Alternative Lixiviants to Cyanide for Leaching Gold Ores

Chapter 27

Alternative Lixiviants to Cyanide for Leaching Gold Ores M.G. Aylmore John de Laeter Centre, Faculty of Science and Engineering, Curtin University, Perth, WA, Australia

1. INTRODUCTION The major impetus in seeking alternative lixiviants to cyanide arises from the environmental hazards posed by cyanide’s toxicity, with numerous environmental groups throughout the world actively pursuing a ban on its use. Approval for any new gold project using cyanide is extremely unlikely in some areas around the world. Elsewhere, increasing regulatory scrutiny of new gold projects and a lowering of acceptable levels of cyanide discharge are of considerable concern to mining companies. The resurgence in evaluating alternative cyanide processes has been driven by the high price of gold and increasing evaluation and treatment of more complex ores. With the commercialization of a thiosulfate-leaching plant at Barrick Gold’s Goldstrike operation, the interest in replacing cyanide by use of an alternative lixiviant will undoubtedly be more rigorously considered in the future. Over the past three decades, a significant amount of literature has examined alternative extraction processes to cyanide for recovering gold from different ores. The chemistry of these alternative processes has been reviewed by Avraamides (1982), Nicol et al. (1987), Hiskey and Atluri (1988), Sparrow and Woodcock (1995), and Aylmore (2005). Several recent reviews have also covered some aspects of alternative lixiviants (Konyratbekova et al., 2015), particularly covering extraction of precious metals from secondary waste materials (Cui and Zhang, 2008; Zhang et al., 2012a,b; Syed, 2012). There are over 1000 references that appear related to the application of alternative lixiviants to cyanide for leaching gold. In addition, there has been work carried out by private companies, research institutions, and commercial metallurgical laboratories that is not readily accessible. In total, 27 possible solvents have been investigated as alternatives to cyanide for leaching gold. These can be grouped into 11 categories as outlined in Table 27.1. Over the past decade, a significant amount of research has gone into improving our understanding of the chemistry and the whole process flowsheet design for thiosulfate-, thiourea-, halide-, and thiocyanate-leaching systems. The oxidative chloride-, sulfide- and ammonia-leaching processes have generally been used for extraction of a wide range of elements, including base metals and platinum group metals (PGM), with gold as a byproduct. Thiocyanate, nitriles, and combined cyanide lixiviant systems contain cyanide or derivatives of it and therefore may not be considered by some to be different to cyanide. Most of the other lixiviants discussed by Sparrow and Woodcock (1995) are of academic interest or have received limited publication. Therefore, this chapter will be concerned with the first eight categories and will concentrate on the most recent literature.

1.1 Stability of Alternative Lixiviants and Gold Complexes Despite the research interest in noncyanide gold lixiviants, many alternative gold processes are still at the developmental stages. A key factor affecting ultimate commercial success is the stability of the lixiviant and the gold complex in solution. In some cases, there is a limited understanding of solution and pulp chemistry. This is partly associated with (1) the difficulties in measuring reliable equilibrium data for various Au(I/III) complexes, (2) the lack of knowledge on

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447

448 PART | II Unit Operations

TABLE 27.1 Alternative Lixiviants to Cyanide 7. Bacteria/natural acids a. Amino acids b. Hyperaccumulating plants 8. Thiocyanate/Fe(III) 9. Nitriles/O2 or Cu(II) 10. Cyanide þ other combination a. Ammonia-cyanide b. Alkali cyanoform c. Calcium cyanamide d. Bromo cyanide 11. Others a. Electrolysis of ore slurry b. CSUT c. DMSO, DMF d. BioD leachant

1. 2. 3. 4.

Thiosulfate (Cu(II)-NH3-S2O3) Thiourea (Fe(III), CS(NH2)2) Halide (Cl2, Br2, I2) Oxidative chloride processes a. Aqua regia b. Acid ferric chloride c. Haber-Platsol d. Intec/N-Chlo e. Kell 5. Sulfide systems a. Sodium sulfide b. Polysulfide c. Biocatalyzed bisulfate d. Bisulfide/sulfur dioxide e. Nitrogen catalyst pressure process 6. Ammonia/O2 or Cu(II)

mixed-ligand complexes, and (3) different kinetic stabilities of Au(I) complexes with respect to disproportionation (Senanayake, 2004). A better understanding of the chemistry has evolved, in particular for thiosulfate (Senanayake, 2012) and thiourea (Li, 2006) leaching systems. The equilibrium data for complex formation, dissolution, precipitation, hydrolysis, and disproportionation reactions of Au(I/III) compounds for a range of noncyanide ligands have been evaluated (Senanayake, 2004). The stability constants (b2 or b4) for various Au(I) and (III) complexes, together with their standard reduction potentials, are shown in Table 27.2. Clearly the cyanide complex is more stable than any of the other reagents with thiosulfate, thiourea and bisulfide several orders of magnitude less stable.

TABLE 27.2 Stability Constants and Standard Reduction Potentials for Gold Complexes at 25 8C Ligand

Au(I) or Au(III) Complex

Log b2 or b4

Eo Au(I or III)/Au (V vs. SHEa)

CN

Au(CN)2e

38.3

0.57

S2O32e

Au(S2O3)23e

28.7

0.17

8e10

CS(NH2)2

Au(NH2CSNH2)2þ

23.3

0.38

<3

Cl

AuCl2e

9.1

1.11

<3

AuCl4e

25.3

1.00

AuBr2e

12.0

0.98

AuBr4e

32.8

0.97

AuI2e

18.6

0.58

AuI4e

47.7

0.69

HSe

Au(HS)2e

29.9

0.25

<9

NH3

Au(NH3)2þ

13

0.57

>9

Glycinate

Au(NH2CH2COO)2e

18

0.632

9

SCNe

Au(SCN)2e

17.1

0.66

<3

Au(SCN)4e

43.9

0.66

Au(SO3)23e

15.4

0.77

e

e

Bre

I

e

SO32e a

Standard hydrogen electrode.

pH Range >9

5e8

5e9

>4

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449

Oxidising

1.5

Hypochlorite Chloride

1.0

Ammonia Amino acids Bromide

Eh (Volts)

Thiocyanate

Iodide

0.5 Thiourea

Ammonium thiosulfate

bisulfide

0.0 Thiosulfate

Sodium cyanide

(ammonia free)

-0.5

-1.0 Reducing

0

2 Highly acidic

FIGURE 27.1

4

6

8 pH

10

12

14

Highly alkaline

EhepH diagram showing typical operating regions for gold lixiviants.

As a result of the wide range of values for the stability constants of the gold complexes, the standard reduction potentials for the different gold ligand species vary by almost 2 V (Ritchie et al., 2001). For many of the ligands, such as thiosulfate and thiourea, oxidation of the ligand occurs at a potential below that of the corresponding Au(I) complex while the reverse is true for ligands SCNe and Cle. Therefore, there is a competing reaction to gold dissolution with most alternative lixiviants, which increases reagent consumption. The presence of Fe(III) catalyst in acid thiourea solutions and Cu(II) catalyst in alkaline thiosulfate solutions also results in rapid oxidation of the ligand. Oxygen itself is often a poor oxidant due to low rates of mass transport and slow rates of reduction on gold surfaces in noncyanide systems. With the exception of the halides, the alternative lixiviants are clearly more complex to operate than cyanide. Most reagents have a small operating window where the alternative lixiviants effectively dissolve gold compared with cyanide (Figure 27.1). The high oxidizing potentials involved with some lixiviants inevitably lead to high reagent consumptions due to reaction with any sulfide minerals as well as oxidation of the reagent itself (Nicol, 1980). This applies particularly to thiocyanate and thiosulfate. Consequently, leaching conditions have to be better controlled than those used for cyanide leaching. Equally important, although not always considered, is the adsorption of reagents and/or precipitation of gold onto some clay and gangue minerals, which will be detrimental to overall gold recovery.

2. THIOSULFATE LEACHING In recent years, thiosulfate has been considered an attractive alternative to cyanide for leaching gold. A process using thiosulfate has been developed and commercialized for Barrick Gold’s Goldstrike operation in Nevada, USA. The primary attraction of thiosulfate is its low toxicity and its potential use on ores that cannot be readily treated by conventional cyanidation. Extensive studies have established the leaching mechanisms and the many issues that reduce gold recovery in thiosulfate solutions compared with cyanide. There are a number of reviews available in the literature that provide extensive references to the work on various ores that have been studied, the thermodynamics, speciation and mechanism of leaching, as well as the stability of thiosulfate and the various gold recovery options (Aylmore and Muir, 2001; Molleman and Dreisinger, 2002; Grosse et al., 2003; Muir and Aylmore, 2005; Senanayake, 2007, 2012; Senanayake and Zhang, 2012). Recent advances in understanding the thiosulfate process are presented in Chapter 28, and therefore this process is only mentioned briefly here for clarification.

2.1 ThiosulfatedProcess Conditions The chemistry of the thiosulfate system is relatively complicated compared with cyanide. However, by maintaining suitable Eh and pH conditions and by controlling reagent concentrations of thiosulfate, oxidant, and oxygen in the leach solution, high gold extractions can be achieved with low reagent consumption for some ores (Wan, 1997). A detailed analysis of anodic oxidation of gold has shown that the anodic oxidation takes place via the adsorption of the MS2O3 ion pair (Mþ ¼ Ca2þ, NH4þ, Naþ, Kþ) (Senanayake, 2005a). The rate of gold dissolution in the thiosulfate oxygen system is

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TABLE 27.3 Various Alternative Thiosulfate-Leaching Systems for Gold Oxidant

Complexing Agent

Lixiviant

Other Additives

pH

Cu2þ

NH4þ

(NH4)2S2O3

d

9.5

Cu

d

CaS2O3

Ore conditioned/pretreated

Neutral

O2

d

Na2S2O3

Pressurized

7e10

EDTA

Na2S2O3

Thiourea

6e7

Oxalate

Na2S2O3

Thiourea

4e5

(NH4)2S2O3

d

9.5

2þ

3þ

Fe

2þ

2þ

Ni /Co

NH4

þ

very low, and dissolution rates comparable with cyanide are only obtained with higher reagent concentrations and require an oxidant or catalyst other than oxygen. Several different oxidants have been proposed for the thiosulfate system, including oxygen, Cu(II) ammine complexes, Co(III) ammine complexes, and various Fe(III) complexes. Of these, copper-catalyzed ammoniacal thiosulfate-leaching system has been most extensively studied, with most work focused on understanding and applying the system to carbonaceous ores and copperegold ores, where gold recovery is poor when using cyanide or cyanide consumption is high. Researchers have also investigated ammonia-free and alternative oxidant systems to mitigate environmental issues associated with the use of ammonia. Barrick Gold has developed an ammonia-free copperecalcium thiosulfate process to commercial stage to treat pressure oxidized residues from their Goldstrike operation in Nevada, USA (Baul, 2013; Choi et al., 2013), which is described in more detail in Chapter 50. The different thiosulfate systems that have been investigated are listed in Table 27.3. The oxidation of metallic gold to the aurous Auþ ion in 0.10 M ammoniacal thiosulfate in the presence of Cu(II) occurs at an Eh of w0 V and can be simply represented by the following reactions: Au þ 5S2 O3 2 þ CuðNH3 Þ4 2þ/AuðS2 O3 Þ2 3 þ 4NH3 þ CuðS2 O3 Þ3 5

(27.1)

4CuðS2 O3 Þ3 5 þ 16NH3 þ O2 þ 2H2 O/4CuðNH3 Þ4 2þ þ 4OH þ 12S2 O3 2

(27.2)

The mechanism involves the formation and absorption of mixed Cu(II)eammoniaethiosulfate complexes on the gold surface with simultaneous oxidation of gold and thiosulfate (Senanayake, 2004, 2005b, 2012). Cu(II) is used as a catalyst for this reaction at concentrations w103 to 104 M (60e6 ppm) (Lam and Dreisinger, 2003), with oxygen acting to reoxidize Cu(I) to Cu(II) according to the simplified Eq. (27.6). In addition, oxidative decomposition reactions of thiosulfate occur, which result in the formation of polythionates such as tetrathionate and trithionate. The form and quantity of these degradation products are dependent on reagent concentrations, dissolved oxygen (DO) concentrations, pH value, Eh, temperature, and mineralogy. Various additives such as phosphates and ethylenediaminetetraacetic acid (ETDA) have been considered to improve the stability of the system, and this is discussed in more detail in Chapter 28. The major advantage of the FeeEDTA or Feeoxalate thiosulfate systems is that the oxidant in both cases is much less reactive toward thiosulfate than Cu(II). However, it is prone to reagent degradation issues associated with minerals such as sulfides. In addition, thiourea is required as a gold oxidation catalyst in the Fe(III)-leaching systems because gold leaching in the absence of thiourea is very slow (Zhang et al., 2005; Chandra and Jeffrey, 2005). The use of CaS2O3 salt instead of Na2S2O3 or (NH4)2S2O3 salt in the leach has beneficial effects on gold extraction, where calcium appears to prevent the formation of thiosulfate degradation products on the gold surface and assists in the maintenance of a constant and high leach rate during prolonged leaching (Feng and van Deventer, 2010). The use of calcium thiosulfate and copper as catalyst without any ammonia is the basis of Barrick Gold’s commercial-scale thiosulfate plant for treating carbonaceous ores.

2.2 ThiosulfatedOptimal Conditions for Leaching Optimal conditions for leaching appear to vary depending on the mineralogy of the ore treated and the deportment of gold. Preferred conditions reported by Muir and Aylmore (2004) for treating oxidized ore are 0.050 M thiosulfate (6.6 g/L),

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0.40 M total ammonia (6.8 g/L), and 60 mg/L Cu(II) at pH 9.5 and 40  C. The Eh is maintained w0.3 V by addition of air to give DO levels between 1 and 3 mg/L. Under the preferred leach conditions, thiosulfate consumption was found to range from 2 to 5 kg/t ore, representing w30% loss over an 8-h period. Ammonia losses were w1 kg/t at 25  C, but it appears that most was absorbed on the ore itself. For heap-leaching carbonaceous ores, a leach solution containing 0.1e0.2 M (NH4)2S2O3, at least 0.1 M NH3, and up to 60 mg/L of Cu(II) has been found suitable (Wan et al., 1994). Thiosulfate consumption was generally w3e5 kg/t ore. Similar reagent concentrations were used by Lakefield Research at 40e60  C and pH 8.0 to leach Barrick Gold carbonaceous refractory ores after pressure-oxidation pretreatment (Fleming et al., 2003). Up to 95% of the gold was extracted from the finely divided gold left in the oxidized residue. The calcium thiosulfateeleaching process has been extensively tested to treat pressure-oxidized double refractory carbonaceous sulfide ore at Goldstrike Nevada USA (Choi et al., 2013). Gold recoveries up to 85% have been reported in solutions containing 0.03 M calcium thiosulfate and 0.8 mM Cu(II), at a pH value of 8 and temperature of 50  C. To achieve high gold recoveries, the addition of a strong-base resin was found necessary; otherwise, gold recoveries were w40%. The process can operate at lower pH value of 8 compared with the conventional copper ammoniacal thiosulfate system. Both ferriceEDTA and ferriceoxalate complexes have been shown to be effective oxidants for the aerobic and anaerobic dissolution of gold in thiosulfate solutions and have been proposed as leaching reagents in the development of an in situ leaching system (Heath et al., 2008). In that study, ferriceEDTA leaching was carried out at a pH value of 7 with 50 mM ammonium thiosulfate (ATS) and between 1 and 10 mM ferriceEDTA. Anaerobic leaching experiments showed that both systems were still active after 7 days leaching, and when 1 mM thiourea was present, there was significant gold dissolution.

