Textural, mineralogical and chemical characteristics of copper reverb furnace smelter slag of the Okiep Copper District, South Africa

1. Textural, mineralogical and chemical characteristics of copper reverb
furnace smelter slag of the Okiep Copper District, South Africa
Abraham Rozendaal ⇑
, Richard Horn
Department of Earth Sciences, University of Stellenbosch, Stellenbosch 7600, South Africa
a r t i c l e i n f o
Article history:
Available online 25 July 2013
Keywords:
Environment
Okiep
Copper-slag
Characteristics
Ore mineralogy
a b s t r a c t
The Okiep Copper District in South Africa has produced more than 110 million tons at a grade of 1.71% Cu
from several small mafic ore bodies. The ore was smelted on site and generated 5 mt of slag. During the
life of mine attempts to recover copper from the slag by flotation had limited success. After mine closure
the challenge of environmental rehabilitation and the possible disposal of the slag, triggered a reinvesti-
gation into the viability of slag as a copper resource. Characterisation of the slag as a contribution to the
potential copper recovery is the objective of this study.
The slags are hard, vitreous with a matrix of Si–Fe–Al–Mg–Ca glass and laths of Mg–Fe–olivine, Fe–Mg–
orthopyroxene and minor Cr-spinel. Copper grade varies between 0.11% and 0.42% with minor nickel,
cobalt, molybdenum, zinc and tungsten. All economic elements are hosted by disseminated spheroidal
prills which consist mainly of the copper sulphides bornite, chalcocite, covellite and chalcopyrite with
exsolved sulphide phases of the minor base metals as well as rhenium and silver. Prills consisting of
metallic copper and alloys are minor constituents. Prill diameter is highly variable with most in the
40–60 lm range and the historically poor copper recovery is attributed to the small prill size. Crushing
of slag to 45 lm as opposed to the previous 75 lm should significantly increase sulphide liberation
and recovery of copper and minor base metal sulphides by conventional flotation.
Provided the operation is economically viable, redistribution of the processed slag to environmentally
acceptable sites will resolve the present pollution and rehabilitation challenge related to the dumps in
the Okiep Copper District. The operation will also have a positive socio-economic impact on this pov-
erty-stricken part of South Africa.
Ó 2013 Elsevier Ltd. All rights reserved.
1. Introduction
The Okiep Copper District is located 600 km north of Cape Town
in the north-western part of the Northern Cape Province, South
Africa and includes the towns of Springbok, Okiep, Nababeep and
Concordia. It is the oldest mining district in the country and has
been a copper producer for more than 150 years. Copper smelters
have been in operation at several localities in the area and pro-
duced slag dumps ranging from approximately 10 000 m3
to more
than 1.5 million m3
. Mining came to an end in 2007 and the district
has been under slow but, systematic rehabilitation ever since
(Fig. 1).
The Okiep Copper District with its widespread distribution of
small mines and extensive mining history has been listed as a
highly polluted area with significant environmental damage
(Hohne and Hansen, 2008). The smelter slags at Nababeep and
Okiep were indicated to be a phosphorous (PO4), lead and chro-
mium risk for the aquatic ecosystem in particular. The windblown
slag dust was also considered a health problem for the towns of
Nababeep and Okiep. Rehabilitation of the copper oxide ore dumps
and extensive slimes dams remains a challenge and as part of the
program, are earmarked for future reprocessing and redistribution.
Rehabilitation of the landscape is required and would include dis-
posal of slag as backfill in old open cast mines (van der Westhuizen
pers comm.). This operation has not commenced too date.
Favourable copper prices and the expected increase in future
demand coupled with the pressure of the rehabilitation, which in-
cluded slag disposal, initiated this orientation study to investigate
the slag dumps as a potential copper resource (Candy, 2011; Hur
and Sedgman, 2013).
When the mine was in operation, limited attempts had been
made to recover copper from the slag by means of conventional
copper sulphide flotation. All these attempts failed because of the
lack of a proper resource estimate and geometallurgical study to
identify the various mineral phases. This resulted in variable grade
of the plant feed and recovery which never exceeded 50% of the
contained copper. This poor recovery was mainly attributed to loss
of elemental copper during sulphide flotation (De Beer, pers
comm.). A flash flotation cell was added to the milling circuit to
0892-6875/$ - see front matter Ó 2013 Elsevier Ltd. All rights reserved.
http://dx.doi.org/10.1016/j.mineng.2013.06.020
⇑ Corresponding author.
E-mail address: ar@sun.ac.za (A. Rozendaal).
Minerals Engineering 52 (2013) 184–190
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2. prevent overgrinding of the sulphides, a possible cause for poor
recovery. Ultimately this had limited effect and did not improve
recovery. Coupled with excessive grinding costs associated with
grinding to 75 lm and the general highly abrasive effect of the
slag on all equipment, as well as decreasing copper prices, these
operations became uneconomical and were abandoned.
