chapter 1 literature review chapter 1

Chapter 1 Literature Review

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Literature Review

CHAPTER 1


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1.1. INTRODUCTION Bioleaching refers to a microbial mediated metal dissolution process to recover

metals of value such as copper, nickel, zinc, uranium and cobalt from sulphide

minerals. These metals become soluble during the bioleaching process. The

solutions are then treated for maximum metal recovery through a solvent

extraction (SX) and electroplating process. Occasionally, the term “biooxidation”

is also used to describe this process. However, there is a difference between the

term “bioleaching” and “biooxidation” [Brierley, 1997].

“Biooxidation” describes the microbiological oxidation of minerals, with the

difference; the metal of value is not solubilized, but remains in the solid residue in

a more concentrated form. The biooxidation process is being used by the gold

mining industry as a pretreatment process (biobeneficiation process) for

removing pyrite and arsenopyrite from refractory or recalcitrant gold ores. These

types of ores are difficult to solubilize with cyanide without removing the sulphide

minerals first. Commercial biooxidation processes have been successfully

applied in South Africa and in other countries for the pretreatment of gold-bearing

concentrates prepared by flotation [Rawlings and Silver, 1995].

Unaware of the microbial benefit, bioleach technology on copper

hydrometallurgical extraction was already implemented in China, 1086 A.D., in

the Northern Sung period (960–1126 A.D.). This technology was used by the

Chinese to produce copper on a commercial scale, many centuries before any

other nation, with a process thought to be developed by medieval Chinese

alchemists [Lung, 1986]. Although the leaching of metals from sulphide minerals

have a very distant historical background [Ehrlich, 2001], the role of

microorganisms during this process was only discovered in the mid twentieth

century. In 1947, Colmer and Hinkle discovered that acid mine drainage (AMD)

was caused by the bacterial oxidation of pyrite in coal seams.


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Modern bioleaching was born when one of the organisms responsible for AMD,

the iron and sulphur oxidizing bacterium, Acidithiobacillus ferrooxidans, was

isolated and described [Temple and Colmer, 1951]. The first patent on

bioleaching was granted in 1958 [Zimmerley et al., 1958]. The patent described

a process where ferric iron (Fe3+) and sulphuric acid (H2SO4) were used for metal

sulphide ore oxidation. After mineral oxidation, the reduced iron or ferrous iron

(Fe2+) was continuously oxidized to ferric iron through iron oxidizing bacteria.

Commercial applications of bioleaching were started with the bioleaching of

copper from submarginal-grade, run-off-mine material. The Kennecott Copper

Corporation in the USA was the first company to implement bioleach technology

in extracting copper from low grade copper deposits [Bryner and Jameson,

1958].

1.2. THE BIOLEACHING OF MINERAL SULPHIDES

During the early days of bioleach research, Acidithiobacillus ferrooxidans, was

considered to be the principal organism in the bioleaching of mineral sulphides

[Lundgren and Silver, 1980]. With research conducted on pyrite (FeS2) oxidation,

using Acidithiobacillus ferrooxidans, Silvermann and Ehrlich [1964] proposed two

mechanisms by which microorganisms catalyze the dissociation of metal

sulphides.

1.2.1. The indirect leaching mechanism

The first mechanism, which was termed “indirect leaching”, described the role of

ferric iron as oxidant in the process of mineral sulphide oxidation. Ferric sulphate

was believed to originate from pyrite oxidation in aerated water according to the

following reactions:

FeS2 + 3.5O2+ H2O FeSO4(aq) + H2SO4(aq)

4 FeSO4(aq) + O2 + 2 H2SO4(aq) 2Fe2(SO4)3(aq) + 2H2O


4

The second reaction proceeded rapidly in the presence of an iron oxidizing

biocatalyst, in this case, Acidithiobacillus ferrooxidans. The ferric iron in solution

oxidized the metal sulphide to corresponding sulphides and elemental sulphur,

solubilizing the metal of interest:

FeS2 + Fe2(SO4)3(aq) 3FeSO4(aq) + 2S0

The ferric iron, in turn, was reduced to ferrous iron. The bacterium catalyzed the

cyclic regeneration of ferrous to ferric to promote continuous leaching of the

sulphide mineral. Before 1980, the classical understanding of indirect pyrite

leaching was generally accepted to proceed according to the following reactions

[Suzuki, 2001]:

FeS2 + 2Fe3+ 3Fe2+ + 2S0 (Chemical oxidation with ferric ions)

3Fe2+ + 0.75O2 + 3H+ 3Fe3+ + 1.5H2O (Bacterial mediated ferrous oxidation)

2S0 + 3O2 + 2H2O 2SO42- + 4H+ (Bacterial mediated sulphur oxidation)

1.2.2. The direct leaching mechanism

The second proposed mechanism, with much debate, was described as a direct

oxidative attack of attached organisms on the metal sulphide’s surface,

independent of ferric iron as oxidant:

MS + 2O2 M(aq) + SO42-

This mechanism was termed “direct leaching” and was believed to involve

intimate microbial and mineral contact with direct enzymatic oxidation of the

mineral under aerobic conditions [Silvermann and Ehrlich, 1964; Silvermann,

1967]. Evidence for the direct leaching mechanism originated from work done

with scanning electron microscopy (SEM), indicating that thiobacilli colonizes the


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mineral’s surface, eventually forming corrosion “pits”, consistent with the size and

shape of the bacteria [Murr and Berry, 1976; Bennet and Tributsch, 1978]. The

attachment of organisms to the mineral’s surface was not debated, but did not

truly establish the existence of a direct leaching mechanism. Furthermore, closer

examination of these corrosion “pits” showed that they have a hexagonal cross

section, typically associated with chemical etching and not microbial activity

[Fowler et al., 2001].

Silverman [1967] redefined his original two mechanisms (direct and indirect) to

include an indirect contact leaching mechanism, termed “contact leaching”.

1.2.3. Contact leaching

In this mechanism, the attached organism oxidized ferrous to ferric iron in an

artificially controlled extra cellular polymeric zone between the organism and the

mineral. This caused the ferrous to ferric iron cycle to occur very rapidly,

increasing the ferric concentration and enhancing the leaching rate of the mineral

[Silverman, 1967; Tributsch, 2001]. In addition to the role of ferric iron during

contact leaching, Tributsch [1976] published SEM scans of etchings where

bacteria were attached. The author concluded that the corrosion “pits” were

formed by a strong oxidizing agent, secreted at the point of microbial attachment.

This process was revised by Tributsch [2001], demonstrating two separate

contact leaching mechanisms, using Acidithiobacillus ferrooxidans (iron and

sulphur oxidizer) and Leptospirillum ferrooxidans (only iron oxidizer) as model

organisms. Leptospirillum ferrooxidans has been described in Rawlings et al.

[1999] as well as Coram and Rawlings [2002]. The basis of this mechanism was

grounded on the argument that during the initial phase of pyrite biooxidation, the

bacteria cannot obtain sufficient energy from chemical species in the surrounding

leach solution alone, but have to produce an additional energy carrier or

artificially increase the chemical oxidant (Fe3+) on the pyritic surface.


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Tributsch and Bennett [1981 A and 1981 B] indicated that the valence bonds of

FeS2 are associated with the metal’s electron orbitals, thus, not contributing to

the chemical bond between the sulphur and iron in the crystal structure.

Consequently, FeS2 was only dissolved through a ferric iron attack and not

through proton (H+) attack, making the mineral insoluble in acid.

Acidithiobacillus ferrooxidans acquired the use of a poly-sulphide forming carrier

molecule that functions with a thiol-group (SH-), originating from the amino acid

cysteine, to dissolve the pyrite in an acidic media. The dissolving pyrite brought

ferrous iron in solution, which is oxidized to ferric iron by the bacteria, which in

turn caused the ongoing leaching of the pyrite. Leptospirillum ferrooxidans,

without the ability to oxidize sulphur or SH-, has adapted to leach pyrite with

electrochemical surface polarization. This organism increased the pyrite’s

surface potential through applying high concentrations of ferric iron to the

sulphide’s surface; thereby the organism used the electron extraction for

depolarization, which electrochemically dissolved the pyrite. The high surface

potential chemically converted the sulphide (S2-) to thiosulphate (S2O32-) and

sulphate (SO42-) [Tributsch, 2001]. Acidithiobacillus ferrooxidans could directly

dissolve pyrite without ferric ions and Leptospirillum ferrooxidans could obtain

energy through ferrous oxidation (originating from the pyrite’s surface) in a high

potential (low ferrous) environment. It is for this reason that Acidithiobacillus

ferrooxidans predominates during the initial phases of bioleaching (lower

potential) and is outgrown by Leptospirillum ferrooxidans as the potential

increases [Rawlings et al., 1999].

Since Silverman and Ehrlich [1964] published their paper, much research and

debate has centered on the direct and contact leach mechanism. No conclusive

results were presented on the mechanism’s existence, or whether it enhances

the rate of mineral sulphide oxidation, above that of purely chemical reactions

with soluble ferric iron [Fowler et al., 1999]. It is now generally accepted that the


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role of microorganisms is to provide chemical oxidants (ferric iron and protons)

and an efficient reaction space for sulphide mineral leaching to take place.

1.2.4. The thiosulphate mechanism Unconvinced with previously described mechanisms of pyrite oxidation,

Schippers et al. [1996] proposed a novel cyclic leaching mechanism which was

an indirect leaching mechanism. The authors primarily focused on intermediate

sulphur species that are formed during pyrite oxidation, catalyzed by

Acidithiobacillus ferrooxidans, Leptospirillum ferroxidans and chemically with

ferric ions. They described a ferric iron mediated pyrite leaching mechanism via

thiosulphate and polythionates, the thiosulphate mechanism. The thiosulphate

(S2O32-) mechanism was exclusively dependent on the oxidative attack by ferric

iron on acid-insoluble metal sulphides i.e. pyrite (FeS2) and molybdenite (MoS2).