2.3 ThiosulfatedCurrent Status The more recent development work and construction of a thiosulfate-leaching plant at Barrick Gold’s Goldstrike operation will undoubtedly help provide some understanding of how effective thiosulfate as an alternative to cyanide will be on a commercial scale. However, the installation and operation of the Barrick plant is specific for the Goldstrike application only, and has alluded to the complex and unstable nature of this method of leaching gold and the significant process design difference of using thiosulfate over conventional cyanide. Further work by the industry is required to continue developing robust overall process flow sheets incorporating gold recovery, reagent recycling, and impurity control for a wider range of ore applications (see Chapters 28 and 50 for more information).

3. THIOUREA LEACHING Interest in the thiourea process for leaching gold occurred mainly during the 1980s and early 1990s. This earlier work was reviewed by Groenewald (1977), Hiskey (1981, 1988), Lan et al. (1993) and Li and Miller (2006). Considerable research has been carried out by Canmet and Mintek for gold extraction in underground (Tremblay et al., 1996) and in-stope applications (e.g., van Staden and Laxen, 1989). In addition, Newmont Mining and Barrick Gold looked closely at thiourea leaching as a potential process for treating refractory ores. Despite the potential carcinogenic properties of thiourea, studies on the oxidation of gold and thiourea in acidic thiourea solution continue to be carried out (Zhang et al., 2001; Li and Miller, 2002; Lapidus et al., 2008; Li and Miller, 2007). There is still laboratory research being carried out in Eastern Europe (e.g., Gönen, 2003; Örgül and Atalay, 2000, 2002; Gönen et al., 2007), the United States (e.g., Miller and coauthors), and Mexico (e.g., Lapidus, et al., 2008). Also, a novel leaching process was patented by Dublin University in Ireland (Kavanagh et al., 2000). A number of reports have also been published by a group of researchers in China on the dissolution of gold electrodes in an alkaline thiourea system (e.g., Chai et al., 1999; Zheng et al., 2006). However, thiourea is generally unstable in alkaline media and its application to ores has not been demonstrated.

3.1 ThioureadProcess Conditions In practice, thiourea leaching of gold is typically performed at thiourea concentrations of 0.13 M (10 g/L), ferric ion concentrations of 0.09 M (5 g/L), pH values of 1e3, and at potentials between 0.4 and 0.45 V (vs. SHE). The overall reaction that describes gold dissolution in thiourea and ferric ion solutions is: þ  Au þ 2NH2 CSNH2 þ Fe3þ ¼ AuðNH2 CSNH2 Þ2 þ Fe2þ (27.3)

452 PART | II Unit Operations

Various oxidants, including ferric ion, hydrogen peroxide, manganese dioxide, mono-per-oxysulfate compounds, and ozone have been examined, but ferric ion is the most common (Sparrow and Woodcock, 1995). The gold-leaching reaction is very sensitive to pH value and redox potential (Kenna and Moritz, 1991). Further, thiourea is intrinsically unstable and decomposes rapidly to substances that are unable to leach gold. Tremblay et al. (1996) concluded that to limit the thiourea decomposition and optimize gold extraction, it was necessary to maintain the leaching potential between 0.42 and 0.45 V (vs. SHE). Thus, careful control of solution potential in a commercial process is necessary. Leaching rates as high as 10 times faster than for cyanide have been reported, with silver and gold reacting differently, which implies that the dissolution mechanisms are not the same. The important kinetic parameters that govern the kinetic response of the gold dissolution in acid thiourea solutions using ferric sulfate as an oxidant in rotating disk technique have been reevaluated by Li and Miller (2007). When the ferric concentration is below 0.1 g/L with an initial thiourea concentration of 4 g/L, pH 1.5, 25  C, the rate of gold dissolution is limited by mass transfer of ferric ion to gold surface, whereas at higher concentrations above 0.2 g/L with an initial thiourea concentration of 4.0 g/L or when the ferric concentration is 0.5 g/L with an initial thiourea concentration of 0.5 g/L, a surface reaction appears to be the rate-limiting step (Li and Miller, 2007). Gold in contact with pyrite or chalcopyrite also exhibits an enhanced gold dissolution rate (van Deventer et al., 1990). However, thiourea also forms strong complexes with some base metals such as copper and, to a lesser extent, lead and zinc (Deschênes et al., 1994; Fang and Muhammed, 1992; Alodan and Smyrl, 1998), which can increase thiourea consumption. High potentials during leaching produce an oxidative degradation of thiourea that proceeds via formamidine disulfide (NH2(NH)CSSC(NH)NH2), which eventually decomposes to thiourea (CS(NH2)2), cyanamide (NH2CN), and elemental sulfur (So). Acid hydrolysis also forms urea (NH2CONH2) and hydrogen sulfide (H2S) as follows: 2NH2 CSNH2 þ 2Fe3þ ¼ NH2 ðNHÞCSSCðNHÞNH2 þ 2Hþ þ 2Fe2þ

(27.4)

NH2 ðNHÞCSSCðNHÞNH2 ¼ NH2 CSNH2 þ NH2 CN þ S

(27.5)

NH2 CSNH2 þ H2 O ¼ NH2 CONH2 þ H2 S

o

(27.6)

Elemental sulfur and hydrogen sulfide are undesirable species during gold leaching. It is believed that both these species cause a decrease in the leaching rate due to surface passivation, and hydrogen sulfide may cause reprecipitation of the gold (Sparrow and Woodcock, 1995; Munoz and Miller, 2000). Formamidine disulfide itself has the capacity to act as an oxidant but is unstable and over time decomposes (Lapidus et al., 2008). For optimum leaching, the oxidant must be added at such levels as to oxidize w50% of the thiourea to formamidine disulfide. Gold dissolution can be accelerated effectively by the presence of formamidine disulfide or by using a higher temperature (Zhang et al., 2001). However, excess oxidant will increase thiourea consumption significantly. In practice, it is necessary to use a stabilizer to convert formamidine disulfide back to thiourea or to complex the oxidant. The role of formamidine disulfide in the mechanism of gold dissolution in acidic thiourea solutions was reevaluated by Yang et al. (2010a) using cyclic voltammetry, electrochemical impedance spectroscopy, and linear sweep voltammetry. The adsorption of formamidine disulfide on the gold surface appears to activate and catalyze the dissolution of gold and inhibits the oxidation of thiourea. A thiourea:formamidine disulfide ratio of 10:1 was found effective to minimize thiourea decomposition and maximize gold dissolution rate. The presence of chloride and sulfate ions in solution may have a negative effect on gold dissolution. There is evidence that these anions co-adsorb with thiourea on gold surfaces and appear to prevent formamidine disulfide formation on the gold surface and promote thiourea oxidation and the formation of sulfur (Parker and Hope, 2008).

3.2 ThioureadStabilizers To combat the loss of thiourea, several workers have used additions of sulfur dioxide or substantial quantities of sulfite and Fe(III) complexing acids with mixed success (see Sparrow and Woodcock, 1995 and references therein). Deng and Liao, 2002; Deng et al. (2001) reported on test work involving the thiourea leaching of a bio-oxidized primary gold sulfide ore to which 4.5 g/L Na2SO3 had been added. The consumption of thiourea decreased significantly from 12 to 3 kg/t and extraction time was shortened from 6 to 1 h at a pH value of 2. Thiourea-substituted compounds such as N,Nʹ-ethylenethiourea (Schulze, 1985) are more stable to oxidation, which minimizes reagent consumption (Kenna and Moritz, 1991). Kenna (1991) patented a thiourea gold extraction method in which di- and tri-carboxylic acids, fluorides, fluosilicic acid, fluosilicate salt, and EDTA and EDTA salts are used to complex the ferric ion to reduce thiourea consumption. Cysteine (HSCH2CH(NH2)COOH) has also been found to be an effective reagent for the stabilization of thiourea

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(Ardiwilaga, 1999). However, cysteine prevented the formation of formamidine disulfide, which is required to maintain a high gold-leaching rate and consequently gold recovery was reduced. Washing with water and with varying concentrations of sulfuric acid prior to thiourea leaching has been found to be advantageous by several workers (e.g., Kavanagh et al., 1994; Tremblay et al., 1996; Lacoste-Bouchet et al., 1998; Deng et al., 2001; Urbano and Lapidus, 2014). The value of this approach varied from ore to ore; however, pretreatment with acid has been shown to not only remove easily leachable metals but also to allow the thiourea system to stabilize more quickly and to prevent unnecessary reagent consumption due to the precipitation of base metals. Activated carbon, some clay, and gangue minerals adsorb thiourea and its gold complex, thus reducing the overall gold recovery (Zhang, 1995).

3.3 ThioureadApplications Applications of thiourea leaching have been demonstrated at laboratory scale after fine grinding (Kusnierova et al., 1993), mechanochemical milling (Balaz et al., 2003), bacterial oxidation (Caldeira and Ciminelli, 1993; Wan et al., 1995; Deng et al., 2001; Deng and Liao, 2002), pressure oxidation (Yen and Wyslouzil, 1985; Bruckard et al., 1993; Murthy et al., 2003), and roasting (Bilston et al., 1990; Moussoulos et al., 1984). The perceived advantage is that the acidic preoxidized sulfide ore can be directly leached with thiourea without a neutralization step that would be required for leaching with cyanide. Thiourea has also been applied to oxide ores with mixed results (Ubaldini et al., 1998; McNulty, 2001). Recently, applications have included recovering gold and silver from spent mobile phones (Gurung et al., 2012, 2013). Most successful applications have been carried out on ores that have a high content of cyanicides such as antimony or sulfide ores that have undergone bacteria oxidation or pressure leaching. A small-scale industrial application of a goldeantimony concentrate was demonstrated at the New England Antimony Mine in Australia in the early 1980s, with gold recovery on activated carbon (Hisshion and Waller, 1984)dsee Chapter 52 for more information. Thiourea concentrations have ranged from 2 to 15 g/L, acid from 1 to 150 g/L (or higher), and pH 1e3 in laboratory and pilot-scale test work. The oxidant has varied from 0 to 20 g/L, where in some cases the iron from the ore has been used. Interestingly, gold extraction has not always been as high as that obtained using cyanide. In general, reagent consumption is dependent on reagent concentration, and values reported in the literature seem to be considerably higher (thiourea consumptions ranging from 4 to 47 kg/t) than those for cyanide, thus resulting in higher extraction costs. The high consumption of thiourea, acid (for pH control), peroxide, and sulfur dioxide (for Eh control) make the overall cost of the process between 1.5 and 2 times the cost of cyanidation for the same material. A 450 t pilot-test heap leach, using acidic thiourea, was conducted by Newmont at the Carlin mine after bio-oxidation (Wan et al., 1995). The bio-oxidation test heap was washed with fresh water following bio-oxidation and drained before thiourea leaching. The thiourea solution at a concentration of 10 g/L was pumped onto the heap at a flow rate of 30e45 L/min, with the whole operation running over a 110-day period at a pH value <2.5. The redox potential was generally in the range of 0.43e0.5 V (vs. SHE). No chemical control was required during the thiourea leach in terms of pH or Eh adjustments. Outcomes of this work were that thiourea yielded poor gold extraction kinetics because of the large particle size used and the cold temperature experienced during the pilot run. A maximum of 29% of the gold was recovered. Gold recovery from the pregnant solution using activated carbon or cation-exchange resin was ineffective. Continuous recirculation of the solution caused elemental sulfur to form, which coated the carbon and resin, impeding the recovery process. Lapidus et al. (2008) developed a method to extract and recover gold and silver from ores by partially electro-oxidizing the thiourea in solution to formamidine disulfide, which was then used as the oxidant in the system. Once the preciousmetal values are extracted from the ore and the solids separated, the pregnant solution is introduced into the cathodic compartment of the same electrochemical reactor for reduction and recovery of the metallic ions. The process has been tested on a variety of minerals, concentrates, and materials but only at laboratory scale. Six successive (0.2 M thiourea, pH 2) cycles on an acid-washed high-grade silverelead concentrate leached 35e40% of the silver. However, successive leaching of the same concentrate yield close to 100% silver recovery.

3.4 ThioureadAlternative Systems A mixed thioureaethiocyanate system has been shown to promote a higher dissolution rate than either lixiviant alone (Yang et al., 2010b, 2011; Zhang et al., 2014). The passivation of gold occurs in a thiourea-only solution, but when thiocyanate is mixed with thiourea, passivation is significantly alleviated. The dissolution rate of gold, in the mixed lixiviant system, increases with increasing thiourea and thiocyanate concentration. Results from linear sweep voltammetry and electrochemical impedance spectroscopy indicated that gold dissolution was controlled by a combination of

454 PART | II Unit Operations

charge transfer and diffusion in the mixed lixiviant system. The optimum concentration for gold dissolution for thiourea and thiocyanate was w5 mM and w0.05 M, respectively. Surface-enhanced Raman spectroscopy results suggested a possible formation of a mixed ligand complex involving the interaction of [Au(NH2CSNH2)2]þ and SCN. Applications to several high-grade gold ores have been carried out using 0.13 M thiourea, 0.78 M ammonium thiocyanate, and 0.028 M ferric sulfate solutions at pH 1.5 (Zhang et al., 2014). While high gold extractions were observed for low sulfur grade ore (95%), high sulfide content greatly reduced gold extraction due to the preferential consumption of ferric ion oxidant with pyrite. Some work has investigated using per-monosulfate as an alternative oxidant to ferric ion in ionic liquids of 1-butyl-3methylimidazolium hydrogen sulfate and chloride media to reduce reagent consumption with mixed results for sulfidic gold ores (Whitehead et al., 2009). Negligible recovery of base metals occurred in the ionic liquid medium, making it highly selective for Au and Ag. The ionic liquid mixture would restrict this type of application to materials such as waste circuit boards or concentrates with high gold grades.

3.5 ThioureadCurrent Status With thiourea labeled as a potential carcinogen, it is difficult to see it providing a replacement for cyanide in the near future. This is despite thiourea being used to treat hyperthyroidism in humans, although both thiourea and cyanide have been linked to inhibited thyroid function at chronic low-level exposure. Extensive investigations on all aspects of leaching have been evaluated and while there would be some potential niche applications, it appears that the sensitivity of leaching conditions would not make it an obvious choice compared with other alternatives. Lower thiourea concentrations would be required to make it economic. Recovery of gold from thiourea solution also requires further development. Despite the complexity of reactions and carcinogen issues, research continues at various academic institutions. The addition of small concentrations of thiourea in thiosulfate and thiocyanate systems has been shown to enhance the anodic oxidation of gold, as discussed in other sections. Hence, the application of thiourea to gold leaching may lie in improving other alternative lixiviant processes where only small dosages are used.

4. HALIDE LEACHING Chlorine, bromine, and iodine are well-known lixiviants for leaching gold, as reviewed by Tran et al. (2001). Chlorination was applied extensively in the late 19th century before the introduction of the cyanidation process. Bromine/bromide for leaching gold from ores was reported as early as 1846. Chlorination was used extensively for pretreating refractory and carbonaceous ores in several plants in the United States in the 1980s (Marsden and House, 2006). Renewed interest in halides as lixiviants for leaching gold occurred in the 1990s after several patents based on the bromine/bromide systems were lodged (see Pesic et al., 1992 and references therein; Tran et al., 2001). In recent years with increasing refractory nature of ores and higher gold price, there has been a resurgence in the development of halide-based gold-leaching processes. For completely oxidized materials, the chloride-based leaching processes have a clear advantage in applications where a high dissolution rate is required, and this is exploited in the Kell process, which applies chlorination to products from a pressure-oxidation step (Liddell and Adams, 2012a; Adams et al., 2015). The use of chlorine is a proven technology in gold-refining and electroplating processes (Feather et al., 1997; Maliarik and Ludvigsson, 2015; Viñals et al., 2006). However, halogens have proved to be very reactive with other ore minerals, especially sulfides. Under typical halideleaching conditions, reagent consumptions would be very high if the ore contained significant sulfide minerals. In recent years, work has gone into establishing ways to effectively recover and recycle reagents to reduce costs.

