2. A previous study found some promising results to extract copper
from chalcopyrite concentrates by electrochemical bioleaching at lower
temperatures (≤50 °C)(Ahmadi et al., 2010a). In that research,
electrochemical and conventional bioleaching experiments were
carried out on a chalcopyrite concentrate using a mixed culture of
mesophilic bacteria at 35 °C and a mixed culture of moderately
thermophilic bacteria at 50 °C, operating at 10% (w/v) pulp density.
The results showed that the control of solution ORP around 425 mV (vs.
Pt, Ag/AgCl), by applying current directly to the slurry, significantly
increases both, the cell concentrations and copper recovery in both
cultures, especially in the presence of moderately thermophilic bacteria.
One of the main limitations of bioleaching processes in stirred
tanks, especially in the case of extreme thermophiles, is the negative
effect of high pulp densities on both extraction rate and final recovery
of metals. Oxygen and carbon dioxide availability, low bacteria–solids
ratio, metabolic stress by high shear stress and abrasive conditions,
inhibition of bacterial attachment, and the build-up of toxic leach
products or other detrimental substances (such as some flotation
reagents) have been reported as the most significant problems for a
successful operation of bioleaching at high solid contents (Bailey and
Hansford, 1993; Acevedo and Gentina, 2007). To overcome these
problems and meet the requirements of industry, microorganisms
must be adapted to high pulp densities (Mishra et al., 2005). However,
to date, there is no information in the literature on the electrochem-
ical control of the bioleaching process at high pulp densities of stirred
reactors (at 10% ; Ahmadi et al., 2010a). For this reason, the present
research work investigates the process of electrochemical bioleaching
at high pulp density and compares its efficiency to that of
conventional and electrochemical leaching processes in the presence
and absence of moderately thermophilic bacteria. The work was
executed by leaching of a Sarcheshmeh chalcopyrite flotation
concentrate in a stirred electrobioreactor.
2. Experimental
2.1. Materials
A mixed culture of moderately thermophilic bacteria supplied by
Sarcheshmeh Copper Mine, Kerman, Iran, was used. The microorgan-
isms were grown at 50 °C on Norris medium (0.4 g/L (NH4)2SO4, 0.4 g/L
K2HPO4, 0.5 g/L MgSO4.7H2O) with the copper concentrate at pulp
densities from 2% to 20% (w/v) replacing the energy source. The
flotation concentrate was obtained from Sarcheshmeh Copper Mine,
and contained 44.02% chalcopyrite (CuFeS2), 23.99% pyrite (FeS2), 6.87%
covellite (CuS), 5.84% chalcocite (CuS2), 13.61% non-metallic minerals
and 4.79% copper oxide minerals as shown by mineralogical analysis.
The chemical analysis of the representative sample is presented in
Table 1. Mineralogical investigations on the feed and solid residues were
performed by optical microscopy using a Leica phase contrast
microscope (DMLP). Since during the quantitative determination of
phases, the iron present in the iron hydroxide precipitates is ascribed to
chalcopyrite and pyrite minerals, the quantitative determination of
these residue analyse are not scientifically reliable; hence in this study
the results are reported only qualitatively. The transformation is
visualised in order to underline our understanding of the mechanism
of dissolution.
The particle size distribution of the concentrate was determined by
wet sieving and cyclosizer and showed that 80% is passing 76 μm
(Fig. 1).
2.2. Electrobioreactor
The leaching experiments were performed in a three-electrode, 2 L
glass electro-bioreactor with 4 baffles, thermostated at the desired
temperature by circulating water from a constant temperature bath
through the double-wall jacket (Fig. 2). The reactor had a medium
volume of 1.3 L with a medium height/diameter ratio equal to 1.1. The
leach slurry was mechanically stirred by a pitched blade impeller
(diameter=5 cm) mounted on a rotating shaft. A titanium–platinum
mesh (15 cm×9 cm×0.1 cm), acting as the cathodic working electrode,
was immersed into the reactor solution. A platinum foil was used as a
counter electrode and was put into a separate small anodic compart-
ment, separated from the cathode chamber by a glass frit. An Ag/AgCl
reference electrode was in contact with the electrolyte in the main
chamber through a Luggin capillary, which ended just short of the
working electrode. Air was supplied through a stainless-steel ring
sparger underneath the impeller.