In addition, finely crushed slag had been used as a low cost
additive to conventional Portland cement because of its pozzolanic
properties. This was added to slimes which acted as cemented
backfill of the mined-out open stopes when the mines were in
operation.
Presently a small operation is recovering copper spills from the
slag and the crushed waste product is used as a low priced sand-
blast cleaning agent on a small scale. Ironically the operators are
penalized for the anomalous copper content which is considered
a contaminant.
The aim of this orientation study is to determine the slag chem-
istry, the distribution, identification, morphology and textures of
the various mineral phases, in particular those with copper and re-
lated elements of economic interest. Recovery potential of copper
from slag is briefly considered.
2. Methodology
In this study three of the larger slag dumps located at Nababeep,
Okiep and Concordia have been sampled. A suite of 30 specimens
have been collected and subjected to a diversity of analytical tech-
niques. Fifty polished thin sections have been prepared for trans-
mitted and reflected light microscopy.
X-ray fluorescence was used for major element determinations
of 19 slag samples. A PANalyticalAxios Wavelength Dispersive
spectrometer fitted with a Rh tube and LIF200, LIF220, PE002,
Ge111 and PX1 analysing crystals was used. Fused glass discs were
prepared from high purity trace element and rare earth element
free flux ((LiBO2 = 32.83%, Li2B4O7 = 66.67%, LiI = 0.50%). Matrix ef-
fects were corrected with SuperQPANalytical software and control
standards used, fit the range of concentration of the samples. The
same fusion discs were used to analyse for a suite of 36 trace
elements by means of LA–ICP–MS. A New Wave 213 nm laser
connected to an Agilent 7500ce ICP–MS was used and three spots
of 110 lm each were ablated per sample. Trace elements were
quantified using NIST 612 for calibration and the SiO2 from XRF
measurement as internal standard. Data was processed using
Glitter software.
The scanning electron microscope (SEM) was used to identify
mineral phases and to determine their chemistry, as well as tex-
tural relationships. Elemental mapping and SEM backscatter
images were used to measure prill diameter and deportment of
elements of interest. For this purpose a Zeiss EVOÒ
MA15 Scanning
Electron Microscope was used and phase compositions were
quantified by EDX analysis using an Oxford InstrumentsÒ
X-Max
20 mm2
detector and Oxford INCA software. Internal Astimex
Scientific mineral standards were used for standardization and
verification of the analyses.
Mineral identification was supported by X-ray diffraction using
a PANalytical XRD powder diffractometer. Cluster analyses were
performed on the scans to determine semi-quantitative phase
abundances using HighScore Plus software.
All instrumentation discussed above is housed and operated by
the Central Analytical Facility of the University of Stellenbosch and
detailed procedures used for the various methods are available
from that facility (http://academic.sun.ac.za/saf/about.htm).
Relative density was determined by means of a Snowrex Preci-
sion NHV-3 density scale and has a maximum capacity of 3000 g
and sensitivity of 0.1 g. It was calibrated with 2 kg, 1 kg, and
500 g metallic weights. Pure vein quartz was used as a reference
sample. Froth-type slag was covered by cling wrap to ensure a rep-
resentative reading.
3. Historical background
Copper mineralisation in the district was first discovered by
Governor Simon van der Stel in 1684 but, it was not until 1836 that
the first mining company was established. This was soon followed
by a multitude of enterprises that operated with mixed success.
The copper boom lasted up to 1855 and left two companies stand-
ing the Phillips Group and King Group, later in 1863 to become the
Cape Copper Mining Company Limited, and the Namaqua Copper
Company Limited. These two entities produced the first significant
quantities of ore for export from mines at Okiep, Spektakel and
Nababeep and Concordia. Records for the period 1852–1926
showed that the entire district produced 2.4 million tons of ore
at a grade of 12% copper (Cornelissen, 1965). During the Great
Depression mining effectively ceased but, was reinstated by New-
mont Mining Corporation who consolidated the mining rights for
the entire district under the newly established O’okiep Copper
Company Limited in 1939. Gold Fields of South Africa took control
of the mines in 1984 and was taken over by Metorex in 1996.
Between 1940 and 1994 105.6 million tons at a grade of 1.71%
copper was mined from 30 small mines in the area (Potgieter,
1996). From 1996 onwards production rapidly decreased and
mining finally ceased in 2007. It is estimated that a total of 2.12
million tons of copper metal was produced from the district during
its entire operational lifespan (Rozendaal, 2011).