4.1 HalidesdProcess Conditions Typical conditions used for leaching gold by halogens are listed in Table 27.4. In all halogen-based leaching processes for gold, high-oxidation conditions are required. The general equation describing the reaction of gold with chlorine or bromine is as follows: 2Au þ 3X2 þ 2X ¼ 2AuX4  where X ¼ Cl; Br

(27.7)

The complex AuCl2e is formed initially and is rapidly oxidized to AuCl4e (Nesbitt et al., 1990; Sun and Yen, 1992). In the iodine system, iodine reacts with iodide in aqueous solution to form I3e ions with Au(I) rather than forming the Au(III)

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TABLE 27.4 Typical Conditions Used in Leaching Gold With Halides Reagent

Ligand

Oxidant

Gold Complex

Typical Leaching Conditions

pH Value

Chlorine

Cle

Cl2 or HClO

AuCl4e

5e10 g/L Cl2

<3

Bromine

Br

Br2

AuBr4

Iodine

Ie

I2

AuI2e

5e10 g/L NaCl e

e

2e5 g/L Br2, 0e10 g/L NaBr

5e8

1 g/L I2, 9 g/L NaI

5e9

complex (Eo for AuI2  < AuI4  ; Tran et al., 2001). The Au(III) complex is not stable because this ion oxidizes iodide ion to iodine and the Au(III) complex is reduced to the Au(I) complex: 2Au þ I3  þ I ¼ 2AuI2 

(27.8)

Unlike gold cyanide, which is very stable and does not decompose easily in most applications, stability of the gold halide complex is dependent on the solution pH value, composition (with respect to halide concentration), solution potential and the presence of reductants (such as metals and sulfidic minerals) in the ores. A residual amount of oxidant is required to maintain a high solution potential to avoid the precipitation of metallic gold from solution (Tran et al., 2001). The gold dissolution rate can be very high and is dependent on solution pH and lixiviant and oxidant concentrations (Sergent et al., 1992; Tran et al., 2001). Overall, the stability of halides is in the order of I > Br > Cl, whereas the rate of reaction is Cl > Br > I, with halide leaching significantly more rapidly than that for cyanide.

4.1.1 Chlorine The fundamental electrochemical kinetics of gold dissolution in chloride media and the chemistry of the chlorine process has been described by Finkelstein (1972), Nicol (1976) and Avraamides (1982). The dissolution of gold in chloride media was evaluated by Yen et al. (1990) and Tran et al. (1992), using hypochlorite as an oxidant. The weight loss of gold strips immersed in different chlorideehypochlorite mixtures (up to 20 mg/cm2 h) was much faster than that achieved by cyanidation under similar test
conditions (2.5 mg/cm2.h for 2 g/L NaCN) (Tran et al., 2001). The stability of the Au(III) e chloride complex AuCl4  is strongly dependent on the solution pH value and requires high chloride and chlorine levels, increased temperature, and high ore surface area. The complex is only stable at pH < 3.0 unless a sodium chloride concentration higher than 100 g/L is maintained. At a high pH value of 8, the hypochlorite ion (ClOe) is the dominant oxychloride species and the dissolution rate of the gold is very low. However, at more acidic pH conditions, hypochlorous acid (HClO) forms, which is a stronger oxidizing agent than ClOe, resulting in faster gold-leaching kinetics (Jeffrey et al., 2001a,b). The form of hypochlorite used appears to have some effect on gold-leaching kinetics. Baghalha (2007) observed that calcium hypochlorite produces slower gold-leaching kinetics, taking twice the time to achieve maximum gold recovery compared with sodium hypochlorite. Other alternative oxidants such as hydrogen peroxide and ozone have been used in gold-refinery processing (Maliarik and Ludvigsson, 2015), whereas chlorine dioxide has considered for treating high-grade enargite concentrates (Moldoveanu et al., 2014). As the dissolved gold complex is unstable and reprecipitates on contact with a reductant such as sulfidic materials or metals, application of the chlorideechlorine systems is limited to extraction of gold from oxidized materials. Attempts have been made to reduce the reactivity of sulfides in halide systems using compounds such as flotation collectors used to coat sulfides (Stace, 1984). This was found effective for some metal sulfides (e.g., copper sulfides), where the coating significantly reduced their reactivity without altering the reaction rate of gold dissolution. Other sulfides such as pyrite were unsuccessful. In addition, strong adsorption of goldechloro complexes on mineral surfaces, such as goethite, quartz, and alumina can occur, with adsorption becoming more pronounced as pH value is increased from 4 to 7 (Machfsky et al., 1991). High silver contents in ores may dissolve slowly under certain conditions in low-chloride solutions because of the formation of a passivating film of insoluble silver chloride (Sparrow and Woodcock, 1995). Consequently, higher concentrations of chloride in solution are required to solubilize the relatively insoluble silver chloride. Therefore, the chloride system may not be ideally suited to treatment of ores in which silver is of primary value.

456 PART | II Unit Operations

There are several chloride-leaching process variations for treating high gold-grade materials, which have been reported on various websites. These include the Clarkdale process and hyperleach process, but limited details are available on the efficiency of such processes for general gold leaching. The Kell process, which applies chlorination to products from a pressure-oxidation step (Liddell and Adams, 2011; Adams et al., 2015), has shown potential for application to a range of gold-bearing materials. Further information on chloride leaching of gold may be found in Chapter 29.

4.1.2 Bromine Bromine-leaching results in rapid gold dissolution at near neutral pH conditions. Due to the high potential required for dissolution (Eo ¼ 0.97 V vs. SHE), compared with the formation of gold cyanide at 0.57 V (vs. SHE), gold bromide is unstable and requires the addition of a strong oxidant such as bromine. Extensive research on the bromideebromine system was carried out in the 1990s. Electrochemical techniques were applied by Pesic and Sergent (1992) and by van Meersbergen et al. (1993) to determine the complex reaction mechanism for gold dissolution in bromineebromide systems. The dissolution of gold was shown to depend on the bromine:bromide ratio and the associated minerals in the ore. The presence of copper, zinc, and aluminum as sulfates has negligible effects on dissolution, but Fe(II) and Mn(II) are oxidized and consume bromine. Alternative oxidants to bromine to eliminate the problems associated with high vapor pressure and corrosive reactions of bromine have been examined with limited success. These are ferric ion, hydrogen peroxide, and sodium hypochlorite (Trindade et al., 1994; Sparrow and Woodcock, 1995). To eliminate problems associated with the high vapor pressure and corrosive nature of bromine, several organic-based bromine carriers (e.g., N-halo hydantoins such as Geobrom 3400) have been developed to stabilize bromine. Albermarle Corporation has tested an undisclosed product called Stabilized Bromine (11e15% active bromine), which stabilizes bromine in solution in a similar way as Geobrom 3400 (Melashvili et al., 2014).

4.1.3 Iodine Of all the halogens, the goldeiodide complex is the most stable in aqueous solution, because of its lower redox potential compared with the other halogens. Hiskey and co-workers have extensively evaluated the leaching behavior and chemistry of gold and silver dissolution in various iodideeiodine mixtures (Hiskey and Atluri, 1988; Hiskey and Qi, 1991; Qi and Hiskey, 1991, 1993; Angelidis et al., 1993). The gold dissolution rate was directly proportional to iodine and iodide concentrations and was not greatly affected by changes in pH over the range of 2e10. Several early patents were granted on the application of this system for in situ leaching (McGrew and Murphy, 1985; Jacobson and Murphy, 1988; Kubo, 1992) or for processing gold from electronic scrap (Falanga and MacDonald, 1982). The iodideeiodine system does not normally oxidize metal sulfides such as chalcocite and pyrrhotite, thereby avoiding excessive reagent consumption in a potential gold-processing plant (Hiskey and Qi, 1991). Consequently, it might be appropriate for treating ores containing sulfidic minerals. In recent studies investigating in situ leaching, low-reagent concentrations (e.g., 10 mM Ie and 5 mM I2), with gold leach rates similar to those obtained using an air-saturated 2.7 mM cyanide solution, were found effective in treating an oxidized ore (Roberts et al., 2010). However, recovery and regeneration of the active species are important, as iodide and iodine are very expensive. One method is to use strong-base anionexchange resins, which have been successfully tested to recover both gold and reagents from iodineeiodide solutions (Zhang et al., 2012a,b).

4.1.4 Mixed Halide Systems Alternative oxidants have been suggested in the iodine system, such as hypochlorite, which permits a lower iodine concentration in solution and so minimizes losses of iodine by evaporation (Davis and Tran, 1992; Davis et al., 1992). The conditions for extraction including oxidant/iodide molar ratio, concentration and pH value have to be optimized to avoid the passivation of gold by gold iodide (AuI) and to maximize the gold extraction rate. In addition, hypochlorite concentration has to be properly optimized as an overdose of this oxidant can destroy the iodide lixiviant used for complexing gold. A maximum gold dissolution rate was achieved at a [OCl]:[I] molar ratio of 0.25 (Davis and Tran, 1992). In this system, the active species dissolving gold is I3e, which is formed from the reaction between hypochlorite and iodide. The same applies for the hypochloriteebromide system, in which the active species dissolving gold is Br3e (Tran et al., 2001). Mixed chloride and bromine systems have been developed, but most use copper as an oxidant and are covered in Section 5.

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457

4.2 HalidesdApplications Pretreatment of sulfidic or carbonaceous ores by roasting or pressure oxidation is normally required prior to chlorine or bromine leaching to render ores relatively inert and consequently reduce reagent consumption (Li et al., 1992; Sparrow and Woodcock, 1995). For example, the selective recovery of gold and silver was carried out by Puvvada and Murthy (2000) from a chalcopyrite concentrate. Gold and silver grades in the concentrate were 11 and 140 g/t, respectively. Laboratoryscale tests were conducted at room temperature on 20% solids slurry containing 25 g/L NaOCl and 0.35 M HCl. Increasing the NaCl concentration increased the rate as well as the extent of gold and silver extraction. Gold and silver recoveries of 42.7% and 45.0%, respectively, were obtained with 200 g/L NaCl. Dissolution of silver was found to be independent of NaOCl concentration. However, pressure oxidizing the copper concentrate and then leaching with NaOCl concentration of 25 g/L, 200 g/L NaCl, and 0.35 M HCl for 1 h resulted in gold and silver recoveries of 90.0% and 92.5%, respectively. A prefeasibility test on hypochlorite-leaching McDonald Gold Mine oxide ore in Montana gave 68% gold recovery compared with 73% obtained with cyanide (McNulty, 2001). Under the conditions used in their bottle-roll test work (pH 6.5, Eh 1.14 V (vs. SHE)), 33% solids (for 96-h leaching time) reagent consumption was 5.55 kg/t NaOCl and 3.25 kg/t HCl compared with 0.15 kg/t NaCN and 0.55 kg/t CaO for the cyanide process. More recent work by Nam et al. (2008) has examined conditions to leach a mine tailing containing gold and silver. Using chlorinated sea water (approximately 0.5 M NaCl, pH 5.5 and Eh of <1.00 V vs. Ag/AgCl), both gold and silver could not be fully extracted due to the formation of gold hydroxide and silver chloride. Nevertheless, the gold and silver extractions reached values of 80% and 50%, respectively. Interest in the use of chloride-based leaching for treating gold from copper anode slime has continued with several studies on its kinetics (Dönmez et al., 1999; Herreros et al., 1999). The technique was tested or practiced in several copper refineries in the early 1990s (see Tran et al., 2001 and references therein). The MinataurÔ process, developed by Mintek in South Africa, comprises oxidative leaching of the feed material, followed by selective solvent extraction of the gold from the leach liquor to reject impurities and precipitation of high-purity gold (Feather et al., 1997). Suitable feeds include silver-refining anode slimes, gold-electrowinning cathode sludge, zinc-precipitation filtrates, gravity-gold concentrates, and the residues from mill liners in gold plants. Impure gold feed material is leached for 2e3 h in 5 M HCl under oxidizing conditions with chlorine continuously added into the leach reactor. The leach solution is then purified by solvent extraction. Gold is selectively extracted into the organic phase (not specified) over silver as well as the platinum-group and base metals, which report to the raffinate. Gold is recovered as a metal powder by direct reduction in sulfur dioxide or oxalic acid from the loaded strip liquor. In addition, the GravitaurÔ process has been developed that incorporates gravity concentrates as the feed, without the need for an intermediate solvent-extraction step to upgrade the gold tenor of the solution. Chinese researchers (Li et al., 1996) have also evaluated a chloridizing process for gold extraction from silver anode sludge, where gold is recovered by reduction of liquor with oxalic acid, possibly similar to the MinataurÔ process. Outotec’s gold refinery hydrometallurgical process uses a solution of hydrochloric acid with either chlorine gas or aqueous solution of hydrogen peroxide as an oxidant to dissolve gold and platinum group elements from gold bullions that have been atomized (Maliarik and Ludvigsson, 2015). A pretreatment process to remove impurities such as Cu, Fe, Se, Te and Bi may be required before leaching takes place. Precipitated silver chloride formed during the oxidative leach is filtered and recovered. Gold in filtrate is selectively reduced to metallic gold by reaction with hydrogen sulfite, filtered, and the filtrate further processed for removal of platinum-group metals. Albermarle Corporation, using their Stabilized Bromine (11e15% active bromine) compound to stabilize the reagent, observed high gold recoveries from oxide ores, with bromine (88e95%) comparable to that achievable with cyanide (90e95%). Bromine consumption was 1.45 kg/t when leaching was conducted at a near neutral pH value of w6 (Melashvili et al., 2014). Higher gold recoveries were achieved with bromine (70%) than with cyanide (23%) when leaching gold encapsulated in sulfides; however, bromine consumption was high (>500 kg/t) owing to simultaneous oxidation of the sulfide minerals. Bromine was also found to be less reactive than cyanide to copper minerals. In situ leaching, using iodine as the lixiviant, has been examined as a method for the extraction of gold from stranded oxidized resources which cannot be mined economically using conventional technologies (Roberts et al., 2010; Martens et al., 2012). Preliminary laboratory work using iodineeiodide as the lixiviant systems returned promising results (80% gold extraction). However, column leaching tests (initial 2.5 mM I3e, 4.3 mM I, pH w3.5, flow rate 49 mL/day for 94 days and then doubled) observed reprecipitation of gold from leach solutions as a result of loss of tri-iodide concentration and the presence of competing reductants in the ore material. This reduced the effectiveness of in situ leaching and the time required for gold recovery.

458 PART | II Unit Operations

Some other recent applications of iodine leaching for gold extraction from printed circuit boards, oxide ore, and concentrates have been reported (Konyratbekova et al., 2015). For the treatment of carbonaceous ore with 1.6 wt% organic carbon content, results showed only 20% gold extraction in a solution containing 20 g/L iodide and 8 g/L iodine. This low level extraction is considered due to adsorption of goldeiodide complexes on organic matter. In contrast, gold extraction from the oxide ore in a solution containing 20 g/L iodide and 4 g/L iodine reached 89% in 24 h (Baghalha, 2012).