2.3. Leaching experiments
Batch experiments were carried out in the electro-bioreactor at 20%
(w/v) pulp density, an initial pH of around 1.5, a temperature of 50 °C, a
stirring rate of 600 rpm, an aeration rate of 1 vvm (volume of air/volume
of slurry/min) and Norris nutrient medium with 0.02% (w/w) yeast
extract addition. The intense agitation is needed both for maintaining a
homogeneous suspension and increasing the rate of mass transfer
(especially of oxygen and carbon dioxide from the gas phase). During
the leaching experiments, the pH of the suspensions was monitored
periodically and adjusted to around 1.5 by addition of H2SO4 (6 M). The
variations of ORP were recorded daily throughout the leaching period.
The pH and ORP values were measured with a Jenway 3540 pH meter and
a Pt electrode in reference to an Ag/AgCl electrode (+207 mV vs. SHE at
25 °C), respectively. Samples were periodically taken from the slurry and
filtered through Whatman No.41 filter paper. After that the filtrate was
used for copper and iron analysis by the atomic adsorption method. The
remaining solids were returned to the reactor. The evaporated liquid was
periodically replaced by adding acidified distilled water (pH=1.5).
The biotic experiments were inoculated with 20% (v/v) of a culture
previously adapted to 20% pulp density. To maintain the ORP in the
desired range (400–430 or 440–480 mV) during electrochemical
bioleaching, the potential of the working electrode was controlled with
respect to the reference electrode with a Solartron Sl 1287 potentiostat/
galvanostat. To keep the ORP in the desired range, the applied potential
was manually set slightly lower than the set point value during the
experiments, however it was always higher than 250 mV.
The initial solutions of the conventional bioleaching, electrochem-
ical bioleaching at 400–430 mV and electrochemical bioleaching at
440–480 mV experiments contained 2.46, 3.15 and 2.35 g/L iron,
respectively, which originated from their inoculum solutions.
Table 1
Chemical analysis of the copper sulphide concentrate.
Elements Cu Fe S Si Al Zn Mg K
(wt.%) 27.50 23.03 14.82 3.87 1.45 0.99 0.40 0.24
Cumulativepassing(%)
Particle size (micrometer)
100
80
60
40
20
0
1 10 100
Fig. 1. Particle size distribution of copper concentrate.
85A. Ahmadi et al. / Hydrometallurgy 106 (2011) 84–92
3. To evaluate the contribution of acid solution to copper recovery, an
abiotic test of chemical leaching was carried out under the same
conditions as the bioleaching test, except the initial composition of their
solutions (in the abiotic tests there was no iron in the initial solution). In
addition, an abiotic electro-leaching test was conducted by applying a
fixed 100 mA current (no ORP control), to investigate the effect of
applied DC current to the slurry on the recovery of copper and iron,
while other conditions were kept similar to the abiotic chemical
leaching test. It should be mentioned that this electro-leaching test was
different from that performed in a previous study (Ahmadi et al., 2010a),
in which the working electrode was set as anode causing different
electrochemical reactions expected here are different from those that
occurred in the previous test. In the abiotic tests of this new study, the
medium was sterilized with 2% (v/v) bactericide (2% (w/w) thymol in
ethanol) added to prevent microbial growth. In order to sterilize the
reactor before each test, it was operated for 2 h at 85 °C and 800 rpm in a
solution of 10% (v/v) bactericide and 5% (v/v) HCl. Microbial growth
was periodically verified by observation under a Nikon optical
microscope (ECLIPSE, TE 2000-U). Free cells in solution were counted
by direct counting using a Thoma chamber of 0.1 mm depth and
0.0025 mm2
area with the optical microscope (magnification=1500×).
3. Results and discussion
In order to examine the influence of ORP and presence of bacteria on
the dissolution of chalcopyrite concentrate, various processes, namely
chemical leaching (control test), bioleaching, electro-leaching and
electrochemical bioleaching were carried out in an electro-reactor at
20% (w/v) pulp density. During evaluating the results and comparing
them with those obtained from the previous research at the lower pulp
density (Ahmadi et al., 2010a), it should be considered that the
experimental conditions of this study are different from those in the
previous work. One of the main differences relates to their different
stirring rate. In that research, because of fear of bacterial adaptation, the
stirring rate was set at 300 rpm in the first 4 days with the result that
during this period, a portion of the concentrate had settled at the bottom
of the reactor and didn't take part in the leachingreactions. This problem
was solved in the high pulp density by consecutive adaptation tests (4
times) at stirring rate of 600 rpm which was constant at the rate during
the main experiments.