The thiosulphate mechanism can be simplified through the following reactions:

FeS2 + 6Fe3+ + 3H2O S2O32- + 7Fe2+ + 6H+

S2O32- + 8Fe3+ + 5H2O 2SO4

2- + 8Fe2+ + 10H+

Within this mechanism the sulphide group (-S2) of the pyrite was oxidized to a

thiosulphate group by ferric iron. Hydrolysis yielded thiosulphate and ferrous iron

in solution. The soluble thiosulphate was oxidized by ferric iron to sulphate via

tetrathionate, disulfane-monosulfonic and trithionate. Experimental data showed

that elemental sulphur was not the main sulphur moiety formed during the

thiosulphate mechanism, but was merely regarded as a by-product. The authors

concluded that the only function of Acidithiobacillus ferrooxidans and

Leptospirillum ferroxidans was to supply ferric ions through ferrous oxidation.

This data is consistent with microbial attachment and ferric iron generation

through contact leaching.


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1.2.5. The polysulphide mechanism

In 1999, Schippers and Sand expanded on the thiosulphate mechanism to

include not only pyrite dissolution, but also mineral sulphides as a whole. They

proposed that mineral sulphides leach via two distinct indirect mechanisms, the

thiosulphate and the polysulphide route. The mechanism (thiosulphate or

polysulphide), according to which a mineral sulphide leached, was depended on

the acid solubility of that mineral. Acid soluble mineral sulphides like sphalerite

(ZnS) and chalcopyrite (CuFeS2) have shared electrons from both the metal and

the sulphur within the valence bands of crystal structure [Tributsch and Bennett,

1981 A; Tributsch and Bennett, 1981 B]. The bacterial leaching of acid soluble

sulphides proceeded via the polysulphide mechanism and was described by the

following schematic diagram and simplified reactions:

MS [H2S*+ HS* H2Sn] S8

Figure 1. A schematic diagram for the leaching of acid soluble mineral sulphides

via the polysulphide mechanism in the presence of ferric iron.

The dissolution of the metal sulphide (MS) was initiated through proton (H+)

attack on the crystal lattice with hydrogen sulphide (H2S) as reaction product.

This oxidative mechanism did not require ferric to leach the mineral. The

reaction pathway could proceed to elemental sulphur without ferric iron in the

system. In the absence of ferric iron, oxygen acted as electron acceptor, and was

essential for the mechanism to proceed beyond the formation of hydrogen

sulphide [Tributsch and Gerischer, 1976].

H2Sn + 3/2O2 H2S2O3 + [(n – 2)/8]S8

2H+ M2+

Fe3+ Fe2+

H+ H+

Fe3+ Fe2+


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Without oxygen or ferric iron the metal was solubilized through the following

reaction:

MS + 2H+ M2+ + H2S

With ferric iron present, the hydrogen sulphide was in turn oxidized to an acidic

hydrogen sulphide radical (H2S*+) by the ferric iron. Ferric iron, normally present

in bioleach systems, was shown to be more potent in attacking the crystal

structure of the mineral than protons, which lead to the hydrogen sulphide radical

being formed without occurring as hydrogen sulphide first [Tributsch and Bennett,

1981 A; Tributsch and Bennett, 1981 B]. The polysulphide formation started with

the decomposition of the unstable acidic hydrogen sulphide radical to form a new

radical species (HS*), which in turn reacted with each other to form polysulphides

(H2Sn). In an acidic environment the polysulphides were converted to elemental

sulphur (S8). In the presence of ferric iron the mechanism did not require

oxygen.

MS + Fe3+ + H+ M2+ + 0.5H2Sn + Fe2+ (n ≥ 2) 0.5H2Sn + Fe3+ 0.125S8 + Fe2+ + H+

Net reaction:

MS + 2Fe3+ M2+ + 0.125S8 + 2Fe2+

The main microbial function within the polysulphide mechanism was to generate

sulphuric acid (proton supply for hydrolysis) from elemental sulphur oxidation and

to generate ferric from ferrous iron oxidation.

0.125S8 + 1.5O2 + H2O SO42- + 2H+ (microbial mediated sulphur oxidation)

2Fe2+ + O2 + 4H+ 2Fe3+ + 2H2O (microbial mediated ferrous oxidation)


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Important observations in the leaching of acid soluble sulphide minerals through

the polysulphide mechanism are thus made according to the discussion above:

• Ferric iron was said to be more efficient in leaching acid soluble minerals

than protons.

• The mechanism did not require ferric iron in order leach the mineral.

Electrons could be extracted from the acid soluble sulphide mineral

through acid hydrolysis.

• Acidithiobacillus ferrooxidans acted as a strictly sulphur oxidizing microbe

in the absence of soluble iron and can enhance the leaching of acid

soluble minerals through sulphur oxidation only (proton generation).

• In the absence of ferric iron, the mechanism could utilize oxygen in order

to proceed to elemental sulphur formation.

1.3. THE LEACHING OF CHALCOPYRITE

The extraction of copper from chalcopyrite (CuFeS2) is an essential process for

the copper mining industry. The majority of the world’s remaining copper

sulphide resources exist as chalcopyrite bearing ore. Chalcopyrite is almost

exclusively processed through the pyrometallurgical roasting of chalcopyrite

concentrates. In the traditional smelting-refining process, chalcopyrite ore is

firstly concentrated in a flotation process and then smelted in a

reverberatory/flash smelter. This process is preferred by the mining industry

because of high copper extraction in short retention times.

A negative aspect of smelting concentrate is that the metal grade of the ore must

be high in order to produce a desirable metal concentrate for smelting. Lower

grade ore can not be processed through smelting. The efficient extraction of

base metals from low grade ore is becoming more important because high grade

ore deposits are on the decrease around the globe.


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The hydrometallurgical process of heap leaching offers an economic solution for

the treatment of chalcopyrite in low and marginal grade chalcopyrite ore. The

leaching of chalcopyrite in these systems can be categorized according to the

lixiviant i.e. chloride, nitrate, amine and sulphate. Sulphate systems are the most

commonly used because of the ease of copper recovery from sulphate media

with solvent extraction and electro winning. Unfortunately, chalcopyrite leaches

slowly in a ferric sulphate system. The reason for the slow and incomplete

leaching (rapidly declining from initial rates) is a topic of much historical research

and debate [Hackl et al., 1995]. The leaching of chalcopyrite in a sulphate

system is dependent on ferric iron, oxygen and protons as oxidants and the

oxidation of sulphur moieties to sulphate or other sulphur intermediates. The

leaching process can be purely chemical or biological of nature. During the past

half century, significant research effort has been devoted to the leaching of

chalcopyrite in a sulphate system.

1.3.1. The basic semiconductor electrochemistry of chalcopyrite

Almost all mineral sulphides, including chalcopyrite, processed through

hydrometallurgical processes are semiconductors. These minerals are

conductors of electricity and the dissolution reactions involve the transfer of

electrons from the mineral to the aqueous reactants. The dissolution of mineral

sulphides can be described according to the molecular band theory [Crundwell,

1988; Osseo-Asare, 1992].

The electrons associated with atoms are distributed among energy levels. The

energy levels of an atom/molecule increase with increasing distance from the

nucleus. These energy levels consist of sublevels or atomic orbitals. When two

atomic orbitals interact they form molecular orbitals, which can be divided in

bonding, antibonding and nonbonding orbitals. The molecular orbitals of a metal

sulphide are formed by combining the atomic orbitals of the metal with those of

the non-metal ligands. The bonding molecular orbitals are filled with electrons


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(electron dense) and the antibonding orbitals are combined atomic orbitals, but

are void of electrons.

When atomic orbitals are not aligned to contribute to bonding, the orbitals are

called nonbonding. An inorganic solid consists of many atoms associated with

each other to form very electron-dense bonding orbitals, which will appear as a

band of molecular orbitals. These closely spaced bonding orbitals are called the

valence band and the corresponding electron empty antibonding orbitals, the

conducting band. The uppermost energy levels within the valence band are

called Ev and lowest energy level of the conducting band is denoted as Ec.

These energy levels do not normally overlap and are separated by an energy

gap, Eg. The width of this energy gap characterizes a solid material as a metal,

semiconductor or an insulator.

Figure 2. The band theory for solid materials.

All metals have overlapping energy bands of occupied and unoccupied (partially

filled) orbitals. Metals have no energy gap and electrons can easily flow between

the valence band and the conducting band. Semiconductors and insulators are

characterized in having an energy gap; Eg > 2eV for insulators and Eg < 2eV for

semiconductors [Osseo-Asare, 1992]. The energy gap for semiconductors is

termed the forbidden energy gap. When an electron is removed from the

valence band, an electron hole occurs within the bonding orbitals of the valence

band. It is the continuous formation of these electron holes (charge/electron

E Conducting

band

Valence band

Conducting band

Valence band

Energy gap

Conducting band

Valence band

Energy gap Ec

Ec

Ev

Ev

Metal Insulator Semiconductor


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transfer) that weakens the covalent bonds between atoms, promoting dissolution

of the mineral [Crundwell, 1988]. For electron transfer to occur at a solid/liquid

interface, the redox potential (Ered, Eox) of the aqueous species should fluctuate

in close proximity of the energy bands of the semiconductor (Ec, Ev).

Figure 3. Direct electron transfer from the valence band (A). The transfer of an

electron via the conducting band (B).