4.3 HalidesdCurrent Status Halide leaching provides greater flexibility than cyanidation because reagent dosages can be controlled to enhance the dissolution rate and this is particularly useful for processing coarse gold in oxidized materials. However, gold halides are unstable under some conditions, and chemical and process control is required during processing to maintain gold in solution. Further, losses of bromide or iodide absorbed on gangue minerals or precipitated as insoluble copper silver or lead salts can lead to high consumptions of these reagents for low-grade ore applications. The technique is suitable for extracting gold from gold-rich materials such as anode slimes and oxidized gravity-gold concentrates. Chlorination processes are being utilized as a gold- and silver-refining technique to replace smelting. In the past the lack of a suitable recovery process to match the cyanide CIP plant technology has been a limitation of the use of chloride systems for treatment of whole ores. Recent developments in ion-exchange resins and solvent extraction may alleviate this problem. Renewed research into the treatment of gold ores with bromine has occurred with the two main providers of bromine from sea water, Albemarle Corporation and ICL, investigating potential applications in the gold mining industry. There has also been renewed interest in applying iodine-iodide system to in situ leaching (solution mining) applications.

5. OXIDATIVE CHLORIDE LEACH PROCESSES Besides chlorine, several other oxidants, such as oxygen, ferric, and cupric ions or nitric acid, dissolve gold in the presence of chloride. Oxidative chloride leaching may have possible applications in treating silicate ores if only small amounts of high gold sulfide minerals are present, such as in the Kell process, which incorporates POX ahead of chlorination. Oxidative chloride leaching can be used as a total dissolution process for treating sulfide ores to recover base and precious metals. Most oxidative chloride leach processes, including ferric chloride leaching, have been mainly used as a pretreatment process rather than a gold-leaching process. The PlatsolÔ Process has been tested for the leaching and recovery of base metals, gold, and PGMs from sulfide ore, as has the Kell Process. The lntec process, Nippon N-Chlo, Nichromet, and Outotec’s Gold Chloride processes have all been tested for the leaching of gold in concentrated chloride solutions at atmospheric pressure.

5.1 Oxidative Chloride LeachdProcess Conditions Some conditions used for leaching gold in chloride systems are listed in Table 27.5. Except for special processing of highgrade materials, such as the refining of precious-metal concentrates, gold bullion, or platinum metals, aqua regia is not considered practical for use at the plant-scale level. Suitable oxidants include oxygen under pressure, ferric salts, hydrogen peroxide, sodium hypochlorite, and persulfates (Sparrow and Woodcock, 1995). Acid ferric chloride solutions have been used primarily as a process to oxidize sulfide concentrates and complex leadezinc sulfide ores and concentrates prior to cyanidation (e.g., Jain and Hendrix, 1996). However, electrochemical investigations by Liu and Nicol (2002) have demonstrated the effectiveness of Fe(III) as the oxidant for leaching of gold at high temperatures under typical pressure-oxidation conditions used for treating refractory gold ores. Measurements of the equilibrium potentials of the Au(III)/Au and Fe(III)/Fe(II) couples over a temperature range from 25 to 200  C in acidic sulfate solutions containing various concentrations of chloride ions has shown that the equilibrium solubility of gold increases with increasing temperature, chloride concentration, and Fe(III):Fe(II) ratio. The application of ozone as an oxidant has been studied by several authors with an ozone-leaching process in dilute hydrochloric acid (Metalozon) been reported for recycling of gold from scrap metal (Viñals et al., 2006). Chlorine dioxide (ClO2) generated from sodium chlorite and hydrochloric acid can potentially be used as an oxidant in hydrochloric acid systems for leaching gold; applications have only been applied to oxidative dissolution of enargite (Moldoveanu et al., 2014). The other processes listed in Table 27.5 are discussed in more detail in the following sections.

5.1.1 PlatsolÔ Process The PlatsolÔ process was originally developed in collaboration with the University of British Columbia, Kane Consultants Ltd., and Lakefield Research in Canada for the treatment of flotation sulfide concentrate for Polymet Mining Company in Minnesota and was tested on similar types of concentrate materials. This process involved dissolution in one step of the

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459

TABLE 27.5 Typical Oxidative Chloride Conditions Used in Leaching Gold Process

Ligand

Oxidant

Gold Complex

Typical Leaching Conditions

pH

Aqua regia

Cle

HNO3

[AuCl4]e

3:1 HCl:HNO3

<0

Ferric chloride

Cl

Fe

[AuCl4]

e

3e6% FeCl3

<2

Ô

Platsol

Cl

O2 (689 kPa)

[AuCl4]

e

5e10 g/L NaCl, T > 220  C

w2

Intec and N-Chlo processes

Cle/Bre

Cu2þ, Fe3þ

[AuCl/Br4]e

280 g/L NaCl, 28 g/L NaBr T 85e95  C Air or O2, 20e60 g/L Cu2þ

<2

Nichromet process

Cle/Bre

Cl2/Br2

[AuCl/Br4]e

275e300 g/L NaCl 2.7e3 g/L NaBr T 35e45  C

1e2

Neomet gold process

Cle

Cu2þ, Fe3þ HNO3 O2 H2O2

[AuCl4]e

T 105e110  C 30e35% HCl

<2

Outotec gold process

Cle/Bre

Cu2þ, Fe3þ

[AuCl/Br4]e

T 80e90 C 225 g/L Cle NaBr 100 g/L 50e100 g/L Cu2þ

1.8

Kell process

Cle

Cl2

[AuCl4]e

3.5 M HCl; T 80 C

<2

e e

3þ

base metals (copper and nickel) as well as the gold and PGMs. This was followed by solideliquid separation, gold and PGM recovery, and conventional Cu SX/EW and recycling of the copper raffinate to the autoclave. The fundamental difference between the PlatsolÔ process and the conventional high-temperature pressure-oxidation processes is that a small concentration of chloride ions is added to the autoclave with w25 g/L sulfuric acid. The chloride favors the oxidation of gold and PGMs and stabilizes them as dissolved chloro-complexes. Grinding the ore with ceramic rather than iron balls was required to prevent cementation of gold chloride (Ferron et al., 2000). The concentrate tested was a flotation concentrate from the Northmet project, USA, assaying 14.7% Cu, 3.05% Ni, 0.14% Co, 26.7% S, 1.4 g/t Au, 2.2 g/t Pt, and 9.9 g/t Pd. Pressure-oxidation conditions were 225  C, pulp density was 11%, retention time was 120 min, and oxygen overpressure was 689 kPa. The ore treated had a P80 of 15 mm. After solideliquor separation, the gold and PGMs were recovered by sulfide precipitation using NaHS or by activated carbon. The copper was recovered using conventional solvent extraction and electrowinning techniques. Overall recoveries were Cu 99.6%, Ni 98.9%, Co 96%, Pd 94.6%, Pt 96%, and Au 89.4% (Ferron et al., 2000). Studies in treating a variety of refractory gold concentrates under optimum conditions (225  C, NaCl 10e20 g/L, 2e6 h O2 at 700 kPa) achieved gold extractions of the order of 90e96% compared with direct cyanidation where gold extraction was less than 20% (Ferron et al., 2003). Examination of a variety of recovery options showed that loading of gold onto carbon from clear liquors and pulps was rapid and did not require prior neutralization. Zadra elution of the loaded carbon recovered more than 90% of the gold; however, further work in investigating carbon regeneration was required. Gold could easily be precipitated from acidic PlatsolÔ leach liquors with NaHS, but minimization of the co-precipitation of impurities, such as copper, needed to be addressed, given the co-leaching of base and precious metals. In addition, some tests on gold recovery by ion-exchange resins showed promise. Gold chloride can be precipitated using a synthetic covellite produced in residual copper recovery process. The PlatsolÔ process has undergone a detailed engineering phase following DFS and FEED studies for the Northmet project (Wardell-Johnson et al., 2009). More details may be found in Chapter 46.

5.1.2 Intec and N-Chlo Processes The Intec Copper process is primarily aimed at processing copper and other base metal sulfides; however, application to the simultaneous extraction and recovery of gold/silver has also been evaluated (Moyes, 1999). In this process, copper sulfide

460 PART | II Unit Operations

feeds are typically leached at 85  C using a chloride electrolyte (280 g/L NaCl and 28 g/L NaBr) containing a chlorineebromine complex (BrCl2e, named Halex) produced from a copper electrowinning cell downstream. Gold and silver are leached by this leachant, which achieves a solution potential (Eh) of w1.2 V (vs. SHE). Silver is recovered by cementation, whereas gold is recovered by precipitation as the solution potential is dropped to <0.80 V (vs. SHE) after leaching (Severs, 1999). High gold extractions have been achieved from a range of refractory gold concentrates at laboratory scale, which has led to a continuous locked-cycle pilot-plant program. The pilot plant operated at >99% availability, with a maximum of 96.5% gold extraction from a concentrate containing 58.6 g/t gold. Some >99% of the dissolved gold was loaded onto carbon at up to 1% w/w, with no loss of carbon activity detected over five loading/washing/ elution cycles. The Intec gold process, an extension of the halide-based Intec Copper Process, has also been tested. Sulfide concentrate was ground to a P80 of 10e40 mm and fed to a leach train where the refractory sulfides were oxidized along with the liberated gold by oxygen from direct air injection. Operating conditions involve temperatures between 85 and 95  C with a 6e8 M chlorideebromide electrolyte containing 20e60 g/L Cu2þ, leaching for up to 10 h. Copper was used as a catalyst to assist the transfer of oxygen from air. Soluble iron was controlled by the addition of limestone into the last leach reactor. The slurry was then sent to the solideliquid separation stage, where the leach residue was washed and discharged. Gold in the clear liquor was adsorbed onto activated carbon or resin-packed fixed-bed columns and subsequently eluted using thiourea, thiosulfate, or cyanide. Gold was subsequently recovered as a metal by electrowinning or cementation. Golddepleted liquor was sent to the purification circuit where byproducts were precipitated with slaked lime. The precipitated solids were separated by filtration where they were washed, and the filtrate recycled to leaching operations. Ferrous and cuprous reaction products were subsequently oxidized by further air sparging. The N-Chlo process uses a similar approach to the Intec copper process where finely ground copper and gold sulfide concentrate are leached in chloride bromine medium (Abe et al., 2009; Takebayashi et al., 2011). Copper is used as a catalyst to assist the transfer of oxygen from air. The slurry is then sent to the solideliquid separation stage, where the leach residue is washed and discharged. Gold is recovered on carbon, whereas solvent extraction is used to recover silver. A 1.3 t/h demonstration plant was operated in Perth, Australia, treating Newcrest Mining’s concentrates from Telfer and Cadia mine operations (Takebayashi et al., 2011).

5.1.3 Nichromet Process The Nichromet Process is a chloride/bromideebased process (Lalancette, 2009; Lemieux et al., 2014). The Nichromet Process uses a pretreatment roasting step to eliminate sulfur. The roasted ore is leached in sodium or potassium chloride using bromide as a catalyst, and the leach liquor is regenerated by electrolysis. The production of hypochlorite and hypobromite as a source of active Cl2 and Br2, respectively, is done via electrolysis of the brine in a standard electrolytic cell. NaOCl/NaOBr solution is added to slurry, which is made acidic by the addition of sulfuric acid to liberate free Cl2 and Br2. Following filtration, gold and silver in leach liquor can be recovered using activated carbon or a reducing agent such as sulfite or sulfur dioxide. The process was piloted on a high-grade calcined concentrate and claimed to yield gold recovery of 96.5%, compared with 85% obtained by cyanidation. Impurities such as iron and aluminum are removed by neutralization with caustic generated during an electrolysis step. In addition to the difficulties in filtrating caustic-generated precipitates, the handling of heavy metals mobilization during leaching would be an issue. Details of reagent consumption and costs associated with operating such a process have not been reported to date. The process is still in the development stage.

5.1.4 Neomet Gold Process The Neomet Gold Process is a chloride-based metals extraction process in which hydrochloric acid is regenerated and iron in solution precipitated as hematite (Harris and White, 2011a,b). The Neomet Process recovers the acid as either as a gas, or as concentrated (30e35%) HCl. The hematiteealumina product formed is claimed to be suitable for sale as a byproduct. Leaching is carried out under atmospheric conditions at 105e110  C in recycled hydrochloric acid, with some recycle of the leach filtrate to increase both the gold and chloride concentration. For ores containing significant pyrrhotite or other reactive sulfide component, the leach is divided into two stages: a primary reducing leach to destroy reactive sulfide minerals, followed by a secondary oxidizing leach to dissolve the gold. The oxidizing leach is carried out at a high (>750 mV) redox potential of the final solution with a suitable oxidant (nitric acid as an oxygen transfer agent, hydrogen peroxide, or chlorine in association with a high concentration of cupric or ferric chloride) depending on the characteristics of the material being treated. Gold is recovered from the leach liquor by passing the solution through an absorbent resin

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(XAD-7). The loaded resin is collected and eluted by dilute hydrochloric acid to give a concentrated (>20 g/L) solution of gold chloride. Pure gold is recovered from this solution by ferrous chloride reduction. Silver present in the leach liquor, together with any gold not recovered by the resin is recovered by adsorption on activated carbon. While there has been some piloting of this process, it is still in the early development stage.

5.1.5 Outotec Gold Process Outotec has developed a chloride-based process for gold materials that combines chloride leaching and solvent extraction to recover gold (Miettinen et al., 2013). The process uses a combination of chloride and bromide under conditions not too different from the Intec process. Gold is primarily oxidized by copper in concentrated chloride solution at 80e90  C, at atmospheric pressure and at approximately pH 1.8 (Miettinen et al., 2013; Lundstrom et al., 2007, 2014). Iron converted to ferric in solution precipitates as goethite or hematite. Following solideliquid separation and washing of the residue, the gold in solution is recovered separately by solvent extraction using an organic reagent OT-307 developed by Outotec. Gold recovery from the organic extractant is carried out in an aqueous solution. Gold-loading capacity of OT-307 has been reported up to 90 g/L. Accumulations of impurities in leach streams (such as Cu, Zn, Pb, Ni, Co, and Mg) are removed by neutralization. Test results indicate that natural copper, pyrrhotite, sphalerite, galena, nickel sulfides, chalcocite, covellite, and copper oxides are rapidly leached. However, chalcopyrite and bornite show moderate leaching kinetics, whereas pyrite, enargite, tennantite, and tetrahedrite are more refractory. This process is also still in its early stage of development.

5.1.6 Kell Process The Kell process is a sequential sulfate and chlorideebased technology, where base metals are removed by sulfuric acid pressure leaching and PGMs and gold are leached with Cl2/HCl as commonly used in PGM refining. Dissolved precious metals are recovered from the pregnant rich solution by adsorption onto ion-exchange resins and precipitation to highpurity products. The process has been patented (Liddell, 2003) and further developed (Liddell et al., 2011; Liddell and Adams, 2012a,b). After the initial POX leach for base metals, a heat treatment may be applied to convert precious metalebearing minerals into forms that are soluble in a subsequent chlorination leach, at 95e99% extraction efficiencies. The process has been subjected to integrated continuous pilot-scale testing on various PGM/Cu-Ni concentrates. Pallinghurst Resources is assessing potential construction of a full-scale Kell plant to extract base metals (copper, nickel, and cobalt) and PGMs as well as gold at its Sedibelo Platinum Mines subsidiary (Seccombe, 2014). The process has been applied to treatment of various refractory gold feedstocks (Adams et al., 2015). Preliminary batch testing of a copperegold concentrate indicated >99% gold and silver extractability by chlorination. Three samples classed as potentially double refractory yielded >90% chlorine-extractable gold. Samples of calcine tailings that had already been cyanide leached yielded 87% and 100% chlorine-extractable gold. KellGold has been proposed as a potential noncyanide treatment option for polymetallic, copperegold, and refractory gold concentrates.

5.1.7 Other Processes There are several other processes such as the Cominco Engineering Services Ltd. process (CESL process) and Hydrocopper (Lundström et al., 2009, 2014) that incorporate chloride in their processes for copper extraction. In these processes gold is extracted into solution as a chloride complex and can be recovered by resin, solvent extraction cementation, or precipitation methods (Mpinga et al., 2015).