3.1. Chemical leaching
Figs. 3 and 4 show the results of copper recovery (Fig. 3) and iron
dissolution (Fig. 4) from the concentrate by the various methods over
Fig. 2. Schematic illustration of thermostated electrobioreactor.
Fig. 3. Copper recovery as a function of leaching time at 50 °C and 20% pulp density for
different experimental conditions: chemical leaching (CL), electrochemical leaching
(ECL), bioleaching (BL) and electrochemical bioleaching (EBL).
86 A. Ahmadi et al. / Hydrometallurgy 106 (2011) 84–92
4. 10 days. As can be seen, the rate of chalcopyrite concentrate
dissolution in chemical leaching process (control test) is significantly
lower than those obtained in the other processes in which final values
of copper recovery and iron dissolution did not exceed 13% and 5%,
respectively. The main portion of this copper extraction is attributed
to the acid leaching of copper oxides and partial leaching of chalcocite
(Eqs. (1) and (4)) and covellite (Eqs. (2) and (5)) minerals. It is
presumed that chemical dissolution of chalcopyrite (Eqs. (3) and (6))
is negligible due to its high lattice energy (Habashi, 1978).
Cu2S þ 2H
þ
→Cu
2þ
þ Cu
o
þ H2S ð1Þ
CuS þ 2H
þ
→Cu
2þ
þ H2S ð2Þ
CuFeS2 þ 4H
þ
→Cu
2þ
þ Fe
2þ
þ 2H2S ð3Þ
Cu2S +
1
=2O2 + 2H
þ
→Cu
2+
+ CuS + H2O ð4Þ
CuS +
1
=2O2 + 2H
þ
→Cu
2+
+ S
B
+ H2O ð5Þ
CuFeS2 þ O2 þ 4H
þ
→Cu
2þ
þ Fe
2þ
þ 2S∘ þ 2H2O ð6Þ
The smell of rotten eggs during the process could be related to the
production of H2S through reactions (1) and (2).
Fig. 5 shows the variation of pH during various processes. The pH
value rises during the chemical leaching of the concentrate, which is
due to acid consumption by the reactions of copper oxide and
sulphide minerals (Eqs. (1)–(6)) as well as gangue minerals. It was
controlled at around pH 1.5 by the addition of H2SO4 (during the first
6 days). Fig. 6 shows that the ORP values remained low in the range of
300–330 mV.
The low extraction ratio of Fe:Cu is probably due to the dissolution
of iron-free minerals i.e. copper oxides, covellite and chalcocite and
the insolubility of refractory iron bearing minerals i.e. pyrite and
chalcopyrite under the conditions studied. Comparison of the
mineralogical analysis of the solid residue (after 10 days) (Fig. 7b)
with that of the feed concentrate (Fig. 7a) clearly reveals that
secondary copper bearing minerals i.e. covellite and chalcocite were
not dissolved.
3.2. Conventional bioleaching
To investigate the efficiency of the mixed culture of moderate
thermophiles at high pulp density, a conventional bioleaching
experiment was carried out in the stirred bioreactor. The optimal
conditions (temperature, 50 °C; initial pH, 1.5; nutrient medium,
Norris; and yeast extract, 0.02% (w/w)) obtained from a previous
study (Ahmadi et al., 2010b) were employed. During preliminary
bioleaching experiments (data not shown), the bacteria were adapted
to a high solid content and high degree of slurry agitation by
increasing the pulp density from 2 to 20% (w/v) and the stirring rate
from 300 to 600 rpm and then maintaining these conditions. The
adapted culture (the 4th generation of a culture maintained at 20%
pulp density and 600 rpm) was used to inoculate the bioleaching
experiment discussed here (inoculation volume=20% (v/v)).