Depending on energy orientations, electrons can be either extracted from the

conducting bands or directly from the valence bands. When the redox potential

of the oxidant/reductant couple is close to the highest energy level of the valence

band, electrons can be transferred directly from the valence band to the oxidant

(Figure 3A). Alternatively, when the redox potential of the oxidant/reductant

couple is close to the lowest energy level of the conducting band, electrons need

to be excited through the forbidden energy gap, from the valence band to the

conducting band. Only then are electrons transferred from the conducting band

to the oxidant (Figure 3B).

An increase in temperature on the system enhances the frequency of this

electron excitation. With continuous electron transfer to the available energy

bands of the aqueous redox couples, a thin film on the surface of the

semiconductor (in contact with the solution) is rendered more positively charged

than the rest of the material. This region is called the “space charged region”.

Ec

Ev

Eox

Ered

Solid Solution

Ec

Ev

Eox

Ered

Solid Solution

e-

e-e-

A B


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The resulting difference in potential between the “space charge region” and the

surrounding matter can cause the energy bands (Ev and Ec) of the semiconductor

to bend up or downwards, an n-type or p-type semiconductor respectively. In

the case of an n-type semiconductor, like chalcopyrite, the more positively

charged “space charge region” causes a rise in the energy levels, (Evs and Ecs) of

the material with respect to the flow of electrons to the aqueous redox couple

(Figure 4). Higher energy levels within the aqueous redox couples are thus

needed for charge transfer to proceed [Crundwell, 1988; Osseo-Asare, 1992].

Figure 4. The energy level increases (ΔE) behavior of an n-type semiconductor

i.e. chalcopyrite, exposed to continuous electron transfer, to the aqueous redox

species.

The exact band energies of chalcopyrite are difficult to determine because of

antiferromagnetism, but studies on similar crystal structure compounds indicated

the Fe 3d orbital as the lower orbital of the conducting band, while the Cu 3d and

S 3p orbital constitute the highest energy orbital from the valence band. The

band gap of chalcopyrite was estimated at approximately 600 mV [Shuey, 1975].

The ability of an aqueous redox couple to donate or accept electrons is given by

the redox potential (volt) and is measured as a reduction potential, (Eh) with

Cha

rge

spac

e

Ec

Ev

n-type semiconductor

Eredox

e-

Solution

Ecs

Evs

n-type semiconductor Solution

+

+ΔE

Continuous electron transfer

Eredox+

e-

ΔE


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reference to a standard hydrogen electrode (SHE). The standard reduction

potential (Eo) of a redox couple is the redox potential measured at standard

conditions, (1 M oxidant/reductant, 1 atm pressure at 25 oC) against a standard

hydrogen electrode. In general, the higher the standard redox potential of a redox

couple, the less energy is required for the reduction reaction (oxidant’s ability to

accept an electron) to proceed.

The standard reduction potential of a redox couple can be described in terms of a

single energy level, Eredox. This energy level is midway between the energy level

of the oxidant (Eox) and the reductant (Ered). The energy level of Eredox is higher in

redox couples with lower standard reduction potentials [Osseo-Asare, 1992]. As

a general rule, aqueous couples with standard redox potentials > 0.5 V are

normally valence band processes and those with Eo < 0.5 V are conductive band

processes [Gerischer, 1960]. The ferrous/ferric redox couple is a very stable and

irreversible couple in an acidic media, and has a standard redox potential of 0.77

V (SHE). This means, that the redox potential (related to an energy level) can be

easily controlled and maintained by varying the ferrous (higher ferrous, Eh < 0.77

V) or ferric (higher ferric, Eh > 0.77 V) concentration.

In other words, for a mineral sulphide to leach, the rest potential of that mineral

should fall below the redox couple of the aqueous species. The rest potential of a

sulphide mineral is the open circuit potential in an aqueous system, that a

sulphide mineral electrode (in reference with a standard hydrogen electrode) will

naturally approach if no external voltage is applied. The rest potential of

chalcopyrite at 20 oC is given as 0.52 V (volt vs SHE) [Venkatachalam, 1998].

1.3.2. Chalcopyrite leaching in ferric sulphate systems

Parker et al. [1981] conducted studies on the reduction rates of various aqueous

oxidants which are in contact with a corroding chalcopyrite electrode. The data

indicated enhanced reduction rates of species with standard reduction potentials


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overlapping with that of the valence and the conducting band energies of

chalcopyrite (Figure 5) [Crundwell, 1988].

Figure 5. The standard reduction potentials of different oxidants with respect to

the energy diagram of chalcopyrite [Crundwell, 1988].

The reduction rates were reported as follow (high to low):

Br2, CuCl2, I3- > FeCl3 > Fe2(SO4)3

The standard reduction potential of ferric iron was well within the forbidden

energy gap of chalcopyrite. This energy configuration caused a lower reduction

rate compared to the other oxidants, whose standard reduction potentials fell

either within the valence or the conducting band energies of chalcopyrite

[Crundwell, 1988]. Parker et al. [1981] also showed that the electron transfer to

the ferric/ferrous couple decreased as leaching progresses, slowing down the

rate in a parabolic manner. This electron transfer phenomena was more

Ec

Ev

600 mV

Fe 3d

Cu 3d S 3p

Cu2+/Cu+

I3-/I-

Br2/Br

Fe3+/Fe2Eredox

O2/H2O


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profound in ferric sulphate systems than in ferric chloride systems. With the

leaching of chalcopyrite in a ferric sulphate or ferric chloride system, the cupric

(Cu2+) concentration increased over time. Cupric, in the presence of chloride,

was a much better electron acceptor than cupric sulphate, ferric chloride or ferric

sulphate. The enhanced electron transfer effect seen in ferric chloride systems

was partly the combination of rate retardation (caused by ferric as electron

acceptor) and rate enhancement, facilitated by the highly reversible

cupric/cuprous (Cu2+/Cu+) couple in the presence of chloride [Parker et al., 1981].

The reversibility of a couple can be described in terms of the rate at which the

reductant (Cu+ or Fe2+) is re-oxidized after reduction. In bioleaching, the

microorganisms act as catalysts for the oxidation of ferrous, enhancing the

reversibility of the couple. In a chloride system, the cuprous (Cu+) is temporally

stabilized by the chloride ions, but is rapidly oxidized in the presence of oxygen.

A Cu2+/Cu+/Fe3+ chloride system was noted to be more effective than a Cu2+/Cu+

chloride system alone, because the ferric increased the activity of the cupric ions.

The combination of redox couples (i.e. Fe3+/ Fe2+ and Cu2+/Cu+) in solution or on

the surface of mineral constituted a combined measured potential, called a mixed

potential. This mixed potential can often lead to incorrect interpretation of

leaching results concerning the leaching of a mineral with only one specific redox

couple. In addition to the cupric/cuprous electron transfer effect, the

ferric/ferrous couple was also more reversible in the presence of chloride than in

sulphate. This increase in reversibility of the ferric/ferrous couple can also

promote electron transfer from the chalcopyrite [Parker et al., 1981]. With the

reported slow leaching kinetics of chalcopyrite in a ferric sulphate system, it

seems logical that the leaching rate can be enhanced with applying higher

concentrations of the oxidant (ferric iron) to the system. Investigations

concerning this matter confirmed that the leaching rate of chalcopyrite was

almost unaffected by ferric iron concentrations exceeding 0.01 M at 90 oC, but

was dependent upon ferric concentrations below this concentration

[Dutrizac et al., 1969]. Munoz et al. [1979] leached chalcopyrite in 0.06 M and


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0.5 M ferric iron at 90 oC and reported essentially identical leaching rates.

Dutrizac [1981] re-investigated the leaching rate of chalcopyrite in a ferric iron

concentration range of 0.01 M to 0.5 M (90 oC). The results indicated a marginal

increase in the leaching rate with respect to increasing ferric iron concentrations.

Although chalcopyrite leaches slower in a ferric sulphate system, the leaching

kinetics is still very dependant on the temperature of the leach system. The

copper recovery rate can be drastically improved by elevating the leach

temperature. The effect of temperature is apparent in bioleach processes, since

different microbes can oxidize ferrous, sulphur and sulphur intermediates at

different temperatures. Berry and Murr [1978] recognized this temperature

benefit by leaching chalcopyrite ore at 28 oC and 60 oC, with a mesophilic and

thermophilic organism respectively. The authors suggested that the higher

copper extraction achieved with the thermophilic organism was not due to the

superior ability of organism to leach chalcopyrite, but merely the combination of

the organism’s ability to generate oxidants and the exposure of the mineral to the

high temperature environment. The advantage of leaching chalcopyrite at

thermophile temperatures was described by several researchers [Brierley and

Brierley, 1973; Marsh and Norris, 1983; Clark and Norris, 1996 A; Konishi et al.,

2001; Petersen and Dixon, 2002; Rodriguez et al., 2003].

Rodriguez et al. [2003] leached chalcopyrite concentrate at 35 oC and 68 oC, with

a mesophile and thermophile culture respectively. The high temperature leach

recovered 56 % copper extraction within 40 days, while the low temperature

leach only recovered 9.5 % copper within the same period. The results are

illustrated in Figure 6.


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Figure 6. The efficiency of bioleaching chalcopyrite at mesophile (a) and

thermophile (b) temperatures [Rodriguez et al., 2003].