3.1. The smelters
During the early stages of mining in the copper district and with
very limited infrastructure such as railway links and proximal deep
ports, it became clear that for the mines to be viable only very high
grade hand sorted ore (35–40% copper) could be exported. Low
grade ore had to be upgraded on site and was the impetus to estab-
lish small smelters. By 1867 the first smelter was in operation as
the Springbokfontein Reduction Works producing regulus for ex-
port from the local mine. It was soon superseded by the O’Okiep
Reduction Works in 1870 when the ore body at Springbokfontein
(Springbok) was depleted. Coal fired German blast furnaces were
used and operated by gentlemen from the German School of Mines.
One furnace would produce 1 ton of copper per day as regulus
Fig. 1. View of Nababeep in the Okiep Copper District showing smelter stack and
slag dumps in the foreground. These dumps pose an environmental problem and
will have to be rehabilitated. (Photo, Isky, 2009).
A. Rozendaal, R. Horn / Minerals Engineering 52 (2013) 184–190 185

3. assaying between 50% and 60% copper (Smalberger, 1975). These
smelters were shut down in 1919 and never restarted. The smelt-
ers produced approximately 200 000 tons of slag. They were later
transformed to sulphur burners of pyrrhotite concentrate for the
production of sulphuric acid for leaching of oxide copper ore. This
continued up to the late 1960s.
Around 1870 reports indicate that smelting works were also
being built at Spektakel mine. Due to the intermitted but, generally
short-lived mining of that area it is uncertain if the smelter oper-
ated for long. The tons of slag produced are unknown.
At Concordia (Tweefontein-Jubilee) a furnace was erected in
1872 and the first regulus and smelted copper was produced in
1873 from water-jacket furnaces. Expansions were made around
1905 and these smelting operations were suspended in 1930. All
mining stopped in 1931. Based on the tons copper ore processed
approximately 60 000 tons of slag was produced (Rozendaal,
2011).
The Nababeep smelting works, with two Frasers and Chalmers’
blast-furnaces had been erected with great haste around 1899
when substantial copper resources were confirmed. It produced
regulus of 40–50% copper. Mining and smelting stopped in 1919.
The tonnage slag produced is unknown for that period. When the
O’okiep Copper Company Limited was established a new modern
smelter with a reverberatory furnace and Peirce-Smith converter
with a ladle tilting quadrant and Walker-type casting wheel, was
constructed at Nababeep in 1939. The smelter was expanded over
the years to increase output capacity, but the pyrometallurgical
route remained the same. The latter is discussed in Appendix A.
This centralized smelter treated the entire concentrate produc-
tion of the district and was dismantled when the mining opera-
tions ceased in 2007. The smelter had a maximum capacity of
30 000 tons of blister copper per year (Cornelissen, 1965; Kinsella
and Goosen, 1996) and produced 1.8 million tons of copper in its
life time. The blister copper ingots had a grade of 99.2% copper
and credits were received for gold (5 g/t), silver (200 g/t) and
nickel (0.5%; Fairfax, 1955). The smelter produced 5 million tons
of slag. Fairfax (1955) reports a copper grade of 0.45% copper
for the slag in those days. Random sampling in the 1980s showed
that the copper grade of subsequent slag production was slightly
lower.
It is estimated that the entire Okiep Copper District has a slag
potential of 5.26 million tons preserved at three different localities
of which Nababeep is the most significant.
4. Geological setting
Copper sulphide mineralisation is hosted by a suite of syn- to
post-tectonic mafic bodies that intruded the meta-volcanosedi-
mentary Mesoproterozoic Namaqualand Metamorphic Complex,
the western end of the Namaqua-Natal Mobile Belt. These bodies
are restricted to an area known as the Okiep Copper District and to-
tal several thousand small intrusives referred to as the Koperberg
Suite (1070 + 20 Ma) of which the most mafic (norites and
hypersthenites) display the best mineralisation and have been
mined (Lombaard et al., 1986).
Mineralogically the host rocks consist of orthopyroxene, plagio-
clase with minor biotite and diopside and traces of apatite and geo-
chemically are essentially silica-, iron- and magnesium-rich with
subordinate aluminium and phosphorus. Opaque minerals include
variable concentrations of magnetite, ilmenite, and a sulphide min-
eral suite which consists of chalcopyrite (CuFeS2), bornite (Cu5-
FeS4), subordinate chalcocite (Cu2S), pyrrhotite, pyrite, galena,
sphalerite and molybdenite. On surface the so-called oxide ore
consists of malachite, azurite and chrysocolla. The copper-bearing
minerals are finely disseminated throughout the mafic host
however two of the mines display massive pyrrhotite–chalcopyrite
ore similar to the classic magmatic mafic to ultramafic copper–
nickel–cobalt bodies. Copper grade of the bodies mined during
the modern era is extremely variable ranging from 0.2% to 4%
copper with traces of nickel, cobalt and sporadic platinum group
elements and molybdenum.
