Namibia, and the large-scale refining of gold by SX at Harmony Gold, South Africa. Several other flowsheets that use SX

technology are currently under commissioning, development, or feasibility study for implementation in this part of the world,

Solvent extraction (SX) has been an integral part

(as a by-product of gold mining in South Africa) was

the first major commercial application of SX tech-

nology in the hydrometallurgical industry. Following

Hydrometallurgy 78 (20of the hydrometallurgists arsenal in southern Africaincluding those for cobalt, nickel, vanadium, tantalum, and niobium.

A review of SX operations in the African subcontinent is presented, with particular attention paid to advances since the turn

of the millennium. Several interesting projects under development are also discussed, along with some innovative concepts in

flowsheet chemistry that should soon reach commercial application.

D 2005 Elsevier B.V. All rights reserved.

Keywords: Solvent extraction; Review; South Africa; Botswana; Uganda; Zimbabwe; Zambia; Namibia; Democratic Republic of Congo;

Madagascar; Copper; Nickel; Cobalt; Zinc; Tantalum; Gold; Precious metals; Uranium

1. Introduction for many decades. In the 1950s, uranium recoverySolvent extraction in southern Africa:

An update of some recent hydrometallurgical developments

Kathryn C. Solea,*, Angus M. Featherb, Peter M. Colec

aAnglo American Research Laboratories, P. O. Box 106, Crown Mines 2025, South AfricabCognis Corporation, P.O. Box 361, Honeydew 2040, South Africa

cMatomo Projects (Pty) Ltd., P. O. Box 9021, Edenglen 1613, South Africa

Received 17 August 2004; received in revised form 8 November 2004; accepted 19 November 2004

Abstract

Southern Africa was the site of one of the first large solvent-extraction (SX) plants built, following smaller plants in the

North American uranium industry and the Ranchers and Bagdad copper plants in Arizona. The copper Tailings Leach Plant at

Nchanga, Zambia, was commissioned in 1973 with a capacity of 2800 m3/h. This was the largest SX plant in the world for more

than a decade and is still operational today. South Africa witnessed the first commercial implementation of SX for the refining

of the platinum-group metals. More recently, southern Africa has seen the implementation of SX for other base metals, precious

metals, and specialty metals. These include the bworld firstsQ of primary production of zinc using SX by Skorpion Zinc in0304-386X/$ - see front matter D 2005 Elsevier B.V. All rights reserved.

doi:10.1016/j.hydromet.2004.11.012

* Corresponding author.

E-mail address: ksole@angloresearch.com (K.C. Sole).05) 5278

www.elsevier.com/locate/hydrometess of copper SXthe commercial and technical succat the smaller Ranchers Bluebird and Bagdad plants

in Arizona during the late 1960s, the construction of

constructed on concrete pillars approximately 8 m

high.

metalthe Nchanga plant in 1973 in the Zambian Copper

Belt marked the beginning of a new era for large-

scale SX operations. This plant was the worlds

largest SX plant for more than a decade and is still

operational today. More recently, other large copper

SX circuits have come on-line in Zambia, but South

American and Arizona installations have overtaken

Zambia in terms of volumes treated and copper

produced.

The first primary uranium producer in southern

Africa was Rossing Uranium, Namibia, which incor-

porates one of the first large-scale SX plants to be

built. Commissioned in 1976, this plant today still

produces some 3200 t/a U3O8. At the other end of

the scale, the commercial refining of the platinum-

group metals (PGMs) by SX was first implemented at

Rustenburg, South Africa, in the early 1980s. In this

case, the solution volumes and equipment are small,

but the process chemistry are complex and elegant,

and the products extremely valuable.

In more recent years, southern Africa has seen the

implementation of this technology for other base

metals, precious metals, and specialty metals. These

include the bworld firstsQ of primary production ofzinc using SX by Skorpion Zinc in Namibia, and the

large-scale refining of gold by SX at Harmony Gold,

South Africa, using the Minataurk Process. Both ofthese processes yield a metal product of purity ex-

ceeding 99.99%a testament to the flexibility and

robustness of modern SX chemistry and engineering

in achieving excellent separations and high-purity

products.

During the last two decades, many diverse hydro-

metallurgical SX installations have been commis-

sioned in southern Africa. Several flowsheets that

use SX processes are currently under commissioning,

recommissioning, development, or feasibility study

for implementation in this part of the world, including

those for cobalt, nickel, vanadium, tantalum, and

niobium.

This review discusses the status of commercial SX

operations in the African subcontinent, with particular

emphasis on some of the more innovative develop-

ments that have taken place in recent years. Several

interesting projects are also presented, along with

some novel concepts in flowsheet chemistry that

K.C. Sole et al. / Hydroshould soon reach commercial application. The re-

view has been structured according to commodity:The combined PLS flowrate is approximately 2800

m3/h. Unusually, all extraction stages are run aqueous

continuous, mainly due to restrictions in organic

pumping capacity, no organic recycle facility, and

the presence of some solids in the PLS from the

agitation leach. All trains have a stripped organicthe metals considered are copper, cobalt, nickel,

zinc, precious metals, uranium, tantalum, and niobi-

um. In cases where a particular operation produces

more than one product via SX (e.g., copper and

cobalt), it is discussed under the category of the

primary product.

2. Copper

2.1. Konkola Copper Mines, Zambia

The Tailings Leach Plant (TLP) at Konkola Copper

Mines Nchanga facility came into production in 1971

to treat both current and stockpiled flotation tailings.

In 1973, an integrated solvent extraction-electrowin-

ning (SX-EW) plant, constructed by Davy Powergas,

was commissioned to replace the copper cementation

process in Kennecott cones. The plant is currently

owned by Zambian Copper Investments Ltd.

The TLP currently produces some 80000 t/a of

copper cathode via the flowsheet shown in Fig. 1.

Flotation tailings are dewatered before being leached

using sulphuric acid and SX raffinate. Liquid/solid

separation is achieved by counter-current decantation

(CCD). Modernisation of this circuit, completed in

December 2003, was expected to increase copper

recovery by 7%.

After clarification in a thickener, the pregnant leach

solution (PLS) is sent to SX. The SX plant comprises

four trains, each with three extraction (3E) and two

strip (2S) stages. All stages have a single mixing

compartment; the extraction mixers have a volume

of 55 m3 and the strip stage mixers are 82 m3. Due

to the age of the plant, the settlers are a little unusual

by modern standards, being of the long, thin type

often seen on uranium SX plants (3612.5 m onextraction and 2712.5 m on strip). The units are

lurgy 78 (2005) 5278 53tank rather than the loaded organic tank that is more

common on recently constructed plants.

para

& Fil

n Le

& Fi

xtra

winn

para

& Fil

para

& Fil

n Len Le

& Fi & Fi

xtraxtra

winnwinn

ach P

metalTreating a PLS containing 3 to 4 g/L Cu, two of the

trains run organic phases containing 11 vol.% LIX

984N (Cognis) in Shellsol 2325 (a partially aromatic

diluent supplied by Shell Chemicals), while the other

two run 12 vol.% Acorga M5774 (Cytec) in the same

diluent. The decision to employ extractants from two

Tailings

Dam

Feed pre

Thickening

Agitatio

Thickening

Solvent E

Electro

Tailings

Tailings

Dam

Tailings

Dam

Feed pre

Thickening

Feed pre

Thickening

AgitatioAgitatio

ThickeningThickening

Solvent ESolvent E

ElectroElectro

Tailings

Fig. 1. Simplified flowsheet of the Tailings Le

K.C. Sole et al. / Hydro54vendors is based on strategic, rather than technical,

considerations.

To ensure the production of a high-quality cathode,

the electrolyte fed to the tankhouse is cleaned of solids

(crud and particulate matter) and entrained organic

phase. A two-stage electrolyte cleaning circuit com-

prises two Cominco flotation columns operating in

either parallel or series, followed by six Natco anthra-

cite filters operating in parallel. The Cominco col-

umns were designed to process 800 m3/h of

electrolyte in either a parallel or series arrangement

and to reduce the solids by 25% and entrained organic

phase by 85%, respectively. The final solids and

organic concentrations in the electrolyte are 20 to 50

ppm and 20 to 100 ppm, respectively. The six Natco

anthracite/garnet/sand filters operate in parallel and

process up to 800 m3/h of advance electrolyte.

The most significant process issue in the SX plant

is the formation of crud. This is attributed to the lack

of adequate clarifying facilities downstream from the

agitation leach: most large copper SX plants use heap

leaching to solubilise the copper. The suspendedsolids concentration in the PLS is typically in the

range of 30 to 50 ppm, compared to b20 ppm in heapleach operations. Both bottom and interfacial crud are

prevalent in the settlers. The crud is treated in a newly

installed Flottweg tricanter centrifuge. This produces

a relatively dry solids stream as well as clear aqueous

Flotation tailings

tion

tration

ach

ltration

ction

ing Copper cathode

Raffinate

bleed

Flotation tailings

tion

tration

tion

tration

achach

ltrationltration

ctionction

inging Copper cathode

Raffinate

bleed

lant (TLP) at Konkola Copper Mines, Zambia.

lurgy 78 (2005) 5278and organic phases. Solids are disposed of and liquids

returned to the SX circuit. Organic losses are pre-

dicted to be significantly lower in the future following

the installation of the centrifuge.

2.2. Bwana Mkubwa Mining, Zambia

The original Bwana Mkubwa copper mine in the

Zambian Copper Belt started operations in the early

1900s. In 1996 it was purchased by First Quantum

Minerals (FQM). A small plant with agitation leach,

followed by SX-EW was constructed and commis-

sioned in 1998. By processing old tailings, this plant

produced approximately 10000 t/a of copper cathode.