Looking at Fig. 3, the copper recovery profile can be divided into
three distinct phases. Initially, approximately 21% of copper is rapidly
leached within the first day; this is mainly associated with the
dissolution of copper oxides by H2SO4 and the first stage leaching of
chalcocite (Eq. (7)). Thehigher initial copper extraction in this biotic test
compared to the chemical leaching test (21.4% vs. 8.1%) is mainly related
to their different initial solution compositions. In the bioleaching test,
the iron concentration in the initial solution is 2.46 g/L (mostly as ferric
iron), originating from the inoculation solution, which acts as a leaching
agent for copper sulphides (Eqs. (7) to (9)) and pyrite (Eq. (10)).
Cu2S þ 2Fe
3þ
→Cu
2þ
þ 2Fe
2þ
þ CuS ð7Þ
CuS þ 2Fe
3þ
→Cu
2þ
þ 2Fe
2þ
þ S∘ ð8Þ
CuFeS2 þ 4Fe
3þ
→5Fe
2þ
þ Cu
2þ
þ 2S∘ ð9Þ
FeS2 þ 8H2O þ 14Fe
3þ
→15Fe
2þ
þ 2SO
2−
4 þ 16H
þ
ð10Þ
Fig. 5. Variation of pH as a function of leaching time at 50 °C and 20% pulp density for
different experimental conditions: chemical leaching (CL), electrochemical leaching
(ECL), bioleaching (BL) and electrochemical bioleaching (EBL).
Fig. 6. ORP variation as a function of leaching time at 50 °C and 20% pulp density for
different experimental conditions: chemical leaching (CL), electrochemical leaching
(ECL), bioleaching (BL) and electrochemical bioleaching (EBL).
Fig. 4. Total iron recovery as a function of leaching time at 50 °C and 20% pulp density
for different experimental conditions: chemical leaching (CL), electrochemical leaching
(ECL), bioleaching (BL) and electrochemical bioleaching (EBL).
87A. Ahmadi et al. / Hydrometallurgy 106 (2011) 84–92
5. The initial copper extraction rate in this research is significantly
higher than that obtained in the previous study (Ahmadi et al.,
2010a), which could be attributed to the higher stirring rate employed
here (which prevents the setting of the solids.
The first phase is followed by another linear increase to about 51%,
between days 1 and 7. This increase is probably associated with the
leaching of chalcopyrite and covellite (natural mineral and the
product of chalcocite leaching in Eq. (7)) according to Eqs. (9) and
(8), respectively. This phase is followed by ceasing the copper
dissolution in the remaining days to the end of experiment.
Comparing the various processes in Figs. 3 and 4, it can be found
that the final values of copper recovery (51.6%) and iron dissolution
(15.9%) are, respectively, around 3.8 and 3.2 times higher than those
of the chemical leaching process. Mineralogical examinations showed
that secondary copper bearing minerals such as chalcocite and
covellite were not found in the bioleaching residue (Fig. 7c), although
they were present in the feed concentrate (Fig. 7a) and in the residue
of the chemical leaching experiment as well (Fig. 7b).
In the bioleaching process, bacteria oxidize the insoluble metal
sulphides, such as chalcopyrite, by indirect and/or contact mechanisms
(Sand et al., 2001). Iron- and sulfur-oxidizing bacteria catalytically
generate the leaching agents of [Fe3+
] and [H+
] according to Eqs. (11)
and (12), respectively, and then these agents, especially ferric iron,
dissolve the sulphide minerals (Eqs. (1)–(10)). In this regard, sulphur
Fig. 7. Mineralogical images of feed (a); solid residues of (b) chemical leaching, (c) bioleaching, (d ) electrochemical leaching, (e and f) electrochemical bioleaching, (Cv=covellite;
Cc=chalcocite; Ccp=chalcopyrite; Py=pyrite).
88 A. Ahmadi et al. / Hydrometallurgy 106 (2011) 84–92
6. oxidizing bacteria remove elemental sulphur, as a passivation layer,
from the surface of sulphide minerals during the acid production
process (Eq. (12)).
4Fe
2þ
þ O2 þ 4H
þ
þ Iron oxidizing acidophiles →4Fe
3þ
þ 2H2O
ð11Þ
S
B
+ H2O +
3
=2O2þ Sulphur oxidizing acidophiles →H2SO4 ð12Þ
Schippers and Sand (1999) proposed that the indirect mechanism
occurs via the thiosulphate route (primarily pyrite) or via the
polysulphides and sulphur route (most other sulphide minerals
such as chalcopyrite). However, in the contact mechanism bacteria
attach to the mineral surface and prepare the medium and then
facilitate the mineral attack through an electrochemical dissolution
involving ferric ions contained in the microbe's extracellular poly-
meric substances (EPS) (Sand et al., 2001). Tributsch (2001)
concluded that in practice suspended bacteria feed on chemical
species released by attached bacteria (cooperative mechanism).