Even though the thermophile condition obtained a higher copper recovery, both

the leaching curves showed the typical leaching plateau (parabolic curve),

frequently observed when treating chalcopyrite in a ferric sulphate system. The

initial slow linear kinetics of chalcopyrite leaching in a ferric sulphate system is

simply that the ferric/ferrous iron couple is not as effective in electron extraction

Plateau

Plateau


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as a cupric/cuprous couple in a chloride system. The question as to why the

initial rate declines as leaching progresses is still unknown. It is mostly accepted

that the parabolic leaching curves are due to a lack of transport/diffusion of

reactants through a solid product layer on the chalcopyrite’s surface. It is this

solid product layer that could be responsible for the “passivation” of chalcopyrite

in a ferric sulphate system. The nature or composition of this diffusion barrier is

controversial, with little agreement among researchers [Rodriguez et al., 2003].

1.3.3. The passivation of chalcopyrite

Elemental sulphur, iron deficient copper sulphides, polysulphides, bornite,

covallite and iron precipitates i.e. jarosite have all been suggested to be the main

culprits in the passivation of chalcopyrite [Dutrizac,1989; Parker et al., 1981;

Warren et al., 1982; Buckley and Woods, 1984; Majima et al., 1985; Hackl et al.,

1995; Sandstrom et al., 2005; Nava and Gonzales, 2006]. The next section is a

brief summary of the different product layers described in literature.

1.3.3.1. Iron deficient copper sulphides, polysulphides and other product layers

Electrochemical studies focused on corroding currents of chalcopyrite electrodes

in copper chloride, ferric chloride and ferric sulphate systems, caused Parker et

al. [1981] to acknowledge the formation of elemental sulphur on corroding

chalcopyrite, but the concept of it being the rate limiting product layer was

rejected. The authors postulated that the rate limiting surface film was a

thermally unstable metal-deficient polysulphide, with semiconductor properties,

different to that of chalcopyrite. Buckley and Woods [1984] conducted X-ray

photoelectron spectroscopy on the oxidation products of chalcopyrite under

various conditions. Their electrochemical studies on the leaching of chalcopyrite

at temperatures above 67 oC showed similar results to that obtained by Parker et

al. [1981]. It was furthermore concluded that the reaction rate was limited by a

semiconducting film rather than a sulphur product layer.


21

It was concluded that this semiconductor film consisted of the same sulphur

crystal lattice as the original chalcopyrite, but with copper and iron (more iron)

vacancies within the original crystal structure. The semiconductor film was also

described as a metal deficient copper sulphide with a chemical composition near

Cu0.8S2.

Anodic polarization studies on chalcopyrite in acidic media at 25 oC revealed the

existence of a passive region (no current increase or decrease with change in

potential) in the potential range of 0.6-0.9 V (SHE) [Warren et al., 1982]. The

potential range of this passive region is well within the potential range of the

ferrous/ferric couple (standard redox potential of 0.77 V) and could well describe

the slow leaching kinetics. It was proposed that the passive region was caused

by the formation of two distinct phases, bornite (Cu5FeS2) and covallite (CuS),

which in turn passivates the mineral. However, it is widely accepted that bornite

and covallite leaches much faster than chalcopyrite and can thus not passivate

chalcopyrite [Hackl et al., 1995].

Hackl et al. [1995] proposed an alternative chemical composition of this

passivation film described above. In their experimental approach, chalcopyrite

was oxygen-pressure-leached at temperatures between 110 oC and 220 oC. The

leached mineral surfaces were studied with Auger electron spectroscopy and X-

ray photoelectron spectroscopy. The results suggested that the chalcopyrite was

passivated by a thin copper rich polysulphide, CuSn (n>2). The rate determining

step was shown to be the slow decomposition of the copper rich polysulphide to

cupric ions and elemental sulphur. The elemental sulphur was described as

porous and not rate limiting. The copper rich polysulphide decomposes quicker

when the leach temperature is increased to 200 oC, whereas the CuSn can no

longer passivate the mineral.

Gomez et al. [1996] tested the electrochemical response of massive chalcopyrite

electrodes in acidic microbial growth media at 25 oC and 68 oC.


22

The experimental conditions were selected to mimic chalcopyrite’s behavior at

mesophilic (25 oC) and thermophilic (68 oC) growth temperatures, when

subjected to an external applied potential range. A “prewave” or passive region

was identified in the anodic potential range between the rest potential and 0.9 V

(SHE) at 25 oC and 68 oC. These results were similar to the findings of Warren et

al. [1982]. The authors concluded that iron is dissolved in preference to copper

and that the resulting Cu-rich and Fe-poor phase is responsible for the passive

region within the ferrous/ferric couple. The Cu-rich and Fe-poor phase can be

associated with sulphides, polysulphides and elemental sulphur.

Anodic current density profiles of chalcopyrite over the passive potential range

(various voltametric sweep rates), indicated higher overall current densities at

68 oC than at 25 oC. This illustrates that within the ferric/ferrous couple, this

passive phenomenon was less prominent at 68 oC than at 25 oC. The different

electrochemical responses at the two temperatures were thought to be the

difference in physical structure of this complex Cu-rich and Fe-poor phase, which

in turn influenced diffusion control and passivation.

1.3.3.2. Elemental sulphur as product layer

Dutrizac et al. [1969] reported that elemental sulphur was the major leach

product formed during the leaching of chalcopyrite in a ferric sulphate system.

The dominant leaching reaction was reported as follows:

CuFeS2 + 4Fe3+ Cu2+ + 5Fe2+ + 2S0

Dutrizac [1989] confirmed his first observation and added that 94 % elemental

sulphur and only 6 % sulphate was formed during the leaching of chalcopyrite in

a ferric sulphate system at 95 oC. This sulphur/sulphate ratio did not change,

regardless of the leaching time (0-70 hours), the ferric iron concentration (0-2 M,

or the particle size.


23

The amount of sulphate formed was described by the following reaction:

CuFeS2 + 8Fe2(SO4)3 + 8H2O CuSO4 + 17FeSO4 + 8H2SO4

In his paper, Dutrizac [1989], described a leaching mechanism that may involve

the direct acid attack on the chalcopyrite with the formation of dissolved

hydrogen sulphide species. The hydrogen sulphide is subsequently oxidized to

elemental sulphur by ferric iron in the following reactions:

CuFeS2 + 4H+ Cu2+ + 5Fe2+ + 2H2Saq

2H2Saq + 4Fe3+ 4Fe2++ 4H+ + 2S0

The net reaction is similar to that of the direct ferric attack on chalcopyrite:

CuFeS2 + 4Fe3+ Cu2+ + 5Fe2+ + 2S0

Schippers and Sand [1999] measured the sulphur based compounds formed

when leaching various sulphide minerals in 10 mM ferric chloride solution at

28 oC. In the case of chalcopyrite they detected approximately 92 % elemental

sulphur and only around 7 % sulphate formation in 24 hours of incubation.

These results corroborated with the data obtained in Dutrizac’s [1989] work and

confirmed that elemental sulphur was the main leach product formed when

chalcopyrite was leached within a ferric sulphate based system. Majima et al.

[1985] illustrated a more dense sulphur layer on chalcopyrite when leached in a

ferric sulphate system compared to a ferric chloride system. The authors

suggested that the dense sulphur product layer significantly influenced the

leaching kinetics by preventing oxidant and product diffusion to and from the

chalcopyrite’s surface. This was not the case in the corresponding ferric chloride

system, where the sulphur layer was shown to be more porous.


24

In leaching chalcopyrite concentrate in a ferric sulphate system at 50 oC for

120 min, Klauber et al. [2001] illustrated that neither metal deficient sulphides nor

polysulphides were formed on the surface during initial leaching. The research

team investigated leached chalcopyrite surfaces using a powerful technique, X-

ray photoelectron spectroscopy (XPS). Elemental sulphur was identified as the

primary surface product layer produced with a ferric sulphate leach. A second

major leach product was identified as a sulphide of sort (S2-2). The cation

associated with the sulphide was not identified, but the absence of a copper 2p

spectrum eliminated possible CuS2 type products.

1.3.3.2.1. Removal of the elemental sulphur layer

Havelik and Kammel [1995] carried out leaching experiments on chalcopyrite

concentrate in acidified ferric chloride solutions at 40 oC and 80 oC. Carbon

tetrachloride (CCl4) was used to dissolve the elemental sulphur formed during

leaching. The results showed improved copper extraction with the addition of

carbon tetrachloride. Leaching at 40 oC showed 9.42 % and 4.42 % copper

extracted within 4 hours, with and without the addition of carbon tetrachloride

respectively. Leaching at 80 oC for 4 hours achieved copper recoveries of

23.15 % and 16.15 %, with and without the addition of carbon tetrachloride

respectively. In removing the sulphur product layer during the lower temperature

leach, it proved to be more beneficial in copper recovery (more than twice),

compared to what was achieved at the higher temperature. The authors

concluded that in a temperature range of below 45 oC, the leaching of

chalcopyrite in a ferric mediated system seems to be depended on diffusion

control. Diffusion restriction concerning elemental sulphur in a ferric system is

not that applicable in leaching chalcopyrite above 45 oC, where the process is

much more dependant on a chemical controlled reaction.


25

In abiotic ferric sulphate systems the passivation of chalcopyrite by elemental

sulphur is a possibility, but in most bioleaching systems, sulphur is almost

completely oxidized to sulphate by sulphur oxidizing microorganisms. Several

researchers believe that the passivation of chalcopyrite in a biotic acidic ferric

sulphate environment is due to jarosite precipitation [Sandstrom et al., 2005;

Kinnunen et al., 2006; Nava and Gonzales, 2006].