5. Okiep Copper District slag characteristics
Slag samples obtained from the three localities at Nababeep,
Okiep and Concordia are all very similar with respect to physical
appearance and textures. The slag has a dark grey to almost black
colour with a massive vitreous to vuggy texture. It has not been
granulated at any of these localities. Top of the slag flow is marked
by large vesicles grading into the massive to banded type (Fig. 2).
The slag is very fine-grained, appears homogeneous, breaks with
a concoidal texture and has a hardness exceeding 7 on the Moh
scale. Grinding and uniaxial compression tests have not been done.
Slag from all the localities is weakly to moderately magnetic. Both
slag types were studied separately to determine the copper distri-
bution in each and to consider the possibility of separation based
on density contrast. On average the massive slag has a relative
density of 3.5 (n = 26) and the vesicular or frothy type is variable
between 2.1 and 2.7 (n = 20). This contrast could allow separation
on density basis if, for example, the froth-type proves to be less
well mineralized than the massive type.
Microscopically slag from all three areas displays the same tex-
tural features consisting of a fine-grained glassy matrix traversed
by variable quantities of crystalline phases. The proportion of glass
to crystalline phases varies, but is never less than 60% glass. Small
discreet spheroidal prills of sulphide and alloy phases are hosted
by this essentially silicate matrix.
5.1. The slag chemistry
A selection of 19 samples from three slag localities have been
selected for major and trace element analyses. The sample suite
was biased towards the Nababeep locality, as it constitutes more
than 99.5% of the slag resource (Table 1).
Major elements: All slags from the district have a typical silica–
iron–magnesium–aluminium–calcium chemical composition with
minor amounts of sodium, potassium, titanium, manganese and
chromium (Table 1; Fig. 3). The Okiep slags are significantly
enriched in aluminium–calcium–magnesium–phosphorous at the
expense of silica. This is considered a function of the mineralogical
composition of the sand that was added during the smelting
process to increase the silica content and aid slag-sulphide melt
segregation. Fairfax (1955) reported that sand was locally collected
and comprised a weathering product of the nearby feldspathic
gneisses.
Fig. 2. Typical Okiep slag consisting of a vesicular froth part and a high density,
massive, vitreous component.
186 A. Rozendaal, R. Horn / Minerals Engineering 52 (2013) 184–190

4. Trace elements: Copper content of the slags varies between
0.11% and 0.42% and compares well with the slag data reported
by Fairfax (1955). Concentrations of up to 1% zinc, 0.36% tungsten,
0.15% cobalt and 137 ppm molybdenum have been reported from
the Nababeep slags (Table 1). Lead content is also elevated com-
pared to Okiep and Concordia, however the former may have sig-
nificant nickel concentrations of up to 0.14%. From an economic
point of view the Nababeep slags are the most enticing followed
by those from Okiep. None of the slag analysed showed any rare
earth element enrichment (Table 1). The platinum group element
potential of the slags has not been determined by means of whole
rock analyses. There is no significant chemical difference between
frothy and massive slag and separation based on density contrast
will serve no purpose.
5.2. The matrix mineralogy
Silica-rich glass generally constitutes more than 60% of the slag
matrix and hosts large crystalline euhedral grains of magnesium–
iron olivine ((Mg, Fe)2SiO4), iron-magnesium orthopyroxene ((Mg,
Fe)SiO3) and minor to rare amounts of ortho-amphibole (gedrite)
(Fig. 4; Table 2; www.sun.ac.za/earthsci/people/rozendaal_e.htm).
The major silicate phases have been confirmed by X-ray diffraction.
An unidentified low silica phase chemically conforms to the iron-
rich olivine iddingsite but, the structure has not been confirmed.
These silicate phases display a spinifex-like texture of radiating
grains. Small isometric grains of the spinel group conforming to
chrome-hercynite (picotite) and hercynite are common and
disseminated magnetite is responsible for the weak magnetic
Table 1
Average and standard deviation (S) of major and trace element chemistry of massive (M) and frothy (F) slags from Nababeep, Okiep and Concordia in the Okiep Copper District.