The rapid increase of ORP to relatively high levels (Fig. 6) and the
pH decrease (Fig. 5) indicate that both, activity and growth of the
bacteria are very favourable under the conditions studied. It can be
seen that ORP increases from 390 to 540 mV (after day 7) with the lag
phase of bacterial growth being less than 1 day, whereas it was
significantly longer in the bioleaching experiment of the previous
study (Ahmadi et al., 2010a).
Furthermore, Fig. 5 shows the variation of pH during the bioleaching
experiment. During the first day of the experiment, pH increases and
sulfuric acid needs to be added to keep the reactor pH around 1.5. After
that the pH decreased gradually to a final value of 1.3, which indicates
that the amount of acid-production is more than that of acid
consumption during the transition phase. It should be noticed that the
initial upward trend of pH is due to the acid consuming reactions such as
the dissolution of copper oxides, chalcocite (Eqs. (1) and (4)), covellite
(Eq. (2) and (5)), chalcopyrite (Eq. (3) and (6)) and gangue minerals as
well as the bacterial oxidation of Fe(II) to Fe(III) (Eq. (11)), while the
subsequent decrease of pH is due to the activity of bacteria to produce
acid (Eq. (12)), the dissolution of pyrite (Eq. (10)) and the hydrolysis of
ferric iron to form jarosite (Eq. (13)). Jarosite is formed under the
conditions of high pH and high ferric iron concentration, as might be
expected from Eq. (13) (Stott et al., 2000).
3Fe
3þ
þ X
þ
þ 2HSO
−
4 þ 6H2O→XFe3ðSO4Þ2ðOHÞ6 þ 8H
þ
ð13Þ
where X+
=K+
,Na+
,NH4
+
and H3O+
.
Despite using a dilute nutrient medium (Norris), due to the high solid
content, the concentration of alkali ions increased in the solution, which
is a favorable factor for precipitation of jarosite at the high solution ORPs
facilitated by iron oxidizing bacteria. Moreover, the precipitation is not
reversible in chalcopyrite systems (Leahy and Schwarz, 2009); hence,
once formed, the later increase of acid concentration doesn't dissolve the
precipitate. Jarosite formation causes the removal of Fe3+
and essential
bacterial nutrients, suchas K+
orNH4
+
, from solution, potentially resulting
in a slowed-down process or even a complete stop. It restricts the flow of
bacteria, nutrients, oxidants and reaction products to and away from the
mineral surface (Hackl et al., 1995). The stoppage of copper dissolution
after day 7 is likely related to the passivation of chalcopyrite by jarosite
precipitates. In our previous study (Ahmadi et al., 2010a), a significant
amount of jarosite was found on the bioleaching residue, which was
confirmed by SEM/EDS analyses (Fig. 8). The low leaching rate in the
chemical leach could also be caused by the passivation of chalcopyrite
surface by polysulphide compounds (Biegler and Horn, 1985). It is likely
that both jarosite and polysulphides passivate jointly the mineral surface.
The increase of the rate of iron dissolution and the concomitant
decrease of the rate of copper dissolution in the final days of the
experiment (Fig. 4) are attributed to the leaching of pyrite and could
be explained by the mechanism described by Petersen and Dixon
(2006), in which the dissolution of chalcopyrite is preferentially
accelerated at low ORPs, while at higher ORPs, pyrite and covellite are
leached much faster than chalcopyrite.
During the analysis of iron dissolution profiles, it should be borne
in mind that these values do not take into account the portion of iron
leached from the concentrate which was subsequently precipitated as
iron hydroxides especially jarosite. This problem occurs to a a
significant extent during the conventional bioleaching test, where
the ORP values surpass 500 mV and jarosite precipitation would
occur. As noted above, this phenomenon was encountered in our
previous study (Ahmadi et al., 2010a). It should be noticed that iron
dissolution reported here represents the minimum iron recovery from
the concentrate.
