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Cyanidation of gold ores containing copper, silver, lead, arsenic and antimony
International Journal of Mineral Processing 95 (2010) 68–77
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International Journal of Mineral Processing j o u r n a l h o m e p a g e : w w w. e l s ev i e r. c o m / l o c a t e / i j m i n p r o
Cyanidation of gold ores containing copper, silver, lead, arsenic and antimony P. Karimi a,⁎, H. Abdollahi a, A. Amini b, M. Noaparast a, S.Z. Shafaei a, F. Habashi c a b c
School of Mining Engineering, University of Tehran, Islamic Republic of Iran Geological Survey and Exploration of Iran, Islamic Republic of Iran Laval University, Quebec City, Canada G1V 0A6
a r t i c l e
i n f o
Article history: Received 11 December 2009 Received in revised form 27 February 2010 Accepted 13 March 2010 Available online 20 March 2010 Keywords: Gold ores Cyanidation Reaction kinetics Pretreatment
a b s t r a c t Cyanidation tests on two ore samples from two different gold deposits are reported. The first sample contained 10.5 ppm gold with high arsenic and antimony. The second sample had a low gold content (2.5 ppm) but a high silver content (160 ppm). The first series of test work focused on the determination of conditions for extracting gold from the samples ground to − 75 μm. The optimum parameters were 4000 mg/L for cyanide concentration, pH = 11.1 and 24-h cyanidation time for the first sample; and 2500 mg/L, pH = 10.5 and 24 h for the second sample. Silver and gold recoveries were 94.91% and 28.2% respectively for the first sample and 92.5% and 93.5% for the second sample. In the second series of test work, H2O2 (0.015 M), air (0.15 L/min) and a mixture of H2O2 and air were used as oxidizing agents to improve the gold extraction kinetics of the first sample. It was found that gold leaching recovery followed first order kinetics and that the injection of air had the maximum beneficial effect on the leaching kinetics. In the third series of test work, acid pretreatment (HNO3 and HCl) and roasting (600– 1000 °C) from 0.5 to 2 h were carried out before cyanidation. Acid pretreatment reduced cyanide consumption by 340 and 210 mg/L respectively and the corresponding gold recoveries increased to 98.87% and 95.11%. Cyanidation results on roasted samples showed that cyanide consumption was drastically reduced by 1150 mg/L and gold recovery increased by 5.2%. Furthermore, arsenic, antimony, cadmium and mercury contents were considerably reduced in the roasted sample (2 h at 1000 °C). © 2010 Elsevier B.V. All rights reserved.
1. Introduction
reactions and the preferential formation of Au (CN)− 2 on the surface instead of AuSx (Senanayake, 2008).
1.1. General Early studies on the dissolution of gold in cyanide solution in the presence of sulfide minerals have shown that heavy metal components, such as Cu, Fe and Zn, significantly increase the consumption of both cyanide and oxygen (Habashi, 1967; Dai and Jeffrey, 2006). In addition, the sulfide component has been shown to have a strong impact on the gold leaching kinetics (Fink and Putnam, 1950; Hedley and Tabachnick, 1968; Dai and Jeffrey, 2006). The leaching behavior of gold in the presence of sulfide minerals depended strongly on both the solubility of the sulfides and the oxygen concentration in the solution (Dai and Jeffrey, 2006). Based on the experiments observation, it is postulated that sulfide ions (formed by the decomposition of sulfide minerals) show detrimental effect on the cyanidation kinetics of gold and silver. However, the detrimental effect of dissolved sulfide on gold cyanidation according to reactions R1 and R2 which diminishes at higher cyanide concentrations, and favor the reverse
⁎ Corresponding author. Tel./fax: +98 21 64592257. E-mail address: [email protected] (P. Karimi). 0301-7516/$ – see front matter © 2010 Elsevier B.V. All rights reserved. doi:10.1016/j.minpro.2010.03.002
−2 R1 M2S + 4CN− ⇆ 2M(CN)− 2 +S − − R2 M2S + 4CN− + H2O ⇆ 2M(CN)− 2 + OH + HS