5.2 Oxidative Chloride LeachdCurrent Status Most of these processes are suitable for high-grade concentrates. For sulfide ores containing a combination of base metals, precious metals, and PGMs, some of these processes may be economical and viable process for polymetallic ores. To overcome issues associated with chlorine consumption in the presence of sulfides, some of the processes use a pretreatment process to remove sulfide components (e.g., generating H2S gas or pressure oxidation), which would add an extra cost to the process plant. To enable recycling of reagent and the recovery of gold, processing requires a solideliquid separation step and the need to use solvent extraction, an absorbent, or precipitation, rather than recovery in pulp. Handling of heavy metals mobilized during leaching may present an issue, particularly for the processes leaching base and precious metals in one step.

462 PART | II Unit Operations

6. SULFIDE/BISULFIDE/SULFITE LEACHING The sulfur-based gold lixiviants other than thiosulfate and thiourea are sulfide, bisulfide, bisulfite (or sulfur dioxide), and polysulfides. Of these possible lixiviants only the bisulfide and bisulfite ions appear to have any practical use under ambient conditions. YES Technologies (Hunter et al., 1998) described a process using bisulfide generated from sulfatereducing bacteria as a lixiviant for gold, although bisulfide ions can be readily generated from hydrogen sulfide gas. Alkaline sulfide lixiviation of sulfur residues from the Nitrogen Species Catalyzed Pressure Leaching process (Anderson, 2003, 2008, 2014; Anderson and Twidwell, 2008) has been demonstrated to leach and recover silver successfully.

6.1 Sulfide/Bisulfide/Sulfite LeachingdProcess Conditions The possible processes investigated using sulfur-based lixiviants for gold are listed in Table 27.6. The mechanism of gold dissolution and precipitation from aqueous sulfide solutions under a range of conditions has received considerable attention in the early geological literature (e.g., Hannington and Scott, 1989; Seward, 1973; Tan and Bell, 1990). Neutral bisulfide solutions dissolve gold as follows: 2Au þ 2H2 S þ 2HS ¼ 2AuðHSÞ2  þ H2

(27.9)

o

Under weak acid conditions, the neutral gold complex Au(HS) species forms, while under strongly alkaline conditions, Au2(HS)2S2 is formed (Seward, 1973). Because of the similar stabilities of gold and silver bisulfide complexes, bisulfide leaching may be suitable for leaching gold ores with significant silver content (Hunter et al., 1998). The YES process uses naturally occurring, sulfate-reducing bacteria for the recovery of gold and silver from ores. A full description of the process is outlined in a patent by Hunter et al. (1998). Conventional bio-oxidation of ore particles is carried out to free the precious metals dispersed or occluded within the ore. A portion of the acidic, base-metal sulfate leach solution produced from bio-oxidation is introduced to an anaerobic reactor. A nontoxic donor such as acetate or methanol (which does not bind to activated carbon) is added to the anaerobic reactor to enrich within it a mixed culture of sulfatereducing bacteria. Bisulfide ions are generated biologically in the process with an electron donor such as acetate by the following reaction; CH3 COO þ SO4 2/2HCO3  þ HS

(27.10)

Additions of acidic sulfate solution may be required to maintain a neutral pH in the reactor. In a second process step, the submerged oxidized ore (to exclude oxygen) is leached by recirculating the neutral bisulfide lixiviant, saturated with H2S. Precious metals are recovered from the pregnant bisulfide solution by contact with activated carbon or other conventional techniques. Since HSe ions and H2S molecules diffuse more slowly than cyanide ions and oxygen molecules, slower gold dissolution can be expected than with cyanide leaching if the same concentrations of reactants are used. Possible problems with regard to obtaining high gold recoveries include passivation and adsorption effects, which are not well understood.

TABLE 27.6 Typical Sulfur-Based Conditions Used in Leaching Gold Reagent

Ligand

Oxidant

Gold Complex in Solution

Typical Leaching Conditions 50 g/L Na2S, pH 12

Sodium sulfide

e

HS

H2S generated

Au(HS)2e

YES technology (bacteria SO42 reducing process)

HSe

H2S generated

Au(HS)2e

Around 2.5 g/L H2S, HS, and S2, pH 6e9

Bisulfite or SO2

HSO3e

O2

Au(HSO3)2e

15e50 kg/t SO2, pH 4e5

2e

o

S

Au/Sx

2 M polysulfides

O2, Cu(II), and NH4þ

Au(S2O3)23,

?

Au/Sx(?)

NOþ, So

AuS5e; Au(S2O3)23e

Polysulfide

Sx

Lime sulfur synthetic solution/phase transfer catalysts

S2O32,

Nitrogen species catalyzed pressure leaching process

S52, S2O32e

Sx

2e

20e175 g/L H2SO4, 620e975 kPa, 125e170  C, 2.0 g/L HNO3

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Early investigations into using polysulfides to leach gold have been carried out by Chen et al. (1996). Gold dissolution can occur as a result of adsorption of polysulfide on gold surfaces accompanied by oxidation of polysulfide via the following (unbalanced) reactions: Au=Sx /AuS þ ðx  1ÞSo þ e

(27.11)

Au=Sx /Au=Sx þ 2e

(27.12)

So þ 2e/S2

(27.13)

Gold dissolution can occur as a result of adsorption of polysulfide on gold surfaces accompanied by oxidation of polysulfide as Chen et al. (1996) reported, with 90% gold extraction from a sulfide concentrate at 50  C without the addition of an oxidant. However, a relatively high polysulfide concentration is required for high gold extraction. Similar processes where polysulfide was generated in situ such as the lime/sulfur synthetic solution and phase transfer catalysts have also been described (Deng et al., 1984; Zhang et al., 1993), but in this case thiosulfate is also present. In mixtures of thiosulfate and polysulfides, the polysulfides only act as a lixiviant when no other oxidant is present. In the presence of copper, polysulfides precipitate with copper to form CuS.

6.1.1 Nitrogen Species Catalyzed Pressure Leaching Process In the nitrogen species catalyzed pressure leach process developed at Montana Technical University, the sulfur residue is converted into a gold-extracting lixiviant (Anderson, 2003; Anderson and Twidwell, 2008). The principles of the reaction of a sulfide mineral with nitric acid in conjunction with sulfuric acid are shown below. It is postulated that the actual reaction species is NOþ, and not NO3e, which reacts with the mineral and oxidizes the sulfide to sulfur at an Eh value of w1.45 V (vs. SHE). 2MeSðsÞ þ 4NOþ ðaqÞ /2Me2þ ðaqÞ þ 2So þ 4NOðgÞ

(27.14)

By partially oxidizing the sulfide to elemental sulfur instead of sulfate, the gold can be accumulated in the elemental sulfur, recovered from the other leached solids and then leached via alkaline sulfide lixiviation, whereby the sulfur containing the gold is dissolved in an alkaline solution. The combination of sodium hydroxide and elemental sulfur results in the formation of sulfide (S2), sodium polysulfide (Na2Sx), and sodium thiosulfate (Na2S2O3) as follows: 4So þ 6NaOH/2Na2 S þ Na2 S2 O3 þ 3H2 O ðx  1ÞSo þ Na2 S/Na2 Sx

ðwhere x ¼ 2  5Þ

(27.15) (27.16)

The gold is leached by polysulfide and thiosulfate as follows: Auo þ S5 2/AuS þ 4S þ e

(27.17)

Auo þ 2S2 O3 2/AuðS2 O3 Þ2 þ e

(27.18)

3

The leaching kinetics and mechanism of alkaline sulfide system have been examined in causticesulfur solutions (4 M So, 9 M NaOH) and sulfideesulfur systems (2 M So, 2 M S2 or NaHS) (Jeffrey and Anderson, 2003). The dominant lixiviant for gold was ascertained in these systems as sulfide with polysulfide as oxidant. The leaching reaction was determined to be chemically controlled and highly dependent on temperature. Gold leaching was observed to be very slow where pure gold samples or when hydrosulfide was used in place of sulfide ions. For the sulfide sulfur system, the optimum ratio of sulfide to sulfur was found to be 1:1, but adequate leach rates were still obtained at a ratio of 1.5. The gold can be recovered by electrowinning, chemical precipitation, cementation, solvent extraction, or ion exchange. The barren alkaline sulfide solution can be recycled for further gold leaching or processed with low-temperature oxidation to produce sodium sulfate. As such, it is claimed (Anderson, 2003; Anderson and Twidwell, 2008) that there is no environmental or toxicological issue in the use of alkaline sulfide gold recovery.

6.2 Sulfide/Bisulfide/Sulfite Leaching e Applications Bisulfide leaching is clearly more applicable to sulfidic ores rather than oxidized ores. Touro and Wiewiorowski (1992) described a process of steam heating to 35e45  C finely-ground ore (50% solids) and then conditioning with 2e5 kg/t of H2S gas to make a goldesulfide complex. A chelating agent such as 0.5 kg/t EDTA may be added to complex the calcium.

464 PART | II Unit Operations

Sulfur dioxide (15e50 kg/t, depending on the carbonate content of the ore) was then injected into the pulp to reduce the pH value to the optimum range of pH 4e5. The pulp was then air agitated to maintain an oxidizing atmosphere for 16e20 h in the presence of an ion-exchange resin to simultaneously dissolve the gold and transfer it to the resin. The loaded resin was finally screened from the pulp and treated to recover the gold and regenerate the resin. Gold recovery obtained was 80% compared with 30% by CIL cyanidation in some ores. Initial bisulfide-leaching tests using the YES process attained 25e31% gold extraction from a bio-oxidized Nevada ore compared with 88% by conventional cyanide leaching (Hunter et al., 1998). At the same time, 39e81% silver was recovered compared to 86% with cyanide. However, tests conducted in a pressure vessel to allow higher bisulfide concentrations gave 75% gold extraction. Some applications of the nitrogen species catalyzed pressure leach process to treat auriferous copper sulfide concentrates have been reported (e.g., Anderson, 2001, 2003, 2008, 2014). Preoxidation of 100 g/L concentrate (ground to a P80 of 10 mm) was carried out with 175 g/L H2SO4 at 620 kPa and 125  C, in the presence of 2 g/L nitrogen species. This was followed by alkaline leaching, resulting in 98.3% gold extraction. Other applications to different refractory ore types have been reported by Anderson and Twidwell (2008 and references therein).

6.3 Sulfide/Bisulfide/Sulfite LeachingdCurrent Status Further fundamental investigations on gold dissolution processes involving sulfur chemistry are still required. A better understanding of the adsorption and precipitation reactions, which reduce gold extraction, also requires further investigation. The advantages the process offers over cyanidation include lower reagent costs and the potential ability to leach preg-robbing ores and other ores not amenable to cyanidation. It may also selectively leach precious metals from basemetal concentrates (Hunter et al., 1998). One of the limitations would seem to be the bio-reduction of sulfate ion using organic substrates for bisulfide regeneration. The bisulfide process is recyclable, but oxidation to sulfate is theoretically possible. A major drawback is that H2S, which is also generated, has an occupational health standard very similar to HCN. Extremely long retention times and a closed system would be required. The nitrogen species catalyzed pressure leach process is less hazardous and has been claimed to have been successfully demonstrated on a number of pilot and laboratory scale for various concentrates. The process is most probably better suited for the extraction of multielemental systems such as base and precious metals.

7. AMMONIA LEACHING Ammonia is most commonly used as an additional reagent in cyanidation for copper-containing ore bodies (see Chapter 43). However, the laboratory use of ammonia alone as a lixiviant for refractory gold ores has been reported at high temperatures (Han, 2001).

7.1 AmmoniadProcess Conditions Ammonia leaching of gold in the presence of an oxidant is carried out at temperatures between 100 and 300  C and 600e1000 kPa of pressure using 5e10 g/L Cu(II) as oxidant, 5.5 M free ammonia, and 0.5 M ammonium sulfate. Leaching times are short, w1e4 h. At ambient temperatures, gold is passivated and gold dissolution in ammonia solutions is observed only above 80  C (Meng and Han, 1993). The most effective oxidant for gold dissolution in ammoniacal solutions is Cu(II), as represented in the following reaction (Guan and Han, 1996): Au þ CuðNH3 Þ4

2þ

þ

¼ AuðNH3 Þ2 þ CuðNH3 Þ2

þ

(27.19)

The dissolution rate of gold increases with increasing copper and ammonia concentrations provided that there is sufficient ammonia to complex copper. Optimum pH conditions appear to be w9.5. Increasing the temperature also increases the gold dissolution rate. Alternative and less-effective oxidants include oxygen, hypochlorite, peroxide, and Co(III). The oxidation of sulfides in the ore produces some thiosulfate that complicates the leaching process. The use of combined halogen and ammonia as a gold-leaching process has also been examined. High gold recoveries are achieved at <100  C with iodine, providing the best oxidant in terms of the rate of dissolution of precious metals (Han, 2001). Iodine (I2) is a very effective oxidant (Peri et al., 2001), where gold can react with ammonia in the presence of iodine as an oxidant to form gold-ammine: þ

2Au þ I2 þ 4NH3 ¼ 2AuðNH3 Þ2 þ 2I

(27.20)

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One of the advantages of such a reaction is the ability to regenerate iodine by reaction with oxygen. However, no published applications on the use of iodine as an oxidant have been found in this review.

7.2 AmmoniadApplications Sulfidic and carbonaceous refractory ores have been successfully treated by the ammonia process (Han and Fuerstenau, 2000) with >95% gold extraction achieved in 2e4 h compared with <70% by conventional cyanidation.

7.3 AmmoniadCurrent Status Capital and operating costs are limiting factors in the implementation of this process, since high pressures and temperatures are required. Consequently, it would probably only be viable in isolated cases (e.g., spent catalysts or very high-grade concentrates).

8. BACTERIAL AND NATURAL ACID LEACHING Gold has been reported to be solubilized in biological or natural systems by a combination of microorganisms and amino acids such as glycine, in the presence of an oxidant. Microorganisms can promote gold solubilization by the excretion of ligands, such as thiosulfate, organic acids, cyanide, and iodideeiodine, which oxidize and form complexes with gold (Kaksonen et al., 2014). High gold recoveries have been reported using both amino and humic acids. However, the rate of gold dissolution is extremely slow and therefore could probably only be used in treating low-grade ores by a heap-leach process (Sparrow and Woodcock, 1995; Oraby and Eksteen, 2015).

8.1 Bacterial and Natural AciddConditions Organic acids (e.g., humic and fulvic acids, amino acids, and carboxylic acids) have been shown to promote the solubilization of native gold or promote the formation of gold colloids (Kaksonen et al., 2014). Amino acids produced by heterotrophic microorganisms, such as Bacillus (B.) subtilis, B. alvei, B. megaterium, B. mesentericus, Serratia marcescens, P. fluorescens, and P. liquefaciens, can enhance gold solubilization by forming goldeamino acid complexes. Typical conditions for the direct leaching of gold with bacteria and natural acids are listed in Table 27.7. Dissolution of gold is enhanced by the addition of an oxidizing agent such as sodium peroxide and by the selective breeding of more active strains of organisms. The reaction of amino acids with gold in the presence of permanganate produces a complex anion containing the salt of the relevant acid linked to Au(I) through goldenitrogen bonds. An example of a gold dissolution reaction using amino acid such as glycine as the ligand is as follows:   4Au þ 8NH2 CH2 COOH þ 4NaOH þ O2 ¼ 4Na AuðNH2 CH2 COOÞ2 þ 6H2 O (27.21) Bacteria, most notably B. subtilis, in the presence of amino acids such as glycine and an oxidant have been shown to dissolve gold from ore samples. Groudev and Groudeva (1990) found that B. subtilis was the most active solubilizing bacterium among 18 varieties tested. The aurous gold complex is an anion containing the salt of the relevant organic acid. Groudev and Groudeva (1990) found that under optimal leaching conditions (5 g/L amino acid, minimum 3e10 g/L potassium permanganate, 1.0 g/L sodium merthiolate, at pH 9.5, Eh > 0.5 V, 30  C), nearly 90% gold extraction was obtained in 3 days by agitation leaching and 70% extraction in 150 days by percolation leaching.