3.3. Electro-leaching
It can be assumed that the electrical charging of semiconducting
metallic sulphide minerals such as chalcopyrite could be done as a
result of periodic electrical contact between mineral particles and the
working electrode. These contact interactions may occur in the
electrochemical bioleaching experiment when current is applied to
control the solution ORP. Hence, to investigate the occurrence of this
phenomenon and its influence on the copper recovery and iron
dissolution from the concentrate, an electro-leaching experiment was
carried out at a current density of 1 mA cm-2
(total direct curren-
t=100 mA) and 20% (w/v) pulp density. Such a low current density
was chosen to minimize the evolution of hydrogen (Eq. (14)) and was
Fig. 8. SEM image and EDS analysis of the solid residue of conventional bioleaching.
89A. Ahmadi et al. / Hydrometallurgy 106 (2011) 84–92
7. in the range of the current passing in the electrochemical bioleaching
tests discussed in the next subsection
2H
þ
þ 2e
−
→H2 ð14Þ
The values of copper recovery and iron dissolution from the
concentrate by the electro-leaching method are also presented in
Figs. 3 and 4. Fig. 3 shows that the final copper recovery by electro-
leaching (~20%) is significantly higher than that in the chemical
leaching process, but substantially lower than those in both the
conventional bioleach and the electrochemical bioleaching processes
discussed below. However, iron dissolution is initially very low,
whereas, the final dissolution is significantly higher than that obtained
in the chemical leachingtest (Fig. 4). As Fig. 6 depicts, duringthe electro-
leaching process, the ORP profile is slightly below that measured in the
chemical leaching process. Because of low ORP, the dissolution of pyrite
as the main iron bearing phase is very slow. Moreover, similar to what
was measured in the other experiments, the solution pH rises initially
due to the initial acid consuming reactions (Fig. 5).
Using the electro-leaching method, some reactions would be expected
to occur as a result of passing direct current across the slurry. As noted by
Warren et al. (1982), electrochemical reactions of a mineral are a direct
result of the thermodynamic properties of the mineral, properties of the
electrolyte, and their interaction at the mineral-electrolyte interface.
The data in Figs. 3 and 4 suggests that during the electro-leaching
process, the extraction ratio of Fe:Cu rises from 0.28 in day 1 to 0.45 at the
end of the test. This increase could be related to the electro-reduction of
chalcopyritetochalcocite(Eqs.(15)and (16))andremovingironfromthe
chalcopyrite lattice as described by Biegler and Swift (1976).
2CuFeS2 þ 6H
þ
þ 2e
−
→Cu2S þ 2Fe
2
þ 3H2S ð15Þ
CuFeS2 þ 3Cu
2þ
þ 4e
−
→2Cu2s þ Fe
2þ
ð16Þ
This result was confirmed by mineralogical analysis (Fig. 7d), in
which chalcocite was found around chalcopyrite particles in the solid
residue (blue-gray region). Chalcocite could also be produced by electro-
reduction of covellite according to the following equation (Elsherief
et al., 1995).
2CuS þ 2H
þ
þ 2e
−
→Cu2S þ H2S ð17Þ
This reduction increases the overall dissolution rate of covellite or
stage (II) of chalcocite leaching (Eq. (8)) as the rate of stage (I)
chalcocite leaching (Eq. (7) is more rapid.
Biegler and Swift (1976) reported that at high current densities
(N10mAcm−2
) metallic-copper could be deposited on the cathode
electrode. This deposition may be as a result of a cathodic reaction
described by Eq. (18) or due to the direct electroplating of Cu2+
in
solution (Eq. (19)).
Cu2S þ 2H
þ
þ 2e
−
→2Cu∘ þ H2S ð18Þ
Cu
2þ
þ 2e
−
→Cu
∘
ð19Þ
It should be noted that to check the formation of copper on the
cathode during the process, a portion of the slurry was occasionally
taken out and returned instantly to the reactor. No copper deposit was
observed at any stage during the test. However, if copper is deposited,
it can be a reductive agent for chalcopyrite leaching according to
Eq. (20) as reported by Hiskey and Wadsworth (1981). Therefore the
deposit could not have been formed.