In 2000, the rights to mine copper ore at the Lonshi

deposit in the Democratic Republic of Congo (DRC)

were secured by FQM. By the end of 2002, the Bwana

Mkubwa plant had been expanded to treat the Lonshi

ore, with the major hydrometallurgical capital expen-

diture being for a second SX train, another EW tank-

house, and a four-stage CCD. Bwana Mkubwa also

has two sulphur-burning acid plants to produce the

acid required for the operation. Excess acid is sold to

other local mining operations. The plant currently

produces approximately 40000 t/a of copper cathode,

and is currently one of the lowest cost copper produ-

cers (Minesearch, 2004).

The flowsheet is illustrated in Fig. 2. The ore is

milled and dewatered before being leached in a cas-

cade of four reactors using sulphuric acid and raffinate

from the high-grade SX plant. The leach solution is

separated from the residue in a thickener and clarified

in a Bateman pinned-bed clarifier before reporting to

the 2E-1W-1S high-grade SX circuit. All of the stages

have two mixing compartments and modern stainless

steel reverse-flow settlers. As the ore treated at Bwana

Mkubwa has a high copper grade, the high-grade PLS

can contain up to 10 g/L Cu. The organic phase

employed is 25 vol.% LIX 984N in Shellsol 2325.

The leach thickener underflow is washed in five

2325, designed to handle a typical PLS composition

of 3.4 g/L Cu at pH 1.7.

2.3. Kansanshi, Zambia

The Kansanshi copper deposit, containing both

oxide and sulphide mineralisation, in the north-west

province of Zambia was aquired by FQM in 2001.

The development of the project is a joint venture

between FQM (80%) and the government-owned

Zambian Consolidated Copper Mines (ZCCM)

(20%). The JV partners are currently constructing a

plant to produce approximately 65000 t/a of copper

cathode by the SX-EW route. The project will also

produce around 25000 oz of gold annually, as well as

a saleable copper sulphide concentrate. The plant is

expected to start producing copper by early 2005.

tion L

icken

trowin

ailing

CCD

U/F

U/F

tion Ltion L

icken

trowintrowin

ailing

CCDCCD

U/F

U/F

K.C. Sole et al. / Hydrometallurgy 78 (2005) 5278 55stages of CCD using raffinate from the low-grade SX

circuit. The washed residue is discarded and the over-

flow from the CCD is clarified in a second Bateman

pinned-bed clarifier before being treated in the low-

grade SX plant. The mixer-settler units in the low-

grade SX plant are the same as the high-grade plant

except that they are constructed of HDPE-lined con-

crete. The plant is operated in conventional series-

parallel mode with a single strip stage. The organic

phase in this circuit is 26 vol.% LIX 984N in Shellsol

Agita

Th

High grade SX

Elec

T

Pinned bed

Clarifier

O/F

O/F

U/F

Raffinate

RaffinateAgitaAgita

Th

High grade SXHigh grade SX

ElecElec

T

Pinned bed

Clarifier

Pinned bed

Clarifier

O/F

O/F

U/F

Raffinate

RaffinateFig. 2. Simplified flowsheet of the copper circ2.4. Mopani Copper Mines, Zambia

During 2004, Mopani Copper Mines (MCM) will

construct and commission two SX-EW plants, one at

Nkana and the second at Mufulira. The Nkana plant

is a small heap leach-SX-EW circuit. Pockets of

copper oxide ore will be mined and leached on

three permanent onoff pads. Copper will be

extracted from the PLS in a single-train 2E-1S SX

plant. Copper cathode will be produced by EW after

each

er

ning Copper cathode

s

Pinned bed

Clarifier

Low grade SX

Ore

O/F

O/F

U/F

Raffinate

H2SO4eacheach

er

ningning Copper cathode

s

Pinned bed

Clarifier

Pinned bed

Clarifier

Low grade SXLow grade SX

Ore

O/F

O/F

U/F

Raffinate

H2SO4uit at Bwana Mkubwa Mining, Zambia.

metalstripping in SX. The initial annual production is

estimated at 3000 t/a of copper cathode. The project

is seen as a test of heap leaching of complete Zam-

bian ores in a high rainfall area.

At Mufulira, MCM will recover copper from their

old underground mining operations by an in situ

leaching programme. The leach produces a PLS of

4 g/L copper at pH 2, with recovery by an SX-EW

route. The SX plant (installed by Sinclair Knight

Merz) is designed for conversion from a conventional

2E-1S configuration to a series/parallel configuration,

allowing greater copper throughput, as more old

stopes are brought into the leaching programme. The

first train of the SX plant has been designed to recover

17500 t/a Cu, at a design volume of 14400 m3/day.

Planning is already underway for expansion of this

project by construction of further SX trains and more

EW tankhouse capacity. Some of the existing electro-

refining tankhouse capacity is being converted to EW

for this project.

The SX-EW facility will also be used to recover

copper produced by vat leaching of other oxide ores.

This may include ores foreign to Mufulira.

At Mopanis Nkana Cobalt Refinery, zinc removal

from the cobalt solution is effected by pH adjustment

of 80% of the ferric-removal thickener overflow and

zinc SX with di(2-ethylhexyl)phosphoric acid

(D2EHPA) for the remainder of the stream. Since

the pH adjustment step contributes significantly to

the overall cobalt loss (presently 20% of the cobalt

from the tankhouse) and the total lime consumption

(30% of the total operating costs), Mopani wish to

discontinue the practice and instead use SX on the

entire overflow stream to remove zinc. The SX plant

is to be sized to have a feed liquor flowrate of 90 to

120 m3/h and will process 150 kg/d zinc.

The recent announcement by MCM of a project

to construct a new smelter and acid plant will ensure

the availability of cheap acid for their leach-SX-EW

projects.

2.5. Zenzele OKiep project, South Africa

A small but novel development is taking place at

OKiep, site of one of the oldest copper mines in

Africa, located near Springbok in the Northern Cape

K.C. Sole et al. / Hydro56province of South Africa. The ore body, an oxidised

deposit of copper carbonates and silicates, has been3. Cobalt

3.1. Kasese Cobalt, Uganda

Kasese Cobalt Company Ltd (KCCL) treats a

cobaltiferous pyrite concentrate stockpiled at the

Kilembe copper mine in Uganda for the recovery of

cobalt, copper, and nickel via a bioleaching route

(Blanchard, 1995; Morin et al., 1996; Fisher and Pav-

lides, 1998). Commissioned in 1999, the plant pro-

cessed approximately 1 Mt/a pyrite until mid-2002,

when production was suspended due to low base metal

prices. The plant was recommissioned in early 2004,

with a production of 1000 t/a cobalt cathode. The

process flowsheet is summarised in Fig. 3.

Following solubilisation of the base metals by

bacterial oxidation, the bulk of the iron is removed

in a two-stage neutralisation circuit. The iron-free

solution is processed through the first SX circuit

where zinc and some manganese are removed using

D2EHPA. After treatment of the raffinate with caustic

soda to remove copper as the hydroxide, the solutionworked out as far as conventional mining is

concerned. An initiative by Zenzele Technology Dem-

onstration Centre (a non-governmental organisation

which assists artisanal and small-scale mining opera-

tions) will extend the life of mine for about 20 years

and benefit the indigenous population at the same

time.

The hand-picked ore, containing about 5% copper,

is crushed in a jaw crusher and then leached in

sulphuric acid to yield a solution containing about

40 g/L Cu. The leach liquor is refined by means of

SX (3E-1S) with 30 vol.% LIX 984N to separate

copper from iron and other impurities. Using technol-

ogy first demonstrated by Zenzele, the loaded strip

liquor (LSL) then becomes the electrolyte in a special

electrochemical cell designed to electroform a variety

of copper artifacts. These include items such as bowls,

ornaments, plaques, and jewellery, which are sold to

the tourist market and for export. This development

provides a unique combination of first-and third-

world technologies, to the benefit of people of both

worlds.

lurgy 78 (2005) 5278passed to a second SX circuit in which cobalt is

selectively extracted from nickel and magnesium

toneing

ch any con

SX

stonerry

toneing

ch any con

SX

stonerry

metalPyritegrinding

Limesgrind

Bioleagravit

ZincCopper hydroxideprecipitation

Limequa

Pyritestockpile

Cu(OH)2

for sale

Pyritegrinding

Limesgrind

Bioleagravit

ZincCopper hydroxideprecipitation

Limequa

Pyritestockpile

Cu(OH)2

for sale

K.C. Sole et al. / Hydrousing CYANEX 272 (di(2,4,4-trimethylpentyl) phos-

phinic acid) (Cytec). This produces an advance elec-

trolyte that reports directly to the cobalt EW circuit to

produce cobalt cathode of N99% purity.The spent scrub liquor from the cobalt SX circuit

and a bleed from the cobalt EW tankhouse contain

significant quantities of cobalt. This combined stream

is treated for cobalt recovery by precipitation of

Co(OH)2 using NaOH. The raffinate from the cobalt

SX circuit is treated for the recovery of nickel as

Ni(OH)2 at pH 6 to 10. Effluents from the zinc SX

and cobalt and nickel precipitation circuits are neu-

tralised with lime and disposed of in a tailings dam.

3.1.1. Zinc SX

The typical feed to the zinc SX circuit comprises

0.012 g/L Zn, 3.5 g/L Co, 0.12 g/L Mn, 0.1 g/L

Cobalt SX

Cobalt EW

Cobalt cathodeconditioning

Cobalt effluetreatmen

Nickel hydroxprecipitatio

Co(OH)2for sale

Cobalt cathodefor sale

Cobalt SX

Cobalt EW

Cobalt cathodeconditioning

Cobalt effluetreatmen

Nickel hydroxprecipitatio

Co(OH)2for sale

Cobalt cathodefor sale

Fig. 3. Flowsheet for the recovery of cobalt, copdc.

NeutralisationpH 2.8

Iron removalpH 5

dc.