Nababeep n SiO2 TiO2 Al2O3 Fe2O3 MnO MgO CaO Na2O K2O P2O5 Cr2O3 LOI H2O– Total
M 7 61.67 0.59 3.50 23.86 0.07 3.82 1.57 0.19 0.32 0.13 0.05 3.47 0.17 99.40
F 3 49.58 1.15 6.05 29.87 0.09 6.30 2.64 0.49 0.48 0.22 0.11 2.70 0.20 99.88
S–M 8.26 0.39 2.31 3.75 0.02 1.92 0.97 0.31 0.14 0.07 0.03 0.69 0.04 0.61
S–F 14.95 0.53 2.74 11.32 0.01 1.62 1.18 0.36 0.19 0.09 0.08 0.42 0.03 1.53
Okiep
M 4 43.76 1.12 9.42 27.78 0.22 7.84 3.64 0.94 0.82 0.63 0.28 2.43 0.08 98.94
F 1 47.78 0.90 12.16 16.88 0.22 10.28 4.22 1.54 0.82 0.56 0.34 1.28 0.10 97.08
S–M 3.97 0.45 1.34 7.59 0.02 2.06 0.32 0.32 0.20 0.31 0.02 0.90 0.02 1.94
Concordia
M 3 66.11 0.63 6.92 14.96 0.11 1.89 2.49 0.85 0.53 0.49 0.17 2.44 0.20 97.90
F 1 66.52 0.68 7.44 12.76 0.12 2.26 2.62 0.94 0.54 0.46 0.16 2.22 0.20 96.92
S–M 14.95 0.53 2.74 11.32 0.01 1.62 1.18 0.36 0.19 0.09 0.08 0.42 0.03 1.53
Nababeep n Ba Cr Co Cs Cu Hf Mo Nb Ni Pb Rb Sc Sr Sn
M 7 241.15 359.84 353.90 0.76 2373.95 1.54 56.84 6.34 140.37 148.17 16.11 31.53 197.56 21.38
F 3 228.15 710.16 369.45 0.95 2601.82 2.05 46.91 11.17 141.54 149.62 22.05 35.02 355.60 25.88
S–M 240.67 148.61 240.80 0.33 559.77 0.52 37.40 3.48 60.62 61.53 5.73 4.48 119.95 11.13
S–F 53.45 434.75 151.38 0.32 889.75 0.26 20.42 5.17 45.18 81.38 8.07 2.72 154.36 16.90
Okiep
M 4 323.64 1815.44 504.98 1.62 2633.54 8.21 7.40 11.66 469.35 339.25 41.60 41.98 293.97 12.00
F 1 315.44 2237.50 120.87 1.18 3870.71 5.39 4.17 8.45 194.52 8.90 37.42 42.31 354.63 16.53
S–M 79.38 120.56 54.21 0.41 287.38 2.57 5.21 3.92 35.41 1.96 11.55 0.58 69.03 3.52
Concordia
M 3 122.82 1045.77 96.80 0.91 2492.33 5.46 10.77 6.19 104.77 8.88 27.24 30.40 207.31 10.65
F 1 132.22 988.51 107.09 1.15 1163.20 5.19 9.54 6.44 66.32 6.81 28.56 31.09 216.04 10.45
S–M 53.45 434.75 151.38 0.32 889.75 0.26 20.42 5.17 45.18 81.38 8.07 2.72 154.36 16.90
Nababeep n Ta Th U V W Y Zn Zr
M 7 0.53 12.21 1.81 103.58 736.05 18.25 2203.81 57.95
F 3 0.83 18.57 2.80 159.16 365.97 28.60 1284.67 76.21
S–M 0.23 6.91 0.70 43.88 1274.69 8.25 3545.34 19.42
S–F 0.33 8.37 0.46 60.20 229.16 6.94 886.63 6.68
Okiep
M 4 2.03 49.14 9.40 453.79 777.18 12.15 5247.26 275.74
F 1 1.33 31.07 3.72 399.05 456.15 3.72 250.91 204.87
S–M 0.71 11.67 1.82 179.13 269.35 1.82 84.02 92.99
Concordia
M 3 1.05 38.70 5.55 338.49 236.28 51.00 156.79 191.58
F 1 0.98 37.55 5.13 363.25 433.38 49.61 146.18 189.26
S–M 0.33 8.37 0.46 60.20 229.16 6.94 886.63 6.68
Nababeep n La Ce Pr Nd Sm Eu Gd Tb Dy Ho Er Tm Yb Lu
M 7 57.17 117.43 13.28 51.52 7.81 0.89 5.88 0.65 3.77 0.69 1.87 0.26 1.80 0.26
F 3 94.81 194.76 22.04 84.92 12.74 1.45 9.35 1.00 5.78 1.06 3.01 0.41 2.59 0.40
S–M 25.79 53.26 6.30 24.51 3.74 0.45 2.69 0.29 1.69 0.30 0.78 0.11 0.84 0.11
S–F 27.83 59.11 6.37 25.70 4.21 0.45 2.29 0.28 1.61 0.25 0.72 0.08 0.64 0.10
Okiep
M 4 77.06 167.21 24.68 69.61 13.00 1.30 11.53 1.54 9.29 1.75 4.84 0.70 4.45 0.64
F 1 69.91 154.70 17.62 70.74 14.01 1.57 11.41 1.49 9.24 1.74 5.14 0.72 5.11 0.67
S–M 17.53 36.38 4.86 20.38 5.13 0.13 4.35 0.52 3.60 0.57 1.69 0.21 1.04 0.13
Concordia
M 3 64.75 136.56 16.66 68.17 13.62 1.18 11.89 1.62 9.96 1.90 5.32 0.76 4.74 0.68
F 1 63.40 133.90 16.11 66.83 14.29 1.13 12.12 1.56 9.47 1.83 4.78 0.71 4.32 0.65
S–M 27.83 59.11 6.37 25.70 4.21 0.45 2.29 0.28 1.61 0.25 0.72 0.08 0.64 0.10
A. Rozendaal, R. Horn / Minerals Engineering 52 (2013) 184–190 187

5. signature of the slag (Table 2; www.sun.ac.za/earthsci/people/ro-
zendaal_e.htm). The anomalous whole rock zinc values also report
to the spinels.