2CuFeS2 þ Cu
o
þ 2H
þ
→Cu2S þ Fe
2
þ H2S ð20Þ
In terms of galvanic interactions between different metallic
sulphide minerals, Mehta and Murr (1983) observed that when
pyrite, chalcopyrite, chalcocite and covellite are in contact with each
other, due to their different rest potentials the rate of covellite
dissolution would be the fastest of all (anodic corrosion), followed by
chalcocite and chalcopyrite, with the rate of pyrite dissolution the
lowest of all (cathodic protection).
Furthermore, it has been reported (Warren et al., 1982; Lu et al.,
2000) that chalcopyrite may be electrochemically converted to other
copper sulphide phases such as talnakhite (Cu9Fe8S16) (Eq. (21)) or
bornite (Cu5FeS4) (Eqs. (22) and (23)), which would be dissolved
faster than chalcopyrite.
9CuFeS2 þ 4H
þ
þ 2e
−
→Cu9Fe8S16 þ Fe
2þ
þ 2H2S ð21Þ
5CuFeS2 þ 12H
þ
þ 4e
−
→Cu5FeS4 þ 4Fe
2þ
þ 6H2S ð22Þ
2CuFeS2 þ 3Cu
2þ
þ 4e
−
→Cu5FeS4 þ Fe
2þ
ð23Þ
H2S produced during the process could be oxidized in the presence
of Fe(III) (Eq. (24)) and/or oxygen (Eq. (25)) or undesirably led to the
precipitation of covellite in the presence of Cu(II) (Eq. (26)).
H2S þ 2Fe
3þ
→S∘ þ 2Fe
2þ
þ 2H
þ
ð24Þ
H2S +
1
=2O2→SB + H2O ð25Þ
Cu
2þ
þ H2S→CuS þ 2H
þ
ð26Þ
3.4. Electrochemical bioleaching
Two electrochemical bioleaching experiments were carried out,
one at an ORP in the range of 400–430 mV and one at an ORP in the
range of 440–480 mV over 10 days.
The values of copper recovery and iron dissolution from the
concentrate by electrochemical bioleaching are shown in Figs. 3 and 4,
respectively. As can be seen in Fig. 3, the highest copper recovery
among the various processes was obtained in the experiment
conducted at the ORP range of 400–430 mV, in which copper recovery
after 10 days reached 77%. When the ORP was kept higher, in the
range of 440–480 mV, Cu dissolution leveled off after the 3rd day
despite showing the highest initial rate of copper extraction in all
tests. The final recovery (66%) was significantly lower than that
obtained during the experiment conducted in the range of 400–
430 mV, however it was still higher than that obtained by conven-
tional bioleaching (~51%), electro-leaching (~20%) and chemical
leaching (~13%). The higher initial copper extraction rate in the first
3 days of electrochemical bioleaching at 440–480 mV could be related
to the higher ORP values in this period which is more favorable for the
leaching of chalcocite and covellite minerals. In these biotic tests, the
initial iron concentrations, originating from the inoculated solution,
were 3.15 g/L (ferric iron concentration=2.21 g/L) and 2.35 (mostly
as ferric iron) g/L in tests conducted at 400–430 mV and 440–480 mV,
respectively. Ferric iron acts as a leaching agent for copper sulphides
and would lead to a high initial rate of copper dissolution from the
concentrate. Fig. 4 illustrates that the dissolution of iron during
electrochemical bioleaching in the range of 400–430 mV increases
linearly up to around 30%, as compared to 23%, 16%, 9% and 5% during
electrochemical bioleaching at the range of 440–480 mV, conventional
bioleaching, electro-leaching and chemical leaching processes, respec-
tively. It should be noted that in electrochemical bioleaching the
dissolution of iron is significantly lower than that of copper (29%
against 77%), similar to what was observed in the other processes
90 A. Ahmadi et al. / Hydrometallurgy 106 (2011) 84–92
8. studied. The main reason for the low Fe:Cu extraction ratio in
electrochemical bioleaching could be the low solubility of pyrite as the
main iron bearing phase in the concentrate at the prevailing potential.
Mineralogical analysis of the solid residue from the electrochemical
bioleaching at 400–430 mV (Fig. 7e) shows that the chalcopyrite was
corroded much more, while the pyrite surface remained almost
unaffected.