1.3.3.3. Jarosite precipitation as diffusion barrier Jarosite is a ferric sulphate based crystalline (detectable by X-ray diffraction)

precipitate commonly associated with ferric sulphate based systems, especially

found in bioleach systems. Previous research indicated that jarosite precipitated

on the mineral surface, creating a diffusion barrier, thus restricting microbial and

reactant interaction with the mineral [Howard and Crundwell., 1999; Stott et al.,

2000; Parker et al., 2003; Sandstrom et al., 2005; Kinnunen et al., 2006; Nava

and Gonzales, 2006]. Jarosite is characterized according to the alkali cation

associated with ferric sulphate hydroxyl complex. Jarosite was rapidly formed in

a high ferric and sulphate environment according the following microbial

mediated reactions during chalcopyrite leaching [Stott et al., 2000]:

4Fe2+ + O2 + 4H+ 4Fe3+ + 2H2O (microbial mediated ferrous oxidation)

CuFeS2 + 4Fe3 5Fe2+ + Cu2+ + 2S0

2S0 + 2H2O + 3O2 2SO42- + 4H+ (microbial mediated sulphur oxidation)

X+ + 3Fe3+ + 2SO42- + 6H2O XFe3(SO4)2(OH)6 (jarosite) + 6H+

(X+ represents cations such as K+, Ag+, Na+, NH4+ and H3O+)

Potassium (KFe3(SO4)2(OH)6) and ammonium jarosite (NH4Fe3(SO4)2(OH)6) are

the most frequently observed jarosite species associated with bioleaching.


26

The potassium and ammonium cations associated with these jarosite complexes

are mainly introduced to the system via the nutrient media, necessary for

microbial growth. Ammonium jarosite is a less stable complex than potassium

jarosite, which is mainly detectable after potassium jarosite precipitation or in

potassium free systems [Stott, 2002].

Gomez et al. [1999] tested the influence of using different nutrient media on the

leaching of a mixed sulphide concentrate (chalcopyrite as copper bearing

mineral) at 30 oC. A mixed population of iron and sulphur oxidizing bacteria were

used in order to minimize sulphur accumulation. The chemical composition of

the different nutrient media is summarized in table 1.

Table 1. The chemical composition of various nutrient media commonly used in

bioleaching reactions.

Nutrient salt 9K (g/L) Norris (g/L) D1 (g/L) D2 (g/L)

(NH4)2SO4 3.0 0.2 0.06 0.01

MgSO4.7H2O 0.5 0.2 0.06 0.01

K2HPO4 0.5 0.2 0.02 0.01

KCl 0.1 - 0.02 0.01

Ca(NO3)2.H2O 0.01 - - -

In the paper by Gomez et al. [1999], the nutrient media from the first column was

referred to as 9K media. For clarity, 9K media consists of additional ferrous

sulphate above that of the nutrient salt composition described in table 1. In order

to minimize excess jarosite precipitation, no ferrous sulphate was added during

this experimental work. The ferrous iron in the system was introduced through

mineral dissolution only. This particular media without ferrous sulphate is actually

termed 0K media. Copper leaching rates (V) from chalcopyrite in the different

media were reported as follows:


27

V9K > VNorris > VD1 > VD2 > Vtap water

Approximate total copper recovery at the end of 20 days:

9K (35 %) > Norris and D1 (25 %)> D2 (12 %) > tap water (9 %)

It is clear from the results that an increase in total nutrient salts within the

respected media promoted the leaching of chalcopyrite at 30 oC over 20 days.

This was due to enhanced microbial growth kinetics and ferrous oxidation (ferric

generation), associated with the media containing higher concentrations of

nutrients, especially 9K media. Pronounced in the case of 9K media, 25 % of the

copper was rapidly leached during the first 8 days with a drastic declined rate for

the remainder of the time. The other media showed, to a lesser extent, this

enhanced parabolic leaching curve, associated with the passivation of

chalcopyrite. The higher concentrations of ferric, sulphate, ammonium and

potassium in the 9K media caused jarosite precipitation (X ray analysis on leach

residue), which was not the case for the other media. The authors concluded

that jarosite precipitation could have been the cause for the rapid declining rate

observed when leaching chalcopyrite in 9K media. The effect of jarosite

precipitation was enhanced with increasing temperature and pH, especially in

bioleaching systems with temperatures exceeding 65 oC and pH values between

1.7 and 2.7 [Margulis et al., 1976].

Duncan and Walden [1972] showed that the removal of soluble ferric iron and

nutrients during jarosite precipitation also influenced the leaching rate. Not only

did jarosite precipitation complex the oxidant, but also essential nutrients, which

in turn influenced microbial growth and the regeneration of ferric iron.


28

1.3.3.3.1. Jarosite precipitation and chalcopyrite surface chemistry

Parker et al. [2003] described an alternative mechanism for the oxidative acid

leaching of chalcopyrite in a ferric sulphate system. Emphasis was placed on the

formation of a ferric sulphate phase on the surface of leached chalcopyrite, which

acted as precursor for jarosite precipitation, and in turn passivation. This work

suggested that jarosite precipitation on the surface of chalcopyrite was not only a

function of the bulk solution chemistry, but also the mineral’s surface chemistry.

Klauber et al. [2001] (section 1.3.3.2) identified a type of sulphide (S2-2) as a

major leach product, but the cation associated with the sulphide was not

identified. Parker et al. [2003] described this sulphide as a “pyritic-like

disulphide”, with the cation being iron. This pyritic-like disulphide was oxidized

by ferric ions to a thiosulphate intermediate, very similar to pyrite oxidation via

the thiosulphate mechanism. Subsequent reactions with thiosulphate and ferric

iron (close to the mineral surface) resulted in the formation of the ferric sulphate

phase. This ferric sulphate phase acted as an initial product layer, which caused

further mass jarosite deposition on the chalcopyrite’s surface.

The link between the passivation of chalcopyrite and jarosite precipitation is

difficult to solve, since jarosite, unlike elemental sulphur, is not removed from the

mineral’s surface with typical oxidative microbial bioleaching reactions. No

conclusive results are available on the complete removal of jarosite, and in doing

so, restoring the initial rapid leaching rate, after passivation has occurred.

Jarosite bioreduction has been attempted [Stott et al., 2000], and is discussed

under the subject of anaerobic bioleaching.

1.3.4. Low potential leaching of chalcopyrite

An obvious solution to jarosite precipitation and chalcopyrite passivation is to

leach the mineral at lower ferric iron concentrations, since the leaching rate of

chalcopyrite is not that sensitive to ferric iron concentrations (section 1.3.2).


29

Sandstrom et al. [2005] performed bioleaching and chemical leaching on

chalcopyrite concentrate in sulphuric acid at a constant low (420 mV) and high

(600 mV) potential. The bioleaching was performed in continuous bioreactors at

65 oC, with the thermophilic microorganism Sulfolobus metallicus. Potentials in

this text referred to the platinum vs. Ag/AgCl electrode instead of the standard

hydrogen electrode (Eh = EAg/AgCl + 207 mV at 25 °C). The low potential during

the bioleach was maintained through the constant addition of sodium sulfite

(SO32-) to the system.

2Fe3+ + SO42- + H2O 2Fe2+ + SO4

2- + 2H+

During the chemical leach the high potential was obtained by the constant

addition of potassium permanganate (MnO4-).

Fe2+ + 1/5MnO4- +8/5H+ Fe3+ + 1/5Mn2+ + 4/5H2O

During the high potential bioleach and chemical leach, large amounts of jarosite

was detected (X-ray diffraction) on the chalcopyrite’s surface when passivation

started to occur. Both the bioleach and chemical leach at a low potential

produced very little jarosite precipitation. Elemental sulphur was the most

prominent leach product formed during the low potential chemical leach. The

bioleaching showed a high degree of sulphur oxidation to sulphate, at both high

and low potentials. The copper dissolution rate from chalcopyrite was much

higher during the low potential bioleach and chemical leach, compared to both

the high potential conditions. The low potential chemical leach showed the

highest copper recovery and resulted in large amounts of sulphur deposition on

the mineral’s surface. The paper concluded that jarosite was the main cause for

chalcopyrite passivation and not elemental sulphur. During the high potential

leach, iron was dissolved preferentially to copper, while during the low potential

conditions, copper leached preferentially to iron [Sandstrom et al., 2005].


30

1.3.4.1. The oxidative acid leaching of chalcopyrite

Ferric iron is generally accepted to be the main oxidant of chalcopyrite in ferric

sulphate based systems, while ferrous iron mainly serves as a source to

generate ferric iron through ferrous oxidation, chemically or microbially. In

contrast, Hiroyoshi et al. [1997] illustrated enhanced copper dissolution from

chalcopyrite in ferrous sulphate, compared to ferric sulphate. Under identical

concentrations of ferrous and ferric iron, the acidified ferrous solution showed

remarkably higher copper recoveries than that of the acidified ferric solution. The

copper dissolution rate increased with increasing concentrations of initial ferrous

sulphate. The low potential leach mechanism was dependent on the acid

concentration (pH) and dissolved oxygen (DO). Lower initial pH values at similar

ferrous concentrations increased the leaching rate. Oxygen and acid were

consumed during the low potential leach, even though most of the iron existed as

ferrous iron. Under nitrogen in the absence of ferric iron, with the addition of

ferrous sulphate, the leaching rate was negligible. The following reactions were

considered to be responsible for the acid, oxygen and ferric consumption.

4Fe2+ + O2 + 4H+ 4Fe3+ + 2H2O (slow during oxidative chemical leaching)

CuFeS2 + 4Fe3+ Cu2+ + 5Fe2+ + S0

If these two reactions proceed in a near equilibrium state, nil or very little ferric

iron will be detected within the system. The ferric iron produced from the first

reaction will be consumed during the second leach reaction.

With additional oxygen consumption work, Hiroyoshi et al. [1997] concluded that

at a low potential chalcopyrite was leached according to the following reaction:

CuFeS2 + O2 + 4H+ Cu2+ + Fe2+ + 2S0 + 2H2O


31

The authors indicated that the oxidative acid leach reaction (third reaction)

proceeded independently from the first two reactions and was not a net reaction.