Glass chemistry is variable and is largely a function of the lo-
cally derived sand composition which was added during the smelt-
ing process, as well as the gangue mineralogy of the mafic host
rock. Glass from the Okiep, as well as Concordia slags is aluminium
enriched indicating a large feldspar component of the added sand
(Table 3: www.sun.ac.za/earthsci/people/rozendaal_e.htm). This
also explains the abundance of Al-spinel in those slags. Glass has
a Si–Fe–Al–Mg–Ca composition in decreasing order of abundance
with minor amounts of sulphur, chromium and traces of copper,
zinc and tungsten.
5.3. The prill mineralogy
Prills are disseminated throughout the various slags and show no
preference for either the massive or frothy types. In general the prills
all have the same spheroidal shape but, mineralogical composition
can be monomineralic or complex, consisting of several sulphide, al-
loy or metallic element phases (Figs. 4 and 5). Prill size has a wide
range between 5 and 1500 lm with an average of approximately
50 lm (Fig. 6). The largest prills occur in Concordia slag.
Most prills consist of either bornite (Cu5FeS4), chalcocite (Cu2S)
covellite (CuS) or combinations thereof and have been confirmed
by X-ray diffraction (Table; www.sun.ac.za/earthsci/people/rozen-
daal_e.htm). These may contain several exsolution phases consist-
ing of chalcopyrite (CuFeS2), cobaltian pentlandite (Ni, Fe, Co)9 S8,
millerite (NiS), marmatitic sphalerite ((Zn, Fe) S), pyrrhotite
(Fe7S8), galena (PbS) and castaingite (CuMo2 (Re) S5). A cubic
Pb–Ni–Cu–Co–S phase is common in the Nababeep slag and a
Re–Ni–Os–Fe–S phase (Rheniite, structure ReS2?) has been
observed at Concordia (Table 4; www.sun.ac.za/earthsci/people/
rozendaal_e.htm). In several prills cubic metallic copper occurs in
the core or may crosscut the prills as small veinlets, indicating sul-
phur deficiency during the segregation process. Prills consisting
entirely of copper metal are rare. Alloys with a Ni–Fe, Re–Mo and
Ru–Rh composition, as well as specs of metallic silver have been
identified. The latter also occurs in solid solution with chalcopyrite.
Molybdenum occurs in all copper, as well as copper–nickel
phases in solid solution, with variable concentrations of up to 6%
molybdenum. Of the platinum group elements rhenium has a close
spatial association with zinc and a Re–Zn–S phase may occur on
the contact between native copper and chalcocite. Osmium has
noticeable concentrations in the copper phases, in particular native
copper. Although the distribution of the platinum group elements
in solid solution in copper phases may be considered with some
reservation, the identification of discrete PGE-rich phases lends
support to their presence.
Distribution of the economically important elements between
the three localities shows that Nababeep prills have more copper,
nickel and Pb–Zn–Re phases and less Ni–Co phases than Okiep.
The Concordia slags are more sulphur-rich and explain the
Fig. 3. Major element chemistry of slags from the Okiep Copper District. The slags
are silica–iron-rich with variable aluminium content. All iron presented as Fe2O3.
Legend: solid triangle Okiep, open triangle Nababeep, open square Concordia.
Fig. 4. Typical back scatter SEM image of Nababeep slag showing distribution of
olivine (ol), orthopyroxene (opx) and spinel (sp) in a glass matrix. Spheroidal prills
consist of bornite (bn). Note the variation of prill diameter.
Fig. 5. A mineralogically complex prill consisting of bornite, native copper and Ni–
Fe alloy hosted by a glass, orthopyroxene matrix.
Fig. 6. Prill size distribution in slags from the Okiep Copper District. The population
is dominated by the 40–60 lm interval. (n = 200).