NeutralisationpH 2.8

Iron removalpH 5

lurgy 78 (2005) 5278 57Cu, 0.2 g/L Ni, and 0.04 g/L Fe (Ellis, 2001; Cole

and Sole, 2003). There are two extraction, one

scrub, and two strip stages. The extractant is 2

vol.% D2EHPA. The extraction is controlled at pH

2.5 to 3.5 by the addition of NaOH, removing zinc to

less than 0. 5 mg/L. The D2EHPA circuit also serves to

control the levels of manganese reporting to the cobalt

circuit. Cobalt losses are minimised by controlling the

scrubbing stage at pH 2.8. The raffinate passes through

a Jameson (flotation) cell and an after-settler to allow

any entrained organic phase to be recovered.

The main problem associated with the D2EHPA

circuit is control of the upstream iron-precipitation

process. The SX operation can handle up to 500

ppm suspended solidssolids present at higher con-

centrations lead to crud formation. Inefficiencies in

the iron-precipitation circuit periodically result in up

ntt

Liquid effluenttreatment

iden

to tailings dam

Ca(OH)2

Ni(OH)2for sale

ntt

Liquid effluenttreatment

iden

to tailings dam

Ca(OH)2

Ni(OH)2for sale

per, and nickel by Kasese Cobalt Co. Ltd.

to 0.3 g/L iron reporting to the zinc SX circuit. As all

iron(III) is co-extracted by D2EHPA, the iron loading

is controlled to 30% to 40% by the addition of sodium

sulphite, which reduces some of the iron to Fe(II)

per day. To recover entrained organic, the raffinate is

passed through an after-settler and Jameson cell. A

Table 1

Typical composition of feed to Kasese cobalt SX circuit

Element Co Cu Fe Mg Mn Na Ni Zn

(g/L) 3.0 b0.001 b0.001 1.5 0.15 0.5 0.2 b0.005

Co PLS

nt scrub

iquor to

H)3 circuit

80 g/l NaOHCo PLS

nt scrub

iquor to

H)3 circuit

80 g/l NaOH

Table 2

KCCL cobalt SX circuit

Extraction Scrub Strip High-acid

strip

Organic flowrate

(m3/h)

20

Aqueous phase PLS Catholyte

diluted 50%

Anolyte 130 g/L

H2SO4Advance O:Aa 0.66 10 1 135

pH pH 5.45.6 pH 3 1.5 g/L

H2SO4

K.C. Sole et al. / Hydrometallurgy 78 (2005) 527858which is not extracted at operating pH values.

3.1.2. Cobalt SX

The PLS to the cobalt SX circuit has the compo-

sition shown in Table 1 (and a temperature of 40 8C).A counter-current flow configuration (Fig. 4) is

employed, using conventional mixer-settler units for

contact of the phases. The mixer settlers are Krebs

units, manufactured of glass fibre and are completely

enclosed to limit the evaporation of the organic phase.

The extractant is CYANEX 272, made up to 7 vol.%

concentration in the diluent. A summary of the circuit

configuration and operating conditions is given in

Table 2.

The extraction circuit produces a consistent raffi-

nate concentration of b0.01 g/L Co. There is a con-siderable concentration of silica in the circuit and it is

necessary to operate the extraction mixers in organic-

continuous mode to minimise crud formation.

The scrub liquor originally comprised 20 to 25 g/L

Co in ~4 g/L H2SO4, however the scrub circuit has

since been taken off-line. The reason for this is that

the spent scrub liquor was not returned to the extrac-

tion circuit, as is common in many circuit configura-

tions, but serves as a bleed from the EW circuit (Fig.

4). This stream contained a considerable quantity of

Spent

electrolyteSpe

l

Co(O

130 g/l

H2SO4

Spent

electrolyteSpe

l

Co(O

130 g/l

H2SO4Advance

electrolyteStripped

organic

Sc1S1S2S3W1

Advance

electrolyteStripped

organic

Sc1S1S2S3W1

Fig. 4. Cobalt SX cicobalt (~30 g/L) along with the co-extracted magne-

sium and manganese. The cobalt was recovered from

the spent scrub liquor by precipitation as cobalt hy-

droxide with NaOH at pH 10. A decision was made to

maximise the cobalt cathode production, albeit at the

expense of purity, and hence the scrub circuit has not

been operating.

The strip liquor (cobalt anolyte) contains typically

47 g/L Co in 5 to 10 g/L H2SO4 at a temperature of

~65 8C. The LSL has a cobalt concentration of 50 g/L.This is passed through a Jameson cell and after-settler

to recover entrained organic phase. The final high-

acid strip (W1) ensures that trace quantities of iron,

copper, and zinc entering this circuit are not permitted

to build up on the organic phase.

The main problem experienced with this circuit is

excessive losses of extractant to the raffinate at the pH

values used. There is a very low salt content (~15 g/L)

in the raffinate, which exacerbates the problem. Re-

plenishment of CYANEX 272 is estimated at 300 to

1000 L per month or about 1 vol.% of the inventory

in S1

a O:A=organic:aqueous volumetric flowrate ratio.Raffinate Aqueous

Organic

E1 E2 E3 E4

Raffinate Aqueous

Organic

E1 E2 E3 E4

rcuit at KCCL.

carbon adsorption column has also been introduced,

which should reduce the total dissolved organic car-

bon content to approximately 0.14 g/L. The carbon

column has an operating life of about 6 months, with

the carbon inventory of 1 tonne. It is proposed to

eliminate the final extraction stage (E4) and use this

for organic recovery, as it is believed that adequate

cobalt recovery can be achieved in three stages of

extraction.

3.2. Chambishi Metals Plc, Zambia

The Chambishi cobalt plant, near Kitwe in Zambia,

was commissioned in 1978 and, as part of the govern-

ments privatisation process, was sold by ZCCM to

Anglovaal Mining in 1998. During 2003, the Kazhak-

Uzbek consortium J&W Holding AG became the new

owners.

Chambishi treats two feed materials: a sulphide

concentrate via a roast-leach process and a cobalt-

The initial flowsheet for the refinery (Aird et al.,

1980; Rao et al., 1993) comprised an extensive series

of precipitation steps employing limestone and quick-

lime slurry to sequentially remove iron, copper, zinc,

and nickel from the cobalt electrolyte. The associated

cobalt losses were high; furthermore, the increasingly

stringent demands on the cobalt cathode purity could

not be met. These problems have been greatly allevi-

ated by the implementation of SX and ion exchange

(IX) steps for the removal of zinc and nickel from the

leach solution (Bailey et al., 2001). Chambishi has

recently explored the use of cobalt SX to improve

overall plant performance and copper SX to deal with

a possible increase in copper throughput (Fig. 5).

3.2.1. Zinc SX

In 1991, SX was incorporated into the refinery

flowsheet to control the zinc concentration by treating

a bleed stream from the iron-precipitation circuit. In

2001, the cobalt refinery was upgraded to accommo-

K.C. Sole et al. / Hydrometallurgy 78 (2005) 5278 59rich copper slag via a smelt-pressure oxidation pro-

cess (Munnik et al., 2003). Nominal production fig-

ures are 18000 t/a copper and 7000 t/a cobalt. Copper

is recovered from the leach liquor in sequential EW

and electrostripping circuits that ensure the delivery of

low copper tenor solution to cobalt refining.

Flotation

Roast

Leach

Cu EW

Ni IX

Co purification

Co EWCobalt

cathode

Copper

cathode

Flotation

Roast

Leach

Cu EW

Ni IX

Co purification

Co EWCobalt

cathode

Copper

cathodeFig. 5. Simplifed flowsheet of the current Chambishi circuit (solid lines

et al., 2003).date the increased throughput from the treatment of

the Nkana slag dumps. This brought with it an in-

crease in zinc concentration. A new zinc SX circuit

(Cole and Sole, 2003) was designed to remove nearly

all the zinc, rendering zinc control by pH manipula-

tion obsolete.

PAL of alloy

Slag reduction

Cu SX

Cu IX

Co SX

PAL of alloy

Slag reduction

Cu SX

Cu IX

Co SX) and proposed expansions (dotted lines). (Adapted from Munnik

technology is being considered for dealing with the

upgraded copper throughput.

The technical feasibility of using copper SX as an

alternative to electrolytic copper removal was inves-

tigated and proved by Chambishi in a continuous

counter-current pilot-scale trial. The economics were

metalThe organic phase comprises 2.5 vol.% D2EHPA in

the aliphatic diluent Shellsol K (Shell Chemicals). The

aqueous feed flowrate is 120 m3/h and the advance

O:A is 3 :1. Following extraction in four stages, loaded

zinc and calcium are stripped with 150 g/L H2SO4 in a

single stage. The stripped organic phase is subjected to

a 180 g/L HCl restrip for iron(III) removal.

The circuit is currently run in a double-stage count-

er-current flow configuration, with fresh organic

phase being fed to the fourth and second extraction

stages. Loaded organic phase from the third and first

stages is sent to stripping and thereafter to the HCl

restrip. The reason for this configuration is to mini-

mise the amount of calcium displaced in the earlier

extraction stages by zinc or iron. A further improve-

ment made to counter gypsum precipitation is the use

of a reverse pH profile over the extraction stages.

Dilute caustic solution (3040 g/L) is used to maintain

the extraction pH values at 3.1 to 3.4. Typical zinc

content of the feed solution varies from 50 to 80 mg/

L; the zinc concentration in the raffinate is consistent-

ly lower than 5 mg/L, averaging around 1 mg/L.

3.2.2. Cobalt SX

Under consideration at Chambishi, although in

more longer-term plans, is to convert the cobalt puri-

fication circuit from precipitation technology to SX to

produce a high-purity cobalt electrolyte suitable for

EW. The perceived benefits of implementing this

flowsheet change include reduced operating costs,

improved cobalt recovery, and higher current efficien-

cies. This route will also enable a greater variety of

feedstocks to be treated, with a wider range of impu-

rities, and the high-purity LSL can be used for the

production of alternative value-added cobalt products

(Cowie, 2002).