TABLE 27.7 Typical Direct Bacterial-Leaching Conditions Used for Gold Reagent

Ligand

Oxidant

Bacteria and natural acids

Amino acids

O2 or KMnO4 or H2O2

Natural acids

Humic acids

KMnO4

Gold Complex in Solution

Typical Leaching Conditions

pH

[Au(CH2NH2COO)2]

3e5 g/L amino acids 3e5 g/L KMnO4 and 1 g/L merthiolate, fermentation fluid obtained from cultivation of Bacillus Subtilis strain III-5

9.5e11

Possibly gold glycine

5 g/L humic acid 10e15 g/L alkali, 2e3 g/L KMnO4

Acidic

e

466 PART | II Unit Operations

Recent work has shown that precious metal dissolution rates in glycine with hydrogen peroxide as the oxidant, increase with increasing glycine concentration, silver content in gold, and increasing pH (up to pH 11) (Oraby and Eksteen, 2015). Leach rates are slower than cyanide leaching and the glycine-leaching solution is very sensitive to temperature and the reaction is a chemically controlled. The addition of Cu2þ to glycine peroxide system enhances gold dissolution. The goldeglycinate complex was also found to be effectively recovered on activated carbon with up to 13.2 g Au/kg carbon reported in 4 h. Humic acid and other naturally occurring organic acids have been studied by the U.S. Geological Survey, as mobilizing agents for gold in acidic swampy environments. Low gold solubility and slow kinetics militate against commercial use of this phenomenon. Mineev and Syrtlanova (1984) reported that a solution containing 5 g/L humic acid and 10e15 g/L alkali in the presence of 2e3 g/L KMnO4 leached 44% of the gold from a quartzecarbonate ore by percolation leaching in 45 days. Leaching the ore in pachuca tanks gave 69% extraction in 96 h. Reagent consumptions were 2 kg/t humic acid, 0.7 kg/t KMnO4, and 8 kg/t NaOH. Estimates from cost data showed that the system is economically viable for ores containing 1e2.5 g/t Au. Fan et al. (1992) also used humic acid and some unspecified additive when column leaching several ores, reportedly obtaining 40e80% gold extraction with humic acid over a 20-day period compared to 60e90% with cyanide.

8.2 Bacterial and Natural AciddCurrent Status The bacterial and natural acid leaching processes for gold are not well understood and consequently no practical process exists. This is an area where further studies should be carried out, particularly as the substances are generally nontoxic. Much work has been carried out on the use of microorganisms in the treatment of minerals and effluents, and this is an area where active research is increasing. The uptake of gold by plants and the mobilization of gold in soils have been the subject of several investigations (e.g., Greene et al., 1986; Wilson-Corral et al., 2012) and could ultimately have limited commercial application for a secondary recovery processes. Some piloting by phytomining of tails has been investigated but relies on the use of other lixivants mentioned in this chapter to transport gold to plant materials. Some processes such as amino acid leaching may have some application to heap-leaching approach; however, studies are still in the development stage.

9. THIOCYANATE LEACHING Thiocyanate has been known for a long time to act as a lixiviant for gold. Early work examined acidic thiocyanate solutions to recover gold and uranium simultaneously from South African gold ores (Fleming, 1986). Several studies on leaching rates, mechanisms, and thermodynamics of the thiocyanate system have been published by Monhemius and coworkers (Barbosa-Filho and Monhemius, 1994a,b,c; Monhemius and Ball, 1995). A thorough investigation looking at the whole flowsheet development on thiocyanate leaching and recovering gold has been carried out both at the University of Utah, USA, and by Newmont (Li et al., 2008, 2012a,b,c,d,e).

9.1 ThiocyanatedConditions Gold can be leached in 0.01e0.05 M thiocyanate at potentials of w0.4e0.45 V at pH 1e3 in the presence of either ferric ions (2e5 g/L) or peroxide as oxidant. The simplified reaction can be written as follows: 

Au þ 4SCN þ 3Fe3þ ¼ AuðSCNÞ4 þ 3Fe2þ

(27.22)

Thermodynamic analyses indicated that at the upper limit of the potential range, Au(SCN)4 is the gold species formed, but with decreasing potential, Au(SCN)2e is formed. It was proposed that intermediate species such as (SCN)2 and (SCN)3e act, like iodine, both as oxidants and complexants. Measurements of the gold electrode surface by Fourier transform infrared (FTIR) spectrophotometry and analysis of the leaching potential indicates that gold dissolution may undergo a two-step reaction at the gold surface. Initially gold forms an insoluble aurothiocyanate (AuSCN) salt as an intermediate. The insoluble intermediate (AuSCN(s)) then undergoes further coordination with other thiocyanate ions to form soluble aurothiocyanate an ion, Au(SCN)2e or Au(SCN)4e. In addition various iron thiocyanate complexes can also be present under certain conditions (Li et al., 2008). Ferric ion forms a series of complexes with thiocyanate: e

ð3nÞ

Fe3þ þ nSCN ¼ FeðSCNÞn

(27.23)

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467

The complexation between ferric and thiocyanate affects the solution potential. Original studies by Barbosa-Filho and Monhemius (1997) showed that thiocyanate can leach gold in the pH range of 1e3 at temperatures up to 85  C. However, increasing the pH value of the thiocyanate-leaching solution has a detrimental effect on the gold dissolution rate (Kholmogorov et al., 2002), presumably because of Fe(III) hydrolysis. Under the oxidation/leaching potentials required for gold dissolution, thiocyanate is thermodynamically unstable and can be oxidized by ferric ion to sulfate, carbonate, and ammonium (Li et al., 2012a,c). Thiocyanate oxidation by ferric ion in acidic solutions is slower at room temperature (25  C) but significant at elevated temperature of 50  C (Li et al., 2012c). The addition of small amounts of iodine or iodide has been shown to increase the rate of gold dissolution in the Fe(III) thiocyanate system (Barbosa-Filho and Monhemius, 1994d). This is most probably due to a synergistic effect as a result of the formation of iodineethiocyanate species (such as I2SCNe and I(SCN)e 2 ) (Monhemius and Ball, 1995). As discussed earlier, the addition of small concentrations of thiourea to thiocyanate-ferric solutions also enhances gold dissolution rates (Yang et al., 2011). Sulfide minerals significantly accelerate the oxidation rate of thiocyanate in the order of pyrite > chalcopyrite > galena (Li et al., 2012c). Some sulfide minerals such as chalcopyrite catalyze the redox reaction between thiocyanate and ferric ions. Besides being destroyed by oxidation during leaching, thiocyanate is known to form complexes with several metal cations. Cupric ions cannot form a stable complex with thiocyanate, and it rapidly oxidizes thiocyanate, with cupric reduced to cuprous ion and forming an insoluble cuprous thiocyanate salt [Cu(CNS)2]. While it may be possible to recirculate the thiocyanate in the iron complexes, the thiocyanate in the precipitated copper complexes would be lost from the circuit. Thus, it can be expected that reagent consumptions would be significant. Gold can be recovered from clarified solution by solvent extraction using a tertiary amine and stripping of gold from loaded organic phase with thiourea at low pH value (1e2) (Li et al., 2012d,e). However, extraction of ferric ions into the organic phase makes gold stripping difficult, and the reagent concentrations therefore have to be controlled to improve stripping kinetics. The recovery of gold by adsorption on carbon favors a pH value in the range of 2.5e3.0 with fast kinetics, relatively good gold loadings, and low thiocyanate consumptions (Li et al., 2008). Activated carbon has been shown to not only catalyze the redox reaction between the thiocyanate and ferric ion but also adsorb free thiocyanate, further contributing to thiocyanate consumption.

9.2 ThiocyanatedApplications The simultaneous leaching of gold and uranium using thiocyanate lixiviant was evaluated by Fleming (1986) on Witwatersrand ores from South Africa. The effects of Eh, pH, thiocyanate concentration, and temperature on the gold extraction rate were examined, and it was found that gold leaching was more sensitive to Eh and thiocyanate concentration temperature, whereas uranium leaching was more sensitive to pH value. Under conditions of 10 g/L NH3SCN, 2.5 g/L Fe3þ, pH 1.5, and ambient temperature yielded percentage gold extractions in high 80s that were higher than cyanide extractions. Gold was recovered on strong-base ion-exchange resins. Gold, silver, and base metals were stripped from the resin with an alkaline thiocyanateecyanide solution. However, thiocyanate ions strongly adsorb onto strong base resins and are not readily removed. Hence, regeneration of the resin was required, which involved a stripping solution of either 1 M ferric sulfate or 0.5 M ferric nitrate solution to displace w90% of the thiocyanate from the resin. Kholmogorov et al. (2002) investigated the use of thiocyanate solutions to extract gold from a chemically treated arsenopyrite concentrate. The concentrate, containing approximately 110 g/t gold, was leached at room temperature for 3e5 h with 39 g/L KSCN and 0.8 g/L Fe(III) at pH 2.2 and a pulp density of 5e10% solids. Gold extractions of 89e93% were achieved under these conditions. Comparative tests using 2.45 g/L NaCN at pH 11.3 resulted in 87% gold recovery in 89e96 h. Test work on the pregnant leach solutions demonstrated that gold could be recovered by adsorption onto activated carbon and ion-exchange resins. Complete desorption of gold from resins was achieved by using sulfuric acidethiourea solutions at room temperature and from activated carbon by using basic thiourea solution at 150  C. In another study, both bottle-rolleleaching tests and column-extraction tests were carried out to assess the effect of thiocyanate and ferric ion concentrations on gold recovery from a bio-oxidized low-grade (2.13 g Au/t) refractory sulfidic ore from Nevada (Wan et al., 2003). In bottle-roll tests conducted on ground bio-oxidized ore, the gold extraction after 24 h was 64% using 2.9 g/L SCNe and 5.6 g/L Fe(III) (pH 2, 20% solids, P80 ¼ 75 mm). However, comparable cyanidation test work gave 69% gold extraction. Increasing the initial ferric ion concentration had little effect on the gold dissolution, and thiocyanate degradation was observed to be rapid in the initial contact with ore. For the column tests, the bio-oxidized ore was crushed to 38.2 mm. After 16 days of thiocyanate treatment, gold extraction was 52% and 10% higher than that obtained with cyanide. However, thiocyanate consumption (0.6e0.8 kg/t of ore) was twice as high as cyanide consumption (0.33 kg/t of ore). Further work looking at all aspects of leaching and recovery has also been evaluated by Newmont (Li et al., 2008).

468 PART | II Unit Operations

9.3 ThiocyanatedCurrent Status The thiocyanate lixiviant may be suitable for most ore types and recycling of the leach solution is possible. Unfortunately, in practice, interference reactions of sulfide minerals and other metal ions may require the use of higher temperatures to achieve satisfactory leach performance and may also cause significant decomposition. The low pH value and higher temperatures may result in high capital and operating costs compared with cyanidation. However, thiocyanate leaching of precious metals in acidic solutions avoids problems related to neutralization and materials handling, especially for slurries after sulfide mineral preoxidation (Wan and LeVier, 2011; Li et al., 2008). The limited availability of thiocyanate is also a restriction, and if thiocyanate needed to be detoxified by oxidation to cyanate and sulfate (e.g., as may be required in, for example, Russia), it would further increase the operating costs.

10. RECOVERY PROCESSES Equally important to leaching processes is the selection of a method to recover gold. One challenge in replacing cyanide with an alternative lixiviant has been identifying an economically suitable option for recovering gold. Much research in this area has been carried out on thiosulfate; thiourea- and halide-leaching systems with gold recovery by cementation, precipitation, solvent extraction, carbon, ion-exchange resin, or electrowinning techniques have been investigated for most of the alternative lixiviants. A comprehensive review of options for gold recovery from noncyanide solutions was first presented by Wan et al. (1993). Possible recovery options from thiosulfate solutions have been reviewed by Grosse et al. (2003) and Aylmore and Muir (2001) and are discussed in more detail in Chapter 28. In general, precipitation methods have been considered for clarified leach solutions, particularly from heap leaching, while carbon and resins have been considered for adsorption from slurries. Solvent extraction can generally only be applied to clarified solutions containing relatively high concentrations of gold and silver.

10.1 Cementation and Precipitation Methods Table 27.8 shows chemical species that are capable of reducing the various gold lixiviant complexes to gold metal. Aluminum can be an effective reductant for chloride, thiocyanate and thiosulfate complexes; however, the purity of precious metals products obtained by this method is generally low. Zinc is unsuitable for acidic solutions because of its high solubility and generation of hydrogen gas, resulting in excessive reagent consumption. Applications of zinc and copper have been effective in the thiosulfate leach system particularly for heap leach applications, but again high concentrations of copper (10e20%) in the product requires further processing. Iron could be used for gold cementation in thiocyanate solutions where ferric iron is used as an oxidant in thiocyanate gold-leaching systems. The barren thiocyanate solutions left after gold cementation can be recycled back to the leaching step and reused (Wang et al., 2007a,b). However, in the cementation process with iron, it was observed that ferric ion hydrolysis occurred following an increase in pH value, from 2 to w4. Detailed kinetic studies by Guerra and Dreisinger (1999) on the copper cementation process showed that increased temperature (30e50  C) and a higher pH/ammonia concentration enhanced cementation performance, whereas the presence of sulfite and copper ions in solution negatively affected cementation performance. Sodium borohydride can also be used as an efficient agent for reducing gold and silver in clarified acidic solutions of thiourea, thiocyanate, or thiosulfate at room temperature. The Au(I) ion is reduced to metallic gold in the form of very fine crystals. A complete reduction of gold can occur by use of a sodium borohydride:gold molar ratio of 0.625 at a pH value of 6 over a 1-h time period (Awadalla and Ritcey, 1991; Groves and Blackman, 1995). However, the presence of ferrous ion, cobalt, nickel, or, in particular, copper in solution decreases the efficiency of borohydride to reduce gold because of extensive coprecipitation of other metals. Gold recovery has also been achieved by sparging or pressurizing thiosulfate or thiourea solutions containing gold with hydrogen; however, a catalyst of nickel or platinum is required and the pressurized equipment used is generally more expensive than other recovery approaches (Deschênes and Ritcey, 1990; Deschênes, 1987). Finally, dissolved gold and silver have been recovered by the addition of a sulfide, bisulfide or hydrogen sulfide solution with regeneration of thiosulfate (Kerley, 1981; Flett et al., 1983). West-Sell and Hackl (2005) successfully demonstrated the application of sulfide precipitation following clarification in the copper ammoniacal thiosulfate tank leach process, which enabled the effective recycling of reagents. The addition of sulfide, bisulfide or hydrogen sulfide is also used in precipitation of gold from clarified chloride leach solutions.

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469

TABLE 27.8 Standard Electrode Potentials for Gold Complexes and Possible Reductant Systems for Gold Precipitation Gold Complex/Gold Metal þ

o

Eo

Au /Au

1.69

AuCl2e/Auo

1.11 0.77

Au(SCN)2e/Auo

0.66

AuI2e/Auo

0.57 0.52

Reduction System

Fe3þ/Fe2þ (iron)

Cuþ/Cu (copper)

Au(NH2CSNH2)2þ/Auo

0.38

Au(S2O3)23/Auo

0.17

SO42/H2SO3 (sulfur dioxide)

0.15

Sn4þ/Sn2þ

0.14

So/H2S

0

Hþ/H2 (hydrogen)

0.39

H2CO3/(COOH)2

0.41

Fe2þ/Fe

0.48

So/S2 (sulfide)

Au(CN)2e/Auo

0.57 0.75

BO33/BH4 (borohydride)

0.76

Zn2þ/Zn (zinc)

1.66

Al3þ/Al (aluminum)

Adapted from Marsden and House (2006).