The presence of galena, as well as both sulfide and added lead(II), also improved the rate of gold dissolution. The dissolution rate was found to be affected by ions such as lead(II), sulfide, and iron(III) released by the sulfide mineral. The significant decrease in the rate of electrochemical and chemical dissolution of gold caused by chalcopyrite, pyrite, and pyrrhotite has been related to the various films formed on the gold surface as well as the galvanic interaction (Lorenzen and van Deventer, 1992; Aghamirian and Yen, 2005; Dai and Jeffrey, 2006; Senanayake, 2008). The dissolution of sulfides results in high cyanide consumption with the formation of cyano complexes (e.g., of the ions Fe+ 2, Zn+ 2, Cu+ 2, Ni+ 2, Mn+ 2) and SCN−. When sulfide minerals are present in gold ores, gold dissolution can be affected in various ways. In one hypothesis (Weichselbaum et al., 1989; Kondos et al., 1995), soluble sulfide (S2− or HS−) generated from mineral dissolution reacts with gold and forms a passive film, which decreases the rate and extent of leaching. In another theory (Lorenzen and van Deventer, 1992; Kondos et al., 1995), a dynamic coupling of reduction forms at the sulfide mineral surface which
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results in oxidation on the gold grains (Kondos et al., 1995). In the leaching of gold using cyanide, copper presents difficulties due to the lack of selectivity of cyanide for gold over copper (Rees and van Deventer, 1999). Cyanide also forms complexes with a number of 2− 2− and Ag(CN)− metal irons: Fe(CN)4− 6 , Ni(CN)4 , Zn(CN)4 2 which decrease the free cyanide concentration in slurry and retard gold leaching(Rees and van Deventer, 1999). The refractoriness of gold ores can result primarily from the inherent mineralogical features with particular reference to the mode of presence and association of gold, and to the presence of carbonaceous matter (Rubisov et al., 1996; Celep et al., 2009). A suitable pretreatment process is often required to overcome the refractoriness and render the gold accessible to the lixiviant action of cyanide and oxygen (Ubaldini et al., 1994; Celep et al., 2009). Roasting, pressure oxidation, biooxidation, ultrafine grinding and modified cyanidation are the pretreatment methods currently practiced for refractory gold ores and concentrates (Sinadinovic et al., 1999; Iglesias and Carranza, 1994; Celep et al., 2009). Roasting, a simple stage process using a fluidized bed roaster at around 650 °C, is one of the most common methods for the treatment of gold-bearing pyrite/arsenopyrite and pyrrhotite concentrates to produce porous calcine with the increased amenability to cyanidation (Roshan, 1990; Dunn and Chamberlain, 1997; Celep et al., 2009). Diagnostic leaching has been also proved to be very useful as an analytical tool for establishing the distribution of gold within different mineral phases of an ore (Celep et al., 2009). It was reported that pre-leaching successfully overcame the effect of pyrite and pyrrhotite on cyanide consumption and the gold leaching kinetics, but it had no beneficial effect when chalcopyrite was present (Dai and Jeffrey, 2006).
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effect by presence of preg-robber, it considerably diminishes in solutions of high concentration of cyanide (Tan et al., 2005). In such refractory gold ores, gold particles may sometimes be occluded or included in the sulfide minerals and pretreatment is necessary to decompose the mineral structure to liberate gold for subsequent recovery. These ores are usually pretreated by some oxidizing process after which gold and silver can be recovered by standard cyanidation process. The main available oxidizing processes for refractory gold ores are roasting (LaBrooy et al., 1994; Angelidis and Kydros, 1995; Qingcui et al., 2009), biological oxidation (Lawrence and Bruynesteyn, 1983; Deng and Liao, 2002) and pressure oxidation (Mason, 1990; Murthy et al., 1992; Qingcui et al., 2009). It has also been demonstrated that the addition of other oxidizing reagents increases the kinetics of gold dissolution (Stoychevski and Williams, 1993; Deschenes and Prud home, 1997). In this study, main cyanidation parameters for two different gold samples (grade and mineralogy) were compared. The studied samples were from two different gold deposits in Iran. The first sample (high gold grade) was taken from “Hirad gold deposit”, situated in Southern Khorasan province, and the second one (low gold grade) was provided from “Latala gold deposit”, located in Kerman province (Karimpour, 2006). After optimization of parameters, and due to the presence of arsenic, antimony, carbonaceous matters and clay minerals, pretreatment processes such as acidic leaching and roasting for the first sample were investigated. Due to the complexity of mineralogy of first sample, gold dissolution kinetic studies were also evaluated by using different oxidizing reagents which was not applied for the second one. 2. Ore preparation and characterization study