Some feasibility studies and piloting trials have

been carried out using Ionquest 801 (Rhodia) and

CYANEX 272 (Cytec) for the cobaltnickel separa-

tion. CYANEX 272 has the advantage of having good

selectivity for cobalt over magnesium, calcium, and

nickel, which are the major impurities in the electrolyte

(Table 3). Iron and zinc are present in small quantities

and extract more strongly than cobalt, so these ions

will remain on the loaded organic phase under mild

stripping conditions that allow cobalt to be stripped;

K.C. Sole et al. / Hydro60these can then be stripped separately under harsher

stripping conditions (Cowie, 2002; Sole, 2003). This

favorable for the installation of this technology based

on 18000 t/a copper. In addition, revenue from copperselective strip enables a pure cobalt electrolyte to be

obtained, contaminated only by copper. The trace

amounts of copper will be removed from the electro-

lyte by IX prior to cobalt EW. The organic phase will

comprise 30 vol.% CYANEX 272, modified by 5

vol.% tri-n-butylphosphate (TBP), in an aliphatic dil-

uent, SSX 210 (Sasol Wax). The extractant is subjected

to saponification prior to extraction to minimise the

necessity for stage-wise pH control. It has been shown

that when more than 50% to 70% of the extractant is

converted to the sodium salt, there is a tendency for

third-phase formation to occur. The use of a phase

modifier can be useful under these circumstances.

Another application of cobalt recovery by SX

using CYANEX 272 is under consideration. Nickel

is currently removed from the cobalt electrolyte by IX

using the Dow M4195 bispicolylamine resin in an

ISEP contactor (Bailey et al., 2001). The cobalt losses

associated with the eluted nickel solution could be

mitigated by inclusion of a small SX scavenger circuit

on this stream.

3.2.3. Copper SX

Since Chambishis current delivery of sulphide

concentrate feed is expected to be discontinued, alter-

native source materials are being sought. Under con-

sideration is a concentrate featuring a relatively high

Cu :Co ratio compared to that currently processed that

will result in a doubling of the copper input to the

plant. Copper removal by electrolysis is uneconomic

under these conditions because of the low current

efficiencies characteristic of the process and the poor

quality copper produced. Capital expenditure for dou-

bling the tankhouse capacity would be exorbitant. SX

Table 3

Expected composition of the feed to the proposed Chambishi cobalt

SX circuit

Element Co Cu Zn Ni Fe Mn Ca Mg

(g/L) 10.2 0.4 0.001 0.05 0.05 0.5 0.4 3.3

lurgy 78 (2005) 5278

sales is significantly increased because SX produces

an advance electrolyte from which LME grade copper

can be electrowon. Two SX operations are necessary

to achieve the desired residual copper concentration

suitable for feeding into the cobalt recovery plant

(b100 mg/L). Studies are presently being undertakento install a copper SX circuit to purify the copper

stream, and then to convert the entire tankhouse to

conventional EW. The two PLS streams will contain

approximately 45 and 22 g/L Cu respectively, and

maximum copper recovery is required. For this rea-

son, the new low-viscosity extractants recently avail-

able from Cognis (Sole and Feather, 2003) are under

consideration for this application.

3.3. Knightsbridge Cobalt, South Africa

Several years ago, Knightsbridge Cobalt Corpora-

tion of South Africa began the operation of a refining

plant to produce purified cobalt carbonate. Feed ma-

terial was oxide ore originating in the DRC. After

leaching with sulphuric acid, classical precipitation

methods were used to remove iron and copper. Al-

though a market existed for the resultant carbonate

product, it was known that higher prices could be

realised if impurities such as manganese and magne-

sium were not present. Other undesired impurity ele-

ments such as nickel and zinc would also report to the

product if feed material containing these elements

were treated. When a decision was made to produce

high-grade cobalt oxide the refinery flowsheet needed

to be altered to deal with the array of impurity elements

expected to be present in a varying feedstock. After

extensive piloting, two SX operations were implemen-

ted to overcome these problems (Cole, 2002). The

schematic flowsheet for the improved cobalt refinery

is shown in Fig. 6. Production at the plant was 1 tonne

of cobalt per day. This refinery ceased operation in

2002, but the plant was purchased by Umicore and

relocated to a new site in Krugersdorp, west of Johan-

nesburg, where operations continue today.

3.3.1. Impurity removal SX

Zinc (50 mg/L), manganese (100 mg/L), and calci-

um are removed from the cobalt (5 to 10 g/L) solution,

derived from the leach/precipitation circuit, using 20

R

Iro

Copp

Impu

CoC

R

Iro

Copp

Impu

CoC

K.C. Sole et al. / Hydrometallurgy 78 (2005) 5278 61Fe waste

Cu product

Solution recycle to leach

Fe waste

Cu product

Solution recycle to leachFig. 6. Schematic flowsheet for the cobalLeach H2SO4

CaCO3

CaCO3

CaCO3

NH4OH

H2SO4

NH4OH

H2SO4

aw material

CoCO3

n precipitation

er precipitation

rity removal SX

Cobalt SX

O3 Precipitation

Leach H2SO4

CaCO3

CaCO3

CaCO3

NH4OH

H2SO4

NH4OH

H2SO4

aw material

CoCO3

n precipitation

er precipitation

rity removal SX

Cobalt SX

O3 Precipitationt refinery at Knightsbridge Cobalt.

vol.% D2EHPA in Shellsol K (an aliphatic diluent

supplied by Shell Chemicals) in three extraction stages

operated at an advance O:A of 1 and pH 2.26 to 2.37

(adjusted using 20% ammonium hydroxide solution).

Co-extracted cobalt is recovered using one scrub stage,

run under integrated steady-state conditions for a

further six weeks to generate design data.

In the selected process, the tailings were leached in

a primary circuit to solubilise copper and cobalt, and

copper recovered by SX/EW. A bleed of this circuit

t Kni

C

K.C. Sole et al. / Hydrometallurgy 78 (2005) 527862and stripping (in one stage) using recycled 6 M HCl

ensured the complete removal of iron.

3.3.2. Cobalt SX

In the cobalt SX plant, the organic phase is 15

vol.% CYANEX 272 in Shellsol K. Cobalt extraction

is achieved in five stages operated at an advance O:A

of 1 and pH 5.0 to 5.3. Scrubbing of co-extracted

magnesium is with 40 g/L cobalt solution in two

stages operated at an O:A of 50. Cobalt stripping is

achieved in two stages using 180 g/L sulphuric acid

and an O:A of 10. A loaded strip liquor pH of 4

ensured that co-extracted zinc and iron did not report

with the cobalt. These metals are stripped in a final

stripping stage operated at pHb1.To illustrate the degree of upgrading achieved,

assays of feed solution to the combined SX circuits

(mother liquor from the precipitation) and product

solution from the SX processes (LSL from the cobalt

extraction) are shown in Table 4.

3.4. Kolwezi Tailings, Democratic Republic of Congo

Congo Minerals Development developed a flow-

sheet to recover copper and cobalt from the King-

anyambo and Musonoi tailings originating over the

past fifty years from the Kolwezi copper flotation

concentrator in the DRC (Alexander, 2001). Over

100 million tonnes of the material exist, averaging

1.5% Cu and 0.32% Co, primarily as malachite, pseu-

domalachite, and heterogenite. Nine different flow-

sheets were evaluated for the processing of this

material to produce high-purity copper and cobalt

cathode during an extensive 16-month piloting of

this circuit at Anglo American Research Laboratories

(AARL). The optimised flowsheet (Fig. 7) was then

Table 4

Assays of feed solution to and product solution from the SX steps aFeed solution (mg/L)(PLS following iron and copper precipitation)

Product solution (mg/L) (LSL from cobalt SX) 8was treated in a secondary circuit to remove iron and

manganese by precipitation with air/SO2, followed by

SX with CYANEX 272 to remove zinc; IX using an

aminophosphonic acid cation exchanger to remove

trace copper and zinc, and finally the cobalt stream

was upgraded by SX with CYANEX 272, enabling

high-purity cobalt cathode to be electrowon.

Interestingly, although D2EHPAwas considered for

the removal of manganese and zinc from the cobalt

bleed stream, the final decisionwas in favour of air/SO2precipitation for manganese removal and CYANEX

272 for zinc removal (Alexander, 2001). Manganese

removal with D2EHPA resulted in unacceptable losses

of cobalt at the pH values required for extraction,

needing enhanced scrubbing requirements. The need

for HCl to remove any co-extracted trace iron(III) and

the production of a dilute waste stream to avoid gypsum

formation also mitigated against SX for this operation.

Zinc SX using CYANEX 272 would avoid possibilities

of extractant cross-contamination with the cobalt SX

circuit, while enabling high extraction efficiencies and

low cobalt losses to be achieved without the use of a

scrubbing section. A small volume, concentrated zinc

LSL is produced, while the option to recover zinc as a

saleable by-product remains open.

Current owners Adastra have recently secured fi-

nance for an 18-month feasibility study and preferred

contractors have been selected. This project is

expected to produce 40000 t/a copper and 7000 t/a

cobalt during a first phase, with possible expansion to

double this capacity.

3.5. Kakanda Tailings, Democratic Republic of Congo

A similar processing philosophy has been proposed

by International Panorama Resource Corporation to

ghtsbridge Cobalt (from Cole, 2002)

o Mn Zn Cu Fe Mg Ca Ni4800 100 50 b1 b1 2300 600 340000 60 b1 5 b1 1400 80 b1

gs m

ary

gs m

ary

metalTailin

PrimH2SO4

Tailin

PrimH2SO4

K.C. Sole et al. / Hydrorecover copper and cobalt from the Kakanda tailings

(Dry et al., 1998). The flowsheet shown in Fig. 8 was

piloted at Mintek during 1998. In a primary circuit,

the tailings are leached to solubilise copper and co-

balt, and copper recovered by SX/EW. A bleed of this

circuit was treated in a secondary circuit to produce

3500 t/a cathode cobalt. In contrast to the Kolwezi

Secondar

Wash

Copper SX2

Iron removal

Zinc SX

CYANEX 272

Copper zinc IX

Purolite S950

Residue disposal

to tailings

Fe, Al, Mn

Zn

LS

LS

SO2

Cobalt SX

CYANEX 272

Cobalt EW

Mg, Ca

Belt filter

Air/SO2

Cu, Zn

Wash water

Cobalt cathode

Partially loaded

Secondar

Wash

Copper SX2

Iron removal

Zinc SX

CYANEX 272

Copper zinc IX

Purolite S950

Residue disposal

to tailings

Fe, Al, Mn

Zn

LS

LS

LS

SO2

Cobalt SX

CYANEX 272

Cobalt EW

Mg, Ca

Belt filter

Air/SO2

Cu, Zn

Wash water

Cobalt cathode

Partially loaded

Fig. 7. Preferred flowsheet for thaterial

leach

Cu PLS

aterial

leach

Cu PLS

lurgy 78 (2005) 5278 63flowsheet, however, iron precipitation was followed

by SX with D2EHPA to remove zinc and manganese.