10.2 Adsorbent Materials Activated carbon and resins have been investigated to allow gold recovery directly from leach pulp for most of the alternative lixiviants. These are summarized in Table 27.9.

10.2.1 Activated Carbon Activated carbon is preferred for gold recovery from cyanide solutions because it can be added to the pulp and avoids soluble losses in tailings. Unfortunately, it has a low affinity for AuðS2 O3 Þ2 3 complex in thiosulfate leach solutions (Gallagher et al., 1989). While some workers have achieved reasonably high gold recoveries, the concentration of gold loading on carbon was too low (5 mg Au/g carbon) to be considered practical or economic (Jiexue and Qian, 1991; Abbruzzese et al., 1995; Yen et al., 1998; Kononova et al., 2001). Activated carbon can also catalyze the degradation of thiosulfate (Aylmore et al., 2014). To circumvent the problem of low loading/low affinity for AuðS2 O3 Þ2 3, gold can be adsorbed from a thiosulfate solution by carbon after adding a small (stoichiometric) amount of cyanide to the system (Lulham and Lindsay, 1991) or by ferricyanide-doped carbon (Parker et al., 2008). Gold recovery from acidic thiourea solutions by activated carbon adsorption is reported to be very high. Deschênes (1986) reported maximum recoveries of 90% after 1 h from a solution containing 27 mg/L gold and a carbon concentration of 20 g/L, with loadings as high as 15e17 wt% possible. However, high losses of thiourea (30%) have been reported in the leach solution due to adsorption by activated carbon (Kavanagh et al., 1994), but can be washed out and collected in the absence of oxygen with hot water (Schulze, 1984). The loading mechanism is similar to that observed in the cyanide system (Fleming, 1987). Formamidine disulfide, an oxidation product of thiourea, strongly adsorbs onto carbon and subsequently further oxidizes to elemental sulfur, which physically deposits in carbon macropores and impedes diffusion and adsorption of the goldethiourea complex. Furthermore, activated carbon catalyzes the oxidation/degradation of thiourea, resulting in the formation of elemental sulfur. Elution of the goldethiourea complex from the loaded carbon is

Adsorbent

Lixiviant

Competing Elements

Elution Process(es)

References

Carbon

Thiourea

Formamidine disulfide elemental sulfur catalyzes the oxidation/degradation of thiourea Thiourea

Sodium cyanide Sodium sulfide Organic solvents

Yen et al. (1994), Wan et al. (1995), Fleming (1987), and Descheˆnes (1986)

Halides Chloride Cle/Bre

Heavy metal chlorides Halides, arsenic

Burn carbon Cyanide/thiosulfate/ Thiourea

Moyes (1999), Severs (1999), Abe et al. (2009) and Takebayashi et al. (2011)

Thiocyanate

Thiocyanate

Cyanide/ethanol

Kholmogorov et al. (2002), Monhemius and Ball (1995), Wan (2003), and Li et al. (2012a)

Amino acids

Oraby and Eksteen (2015)

Thiosulfate

Base metals (Cu, Ni, Hg) Polythionates

Thiocyanate Trithionate Sodium nitrate Sulfite salt mixture

Fleming et al. (2003) Nicol and O’Malley, (2001) Jeffrey (2007)

Thiocyanate

Thiocyanate Ferric ion Base metals

Alkaline thiocyanate/ cyanide

Kholmogorov et al. (2002) Fleming (1986)

Iodine

Triiodide

Sulfite salt mixture

Zhang et al. (2012a,b)

Cation-exchange resins;

Thiourea

Base metals (Cu, Ni, Hg) Sulfur coatings

Ammonium thiosulfate

Simpson et al. (1984) Nakahiro et al. (1992)

Polyacrylic ester adsorbents

Chloride

High selectivity for gold Chloride

Alkaline solution Dilute HCl

Wan et al. (1993) Harris and White (2011a,b)

Strong-base resins

470 PART | II Unit Operations

TABLE 27.9 Adsorbent Materials Used for the Recovery of Gold from Various Lixiviant Systems

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471

found to occur readily. About 90% of the gold can be eluted from the carbon with inorganic stripping solutions (e.g., sodium cyanide or sodium sulfide) or solutions with organic solvents (e.g., butanol, acetonitrile, diethyl ether, ethanol, acetone in decreasing effectiveness, Yen et al., 1994; Wan et al., 1995). All elution methods generally recover precious metals, and a further scrubbing step is necessary if adsorbed impurities have to be removed to regenerate the capacity of activated carbon recycle on the efficiency of gold recovery. Carbon was used at the New England mine in NSW, Australia, to recovery gold and copper from thiourea solutions (Hisshion and Waller, 1984). The carbon product was sold to a smelter. Gold recovery from halide solutions using activated carbon has two major issues. First, reduction of gold by activated carbon makes conventional stripping difficult. Second, the finely precipitated metallic gold can be abraded from the activated carbon surface and lost in the tailings. Metallic gold deposits superficially on the carbon surface, where the uptake of gold by activated carbon follows a mechanism, whereby Au(III) is reduced to the metallic state (Siegel and Soto, 1984; McDougall and Fleming, 1987; Hiskey et al., 1990; Greaves et al., 1990; Wan et al., 1995). The goldechloride complex loads to very high levels (up to w100 kg/Au/t) onto activated carbon. Loading of w50 kg Au/t was also achieved on Australian char (with cost of char w1/10th of the cost of activated carbon). At these loadings, the carbon (or char) can be economically burnt to allow recovery of the gold bullion (La Brooy et al., 1994). Investigations into the uptake of gold from iodide solutions showed partial reduction by the carbon and a significant amount of metallic gold observed to be disseminated inside the carbon matrix (Hiskey and Qi, 1993; Teirlinck and Petersen, 1995, 1996). Work by Cashion and coworkers at Monash University have developed and demonstrated carbons with properties that load gold from halide systems as the halide complex and not as metallic gold (Cashion et al., 1997). Unfortunately, the carbon used was soft and therefore not abrasion resistant and, hence, impractical in carbon-in-pulp applications. In addition, no research in real pulp slurries or on the effect of impurities and recycling of carbon has been reported. In the Intec and N-Chlo Process pilot plant trials where the leach solutions contain bromine chloride, gold was successfully loaded onto carbon at up to 1% w/w, with no loss of carbon activity detected over five loading/washing/elution cycles. The adsorption behavior of gold in solutions containing the chlorideebromide mixed complex changes the loading characteristics compared with direct chloride solutions where metallic gold is precipitated on carbon surfaces. Some carbon adsorption and elution tests have been reported for thiocyanate (Kholmogorov et al., 2002; Monhemius and Ball, 1995; Li et al., 2008, 2012e). Adsorption of gold onto activated carbon has been observed in some cases to be very rapid, with little iron adsorption. In contrast, the rate of adsorption of goldethiocyanate complex onto activated carbon has elsewhere been reported to be very slow (Wan et al., 2003). Presumably other factors play a role, such as the presence of other anions present in solution, which may compete with gold thiocyanate adsorption. Significant reduction of ferric ion occurs in the presence of carbon and has been attributed to activated carbon catalyzing the degradation of thiocyanate. Attempts to strip the gold from carbon in 3e11% thiourea and 3e4% H2SO4 solutions were unsuccessful at room temperature, with only 15% of the gold desorbed, but increased to 94% at temperatures up to 150  C (Kholmogorov et al., 2002). However, conventional cyanideeethanol stripping of the loaded carbon was possible.

10.2.2 Ion-Exchange Resins Cation-exchange resins such as AG-50W-X8 or Amberite 200 have been tested by several investigators (Simpson et al., 1984; Nakahiro et al., 1992) for gold recovery from acidic thiourea solutions. However, the dissolution and loading of base-metal cations competes with gold for sites on the cation exchange resin and therefore substantially reduces the goldloading capacity. In addition, elemental sulfur formed from thiourea degradation products coats the resin and significantly reduces resin capacity. Further work is required to develop a selective cation-exchange resin for gold in acidic thiourea solutions before it can become a viable process alternative. Application of resins for recovering gold from thiosulfate leach pulps have been show to be a challenge, particularly in control of the loading, as well as the competitive nature of thiosulfate degradation products (Nicol and O’Malley, 2001; Fleming et al., 2003). The development of a suffice-based resin elution process has enabled the elution and subsequent electrowinning process to be more effective in reducing reagent consumption costs and improving resin recycling qualities (Jeffrey, 2007). Copper and other metal impurities are eluted first. The copper containing eluent is then returned to the leach circuit. For gold elution a mixture of a salt, such as sodium chloride, and sulfite is used to strip the gold. The addition of sulfite forms a mixed complex with gold and allows elution using lower ionic-strength salt solutions (Nimal et al., 2005; Perera et al., 2005). In the Barrick Goldstrike thiosulfate gold plant process, the combination of trithionate and sulfite is used effectively to elute gold from resin and allows for effective recycling of reagents. Both strong and weak-base resins can load gold from acidic chloride solutions; however, most commercial resins are not stable under the oxidizing conditions, with the extent of gold loading a function of pH value. Gold adsorbed from

472 PART | II Unit Operations

acidic solutions on to resin can be removed by stripping with an alkaline solution, such as thiosulfate or cyanide. Consecutive loading and stripping studies carried out over 16 days showed no degradation of the resin. Polyacrylic ester adsorbents have also provided a high selectivity for gold with respect to base-metal ions (Wan et al., 1993). Gold is reportedly successfully recovered on resin in the Intec process from a chlorideebromide solution; however, details of the process have not been published in the public domain. Recovery of gold from iodineeiodide solutions using a strong-base ion-exchange resin (Purolite A500/2788) has been investigated (Zhang et al., 2012a,b). The goldeiodide complex can be effectively loaded on the resin provided that the resin is not heavily loaded with triiodide. The loading of triiodide is also found to be extremely strong due to the dissociation of the loaded triiodide to iodide and iodine. The iodine is deposited on the resin by physisorption, fouling the resin surface, and can therefore potentially be detrimental to the gold-recovery process. Column loading was investigated with a solution containing 5.35 ppm Au, 5.51 mM I3 and 7.94 mM free I. The loading of gold achieved 1.5 g/dm3 (resin) at breakthrough (4500 g per dry metric ton of resin), despite the relatively high concentration of I3 in the feed solution. Sodium chlorideebased eluant solution containing sulfite was effective for eluting both gold and iodine from the resin. Early work by Fleming (1986) evaluated the use of a strong-base resin (A101Du supplied by Sentrachem Pty Ltd.) to recover gold and uranium from acidic thiocyanate solutions. Selectivity for silver and gold over all base metals was high except for copper. However, extraction of gold was poor at low resin flow rates. Selectivity of gold (and uranium) increases with decreasing thiocyanate concentration. Thiocyanate ions are strongly adsorbed on to strong-base resins and are not readily removed. Fleming (1986) patented a process for regenerating process involving a stripping solution of either 5 BV of 1 M ferric sulfate or 0.5 M ferric nitrate solution to displace w90% of the thiocyanate from the resin.

10.2.3 Plant and Other Adsorption Media Other adsorption media such as lignin-based adsorption gels derived from wood have been investigated and shown on a laboratory scale to be potentially effective adsorbents of precious metals in hydrochloric acid solutions (Parajuli et al., 2006). Furthermore, treating the cellulose with sulfuric acid followed by chemical modification with guanidine or N-aminoguanidine functional groups improves the selectivity and adsorption capacity of Au(III) but with relatively slow kinetics at ambient temperature (Pangeni et al., 2012a,b). The adsorption of Au(III), Pd(II), and Pt(IV) chloroanionic species in hydrochloric acid medium occurs due to ion pairing at the positively charged quaternary nitrogen atoms of the adsorbent. The Au(III) is then reduced by the adsorbent to the elemental form, yielding gold aggregates. The ability of plants to take up gold has been long recognized (Anderson et al., 1998a,b; Girling and Peterson, 1999; Gardea-Torresdey et al., 1999; Gamez et al., 1999, 2000). A review including economic considerations has been presented by Wilson-Corral et al. (2012). Most work has evaluated the potential of extracting residual gold from tailings or waste produced by mining. Gold can be taken up actively, by using the plant metabolism, or passively, by means of the carbonyl functional groups on plant tissues. Researchers have investigated the effect of the addition of various lixiviants (thiosulfate thiocyanate thiourea, iodine, as well as cyanide) to solubilize gold, with uptake by plants on tailings deposits. Eh and pH conditions of the substrate had to be controlled for both the plant to survive and goldelixiviant complex to be stable. The soil pH value may require modification by applying soil amendment reagents, such as lime- or ammonia-containing fertilizers. Various plants (e.g., field mustardsdBrassica juncea, Brassica campestris, sunflowerdHelianthus annuus) have been proposed, with accumulation of between 200 and 300 mg of gold per kilogram of plant dry weight reported in silica sands to which w1 g of cyanide, thiocyanate, or thiourea per kilogram of substrate has been added (Wilson-Corral et al., 2012).

10.3 Solvent Extraction Solvent extraction can only be economically applied to clarified solutions containing relatively high concentrations of gold and silver. Alkyl phosphorus esters and primary, secondary, and tertiary amines have been applied to thiosulfate liquors (Zhao et al., 1997, 1998a,b). In general, the separation of gold from silver, copper, zinc, and nickel in thiosulfate solutions by primary amines has proved to be difficult. However, the addition of ammonia to the thiosulfate solution increased the separation of gold from other metals in solution. Solvent extraction is widely used to recover gold in refinery processing and has been extensively investigated in chloride-leaching processes. Outotec have developed a solvent extraction organic reagent OT-307, which is claimed to be a selective and nontoxic extractant with a high flash point and low water solubility (Lundström et al., 2012). Gold-loading capacity of OT-307 up to 90 g/L has been reported, with stripping by an aqueous solution.

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473

Solvent extraction has been effective in recovering gold in thiocyanate systems (Li et al., 2012d,e). Gold can be recovered from clarified solution by solvent extraction using a reagent mixture of tertiary amine Alamine 336 (0.05 M) and 2e5% decanol (O/A 1:5, pH  3) and stripping of gold from loaded organic phase with thiourea at low pH (1e2). However, extraction of ferric ion into the organic phase makes gold stripping difficult, and the reagent concentrations have to be controlled to improve stripping kinetics.