Thus, in an acidified ferrous sulphate based system with little ferric iron (low

potential), chalcopyrite was mainly leached with dissolved oxygen and protons.

Under the experimental conditions described in this test work, low potential

induced oxidative acid leaching was reported to be more effective in leaching

chalcopyrite than ferric iron in a high potential environment [Hiroyoshi et al.,

1997].

1.3.4.2. Chalcopyrite leaching by ferrous iron in acidic ferric iron sulphate solutions

Hiroyoshi et al. [2001] continued with further investigations concerning the role of

ferrous iron during chalcopyrite leaching. To clarify the role of ferrous iron during

chalcopyrite oxidation with ferric iron, all experimental work was conducted under

nitrogen. Anaerobic conditions prohibit the oxidative acid leaching of

chalcopyrite and provide a suitable environment to study ferrous iron promoted

ferric iron oxidation on the mineral. Unexpected results indicated that the

oxidation of chalcopyrite with ferric iron is enhanced by high concentrations of

ferrous and cupric ions. However, when the cupric ion concentration was low,

high ferrous iron suppressed copper dissolution. In order to interpret the results,

a two step reaction model was proposed. Anaerobically, in the presence of

sufficient concentrations of cupric and ferrous, chalcopyrite was reduced by

ferrous iron to chalcosite (Cu2S) according to:

CuFeS2 + 3Cu2+ + 3Fe2+ 2Cu2S + 4Fe3+

The chalcosite was then oxidized by ferric iron.

2Cu2S + 8Fe3+ 4Cu2+ + S0 + 8Fe2+


32

In the case of low cupric and ferrous iron concentrations, chalcopyrite was

directly oxidized by ferric iron and did not form the chalcosite intermediate. The

copper recovery rate was slower during the direct ferric iron leach, compared to

the mechanism where chalcopyrite was reduced to chalcosite first and then

oxidized by ferric iron. The critical solution potential “window” (ferrous/ferric

ratio), at which the described mechanism took place was set between 520 and

610 mV vs. SHE. Hiroyoshi et al. [2001] stated that the slow copper dissolution

rate observed at solution potentials above 610 mV (direct ferric attack) was not

typical to mineral passivation due to diffusion barriers, but rather an

electrochemical phenomenon.

1.3.4.3. Controlling the redox potential

Third et al. [2001] obtained similar results, which indicated that chalcopyrite

leaching is approximately three times faster in 0.1 M ferrous ions, compared to

0.1 M ferric ions. Without the addition of ferrous or ferric (acidic water), the

leaching rate was slightly lower than that of the ferric iron (high potential) leach.

Therefore, faster leaching rates can be obtained in restricting the microbial

mediated and chemical oxidation of ferrous to ferric during chalcopyrite leaching.

In the same paper the researchers used a computer controlled bioreactor to

control the redox potential at a specific set point. A constant low redox potential

of 380 mV (Ag/AgCl or 500 mV vs. SHE) was maintained by restricting air supply

to the reactor when the solution potential exceeded the 380 mV set point. The

redox controlled mechanism is illustrated in Figure 7.


33

Figure 7. The oscillation of the redox potential over and above the 380 mV set

point. At high oxygen concentrations ferrous oxidation out-competes ferric

consumption by the mineral and causes an increase in potential. The opposite is

achieved at low oxygen concentrations.

A high potential bioleach (sulphur and iron oxidizing mesophiles) with continuous

aeration served as control condition. The low redox controlled system (no

inoculation) achieved twice the final copper recovery than that of the high

potential bioleach. The addition of sulphur and iron oxidizing bacteria to the

redox controlled reactor caused an improvement in copper recovery compared to

the abiotic system. It was concluded that the microbial mediated benefit could be

due to the oxidation of elemental sulphur to sulphuric acid, providing additional

protons and minimizing passivation. Sandstrom et al. [2005] illustrated that this

is not the case and that elemental sulphur is too porous to form a diffusion

barrier.

With all the results presented on the concept of chalcopyrite passivation in a

ferric sulphate system (discussed in sections 1.3.3 to 1.3.4.3), some reasonable

assumptions can be made.

Time

Red

ox p

oten

tial (

mV

vs.

Ag/

AgC

l)

378

380

382

384

386

388 4Fe2+ + O2 + 4H+ 4Fe3+ + 2H2O (1)CuFeS2 + 4Fe3+ Cu2+ + 5Fe2+ + S0 (2)

dFe3+/dt (1) > dFe3+/dt (2)

dFe3+/dt (1) < dFe3+/dt (2)

High [O2] Low [O2]


34

• In an abiotic high potential sulphate system chalcopyrite can be passivated

by either jarosite or elemental sulphur.

• In a biotic high potential sulphate system, with the addition of sulphur

oxidizing microbes, it is likely that chalcopyrite is passivated by jarosite

and not elemental sulphur.

• Low potential leaching seems to be a solution to jarosite precipitation and

passivation. However, it is uncertain whether the enhanced leaching rate is

due to the absence of jarosite or because of a different electrochemical

leach mechanism.

• It is uncertain whether elemental sulphur passivates chalcopyrite during

low potential leaching.

A way to resolve the issues surrounding the passivation of chalcopyrite is to

remove the diffusion barrier after passivation has occurred without the use of

methods that could influence the surface properties of the mineral. In removing

the diffusion barrier the passivation effect should be eliminated and an increased

leach rate re-established under similar redox potential conditions.

Several authors described the solubilisation of elemental sulphur and jarosite

within various ferric reductive type metabolisms associated with acidophilic

organisms. These types of metabolisms occur under anaerobic conditions and

could be a powerful tool in the removal of sulphur and jarosite from leached

chalcopyrite and hence increase the understanding of the leaching mechanisms.

The next section provides an overview on anaerobic leaching with different ferric

reductive metabolisms.

1.4. ANAEROBIC LEACHING The first acidophilic organism in which a ferric reductive metabolism was

identified was the chemolitho-autotrophic and acidophilic mesophile

Acidithiobacillus ferrooxidans. It was believed that the organism derives its


35

energy from oxidative respiration only, involving the oxidation of ferrous, sulphur

and other sulphur intermediate species in the presence of oxygen and carbon

dioxide as carbon source.

1.4.1. Oxidative respiration based on sulphur oxidation

The proposed mechanism for oxidative respiration based on sulphur oxidation

and oxygen as electron acceptor by Acidithiobacillus ferroxidans can be

described by the following reactions:

S0 + O2 + H2O H2SO3

H2SO3 + H2O SO42- + 2e’ + 4H+

Net Reaction Equation: S0 + O2 + 2 H2O SO42- + 4H+

In the oxidative respiration of Acidithiobacillus ferroxidans, the sulphur-oxidizing

enzyme requires reduced glutathione (GSH) to open up the sulphur octet ring to

produce sulfite for further oxidation to sulphate (Silver and Lundgren, 1968;

Vestal and Lundgren, 1971).

1.4.2. Anaerobic respiration based on ferric iron reduction

In 1976, Brock and Gustafson reported that Acidithiobacillus ferrooxidans can

reduce ferric to ferrous in the presence of elemental sulphur in an oxygen limiting

environment. At first the ferric reduction reaction was not recognized as a

respiratory system since the bacteria did not grow under these ferric reducing

conditions [Sugio et al., 1987]. This observed lack of growth could have been due

to carbon dioxide limitations during anaerobic conditions, typically found when

operating a system under nitrogen only. Carbon dioxide limitation effects were

not mentioned in the research paper.

GSH


36

Later research indicated that Acidithiobacillus ferrooxidans grows chemolitho-

autotrophically via the oxidation of sulphur by ferric iron with sufficient carbon

(CO2) under anaerobic conditions [Ohmura et al., 2002]. The proposed

mechanism for anaerobic respiration based on sulphur oxidation and ferric

reduction by Acidithiobacillus ferrooxidans can be described according to the

following reactions:

S0 + 2(reduced glutathione) H2S + 2(oxidized glutathione)

H2S + 3H2O + 4Fe3+ SO32- + 4Fe2+ + 8H+ (sulphur: ferric oxidoreductase)

SO32- + H2O + 2 Fe3+ SO4

2- + 2Fe2+ + 2H+ (Ferric dependent sulfite oxidase)

Net Reaction Equation: S0 + 6 Fe3+ + 4 H2O SO42- + 6 Fe2+ + 8 H+

Elemental sulphur reacts with reduced glutathione (GSH) to produce hydrogen

sulphide. The hydrogen sulphide is oxidized to sulfite (catalyzed by sulphur: ferric

oxidoreductase), which in turn is oxidized to sulphate by the enzyme ferric

dependent sulfite oxidase (Sugio et al., 1987; Sugio et al., 1989).

Anaerobic respiration based metabolism of acidophilic organisms was also found

in archaea and other thermophilic temperature bacteria. Thermophilic archaea

(optimal growth temperatures above 60 oC) do not seem to have the capability of

direct anaerobic ferric reduction, as found in the mesophilic bacterium,

Acidithiobacillus ferrooxidans.

Acidianus species from the group Sulfolobales are true chemolithotrophs and

facultative anaerobes growing either anaerobically by sulphur reduction to form

hydrogen sulphide (hydrogen gas as electron donor) or aerobically through the

oxidation of elemental sulphur to sulphate with oxygen as electron acceptor.

Some of these high temperature facultative anaerobic strains from the species

Acidianus are also capable of anaerobic sulphur oxidation (forming sulphuric

acid) in the presence of molybdate as electron acceptor. These strains were


37

isolated from the Solfatara fields where molybdate is the only minor metal

present (Segerer et al., 1985; Segerer et al., 1986).