The amount of current passing through the reactor varied from
100 to 450 mA. As has been shown, both, applying current directly
into the slurry and controlling the ORP, have a positive effect on both
the chemical and biological sub-systems. In this regard, the increased
copper extraction rate is postulated to be due to the three following
reasons: firstly, by electrochemical reduction of ferric iron, the
concentration of ferrous iron which is a source of energy for bacteria,
increases, so their growth and activity would be enhanced. Enumer-
ation of bacterial populations in the final solutions showed that the
cell density increased from about 4×108
cells/ml for conventional
bioleaching to about 9×108
cells/ml for electrochemical bioleaching
at 400–430 mV. Previous work done by the authors (Ahmadi et al.,
2010a) also showed that after 5 days the cell concentrations of both,
mesophilic and moderately thermophilic cultures, were about 3–4
fold higher in electrobioleaching slurries than those counted in the
related conventional bioleaching processes. Natarajan (1992) and
Nakasono et al. (1997) also found that applying a current into a
solution containing Acidithiobacillus ferrooxidans increases the cell
concentrations substantially. Industrially, this result is very important,
because low bacteria–solids ratios have been cited as one of the main
problems of the bioleaching processes at high pulp densities (Bailey
and Hansford, 1993). The second reason is that the highest dissolution
rate of chalcopyrite occurs at low solution ORPs around 400 mV which
has been explained by Hiroyoshi and co-workers (Hiroyoshi et al.,
1997, 2000, 2001). This effect has also been confirmed by other
authors (Pinches et al., 2001; Third et al., 2002; Sandström et al., 2005;
Cordoba et al., 2008; Gericke et al., 2010). Moreover, SEM/EDAX
examinations of our previous research (Ahmadi et al., 2010a) showed
that electrochemical bioleaching of chalcopyrite concentrate at
around 425 mV, significantly reduces the amount of jarosites on the
chalcopyrite surface. Conner (2005) also demonstrated that applying
DC current into the bioleaching solutions significantly reduces the
formation of jarosite in the solid residues. The third reason is that, as
observed in the electro-leaching experiment, periodic electrical
contact between chalcopyrite and the working electrode electro-
chemically reduces chalcopyrite to more soluble minerals such as
chalcocite and covellite. Mineralogical observations of the residue of
electrochemical bioleaching at 400–430 mV clearly visualized the
coverage of chalcocite and covellite minerals on the surface of not
leached chalcopyrite which is attributed to the reduction of
chalcopyrite to these secondary copper sulphides through both direct
electron transfer to the mineral and indirect reduction by Fe(II) at the
low solution ORP. This reason was verified by electrochemical
analyses (Ahmadi, 2010c) and was also reported previously by
Biegler and Swift (1976).
4. Conclusions
Conventional and electrochemical (bio)-leaching were explored to
extract copper from Sarcheshmeh chalcopyrite flotation concentrate
at high pulp density using a mixed culture of moderately thermophilic
bacteria adapted to 20% (w/v) pulp density. Leaching of copper was
found to be most efficient during bioleaching with the ORP
electrochemically controlled between 400 and 430 mV. The final
copper recovery of electrochemical bioleaching at 400–430 mV (77%)
was higher by a factor of 5.9, 3.9, 1.5 and 1.17 relative to that of
chemical leaching (control test), conventional bioleaching, electro-
leaching and electrochemical bioleaching at 440–480 mV, respectively.
Mineralogical observations of the sold residues of electrochemical
processes showed that chalcopyrite is converted to chalcocite and
covellite minerals which is attributed to both direct electron transfer to
the mineral and indirect reduction by Fe(II) at the low solution ORP.
Prevention of formation of passive layers such as jarosites as a
result of low slurry ORP, increase of the bacterial concentration and
the electro-reduction of chalcopyrite to less refractory minerals such
as chalcocite and covellite are considered to be the main reasons for
enhancing both, the dissolution rate and final copper recovery of the
concentrate in the electrochemical bioleaching process.
From the results of this research work, it can be concluded that the
electrochemical bioleaching could be considered as one of the most
promising alternatives to extract copper from high grade chalcopyrite
concentrates. In this regard the development of the process by
establishing a continuous system and the assessment of economical
parameters are needed to justify conducting the process in larger
scales.
Acknowledgments
The authors would like to appreciate the support of the National
Iranian Copper Industries Company (NICICO) especially Mr. Reza
Atashdehghan, Mr. Saeid Ghasemi and Mrs. Zahra Manafi.
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