The cobalt stream was first upgraded by SX with

CYANEX 272, and then trace copper and zinc re-

moved by IX ahead of cobalt EW (Preston et al.,

1999; Feather et al., 2000a). This project is currently

seeking finance to proceed to the next stage.

y leach

Copper SX1

Copper EW

Copper cathode

Raffinate

organicStripped

organic

y leach

Copper SX1

Copper EW

Copper cathode

Raffinate

organicStripped

organic

e Kolwezi Tailings project.

would be treated at an existing agitation leach facility

at Chibuluma in Zambia. The process flowsheet is not

nese

removal

alt E

lt cat

Residue

LSL (to waste)

nese

removal

alt E

lt cat

Residue

LSL (to waste)

thode

metallurgy 78 (2005) 5278yet finalised, but may include recovery of copper and

cobalt by SX-EW processes.

4. Nickel

4.1. Tati Nickel, Botswana

Lionores Tati Nickel operates the Phoenix Nickel

Mine and Tati Nickel Concentrator in Botswana. The

concentrates produced are treated offsite and Tati are3.6. Etoile, Democratic Republic of Congo

Another copper-cobalt project experiencing a re-

vival of interest is Etoile, located near Ruashi, DRC.

This deposit is now controlled by Metorex. A feasi-

bility study is currently underway to recover a copper/

cobalt oxide concentrate by flotation. The concentrate

Leach

Copper SX

Copper EW

Manga

Iron

Cob

CobaCopper cathode

Tailings

Leach

Copper SX

Copper EW

Manga

Iron

Cob

CobaCopper cathode

Tailings

Fig. 8. Proposed flowsheet for the recovery of copper and cobalt ca

K.C. Sole et al. / Hydro64currently evaluating a proposed expansion at the

Phoenix operation, which will include the incorpora-

tion of a hydrometallurgical refinery to produce

17000 t/a nickel metal, 8000 t/a copper metal, 1240

t/a cobalt carbonate, and a PGM concentrate. Initial

piloting was carried out in Australia by Western

Minerals Technology and SGS Lakefield Oretest. A

$10 million demonstration plant at 1 /170 scale is

currently operating on site in Francistown, Botswana,

treating 200 kg/h of concentrate and producing 100 t/a

Ni and 60 t/a Cu. This plant is expected to run for the

next three years, during which time all technical pro-

blems can be ironed out, long-term impurity and

degradation effects determined, and optimisation of

equipment, reagent selection, and operating condi-tions undertaken. The plant will also be used for

training of operators for the full-scale plant.

Following Activox leaching of the sulphide flota-

tion concentrate, copper is recovered by SX-EW. A

50% bleed of the copper SX raffinate is treated for

iron removal in two stages using limestone. The first

stage at pH 3.5 provides partial iron removal but

essentially no loss of nickel or cobalt, and the second

stage at pH 4.5 completes the removal of iron but the

precipitate contains significant amounts of cobalt and

nickel that is recycled to the leach. The liquor is then

purified by cobalt SX with CYANEX 272, then nickel

recovery by SX with neodecanoic acid followed by

EW. The cobalt product is initially expected to be

CoCO3, with electrowinning of metal as a later option.

4.1.1. Copper SX

The copper circuit currently employs a 2E-1W-2S

configuration. The wash stage is necessary because of

the addition of chlorides to the leach to assist with

SX

W Copper zinc IX

Cobalt SX

hode Impurities

SX

W Copper zinc IX

Cobalt SX

hode Impurities

from Kakanda dump tailings material (from Feather et al., 2000a).achieving copper extractions under the low tempera-

ture, low pressure Activox conditions. Reagents from

both Cytec and Cognis are under evaluation.

4.1.2. Cobalt SX

The composition of the feed to the Tati cobalt SX

circuit is shown in Table 5. This cobalt SX circuit

comprises three extraction, one scrub, and two strip

stages. The organic phase is 5 vol.% CYANEX 272 in

Table 5

Composition of feed to the cobalt SX circuit for the Tati pilot plan

Element Ca Co Cu Mg Ni

(g/L) 0.5 0.2 0.002 0.3 7.1t

Shellsol D70 (an aliphatic diluent supplied by Shell

Chemicals). The pH profile is optimised to enable a

cobalt recovery of N98.5% to be achieved, with b3.5mg/L Co in the raffinate, while facilitating the rejec-

tion of Ni, Ca, and Mg in the extraction circuit.

Scrubbing is carried out with a solution containing

0.9 g/L Co at pH 2.

4.1.3. Nickel SX

The cobalt SX raffinate is passed through an after-

settler followed by a diluent wash and an activated

carbon step to ensure that no CYANEX 272 leaks into

the nickel circuit. The nickel SX circuit comprises

five extraction, two scrub, and two strip stages. A

Versatic 10 (Shell Chemicals) concentration of 20

vol.% in Shellsol D70 is used. The nickel concentra-

tion in the feed liquor averages between 6 and 8 g/L.

The pH profile across the extraction bank is con-

trolled from pH 7.2 in E1 to pH 6.5 in E5. These

higher pH values allow calcium loading to take place

in the last two extraction stages, and then be scrubbed

off in the first three stages. Nickel recoveries of

98.8% are measured, with raffinate concentrations of

b0.01 g/L Ni.

4.2. Nkomati, South Africa

Nkomati, a nickel sulphide complex in Mpuma-

langa, South Africa, is under development by African

Rainbow Minerals. The preferred downstream flow-

sheet (Fig. 9), piloted by Mintek in 2000 (Feather et

al., 2002a), will be similar to that proposed for Tati. A

feasibility study has been completed for 375000 t/m

run-of-mine ore producing 16500 t/a nickel metal,

Ni/Co solution from

Cu SX raffinate bleed

Iron removal

Cobalt SX

LS

Ni, Mg, CaCobalt strip

25 g/l Co

pH 2.8

CoSO4 product

scrub

H2O

EW

K.C. Sole et al. / Hydrometallurgy 78 (2005) 5278 65Organic

removal

pH adjustment

Nickel SX

Raffinate

NH4OH

scrub

Mg, Ca

NickelNickel cathodeFig. 9. Flowsheet for the recovery of cobalt and nickel in solution

Nickel strip

NiSO4 advance

electrolyte

Organic

removal

spent

electrolyte

Organic flow

Aqueous flowthe Nkomati process (from Feather et al., 2002a).

7100 t/a copper metal, and 940 t/a cobalt as carbonate.

The project is expected to go to EPCM phase in the

second quarter of 2005.

4.2.1. Copper SX

K.C. Sole et al. / Hydrometal66The copper is recovered using Acorga M5640 in a

2E-1W-2S copper SX circuit. The PLS is passed

through sand filters prior to SX to remove solids

and minimise crud formation. The raffinate is split

into two streams, with 80% returning to the leach and

the remaining 20% passing through a third copper

extraction stage to further reduce the copper content

of the solution prior to cobalt and nickel recovery.

After-settlers and co-matrix filters ensure maximum

organic recovery and minimal organic loss to the

leach and EW circuits.

4.2.2. Cobalt SX

The composition of the feed to the cobalt SX

circuit is shown in Table 6. The pilot-plant cobalt

circuit included five extraction, three scrub, and

three strip stages. The organic phase comprised 7

vol.% CYANEX 272 in a paraffin diluent. The pH

of the extraction circuit was controlled between 5.5

and 5.65, with the pH raised towards the end of the

extraction circuit to ensure maximum extraction of

cobalt. Because calcium is present at saturation con-

centrations, it was necessary to minimise co-extrac-

tion of calcium, as its loading and subsequently

stripping would lead to gypsum formation. Co-

extracted nickel, calcium, and magnesium were

scrubbed from the loaded organic phase using a por-

tion of the LSL (~25 g/L Co, pH 2.8) with the pH

controlled between 4.6 and 5.1. Cobalt extraction

efficiencies in excess of 99.5% were measured on

the pilot plant, reducing the cobalt concentration in

the nickel liquor from 1.8 to b0.01 g/L. The overallco-extraction of nickel was minimised to b0.1%, andthe Co :Ni ratio in the LSL upgraded to N1500.

In the full-scale design, Bateman pulsed col-

umns (BPC) are to be used for cobalt extraction,

Table 6

Composition of feed to the cobalt SX circuit for the Nkomati pilot

plant

Element Ca Co Cu Mg Mn Ni Zn(g/L) 0.5 1.9 0.01 3.6 0.3 32.7 0.12mixer settlers for scrubbing, and a further BPC for

stripping.

4.2.3. Nickel SX

Nickel was then extracted from the calcium-satu-

rated solution (cobalt SX raffinate) using 30 vol.%

Versatic 10 (Feather et al., 2002a). The optimised

circuit comprised five extraction, three scrub, three

strip stages, and a single reclamation stage for the

recovery of dissolved versatic acid. Co-extraction of

calcium was minimised by tight control of the pH in

the extraction. The optimised profile ranged from pH

6.4 in E1 to pH 6.0 in E5. Co-extracted calcium and

magnesium were scrubbed from the loaded organic

phase using a portion of the nickel LSL diluted to a

nickel concentration of ~3 g/L under conditions of

controlled pH (pH 5.9). Stripping was carried out

using spent nickel electrolyte (60 g/L Ni, 50 g/L

H2SO4), and nickel recovered using standard divid-

ed-cell EW technology. The nickel SX circuit was

optimised to recover 99% of the nickel, reducing the

PLS concentration from 32 to b0.3 g/L in the raffi-nate, and producing a LSL suitable for nickel EW.