11. ECONOMIC EVALUATION To justify a mining project using an alternative lixiviant to cyanide, a complete financial analysis covering capital investment, operating expenses, revenue, water treatment, and monitoring is required. However, it is difficult to obtain a complete picture of how different reagents respond to different ores due to the lack of detailed information in the published literature and limited reported pilot-plant data. The recent development of a thiosulfate leaching plant at Barrick Gold’s Goldstrike operation has alluded to the complex and unstable nature of leaching gold using this system and the significant process design difference of using thiosulfate over conventional cyanide (Braul, 2013; Choi et al., 2013; La Brooy and Smith, 2013). Table 27.10 shows comparative investigations carried out by several authors illustrating the variation in conditions, reagent consumptions, and gold extractions observed in treating specific ores with different reagent systems. The higher reagent concentrations associated with the use of alternative lixiviants result in higher reagent consumptions compared with cyanidation. In cases where cyanide consumption is high, such as the reaction of cyanide with copper minerals, alternative lixiviants, such as thiosulfate leaching, can be more favorable than cyanide for leaching gold (Monhemius and Ball, 1995; Dai et al., 2013). Although some alternative reagents can give higher gold extractions with carbonaceous or refractory gold ores (Wan et al., 1995; Choi et al., 2013), cyanidation often achieves the highest overall gold recovery. For high-grade ores and concentrates, the economic benefit of increased gold recovery usually outweighs the reagent cost and increased recovery is the main driver for choosing a particular system. Table 27.11 shows the author’s estimate of reagent cost for leaching a low-grade oxide ore (3.2 g/t Au, 0.3 g/t Ag, 0.5% pyrite in quartz/albite/clinochlore/biotite) in the Kalgoorlie region of Australia using cyanide and three other lixiviants based on optimum reagent compositions and estimated ranges of reagent consumption during a study in 2001. Without taking into account any differences in gold extraction, the reagent costs for thiosulfate are about double those for cyanide. While actual costs may vary and some costs have fluctuated over the years, reagent costs for thiourea and chloride are estimated to be an order of magnitude higher than for cyanide or thiosulfate. These estimates indicate that in treating a lowgrade ore, alternative reagent costs can be more expensive than the value of the gold, unless reagents can be suitably recycled in the process. In many cases, reduced reagent consumption is achieved when a pretreatment process has been applied. Pressure leaching followed by thiosulfate leaching on Goldstrike carbonaceous ores appears economical (Choi et al., 2013), although the variability of the ore can cause significant differences in gold recovery (Fleming et al., 2003). Chlorideleaching processes invariably require preoxidation of sulfide minerals prior to leaching to reduce reagent consumption (Lemieux et al., 2014; Harris and White, 2011a,b; Liddell and Adams, 2012a). High reagent consumptions can be tolerated where ores are refractory to cyanide, such as preg-robbing carbonaceous ores. Gold recoveries on the order of 60e80% using thiosulfate can be achieved, whereas cyanide extractions range from 0% to 15% (Wan et al., 1994; Aylmore, unpublished; West Sells and Hackl, 2005; Choi et al., 2013). An economic evaluation of all processes considered by McNulty (2001), which took into account the transport of reagents required for the leaching operations, revealed cyanide as the only option where a profit could be made by heap leaching at the McDonald gold mine. In addition, from the experience of Newmont Mining, additional costs would be associated with continuously shifting the heap to maintain chemical control for alternative lixiviants, which would increase both capital and operating costs. By comparison, consecutive heaps can be placed on top of one another in cyanide heapleaching operations. Capital costs associated with construction materials can be significantly higher than those used in a conventional cyanide plant. The highly corrosive nature of solutions used in chloride-leaching systems requires titanium impellers and corrosionresistant pumps. In thiosulfate and thiourea leaching, some metal surfaces may need to be coated to prevent cementation of gold from gold-bearing solutions. For the Barrick Gold Goldstrike thiosulfate plant, the leach tanks have been constructed out of 2205 stainless steel because of the corrosive nature of thiosulfate in the leaching system. However, for small plant designs, tanks can be made from fiber-reinforced plastic and polypropylene pipes and fittings can be used, which would reduce the capital costs significantly. Hence, design of a process plant needs to be considered on a case-by-case basis.

474 PART | II Unit Operations

TABLE 27.10 Comparison of Leach Investigations on Selected Ores Reagent

pH

Concentrations (g/L)

Reagent

Au, Ag

Consumption (kg/t)

Extraction (%)

NaCN ¼ 0.15

73, 23

Oxide ore (0.9 g/t Au, 5 g/t Ag) bottle-roll tests (McNulty, 2001) Cyanidation

NaCN ¼ 0.5

10.5e11.0

CaO ¼ 0.55 Thiourea

CS(NH2)2 ¼ 2

1.1e1.3

CS(NH2)2 ¼ 3.05

57, 22

Fe2(SO4)3 ¼ 9.0 H2SO4 ¼ 48 Thiosulfate

(NH4)2S2O3 ¼ 15

9.5

NH4OH ¼ 3.5

(NH4)2S2O3 ¼ 14

37,16

NH3 ¼ 2

CuSO4$5H2O ¼ 0.06 Chloride

NaOCl ¼ 1

6.4e6.5

NaCl ¼ 100 Bromide

Br2 ¼ 1

NaOCl ¼ 5.55

68, 22

HCl ¼ 3.25 1.3e2.0

Br2 ¼ 2.85

57, 13

H2SO4 ¼ 6.8

NaBr ¼ 10 High-grade ore (68.2 g/t Au, 2 g/t Ag) (Monhemius and Ball, 1995) Cyanidation

NaCN ¼ 5

10.5

NaCN ¼ 7.80

Thiourea

CS(NH2)2 ¼ 3.8

1.5

e

<60

2

NaSCN ¼ 1.3

94

97.5

86

CaO ¼ 2.0 Fe2(SO4)3 ¼ 10 Thiocyanate

NaSCN ¼ 8.1 Fe2(SO4)3 ¼ 11 I2 ¼ 0.5

Low-grade ore (4.8 g/t Au, 2 g/t Ag) (Monhemius and Ball, 1995) Cyanidation

NaCN ¼ 5

10.5

NaCN ¼ 7.42

Thiourea

CS(NH2)2 ¼ 40

2.5

CS(NH2)2 ¼ 12.8

83

2

NaSCN ¼ 0.54

95

CaO ¼ 5.0 Fe2(SO4)3 ¼ 10 Thiocyanate

NaSCN ¼ 8.1 Fe2(SO4)3 ¼ 11 I2 ¼ 0.5

Sulfide ore (7.8 g/t Au, 13.4 g/t Ag, 11.2%S) (Munoz and Miller, 2000) Cyanidation

NaCN ¼ 0.98

11

NaCN ¼ 0.36

93.7

Thiourea

CS(NH2)2 ¼ 11.4

2

CS(NH2)2 ¼ 38.3

27.6

2

NaSCN ¼ 1.1

49.5

Fe2(SO4)3 ¼ 4 Thiocyanate

NaSCN ¼ 8 Fe2(SO4)3 ¼ 16

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TABLE 27.11 Comparative Reagent Costs for Treating a Kalgoorlie Oxide Ore (3.2 g/t Au, 0.3 g/t Ag, 0.5% Pyrite in Quartz/Albite/Clinochlore/Biotite) Estimated Costs Range Based on Reagent Consumption (A$/t ore)b

Lixiviant

Reagent Concentrationsa

Cyanide

300e350 mg/L NaCN, (0.54 kg/t) Lime 9.1 kg/t O2 15 mg/L pH 10.0

1.23

Thiosulfate

6.6 g/L S2O32e 64 mg/L Cu as CuSO4 Lime 6.8 g/L total NH3/NH4þ pH 9.5

2.5e3.6

Thiourea

3.7e8 g/L CS(NH2)2 12e52 kg/t H2SO4 Fe from ore 0e5 kg/t NaHSO3 pH 3.0

14.4e39.2

Chlorine

30 g/L NaCl 3.1 g/L HCl 3.15 g/L NaOCl pH 5e6

29.2e66.7

a

Leaching (40% solids), temperature ambient. Costs based on transporting chemicals to Kalgoorlie in 2001; assuming mining/grinding costs fixed. After Aylmore (2005).

b

Further, extra capital costs associated with nonstandard monitoring equipment and water treatment plants may be required, as was found necessary in the case of the Barrick process plant. A process modeling study has been carried out to generate an estimate of the capital and operating costs associated with a bromine-leaching process using Albemarle Corporation’s stabilized bromine technology and comparing the costs to that for a cyanide circuit processing the same ore (43.6 Au, 10.5 Ag 0.23% Cu, 1.35% As, 0.4% S) (Dry et al., 2015). The bromine process consisted of a leaching and resin-in-pulp circuit to leach and recover gold and silver from the ore and a countercurrent decantation circuit to recycle bromine reagents back to the leach via a regeneration circuit. While capital costs were estimated to be higher for the bromine circuit, the operating costs were lower than those for cyanide leaching, which incorporated a cyanide destruction circuit to eliminate cyanide from the tailings. However, while the preliminary model results looked promising for that particular ore, further test work is required to validate operating assumptions. Bleed streams can create significant issues at plant scale and developers tend to overlook the impact of bleed streams in the early stages of process development (La Brooy and Smith, 2013). In addition, the impact of impurities such as mercury containment can provide extra costs to a process that are not always taken into account. A full process flow sheet needs to be tested on a pilot-plant scale operation for any new process design and process models to be validated, taking into account both metallurgical and engineering aspects. This approach can highlight issues (e.g., scaling, impurities buildup) that may be unforeseen based on bench-scale test work results (La Brooy and Smith, 2013). Overall, the effectiveness of an alternative lixiviant to be economic over cyanide leaching needs to be considered on a case-by-case basis.

12. ENVIRONMENTAL CONCERNS Environmental impacts of alternative lixiviants to cyanide have also been considered in a number of articles (e.g., Avraamides, 1982; Swaminathan et al., 1993; DeVries and Hiskey, 1992). However, generally the process development work involving alternative reagents to cyanide has mostly lacked any environmental focus in terms of consideration of regulatory factors, employee health and safety, environmental protection, proper disposal of wastes, and sustainable development, all issues that mining companies would have to address in any practical implementation. DeVries and Hiskey (1992) first reviewed environmental implications of some of the alternative reagents to cyanide. The worker and environmental risks of the reagents used as alternatives to cyanide have been evaluated in detail by Gos and Rubo (2000). De Voto and McNulty (2001) have also emphasized the less than favourable environmental aspects of

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several of the alternative lixiviants. Many issues relating to environmental concerns with respect to the use of any of the lixiviants, including cyanide, will be based on local climatic conditions. Acid production in tailings dams (especially after decommissioning) due to oxidation of residual sulfide minerals is possibly the real threat to the environment, irrespective of leaching process used. While numerous studies have investigated the effect of cyanide in the environment, studies on the effect of alternative lixiviants on the environment are limited. This is primarily as a result of the fact that there have been no commercial alternative processes until relatively recently. However, sufficient investigations on other mineral processes are available, which provide enough information to indicate that some of these alternatives, such as those using ammonia, may be more of an occupational health and safety as well as an environmental concern than cyanide. This has been illustrated by Gos and Rubo (2000). Only recently has some attempt been made to consider the effect of some lixiviants in tailings systems. Ebbs et al. (2011) reported that in separate tests 75% of the initial thiocyanate and >50% of thiosulfate and cyanide concentrations added (1 g/kg ore) in contact with an oxide tailings residue, over a period of 24 h, was lost through bacterial oxidation or physical and chemical reactions within the ore. A major aspect of minimizing the worker and environmental risk would be the selection of a process that minimizes the quantity of chemicals used. Since several of the alternative lixiviants use w50 times the concentration used in cyanide leaching, it is important to be able to recycle as much reagent as possible, not only to reduce costs to the leaching process, but also to prevent buildup of dissolved and precipitated solids in tailings dams. Methods for reagent recycling or destruction of cyanide have been demonstrated to be successful on economic and environmental grounds, and many have been used in the mining industry, as detailed in Chapters 35 and 36 of this volume. However, while some alternative reagents to cyanide can be recycled, such as through adsorption on carbon or resins or other means, further investigations are required. Membrane-treatment process plants can be used to provide clean water for the comminution circuit and also concentrate lixiviants for recycling to the leach circuit, as used in the Barrick Gold strike thiosulfate-leaching plant (La Brooy and Smith, 2013). Mechanical vapor-compression falling film evaporator and crystallizer systems can be used to treat bleed streams, especially if environmental pressures or climate constraints preclude evaporation ponds.

13. CONCLUSIONS Of all the processes available, thiosulfate and chloride leaching appear to be the most favourable options to replace cyanide. Extensive studies have established the leaching mechanisms and the many issues that may reduce gold recovery in these noncyanide lixiviant solutions compared with cyanide. Extensive investigations of thiosulfate or thiourea as alternative lixiviants to cyanide at semicommercial scale have been carried out by both Newmont and Barrick Gold not because of any particular concerns with the environment or health and safety concerns while using the cyanide system, but because cyanide is unable to effectively extract gold from some of their carbonaceous ore resources. Oxidative chloride-leaching and chlorination processes have been used as pretreatment processes to oxidize refractory or carbonaceous ores prior to conventional cyanidation and carbon-in-pulp technology, as well as a gold-leaching process, mainly in refineries. With increasing requirements to treat more complex and refractory ores, some of the chloride-leaching processes will have an application where other metals are extracted along with gold. Despite pilot-plant trials on many of the alternative options to cyanide, only the thiosulfate-leaching process has been commercialized (Choi et al., 2013). Many of the pilot-scale studies have highlighted problems associated with scaling up a new technology from laboratory experiments to pilot-plant or commercial stage: l

l

l

l

l

l

l

Halides, thiourea, and thiosulfate leaching are very susceptible to variations in mineralogy of the feed being leached and may require constant adjustments in regulating reagent concentrations to maintain optimal leaching conditions. Except for halogens, most alternative reagents do not exhibit fast leach kinetics compared with cyanide at similar concentrations. Shortfalls in reagent recovery and gold recovery may be evident when running a continuous leaching operation due to adsorption or reaction with other ore minerals. Conditions in heap-leach operations on a large scale are more difficult to maintain compared with cyanide and are not easily scaled up from column leach experiments in the laboratory. Higher reagent concentrations compared to those used in cyanide leaching make recycling important to make gold extraction by alternative lixiviants economical. The complexity of the process may be substantially greater than in cyanide leaching, potentially requiring nonstandard monitoring equipment and specialized training of staff. The management of bleed streams and stabilizing tailings residues for long-term storage is important

Alternative Lixiviants to Cyanide for Leaching Gold Ores Chapter | 27

477

In developing an alternative gold-leaching process, however, continued work on lowering reagent consumptions and improving gold recovery from solution are required. Feasibility studies into the worker and environmental risks associated with the process will undoubtedly have to be taken into account before commercialization, particularly in environmentally sensitive areas. Further work by the industry is required to continue to develop robust overall process flow sheets, incorporating gold recovery, reagent recycling, and impurity control for a wider range of ore applications.

ACKNOWLEDGMENTS The author acknowledges the input in the past of former colleagues at CSIRO Minerals. Part of the early review was compiled while the author was on secondment to WMC Resources Ltd.

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Mark G. Aylmore currently holds the position of Digital Mineralogy Hub Facility Leader at John de Laeter Centre, Curtin University. He earned his Ph.D. at the School of Applied Chemistry at Curtin University, which involved work evaluating the distribution of gold and processes concerned with the agglomeration of gold in sulfide minerals. Mark has more than 25 years’ experience in fundamental research, practical industrial developments, and the educational promotion of a wide range of mineral chemistry, mineralogy, and extractive metallurgical techniques. He has held senior roles in developing metallurgical programs, implementing and managing programs involving gold and, to a lesser extent, base metals and uranium. As principal process engineer with Tenova Mining & Minerals (Bateman Engineering), his work involved reviewing and developing test work programmes and carrying out engineering studies largely for gold refractory ores and various projects on base metal, PGMs, manganese, rare earths, and uranium extraction. He was lead process engineer in the design of the Nippon chloride demonstration plant process for treating copper, gold, and silver sulfide concentrate. As principal scientist with BHP Billiton, he managed uranium leaching enhancement projects concerned with developing and improving the processing of Olympic Dam coppereuraniumegold ore deposits. As a senior metallurgist for Placer Dome Inc and Barrick Gold Corporation, he was involved in implementing and managing projects concerned with developing and refining techniques for gold extraction. These included piloting the Blinder cyanide resin in leach and thiosulfate leaching processes, evaluating and developing process options for treating complex carbonaceous ore, refinement of an acid solvent extraction process, and preliminary examination of seafloor ore processing options. Prior to that he was a senior research scientist with CSIRO Minerals for 7 years, working on research and managing projects concerned with developing techniques for the extraction of gold by noncyanide methods from ore. Before joining CSIRO he worked as a research leader at the SRCAMMP at the University of Western Australia, investigating mechanochemical milling processes as a mineral processing option in extraction of precious and base metal sulfides.