188 A. Rozendaal, R. Horn / Minerals Engineering 52 (2013) 184–190

6. presence of large pyrrhotite prills poikilitically enclosing magnetite
and silicate glass.
6. Discussion
Slags from all three localities studied have the same basic prop-
erties with relatively minor differences in the glass and silicate
mineralogy. Effectively all economically important elements such
as copper, nickel, molybdenum, cobalt, lead, zinc and some plati-
num group elements have been scavenged by the remaining sul-
phur in the silicate-rich slag and as a result report to the discreet
sulphidic prills hosted by the slag matrix. Copper is mainly present
as sulphides with less than 10% as elemental copper. Bornite is the
dominant copper sulphide phase and accounts for more than 80%
of the total copper sulphides. Chalcocite, minor covellite and traces
of chalcopyrite constitute the remaining copper sulphides.
Other minor to trace elements that may contribute to the value
of the slag include nickel mainly present as pentlandite, molybde-
nite in solid solution with copper sulphides and possibly rhenium
also in solid solution with nickel–copper and sulphides and sphal-
erite. Some of the abundant zinc in Nababeep slag also reports to
the glass, silicate minerals and spinel phases and is refractory
and not considered for extraction.
Although prill size is variable, the dominant 50 lm fraction
and essential sulphide composition of the prills lends itself to
extraction by conventional flotation by grinding to 45 lm. The
poor recovery reported from previous attempts could have been
a function of grinding to 75 lm which resulted in poor liberation
and limited surface exposure of the prills. It is therefore anticipated
that grinding to 45 lm may significantly improve recovery of the
copper sulphide phases which account for most of the copper dis-
tribution. It is also anticipated that the minor exsolved sulphide
phases such as pentlandite may be recovered, as well as molybde-
num and rhenium present in solid solution with complex copper
sulphides. Metallic copper and the metal alloy prills will be sup-
pressed during sulphide flotation unless finely intergrown with
copper sulphides. The results of this study show that the metallic
and alloy phases are of minor concentration (10%) and will have
a limited effect on the overall copper recovery.
The very fine-grained prills (20 lm) are expected to show a
poor recovery, but in terms of their contribution to the overall con-
tained weight per cent copper, this loss will also be limited.
From the mineralogy of the copper-bearing phases and prill size
distribution it can be concluded that conventional sulphide flota-
tion would be the most appropriate method to beneficiate the cop-
per slags. Copper losses will be related to the fine prill size fraction,
elemental copper and metal alloy distribution and are expected to
be relatively limited. Theoretically it should be possible to signifi-
cantly improve upon the historic recovery of less than 50% copper
by grinding the slag to 45 lm. This should however be confirmed
by laboratory flotation tests. To optimise the potential revenue of
the slag resource, sulphide concentrates should be regularly as-
sayed for nickel, molybdenum and rhenium in addition to copper.
Several papers have recently appeared on the hydrometallurgi-
cal extraction of base metals from copper and also zinc smelter slag
in particular from research done in China (Yang et al., 2010; Jiang
et al., 2012). Although this new technology could be considered to
improve copper extraction, the advantages will have to signifi-
cantly outweigh the traditional sulphide flotation route and appear
unlikely at this stage.
7. Conclusions
An orientation textural, mineralogical and chemical study of
smelter slag from three localities in the Okiep Copper District in
South Africa have shown that the slags are significantly enriched
in copper with minor concentrations of zinc, tungsten, cobalt, nick-
el and molybdenum. Effectively all metals of economic significance
are hosted by discreet spherical prills disseminated throughout the
massive and vesicular iron–magnesium–silicate slag. The average
drill diameter is 50 lm but, should be statistically quantified by
a representative follow-up study of this more than 5 million ton
slag resource.
Mineralogically the prills consist predominantly of copper sulp-
hides with exsolved zinc, nickel, molybdenum and lead sulphide
phases, metallic copper and base metal alloys. Previous attempts
to extract copper from the slag by means of conventional flotation
have had limited success, however this study suggests that crush-
ing to 45 lm as opposed to 75 lm could significantly improve
recovery of the sulphide phases by flotation. This should be con-
firmed by a follow-up study. Copper will be the main earner of rev-
enue and could be supported by sweeteners from the minor base
metals and possibly traces of platinum group elements such as
rhenium.
Provided the extraction process is economically viable, the pro-
cessed slag can be redistributed to environmentally acceptable
sites and resolve the present pollution and rehabilitation chal-
lenge. From a socio-economic perspective the copper-from-slag
initiative could revitalize albeit on a small scale, the economy of
a poverty-stricken part of South Africa. The slag contains a poten-
tial resource of 15 000 tons of copper metal and processing
500 000 tons slag per annum a possible life of mine of 10 years is
envisaged.