Overall co-extraction of calcium was limited to ~3%.

In the full-scale design, IX using Purolite S950 is

anticipated for complete recovery of nickel from the

raffinate prior to ammonia recovery using a lime boil.

Two different philosophies to avoiding gypsum

formation were adopted in the Tati and Nkomati

flowsheets. In the former, the solutions were diluted

such that calcium saturation did not occur; in the

latter, calcium was prevented from loading by close

and accurate control of the pH profiles of the extrac-

tion and scrub circuits. There is obviously a penalty in

the capital cost of the Tati option, since the equipment

will need to be much larger to accommodate the

equivalent nickel throughput.

4.3. Anglo Platinum Rustenburg Base Metals Refin-

ery, South Africa

Anglo Platinum is the worlds largest producer

of precious metals. The sulphide ore, from the

Merensky and UG2 reefs in the Rustenburg area

of South Africa, is rich in both PGMs and base

metals. This is smelted to give a PGM-containing

lurgy 78 (2005) 5278matte. The non-magnetic component of the matte

contains nickel and copper which, together with the

metalliquor produced in the leaching of the PGM-rich

magnetic component of the matte to dissolve resid-

ual base metals, is the feed material to the Rusten-

burg Base Metals Refinery (RBMR) (Hofirek and

Kerfoot, 1992).

Cobalt has been produced at the refinery using

D2EHPA since 1979 (Clemente et al., 1980). This

was one of the first commercial operations to use

SX for cobalt refining (Ritcey et al., 1975). Although

more selective extractants are now available, the in-

expensive and efficient original process design is still

used. The leach liquor is treated for lead removal

using Ba(OH)2 before cobalt removal using the

Outokumpu nickelic hydroxide process. The purified

solution advances to nickel EW. Cobalt is recovered

from the precipitate by dissolution of the cobalt cake,

removal of residual iron by precipitation with NaOH,

and copper by sulphide precipitation using BaS. The

cobalt solution, which has now had most base-metal

impurities removed, is purified and concentrated by

SX. The aim is to produce a cobalt sulphate solution

with low impurity content which is suitable for crys-

tallisation of the salt as a saleable CoSO4 product.

The SX plant comprises 19 mixer settlers, operated

in a counter-current configuration. Each mixer settler

has a mixing compartment with a volume of 0.8 m3

and a settler of 5 m3 capacity. There are seven extrac-

tion, six scrubbing, and three stripping stages. The

final two stages are used for the removal of trace

impurities from the stripped organic phase and regen-

eration of the extractant. The first mixer settler is used

as a settler only, and ensures that organic entrainment,

and thereby organic losses, in the aqueous phase are

minimised.

The feed to the SX circuit (1518 g/L Co, 58 g/L

Ni, b1 mg/L Cu and Pb, 5 mg/L Fe, and 50 mg/Leach of Mg, Mn, and Ca) is passed through in-line

filters prior to entering the circuit to ensure a low

value of total dissolved solids, thereby avoiding

crud formation. The organic phase comprises 15

vol.% D2EHPA and 5 vol.% TBP in SSX 210. To

control the pH of extraction more effectively and

minimise dilution of the aqueous phases by the

addition a neutralising solution, D2EHPA is partially

(50% to 70%) converted to the sodium form prior to

entering the extraction circuit, and there is no direct

K.C. Sole et al. / HydropH control. Since it is well known that the separa-

tion between cobalt and nickel is enhanced at ele-vated temperature, extraction is carried out at 40 to

45 8C.The organic flow rate is kept constant at 150 L/

min, and optimisation of the circuit performance is

achieved by altering the aqueous flow rates. The

extraction O:A typically varies from 2.7 and 7.5,

and the raffinate typically contains b0.5 g/L Co atpH 5.4. Any magnesium in the feed solution is co-

extracted with the cobalt, and eventually reports to the

cobalt product. Some nickel is co-extracted by

D2EHPA under the pH conditions at which quantita-

tive cobalt extraction occurs. This is scrubbed from

the loaded organic phase using cobalt sulphate solu-

tion (32 to 36 g/L, pH 6.0) and an O:A from 8 to 75,

depending on the extent of co-extracted nickel. The

greater stability of the cobalt complex causes the

nickel to be dsqueezed offT the organic phase as theloading capacity of the extractant is approached.

Cobalt is removed from the scrubbed organic phase

by stripping at an O:A of 1 :8 to 1 :12 with 10%

H2SO4, regenerating the extractant to its acidic form.

A portion of the LSL produced is diverted back into

the circuit as the scrub liquor, while the remaining

liquor is passed through carbon columns for organic

removal and then evaporated under vacuum in a

crystalliser to produce CoSO4d 7H2O crystals. Thestripped organic phase is contacted with 20% H2SO4at O :A=9 in the final stripping stage to remove the

remaining co-extracted trace amounts of Mg, Mn, Fe,

and Ca. The organic phase is regenerated to the

sodium form by contact with 780 g/L NaOH solution

at an advance phase ratio of ~75.

Cobalt recovery across the SX circuit is better than

98%. Upgrading from a Co :Ni ratio of 2 :1 in the feed

solution to 20000 :1 in the LSL is achieved. Organic

losses in the SX circuit due to solubility of D2EHPA

are very low, typically b0.01 g/L, with an annualreplacement of ~1 m3 on a total organic inventory

of 64 m3.

In 2000, Anglo Platinum announced a major ex-

pansion, aimed at increasing its annual PGM produc-

tion capacity to 3.5 million oz. Associated with this,

the RBMR was to increase its annual nickel produc-

tion from 21000 to 40000 tonnes. In the flowsheet

proposed for this expansion, SX with CYANEX 272

was to be used to replace the nickelic hydroxide

lurgy 78 (2005) 5278 67circuit for the bulk removal of cobalt from the nickel

electrolyte. Such a step would see manganese report-

ing to the cobalt circuit, and the cobalt purification

circuit was therefore redesigned to accommodate a

feed containing 80 g/L cobalt with very little nickel

and manganese. Manganese SX followed by cobalt

SX in separate circuits, both using D2EHPA, were

envisaged to ensure continued production of a high-

grade CoSO4. Although a 5-week piloting campaign

using Bateman pulsed columns successfully demon-

strated the technical viability of this approach (Nagel

et al., 2002), this project has not progressed further.

4.4. Hartley Platinum, Zimbabwe

The Hartley Platinum Project at Selous, Zim-

babwea joint venture between BHP Company Ltd

and Delta Gold NLbegan operation in 1997 and

ceased operation in 2000 (Holohan and Montgomery,

1997). Platiniferous ore was concentrated and smelted

to produce a matte (42% nickel, 34% copper, 1% iron

and 0.4% cobalt) that was the feed to the Base Metal

Refinery (Fig. 10).

In the Hartley flowsheet, nickel was leached in two

currently. The three stages ensured that more than

99.5% of the nickel, cobalt, and iron were dissolved

with no dissolution of copper. Iron was removed from

the nickel pressure leach liquor and copper was re-

moved from the nickel sulphate solution that ad-

vanced to nickel EW by cementation onto the

incoming matte in the first leach stage.

Cobalt was removed by SX prior to nickel EW.

This was the ninth commercial-scale plant in the

world to use CYANEX 272 for nickel and cobalt

separation, and the first to use this reagent in a

PGM flowsheet. Since copper, lead, zinc, and iron

are all extracted more strongly than cobalt, the SX

step ensured that these impurity species were reduced

to N1 mg/L in the advancing nickel electrolyte. Theorganic phase comprised 3 vol.% CYANEX 272 dis-

solved in Kerosol 200 (Sasol). Cobalt was extracted

(PLS 80 g/L Ni, 500 mg/L Co) in five stages. Nickel

co-extraction was minimised using three scrubbing

stages. Jameson flotation and Spintek carbon filters

were used to reduce organic carry over into the nickel

tankhouse. Cobalt was stripped in three stages using

ach

trate

alt

ium

ium

val

C

Nic

ach

trate

alt

ium

ium

val

C

Nic

K.C. Sole et al. / Hydrometallurgy 78 (2005) 527868stages of atmospheric and one stage of pressure leach-

ing where the solids and liquids were fed counter-

Nickel

pressure

leach

Atmospheric

copper

removal

Matte

Final le

concen

Nickel

atmospheric

leach

Formic

reducing leach

Copper

pressure

polishing leach

Pressure

iron

removal

Cob

SX

Selen

tellur

remo

Iron residue

Nickel

pressure

leach

Atmospheric

copper

removal

Matte

Final le

concen

Nickel

atmospheric

leach

Formic

reducing leach

Copper

pressure

polishing leach

Pressure

iron

removal

Cob

SX

Selen

tellur

remo

Iron residueFig. 10. The Hartley Platinum Bas150 g/L H2SO4 and an impure cobalt carbonate was

precipitated from the strip liquor.

Copper

electrowinning

Nickel

electrowinning

/

Nickel cathodes

Copper cathodes

obalt carbonate

Sulphate

removal

circuit

Sodium sulphate

Sodium carbonate

kel carbonate

Solids

Solutions

Copper

electrowinning

Nickel

electrowinning

/

Nickel cathodes

Copper cathodes

obalt carbonate

Sulphate

removal

circuit

Sodium sulphate

Sodium carbonate

kel carbonate

Solids

Solutions

Solids

Solutionse Metal Refinery flowsheet.