1.4.2.1. Ferric iron reduction by moderate thermophiles

Moderate thermophilic acidophiles were classified according to their optimal

growth temperatures (45 oC to 55 oC), which were typically above and below the

growth optima of mesophiles and thermophiles, respectively [Norris, 2006].

Various different metabolisms of moderate thermophiles were described

previously [Golovacheva and Karavaiko, 1979; Wood and Kelly, 1985; Hallberg

and Lindström, 1994; Norris et al., 1996; Clark and Norris, 1996 B]. Some of the

different types of metabolism include:

• Chemolitho-autotrophic growth (media containing ferrous and reduced

sulphur species with CO2 as sole carbon source).

• Heterotrophic growth (media containing only yeast extract).

• Mixotrophic growth (media containing ferrous iron and yeast extract, in

which the ferrous acts as inorganic energy source and the yeast extract as

sole carbon source).

The metabolisms mentioned above are all based on the oxidation of ferrous iron

or sulphur based species by using oxygen as electron acceptor. As with

Acidithiobacillus ferrooxidans and some thermophilic archaea, moderate

thermophiles also have the ability to proliferate under anaerobic respiration.

Anaerobic respiration based on the reduction of soluble ferric and reductive

dissolution of ferric containing minerals by several moderate thermophiles was

described by Bridge and Johnson [1998]. The moderate thermophilic ferrous

oxidizing bacteria (Sulfobacillus thermosulfidooxidans, Sulfobacillus acidophilus,

and Acidimicrobium ferrooxidans) were capable of anaerobic growth, reducing

ferric iron to ferrous iron. The amount of growth was directly proportional to the

amount of ferric iron reduced. The iron reduction rate was optimal when the


38

isolates were grown as heterotrophs, with an organic carbon as electron donor.

Under anaerobic conditions, some strains were also able to oxidize tetrathionate

(inorganic electron donor) by reducing the ferric iron. A specific strain of

Sulfobacillus acidophilus was even capable of anaerobic reductive dissolution of

three ferric iron containing minerals, ferric hydroxide, jarosite and goethite. In this

reaction glycerol served as both electron donor and carbon source.

1.4.2.2. Jarosite bioreduction with moderate thermophiles

Stott et al. [2000] investigated the bioreduction of jarosite from the surface of

leached chalcopyrite concentrate after passivation occurred, and whether initial

rates could be restored after the process. The chalcopyrite was initially leached

under oxidative conditions with the iron and sulphur oxidizing moderate

thermophile bacterium, Sulfobacillus thermosulfidooxidans. The leaching was

continued until the copper dissolution rate drastically declined, which coincided

with jarosite precipitation on the mineral surface (confirmed with X-ray

diffraction). The passivated chalcopyrite was subjected to anaerobic media

(glycerol both as energy and carbon source) containing combinations of several

moderate thermophiles, Sulfobacillus thermosulfidooxidans, Acidimicrobium

ferrooxidans and Sulfobacillus acidophilus. During the bioreduction stage with all

three organisms present, the jarosite was significantly reduced to ferrous iron.

Together with the jarosite and ferric iron reduction, sulphur was also reduced to

hydrogen sulphide, which precipitated the soluble copper as copper sulphide.

In the presence of only Acidimicrobium ferrooxidans and Sulfobacillus

acidophilus the ferrous generation and sulphur reduction were much slower than

with the combination of all three. It was concluded that Sulfobacillus

thermosulfidooxidans, and to a lesser extent Acidimicrobium ferrooxidans and

Sulfobacillus acidophilus, catalyze jarosite reduction to ferrous iron and

elemental sulphur to hydrogen sulphide according to the following reactions:


39

Fe3(SO4)2(OH)6 + 6H+ 3Fe3+ + 2SO42- + 6H+

4Fe3+ + organic + 2H2O 4Fe2+ +4H+ + CO2

Organic + S0 H2S + CO2

Cu2+ + HS- CuS(s) + H+

Jarosite, ferric iron and elemental sulphur were used as terminal electron

acceptors instead of oxygen, in the anoxic environment. Even though a large

quantity (approximate 70 %) of the jarosite was reduced in 700 hours, the copper

extraction rate did not increase significantly above the untreated controls. The

remaining jarosite could not be reduced, not even after 1700 hours of incubation.

Stott et al. [2000] concluded that the remaining jarosite constitutes a thin, tightly

bound surface layer, which could not be detached by the bacteria and continues

to passivate the mineral.

Anaerobic ferric iron reduction metabolisms were also identified within isolates

belonging to the genus of Ferroplasma. The specific Ferroplasma isolates were

capable of facultative anaerobic growth, using ferric iron as electron acceptor, in

the presence of yeast extract and other inorganic electron donors [Dopson et al.,

2004; Hawkes et al., 2004]. The next section provides a summary on the recently

discovered genus of Ferroplasma within the kingdom of archaea.

1.4.3. The genus of Ferroplasma

Golyshina et al. [2000] isolated the first strain of Ferroplasma (strain YT) from an

arsenopyrite bioleach pilot plant in Russia, operated at 30 oC. With the

organism’s distinct phenotypic characteristics and 16S rRNA sequence it could

not be assigned to an existing genus and was classified as a new species in a


40

new genus, within a new family, under the description of Ferroplasma acidiphilum

fam. nov., gen. nov., sp.nov.

The phylogenetic lineage was described as follows:

o Thermoplasmata

o Thermoplasmatales

• Ferroplasmaceae (new family) Ferroplasma (new genus)

Ferroplasma acidiphilum

• Picrophilaceae

Picrophilus

Picrophilus osimae

Picrophilus torridus

• Thermoplasmataceae

Thermoplasma

Thermoplasma acidophilum

Thermoplasma volcanium

The genus Picrophilus and Thermoplasma is more comprehensively described in

Schleper et al. [1995] (Picrophilus), Darland et al. [1970], Segerer et al. [1988]

and Yasuda et al. [1995] (Thermoplasma). The differences in characteristics of

archaea within the order Thermoplasmatales are summarized in Table 2. The

characteristics of the species within the genus of Picrophilus (Picrophilus osimae

and Picrophilus torridus) are almost identical, similar with Thermoplasma

(Thermoplasma acidophilum and Thermoplasma volcanium).


41

Table 2. Comparative characteristics of species within the order

Thermoplasmales.

Characteristic Picrophilus spp. Thermoplasma

spp.

Ferroplasma

acidiphilum

Morphology Irregular cocci Pleomorphic Pleomorphic

Flagella + + -

Autotrophy - - +

Heterotrophy + + -

Fe2+ oxidation - - +

So oxidation - - -

Aerobic growth + + +

Anaerobic growth - + -

Temperature

optimal (oC) 60 60 35

Temperature

range (oC) 45-65 33-67 15-47

Optimal pH 0.7 1.2 1.7

pH range 0.1-3.5 1-4 1.3-2.2

S-layer + - -

DNA G+C content 36 46 36.5

1.4.3.1. The initial genus description of Ferroplasma

The genus was described according to the single species, Ferroplasma

acidiphilum [Golyshina et al., 2000]. The cells are pleomorphic cocci, spherical

to filametous and forms duplex and triplex forms. The organism is acidophilic,

strictly aerobic and strictly chemolitho-autotrophic. The organism fixes carbon

dioxide (CO2) as sole carbon source and did not grow on any organic carbon

source alone. The organism was shown to oxidize ferrous iron as primary

energy source with yeast extract (0.02 %) being essential for growth. It was


42

claimed that the yeast extract served as a nutrient source and not a carbon

source. The yeast extract became inhibitory for growth above a concentration of

0.2 % (w/v). The principal lipids were identified as archaetidic acid and

archaetidyl glycerol.

Shortly after the genus description was published, Edwards et al. [2000] isolated

a second organism within the genus Ferroplasma, Ferroplasma acidarmanus

(strain Fer1T). The isolate is phylogenetically identical to Ferroplasma

acidiphilum, but is physiologically different. Ferroplasma acidarmanus was able

to grow organotrophically on yeast extract as sole energy source, whereas

Ferroplasma acidiphilum could not [Edwards et al., 2000; Golyshina et al., 2000].

The isolate was also capable of growth within extremely low pH environments

(pH 0 - 2.5, optimal pH 1.2), whereas Ferroplasma acidiphilum could only grow

between pH 1.3 and 2.2 (optimal pH 1.7). The optimal growth temperature for

Ferroplasma acidarmanus was 42 °C, whereas Ferroplasma acidiphilum grew

between 15-45 °C with an optimum growth temperature of 35 °C [Golyshina et

al., 2000].

Dopson et al. [2004] conducted a comparative analysis on three Ferroplasma

isolates (Fer1T, MT17 and DR1) phylogenetically similar to Ferroplasma

acidiphilum strain YT. The original Ferroplasma acidiphilum strain YT was also

included in the test work. The results showed that all four isolates were able to

grow chemoorganotrophically on yeast extract or various sugars and

chemomixotrophically on ferrous iron oxidation and yeast extract or sugars.

All four isolates were described as facultative anaerobic, reducing ferric iron to

ferrous iron in the presence of yeast extract. The authors suggested that

Ferroplasma acidiphilum strain YT cannot be described as chemoautotrophic, as

described by Golyshina et al [2000], but rather chemomixotrophic, utilizing

carbon dioxide and organic carbon.


