flotation, an incremental recovery opportunity may be realised. Variability testing on other samples of these geomet units is an obvious next step. Using the Milestone Flowsheet, new samples taken from particularly the first five to ten years of planned mining would be taken and characterised.

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N.O. Lotter et al. / Minerals Engineering xxx (2013) xxx–xxx

80

Cum. Grade % Cu

70 60 50 40 30 Hypogene

20

Supergene

10 0 0

10

20

30

40

50

60

70

80

90

100

Cum. Recovery % Cu Fig. 11. Grade-recovery curves for hypogene and supergene geomet units using the Milestone Flowsheet.

5. Conclusions This work has described and demonstrated modern flowsheet development practice for the Kamoa project, and has shown that this practice delivers a single flowsheet that successfully treats both geomet units known in the resource. Grades and recoveries of smeltable concentrates for both geomet units have been demonstrated and reported. The use of mixed collectors has been key to attaining a suitably high MF2 platform recovery by flotation of liberated and middling copper sulphide particles. Further work to advance the primary MF2 recovery has been identified. 6. Further work Further work to advance this flowsheet further has been identified and scoped. This includes: 1. Deeper study of the mixed collector system with emphasis on the effect(s) on successful middling particle flotation as a result of changes to the collector mixture. 2. Investigation of ways and means of reducing the existing topsize in the Scavenger Tailings in the secondary grinding stage so as to reduce the copper losses to that stream. 3. Variability testing has been planned and provided for so as to align the testing programme with the first five to ten years of planned mining to characterise both comminution and flotation responses. This will focus on the open pit, Kansoko Centrale, Kamoa Sud, and Kamoa Nord. A second phase will sample and test all remaining areas of the resource.

Acknowledgements The authors would like to sincerely thank the management of Ivanplats Ltd and Xstrata Process Support for permission to publish this paper. The technicians who performed the flotation testwork, Mr. Scott Holden, and Ms. Tanya Rainville, are recognised and thanked for their excellent contributions. Ms. Patricia Stack, technician, played a valuable role in the polished section preparation and subsequent interface with QEMSCAN. References Adkins, S.J., Pearse, M.J., 1992. The influence of collector chemistry on the kinetics and selectivity in base metal sulphide flotation. Minerals Engineering 5 (3–5), 295–310. Allison, S.A., Goold, L.A., Nicol, M.J., Granville, A., 1972. A determination of the products of reaction between various sulfide minerals and aqueous xanthate solution, and a correlation of products with electrode rest potentials. Metallurgical Transactions 3 (1972), 2613–2618.

Bhaskar Raju, G., Khangaonkar, P.R., 1984. Electroflotation of chalcopyrite fines with sodium diethyldithiocarbamate as collector. International Journal of Mineral Processing pp. 211–221. Box, G.E.P., Hunter, W.G., Hunter, J.S., 1978. Statistics for Experimenters. Wiley. Bradshaw, D.J., 1997. Synergistic Effects Between Thiol Collectors in the Flotation of Pyrite, (Unpublished doctoral thesis). University of Cape Town, Cape Town, South Africa. Bristette, M., Roman, E., 2012. Concentrate grade and regrind size improvement on the east pit hypogene ore at Kemess mine. In: Proc. Canadian Mineral Processors, Ottawa, pp. 27–38. Bulatovic, S.M., Wyslouzil, H., Kant, C., 1998. Operating practices in the beneficiation of major copper/molybdenum porphyry copper plants from Chile: a review. Minerals Engineering 11 (4), 313–331. Chander, S., Fuerstenau, D.W., 1974. The effect of potassium diethyl dithiophosphate on the electrochemical properties of platinum, copper and copper sulfide in aqueous solution. Journal of Electroanalytical Chemistry and Interfacial Chemistry 56, pp. 217–247. Deeplaul, V., Bryson, M., 2004. Mintek, a National resource of minerals processing expertise for platinum ores. In: Proc. International Platinum Conference ‘‘Platinum Adding Value’’, S. Afr. Inst. Min. Metall., 2004, pp. 9–14. Finkelstein, N.P., and Goold, L.A. The reaction of sulfide minerals with thiol compounds, Mintek report No. 1439, 1972. Fuerstenau, M.C. Thiol collector adsorption processes, Journal Unknown, 1990. Friedland, R., Broughton, D.W., 2009. Keynote address, 8th World Copper Conference, Santiago, Chile, April 2009. Fuerstenau, M.C., 1978. Adsorption Phenomena – Sulfhydryl Collectors. In: King, R.P. (Ed.), Principles of Flotation. Preprint: SAIMM, 1978, pp. 431–466. Grubbs, F.E., 1969. Procedures for detecting outlying observations in samples. 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Lotter, N.O., and Munro, H.C., 1994. The Development of High-Confidence Flotation Testing at Rustenburg Platinum Mines Ltd., Min. Met. Managers Association, Circular 1/94, pp. 29–50. Lotter, N.O., Whittaker, P.J., Kormos, L.J., Stickling, J.S., Wilkie, G.J., 2002. The development of process mineralogy at Falconbridge Ltd., and application to the Raglan Mill. CIM Bulletin 95 (1066), 85–92. Lotter, N.O., Monnapula, R., Oliveira, J., Fragomeni, D., Bradshaw, D.J., 2011. Formulation and plant trial of a mixed collector for Eland Platinum. In: Proc. Canadian Mineral Processors, Ottawa, January 2011, Paper No. 10., pp. 161–183. Nel, E., Theron, J.P., Martin, C.J., Raabe, H. PGM Processing at Impala’s UG2 concentrator in Rustenburg, South Africa. In: Proc. Canadian Mineral Processors, 2004, Paper No. 8. Pease, J.D., Curry, D.C., Barnes, K.E., Young, M., Rule, C., 2006. Transforming flow sheet design with inert grinding – the IsaMill. In: Proc. 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Please cite this article in press as: Lotter, N.O., et al. Flowsheet development for the Kamoa project – A case study. Miner. Eng. (2013), http://dx.doi.org/ 10.1016/j.mineng.2013.02.014