4.5. Impala Platinum Base Metal Refinery, South

Africa

Impala Platinums base metal refinery in South

Africa presently produces 17000 t/a nickel, 9000 t/a

copper, and 140 t/a cobalt as by-products from a

PGM-containing matte using a process developed by

Sherritt Gordon (Kerfoot and Berezowsky, 1991). The

matte is pressure leached with return copper electro-

lyte to solubilise the nickel, cobalt, and iron. The

leach residue is further leached with sulphuric acid

to produce copper sulphate for copper EW and the

residue is sent for PGM recovery. The nickel solution

is treated with nickel scrap to cement copper and then

iron is precipitated. After conversion to the nickel

ammine using return ammonium sulphate solution

and anhydrous ammonia, most of the nickel is precip-

itated using hydrogen reduction. The remaining cobalt

and nickel are precipitated as a mixed double salt.

This is leached under oxidising conditions in ammo-

nia and, after various steps of purification, cobalt is

precipitated by hydrogen reduction.

A new refinery flowsheet (Fig. 11) was developed

to accommodate a mixed nickel cobalt sulphide con-

centrate that was to be produced from the Philnico

laterite deposit in the Philippines (Sole and Cole,

2001). Metal production was to be substantially in-

creased to 60000 t/a Ni and 4300 t/a Co (Anon.,

2000). Although the Philnico prospect is no longer

an option for Impala, the flowsheet demonstrates an

Sulphide leach

Mn, Fe removal

H2 re

So

purifi

Co m

briqu

Metal sulphide

strip residue

Mixed Ni Co

sulphide

NH3

NaHS

H2SO4

O2

Cu, Fe removal Cu product

NH3SO2air

Sulphide leach

Mn, Fe removal

H2 re

So

purifi

Co m

briqu

Metal sulphide

strip residue

Mixed Ni Co

sulphide

NH3

NaHS

H2SO4

O2

Cu, Fe removal Cu product

NH3SO2air

K.C. Sole et al. / Hydrometallurgy 78 (2005) 5278 69Zn SX

CYANEX 272

Co SX

CYANEX 272

Solution

adjustment

H2 reduction

Ni metal

briquettes

N2 N2

H2H2

Ni powder

Ni diammine

NH3

Purified liquor

from matte leach

NH3

H2SO4

NiSO4

(NH4)2SO4

Zn SX

CYANEX 272

Co SX

CYANEX 272

Solution

adjustment

H2 reduction

Ni metal

briquettes

N2 N2

H2H2

Ni powder

Ni diammine

NH3

Purified liquor

from matte leach

NH3

H2SO4

NiSO4

(NH4)2SO4Fig. 11. Proposed new flowsheet for the expanded Impala base metal refin

(From Sole and Cole, 2001.)duction

lution

cation

etal

ettes

Co powder

Co diammine

Metal strip

Ammonium sulphate

CoSO4

Zn product and

residual Fe

(NH4)2SO4

duction

lution

cation

etal

ettes

Co powder

Co diammine

Metal strip

Ammonium sulphate

CoSO4

Zn product and

residual Fe

(NH4)2SO4ery to include mixed sulphide feed. (Courtesy of Impala Platinum.)

innovative use of SX technology to fit in with the

existing process and may still be considered for other

laterite opportunities.

The proposed SX steps were successfully tested

and proven in pilot-plant trials carried out by Mintek

at the refinery during 2000. The major advantages of

the new process are:

! the overall recovery of cobalt across the refinerywas improved substantially;

! the new circuit was sufficiently flexible to handlefeedstocks containing impurities such as zinc that

cannot currently be treated;

(Dynatec, 2004a,b). The bankable feasibility study

will be complete in mid-2004 and the project is

currently seeking finance to proceed (Dynatec, 2003).

Continuous piloting has shown that the low-mag-

nesium ore is amenable to pressure acid leaching

with good kinetics and moderate acid consumptions,

giving Co and Ni recoveries exceeding 96%. Good

solidliquid separation was also achievedtypically

a major cost factor in the processing of laterites. The

metallurgy is very similar to that at Moa Bay, Cuba,

and it is understood that the flowsheet will be

similar to that of Murrin Murrin, Australia. SX

using CYANEX 272 will be employed for cobalt

quiib

K.C. Sole et al. / Hydrometallurgy 78 (2005) 527870! the new circuit would operate using only minimal-ly more staff than are currently employed, enabling

expected unit costs to be in the lower quartile for

nickel;

! the current separation of nickel from cobalt usingthe Sherritt process requires the difficult control of

the molar ratio of the diammine formation, whereas

the cobalt SX step will give far superior separations

more easily.

4.6. Ambatovy, Madagascar

Another interesting nickel project in this part of the

world is Ambatovy, a saprolytic laterite deposit locat-

ed some 130 km east of Antananarivo. Under joint

development by Phelps Dodge and Dynatec, this re-

source contains 190 Mt grading 1.11% Ni and 0.1%

Co. The project is expected to produce 60000 t/a Ni

and 4000 t/a Co for 20 years, and is predicted to

become one of the lowest cost nickel producers

0

20

40

60

80

100

4 5 6

E

Extr

action (

%)

NiFig. 12. The extraction of nickel and calcium (0.05 M, separately) from 1.0

mixture with 0.50 M neodecanoic acid in combination with the Mintek synnickel separation.

4.7. Mintek synergistic nickel extractant

Disadvantages of the use of neodecanoic acid as an

extractant for nickel include its poor selectivity over

calcium, the high pH at which nickel extraction

occurs, and the high solubility of the extractant in

the aqueous phase, leading to unacceptable reagent

losses and the need for significant organic-recovery

operations. Laboratory and mini-plant test work at

Mintek have shown that the use of nitrogen-donor

compounds in synergistic combination with Versatic

10 can overcome some of the disadvantages of car-

boxylic acids and improve their selectivity for nickel

(Preston and du Preez, 1994a,b, 2000).

Until recently, the commercial development of

these reagents was hindered by manufacturing issues.

Fig. 12 shows the pH dependence of the extractions of

nickel and calcium by Versatic 10 and by Versatic 10

7 8 9

rium pH

Ni Ca

Ca0 M NaNO3 by 0.50 M neodecanoic acid (white symbols) and by its

ergist in xylene (black symbols). (From Du Preez and Preston, 2004.)

in combination with the (proprietary) bMinteksynergistQ. The separation is improved markedly inthe latter case. In addition, nickel extraction can be

carried out at much lower pH, reducing solubility

losses of the extractant and minimising the tendency

to form the hydroxide precipitate.

Extensive continuous piloting of this organic system

on various nickel-containing liquors was undertaken in

2004, with a view to commercial implementation of

this system in the near future.

5. Zinc

5.1. Skorpion Zinc, Namibia

Anglo Americans Skorpion Zinc refinery, located

near Rosh Pinah in southern Namibia, produced its

first metal in May 2003 and is currently ramping up

to full production. Involving a capital investment of

US$ 454 million (Anglo American Corp., 2000), the

process flowsheet includes the first commercial ap-

plication of zinc SX for primary zinc processing, and

represents a radical departure from classical zinc

refineries that rely on roast-leach-electrowin technol-

ogy (Bachmann, 2004). The oxide, silicate, and car-

bonate-based zinc ores which are not amenable to

treatment by conventional processes can be viably

treated in a purely hydrometallurgical processing

route. A key feature is that special high-grade

(SHG) zinc cathode (N99.995% Zn) is produced atthe mine site: this is rarely seen for sulphide ore

processing (Martn et al., 2002).

A simplified flowsheet of the process is given in

Fig. 13 (Sole, 2001). Following an atmospheric leach

in sulphuric acid, iron, aluminium, and silica are

removed from solution by precipitation. Zinc is then

selectively extracted by SX with D2EHPA, enabling

the electrowinning of SHG zinc. The selection of SX

as the purification step serves several purposes. The

ore is an oxidised silicate containing soluble chloride

and fluoride minerals, with an average grade of

10.6% Zn. The choice of a cation exchanger ensures

rejection of the halides as well as the base metals that

ate o

nutio

ric lea

sation

ba

Bleed

ate o

nutio

ric lea

sation

ba

Bleed

ate o

nutio

ric lea

sation

ba

Bleed

K.C. Sole et al. / Hydrometallurgy 78 (2005) 5278 71Thickening

Zn SX

D2EHPA

Zn EW

SHG Zn cathode

H2SO4

O/F

U/FThickening

Zn SX

D2EHPA

Zn EW

SHG Zn cathode

H2SO4

Thickening

Zn SX

D2EHPA

Zn EW

SHG Zn cathode

H2SO4

O/F

U/FZinc silic

Commi

Atmosphe

Neutrali

CaCO3

H2SO4

Fe, Al, Si

Zinc silic

Commi

Atmosphe

Neutrali

CaCO3

H2SO4

Fe, Al, Si

Zinc silic

Commi

Atmosphe

Neutrali

CaCO3

H2SO4

Fe, Al, SiFig. 13. Simplified process flowsheet for the recoverre

n

ch

Filtration and

washing

Precipitation of

sic zinc sulphate

to effluent treatment

Reacidification

CaCO3

H2SO4

Residue

(ZnO)3ZnSO4

Secondary filtrate

Primary filtrate

Cementation

of impurities

Zinc dustre

n

ch

Filtration and

washing

Precipitation of

sic zinc sulphate

to effluent treatment

Reacidification

CaCO3

H2SO4

Residue

(ZnO)3ZnSO4

Secondary filtrate

Primary filtrate

Cementation

of impurities

re

n

ch

Filtration and

washing

Precipitation of

sic zinc sulphate

to effluent treatment

Reacidification

CaCO3

H2SO4

Residue

(ZnO)3ZnSO4

Secondary filtrate

Primary filtrate

Cementation

of impurities

Zinc dusty of zinc at Skorpion Zinc (from Sole, 2001).