Appendix A
Pyrometallurgical. route Nababeep smelter Okiep Copper District
Copper sulphide concentrate is obtained by flotation of the
crushed primary ore and assays between 20% and 40% copper.
The concentrate is mixed with high silica sand that serves as a flux
and fed into the reverberatory furnace for smelting. The slag is
skimmed and matte tapped and fed to the Peirce-Smith converter.
The temperature of the matte is around 1115 °C and depending on
the mineralogy of the concentrate it assays between 60% and 65%
copper for bornite-rich ore and 31–40% copper if dominated by
chalcopyrite. The slag is discarded to the slag dump and formed
the focus of the present study.
Chemistry of this process can be expressed as:
2CuFeS2ðsÞ þ 3O2ðgÞ ! 2FeOðsÞ þ 2CuSðsÞ þ 2SO2ðgÞ
FeOðsÞ þ SiO2ðsÞ ! FeO SiO2ðlÞ
2FeSðlÞ þ 3O2ðgÞ þ 2SiO2ðlÞ ! 2FeO SiO2ðlÞ þ 2SO2ðgÞ
Sand is added as flux to the converter matte. Converter slag pro-
duced assays around 6% copper and is returned to the reverbera-
tory furnace. The copper smelt is cast into blister copper bars or
1 ton blocks which assay more than 99% copper.
Chemistry of this process can be expressed as:
CuSðlÞ þ O2ðgÞ ! CuðlÞ þ SO2ðgÞ
2FeSðlÞ þ 3O2ðgÞ þ 2SiO2ðlÞ ! 2FeO SiO2ðlÞ þ 2SO2ðgÞ
Pulverized coal is used as fuel. (Fairfax, 1955; Kinsella and Goosen,
1996).
References
Candy, G., 2011. Copper Prices to Rise in Q1 2012 but Supply Concerns Remain,
December 2011. www.Mineweb.com.
A. Rozendaal, R. Horn / Minerals Engineering 52 (2013) 184–190 189

7. Cornelissen, A., 1965. A History of Copper Mining in Namaqualand. Privately
published, p. 120.
Fairfax, E., 1955. O’okiep smelting capacity keeps pace with expanded mine output.
Special O’okiep volume. Min. World 17 (06), 52–78.
Hohne, S., Hansen, R.N., 2008. Preliminary conceptual geo-environmental model of
the abandoned copper mines of the Okiep Copper District, Namaqualand,
Northern Cape. Council for Geoscience, Report Number 2008-070, April 2008, p.
194.
Hur, J., Sedgman, G., 2013. Morgan Stanley says Copper Price to Rise 7.6 percent in
2013 on demand, January 2013. www.Bloomberg.com/news.
Isky, G., 2009. O’Okiep Copper Mine at Nababeep Northern Cape SA 3. http://
www.iskyfoto.com/index.php?PhotoID=320KeywordID=150.
Jiang, K., Guo, Z., Xiao, X., Zhang, L., 2012. Extraction of metals from a zinc smelting
slag using two-step procedure combining acid and thylene diaminetetraacetic
acid disodium. J. Cent. South Univ. 19, 1808–1812.
Kinsella, B., Goosen, G.J.S., 1996. Shutdown of the Smelter at O’okiep Copper
Company Limited. Hidden Wealth, Johannesburg, 1996. The South African
Institute of Mining and Metallurgy, pp. 37–43.
Lombaard, A.F., Staff of the O’okiep Copper Company Limited, 1986. The copper
deposits of the Okiep District, Namaqualand. Mineral deposits of Southern
Africa. S. Afr. J. Geol. II, 1421–1445.
Potgieter, J.E., 1996. Exploration in the Okiep Copper District, Northern Cape
Province, South Africa: an overview. S. Afr. J. Geol. 99 (2), 209–220.
Rozendaal, A., 2011. Historic Resource Information on O’okiep Copper Company.
Unpublished Research Report, 2011, University of Stellenbosch, p. 10.
Smalberger, J.M., 1975. A History Of Copper Mining in Namaqualand. Publ. C Struik,
1975, Cape Town, p. 152.
Yang, Z., Rui-lin, M., Wang-dong, N., Hui, W., 2010. Selective leaching of base metals
from copper smelter slag. Hydrometallurgy 103 (1–4), 25–29.
Personal communication
Mr. S. van der Westhuizen. Director Siteplan Environmental Consultants. P.O.
Box 241 Strand, South Africa.
Mr. A. de Beer Manager O’okiep Copper Company 2001–2007, now retired.
Hermanus, Western Cape Province, South Africa.
190 A. Rozendaal, R. Horn / Minerals Engineering 52 (2013) 184–190