43

1.4.3.2. An amendment to the description of the Ferroplasma genus. The genus includes Ferroplasma acidiphilum strain YT and three additional

isolates, namely DR1, MT17 and Fer1T. These organisms grow

chemoorganotrophically and chemomixotrophically. Chemoautotrophic growth is

indecisive. The organisms are facultative anaerobic, oxidizing organic carbon

(yeast extract) with ferric iron as electron acceptor. The G+C content is between

36.6 % and 37 %. The isolates are mesophilic (optimal growth < 45 oC) and

acidophilic (optimal pH of 1-1.7) [Dopson et al., 2004].

Hawkes et al. [2004] conducted a comprehensive investigation into the

microbiology of the MICCL Monywa chalcocite heap bioleaching operation in

Australia. A new moderate thermophilic Ferroplasma “like” strain was isolated

from the heap leach samples. The proposed strain, Ferroplasma

cyprexacervatum strain BH2 ("cyprus" L.n. meaning copper; "exacervo" L.v.

meaning "to heap up") is phylogenetically related to Ferroplasma acidiphilum

(16S rRNA gene similarity of 95 %). The morphology of the cells is non-motile

and pleomorphic cocci. The organism grew chemomixotrophically on ferrous

oxidation in the presence of yeast extract. Growth did not occur aerobically on

yeast extract alone (in the absence of ferrous sulphate). The organism was

shown to be facultative anaerobic, growing anaerobically on ferric iron in the

presence of potassium tetrathionate and yeast extract as electron donors.

Growth occurred between 14 °C to 63 °C, with an optimum temperature of

55.2 °C. Growth occurred between a pH of 0.4 and 1.8, with a pH optimum

between 1.0 and 1.2.

With the addition of Ferroplasma cyprexacervatum, it is clear that the genus of

Ferroplasma represents a diverse group of microorganisms. The growth

conditions of the various isolates are summarized in Table 3.


44

Table 3. The growth conditions of various species under the genus of

Ferroplasma.

Ferroplasma isolates Characteristic Fer1T MT17 DR1 YT BH2

Morphology Pleomorphiccocci

Irregular cocci

Irregular cocci

Pleomorphic cocci

Pleomorphiccocci

Chemo- autotrophic - ± ± ± NA

Chemo- mixotrophic + + + + +

Chemo- organotrophic + + + + -

Anaerobic growth + + + + +

Temperature range (oC) 23-46 32-51 32-51 15-45 14-63

Optimum (oC) 42 42 42 35 55.2 pH Range <0-1.5 0.35-3 0.35-3 1.3-2.2 0.4-1.8 Optimum 1.2 1.2 1.2 1.7 1-1.2 DNA G+C content %

36.8 36.5 37 36.5 NA

Sulphur metabolism - ± - - +

(+) growth, (-) no growth, (±) possible, but results indecisive.

To date, no species within the order Thermoplasmatales has been reported with

the ability to oxidize elemental sulphur or sulphur intermediates (thiosulphate or

tetrathionate) aerobically. Ferroplasma strain MT17, isolated from a pilot scale

bioreactor in South Africa, was described as being capable of chemomixotrophic

growth on organic carbon with either ferrous iron or tetrathionate (reduced

sulphur component) [Okibe et al., 2003]. In contrast to these results for MT17,

Dopson et al. [2004] could not obtain tetrathionate oxidation through autotrophic

or mixotrophic growth for the isolates MT17, Fer1T, DR1 and YT. Hawkes et al.

[2004] reported that Ferroplasma cyprexacervatum could oxidize tetrathionate

(S4O62-) anaerobically, using ferric iron as electron acceptor. No indication of

aerobic sulphur/sulphur intermediate oxidation was reported.


45

1.5. AIM OF STUDY

1.5.1. Literature summary

The moderate thermophilic archaeon Ferroplasma JTC 3 is found to abound in

bioleaching environments at JTC (BHP Billiton Johannesburg Technology

Centre) between 50 oC and 60 oC [Minnaar and Rautenbach, 2006]. It is evident

that the most applicable growth characteristic of Ferroplasma isolates,

concerning bioleach operations, is the oxidation of ferrous to ferric iron.

Ferroplasma strains do not seem to contribute to aerobic sulphur oxidation

[Golyshina et al., 2000; Dopson et al., 2004; Hawkes et al., 2004].

The Ferroplasma isolates mentioned in the review were capable of facultative

anaerobic growth, reducing ferric to ferrous iron (electron acceptor) in the

presence of yeast extract (electron donor) [Dopson et al., 2004; Hawkes et al.,

2004]. Ferric iron reduction related metabolisms are not commonly employed in

bioleaching, given that the oxidation of ferrous to ferric iron is generally accepted

to be the major aspect in the bioleaching of sulphide minerals, within a sulphate

system.

Hiroyoshi et al. [1997], Third et al. [2001] and Sandstrom et al. [2005]

demonstrated the benefit of leaching chalcopyrite in a ferrous iron promoted low

potential environment. The critical solution potential “window” (ferric/ferrous iron

ratio), at which low potential leaching was favoured in a sulphate system, was

identified between approximately 310 and 400 mV (mV vs. Ag/AgCl) [Hiroyoshi et

al., 2001].

1.5.2. Difficulty in controlling solution potential

The major drawback of low potential leaching is controlling the solution potential

below the critical upper limit of the “window” for prolonged periods of time.


46

It is especially difficult in low pH environments (pH < 2,), wherein ferric iron is

stable in a soluble state. The reason for the solution control problem is the slow

chemical oxidation of ferrous to ferric iron in the presence of oxygen and protons,

which escalates at increasing oxygen concentrations and a decreasing pH, thus

slowly increasing the redox potential within the leach system.

Third et al. [2001] employed a computer controlled bioreactor to control the

potential at a specific set point below the upper limit of the low potential

“window”. A constant low redox potential of 380 mV (Ag/AgCl) was maintained

by restricting the air supply to the reactor when the solution potential exceeded

the 380 mV set point (section 1.3.4.3).

1.5.3. Research objective - Combining the metabolic capacities of Ferroplasma JTC 3 with an aerobic/anaerobic solution potential control system

Experimental work conducted in this study illustrated that Ferroplasma JTC 3

also has the metabolic capacity for anaerobic ferric iron reduction, demonstrating

facultative anaerobic growth. The objective of the study was to use the anaerobic

ferric iron reductive metabolism and ferrous iron oxidation capability of

Ferroplasma JTC 3 to study the leaching of chalcopyrite in a ferrous iron

promoted low potential sulphate system. The experimental design was directed

towards controlling the solution potential of the leach system within the critical

low solution potential “window”, by means of an aerobic and anaerobic electronic

air and nitrogen gas flow controlled bioreactor, combined with the duel metabolic

capability (Fe2+ oxidation/Fe3+ reduction) of Ferroplasma JTC 3.

1.5.3.1. Experimental approach

A bioreactor fitted with a programmable electronic gas control system, capable of

switching between air and nitrogen, was used to create either an aerobic or


47

anaerobic environment within the bioreactor. Depending on the redox potential

within the leach solution, the controller allowed either air or nitrogen flow to the

bioreactor. The redox controlled mechanism is illustrated in Figure 8. The

experimental setup was different from the potential controlled bioreactor

described by Third et al. [2001]. Instead of using the principle of air restriction in

order to control the potential in the close vicinity of a single set point, the design

was changed to a strictly two phase aerobic and anaerobic gas flow mechanism,

managed between two set points.

Figure 8. The aerobic and anaerobic redox potential control system employed to

control the leach solution potential within the critical low solution potential

“window”.

The design of the gas flow control unit allows for the programming of two specific

redox potential set points, a lower and upper set point i.e. 310 mV and 400 mV

(mV vs. Ag/AgCl), respectively. During the first leaching stage (after reactor start-

up), the reactor is aerated, creating an aerobic environment. Within the aerobic

environment, ferrous iron can be oxidized to ferric iron (chemical or with

Ferroplasma JTC 3) until the redox potential reaches the upper potential set-

point (i.e. 400 mV), at which point the air flow to bioreactor would be switched to

nitrogen. The anoxic environment would be maintained until the redox potential

Air

Time

Red

ox p

oten

tial (

mV

vs.

Ag/

AgC

l)

310

400

Aerobic Anaerobic Aerobic

Air switched to N2

N2 switched to air

Upper set point

Lower set point


48

decreases to the lower set point (i.e. 310 mV), at which point the nitrogen feed

would be discontinued and the air switched back on.

During the anaerobic phase no ferric iron would be generated due to absence of

oxygen. Ferric iron, generated with the aerobic environment, can be reduced

(decrease in potential) via the combination of chalcopyrite oxidation and the ferric

iron reductive metabolism for Ferroplasma JTC 3. The solution potential can thus

be controlled via the oscillation of the potential between the upper and lower set

point. An important part of study was to directly evaluate the described

aerobic/anaerobic controlled low potential leach system against conventional

high potential leaching in terms of copper extraction from chalcopyrite.

In order to reach the objectives of this study the experimental work was

approached according to the following;

• THE IDENTIFICATION OF FERROPLASMA JTC 3. Denaturing gradient

gel electrophoresis (DGGE), molecular cloning and nucleotide sequencing

were used for the identification and phylogenetic classification of

Ferroplasma JTC 3.

• THE GROWTH CONDITIONS AND METABOLIC CAPABILITIES OF

FERROPLASMA JTC 3. This section focussed on the isolation, basic

metabolism and growth conditions of Ferroplasma JTC 3, specifically

directed towards the chalcopyrite leaching related experimental work.

• THE LOW POTENTIAL LEACHING OF CHALCOPYRITE. The metabolic

capabilities of Ferroplasma JTC 3 in combination with an

aerobic/anaerobic solution potential control system were employed for

studying the leaching of chalcopyrite in a ferrous iron promoted low

potential sulphate system.

chapter 1 literature review chapter 1

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