Raffinate

PLS

Zn EWH2O

E1 E2 E3 W1 W2 W3 S1 S2 R

Stripped

organic

6 M HCl

Stripped

organic

Aqueous

Organic

Raffinate

PLS

Zn EWH2O

E1 E2 E3 W1 W2 W3 S1 S2 R

Stripped

organic

6 M HCl

Stripped

organic

Aqueous

Organic

Aqueous

Organic

onfigu

K.C. Sole et al. / Hydrometallurgy 78 (2005) 527872are deleterious to zinc EW. SX also successfully

upgrades the zinc from the rather dilute leach liquor

(30 g/L), produced as a consequence of the leach

conditions dictated by the elevated silica content

(~26%) of the ore, to an advance electrolyte contain-

ing 90 g/L Zn that is suitable for EW. Soluble losses

of zinc in the filtration step are minimised by employ-

ing dilute leach liquor, and the problematic formation

of silica gel is avoided. The use of 40 vol.% D2EHPA

in Escaid 100 (a partially aromatic diluent) allows

high zinc transfer in the extraction circuit without

the need for neutralisation. This ensures that the

acid generated by the extraction reaction is available

for leaching on recycle of the raffinate and minimises

co-extraction of calcium.

Tecnicas Reunidas were responsible for the provi-

sion of the zinc SX technology, which is based on the

Modified Zincex Process developed for the treatment

of secondary materials (Daz et al., 1994, 1995; Gar-

Fig. 14. Skorpion Zinc SX circuit cca et al., 2000). The plant is the largest yet built for

zinc SX, with an aqueous feed flowrate of 960m3/h and

Table 7

Specification for special high-grade zinc and advance electrolyte

Element Permitted concentration

Advance electrolyte (mg/L) Zinc cathode (%)

Zn N90000 99.995Mn 2000

Cd b0.05 0.0015Co b0.05Ni b0.05Sb b0.02Ge b0.02Fe b5.0 0.001an annual cathode production of 150000 tonne. The

SX circuit comprises three extraction stages, three

washing stages, two stripping stages, and an organic

regeneration stage (Fig. 14). Zinc transfer of 20 g/L

across the extraction circuit is achieved. The first two

stages of the washing circuit use demineralised water

to remove physically entrained impurity species; di-

luted spent electrolyte is employed as a scrub liquor

in the third wash stage to remove co-extracted impu-

rity species from the loaded organic phase by means

of crowding by zinc and the reversal of equilibrium

by the high-acid strength. Iron build-up on the organ-

ic phase is controlled by treating a bleed stream with

6 M HCl.

A conventional mixer-settler design has been

employed, with settlers of area 2525 m2. Entrainedorganic phase is removed from the raffinate and ad-

vance electrolyte by a combination of flotation and

carbon adsorption. Because the raffinate is recycled to

> 99.995% Zn> 99.995% Zn

ration (from Cole and Sole, 2002).the leaching circuit, a bleed for removal of impurities

is required. Base metals such as copper, cobalt, nickel,

Element Permitted concentration

Advance electrolyte (mg/L) Zinc cathode (%)

Pb 0.0015

Cu 0.001

Se b0.1Cl b100F b20Sn 0.001

Al 0.001

and cadmium that co-precipitate with zinc are removed

from the primary filtrate by cementation with zinc

dust. The halides and magnesium exit in the secondary

filtrate that is treated first for zinc recovery by the

precipitation of basic zinc sulphate ((ZnO)3.ZnSO4).

This zinc is returned to the main circuit by recycling

the precipitate as a neutralising agent.

The flexibility of the SX operation in handling the

impurities present in the PLS is indicated by Table 7,

which shows the permitted concentrations of various

metals in solution to ensure the production of SHG

zinc (British Standard, 1996).

The feasibility study showed that Skorpion Zinc

will be one of the lowest cost zinc facilities in the

world, with an expected production cost of US$

0.25/lb (Garca et al., 2000).

6. Precious metals

6.1. Harmony Gold, South Africa

lised by Mintek, South Africa. Gold-bearing materials

are leached in HCl, and then SX is employed for the

purification of the leach liquor. Gold recovery from

solution is via precipitation with SO2. Gold of purity

99.99 or 99.999% can be achieved, depending on the

operating conditions. A variety of gold-containing

sources are amenable to treatment in this manner.

The first of these refineries was established at

Harmony Gold Mine in Virginia, South Africa, in

1997, processing gold slimes from the electrowinning

circuit that typically contained ~80% Au and 8% Ag

(Fig. 15) (Feather et al., 1997; Sole et al., 1998). The

circuit was designed to produce 24 t/a high-purity

gold. Following the significant commercial success

of this project, the ease of operation of the process,

and the forgiving nature of the technology, a new

refinery was commissioned in 2001 which currently

produces up to 400 kg/d of gold powder (Feather et

al., 2002b). Some of the performance parameters of

the SX circuit are presented in Table 8, while the

purification capabilities of the SX operation are

rubbi

stage

ue

ase

Scru

rubbi

stage

ue

ase

Scru

K.C. Sole et al. / Hydrometallurgy 78 (2005) 5278 73Sc

(5

Leach

AgCl/SiO2 resid

Extraction

(3 stages)

Cl2

PLS ~ 65 g/l Au

L

S

HCl

Raffinate

Lixiviant

make up

Organic ph

Sc

(5

Leach

AgCl/SiO2 resid

Extraction

(3 stages)

Cl2

PLS ~ 65 g/l Au

L

S

HCl

Raffinate

Lixiviant

make up

Organic phThe Minataurk (Mintek Alternative Technologyfor Au Refining) Process is a novel route for the

refining of gold that was developed and commercia-

Cathode sludge from EW

(50 85 % Au)

Cathode sludge from EW

(50 85 % Au)Fig. 15. Schematic of the gold refining process at Harmony Goldshown in Table 9.

The production of this high-purity gold on site at

the mine has enabled a variety of other value-added

products to be manufactured, including ten-tola bars

ng

s)

99.99% Ag

Reduction

Silver recovery

and refining

Stripping

(4 stages)

Loaded strip liquor

~ 80 g/l Au

99.99% Au

Barren

solution to

CIP circuit

SO2

b liquor Strip liquor

ng

s)

99.99% Ag

Reduction

Silver recovery

and refining

Stripping

(4 stages)

Loaded strip liquor

~ 80 g/l Au

99.99% Au

Barren

solution to

CIP circuit

SO2

b liquor Strip liquor(adapted from Sole et al., 1998 and Feather et al., 2002b).

(99.9% Au), gold granules, and gold potassium cya-

nide. Harmony has now established its own brand

name, Harmony Pure Gold.

While Harmony still remains the flagship of the

Minataurk process, more recently other refinerieshave been established by Agenor in Algeria (25 kg/d

Au) for the processing of dore bullion and by Al

Ghurair, a private gold trading concern in Dubai (100

t/a on a 12-h shift per day), which uses jewellery and

electronic scrap as the feed material (Scott and Match-

ett, 2004).

1979; Charlesworth, 1981; Benner et al., 1991; Mooi-

man, 1993; Al-Bazi and Chow, 1984; Harris, 1993).

180 tonne haultrucks, whereby it is delivered to the

primary crushers which reduce the rock to an aver-

age size of 16 cm. It is further reduced to sand grain

size in three additional crushing and milling stages.

Uranium is extracted by leaching in sulphuric acid.

Following solidliquid separation by a combination

of rotoscoops and CCD, uranium is recovered from

solution by IX using Duolite A101-DU resin. The

K.C. Sole et al. / Hydrometal746.2. Anglo Platinum Precious Metals Refinery, South

Africa

Anglo Platinums Precious Metals Refinery at Rus-

tenburg has been in the forefront of SX developments

for the refining of the PGMs. The technologies

employed today were largely developed by Matthey

Rustenburg Refiners in the UK, and production

started at the Rustenburg site in 1989. Today, this is

the largest PGM refinery in the world, with annual

production of 116000 oz/a Au, 2.3 million oz/a Pt,

and 1.2 million oz/a Pd. Ruthenium, rhodium, iridium,

and osmium are also produced with purities varying

from 99% to 99.99%.

The PGM-bearing concentrates are leached under

aggressive conditions in a chlorine/HCl medium.

Gold is removed first from the chloride leach liquor

by SX using methylisobutylketone (MIBK), an oxy-

gen-donor solvating reagent where selectivity is large-

ly based on the chargesize ratio of the chloroanion.

The liquor is then processed sequentially for the re-

covery of palladium by SX using a b-hydroxyoxime,

Table 8

SX performance in the refining of gold by SX using the Minataurkprocess at Harmony Gold Mine (from Feather et al., 2002b)

Parameter Value

Extraction

Extraction efficiency for gold (%) N99Gold concentration in leach solution (g/L) 65

Organic loading of gold (g/L) 64

Gold concentration in raffinate (g/L) b0.1

Stripping

Strippping efficiency (%) N99.7

Gold concentration in loaded strip liquor (g/L) 82

Au: impurities in loaded strip liquor (%) N99.977. Uranium

7.1. Rossing Uranium, Namibia

Part of the Rio Tinto Group, Rossing is one of the

largest open-cast uranium mines in the world. It is

situated in Namibia, 65 km inland from the coastal

town of Swakopmund in the Namib Desert. Annual

mined tonnage is approximately 21 million tonnes,

with U3O8 production of 3200 t/a and reserves avail-

able for another 16 years (Rossing, 2004).

The uranium-bearing ore body is mined by blast-

ing and loading the rock with electric shovels ontoplatinum by SX using an amine extractant, ruthenium

removal by distillation, iridium SX using a novel

amide extractant, and finally rhodium recovery by

IX. In each case, a pure solution of the respective

metal is obtained, which is then subjected to a reduc-

tion to produce the pure metal powder or sponge. The

final products are obtained by melting to obtain

ingots, granules, or good delivery bars.

While specific details of the SX refining steps

remain closely guarded, several publications in the

open literature give interesting insight to the clever

chemistry and novel ideas that have been employed in

the development of these processes (Cleare et al.,

Table 9

Upgrading capababilities of the Minataur process (fro