1.2. Factors affecting gold leaching 2.1. Sample preparation Cyanidation has been used for over 100 years to extract precious metals from sulfide ores. Despite this fact, the reactions involved are not fully understood. The chemical reactions that take place during the cyanidation of concentrate for ores can be very complex (Luna and Lapidus, 2000). Since a large proportion of gold ores contain sulfide minerals, the effects of these minerals on gold dissolution in cyanide solution have interested many researchers. The initial rate of gold dissolution is largely controlled by the factors such as cyanide and oxygen concentrations, pH, solid–liquid interface, Eh of the slurry, alkalies, particle size and temperature. The presence of other catalytic ions in solution and salinity of water also affect the rate of leaching (Habashi, 1970; Nicol et al., 1987; Ellis and Senanayake, 2004). From the electrochemical studies, the rate of gold leaching in air-saturated solutions increases with increasing concentration of cyanide but becomes independent of cyanide concentration when it exceeds 0.075% KCN. Clearly the high concentrations of oxygen and cyanide improve initial gold leaching kinetics but high cyanide concentrations also increase cyanide consumption (Ellis and Senanayake, 2004). The use of oxygen or an oxidizing reagent is essential for the dissolution of gold under normal conditions. Such oxidizing reagents as sodium peroxide, potassium permanganate, bromine, and chlorine have been used with more or less success; but they are not used now, due to their cost and complications involved in their handling (Yannopoulos, 1991; Guzman et al., 1999; Marsden and Lain House, 2006). Generally, gold ores can be classified as “free milling” and “refractory” depending on their response to cyanide leaching (Brooy et al., 1994; Celep et al., 2009). While high gold recoveries (N90%) from free milling ores can be readily achieved, refractory gold ores are often characterized by the low gold extractions (50–80%) within a conventional cyanide leaching process (Rubisov et al., 1996; Adams, 2005; Celep et al., 2009). The presence of preg-robber components such as carbonaceous matter, sulfide minerals and aluminosilicate have more influence in decreasing of gold and silver dissolution rate and recovery. In despite of negative
Two representative samples were supplied by using a systematic sampling procedure and about 400 kg from each sample were used for experiments. After four crushing stages (two jaw crushers, cone crusher and roll crusher), primary sample size reached to −2830 μm. In the next step, 2 kg samples were prepared by riffling and cone and quartering method. 2.2. XRD, XRF and chemical analysis For ore characterization, optical mineralogy, using the prepared polished and thin sections, X-ray diffraction (XRD) studies were Table 1 Mineralogical compositions of two gold samples. Gold ore type
Mineral name
Common mineral formula
Composition percentage
Sample 1
Quartz Illite
SiO2 (K,H3O)(Al,Mg,Fe)2(Si,Al)4O10[(OH)2, (H2O)] CaSO4.2H2O Fe2O3 FeAsS KAlSi3O8 CaCO3 PbS ZnS CuFeS2 FeS2 Cu2(CO3)(OH)2 Cu3(CO3)2(OH)2 SrSO4 SiO2 FeO(OH) CaCO3 KAlSi3O8
61.6 11.1
Sample2
Gypsum Hematite Arsenopyrite Feldspar Calcite Galena Sphalerite Chalcopyrite Pyrite Malachite Azurite Celestite Quartz Goethite Calcite Feldspar
9.5 5.2 3.6 3 3 1 0.6 0.5 0.2 0.2 0.2 0.2 73 21.6 2.7 2.4
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2.3. Microscopic mineralogy
Table 2 Qualitative analysis by XRF method. Components (wt.%)
First sample
Second sample
Components (wt.%)
First sample
Second sample
SiO2 Al2O3 Fe2O3 P2O5 CaO SO3 K2O
59.05 17.20 5.47 0.10 4.12 0.60 3.91
68.91 8.34 14.06 0.10 0.44 0.86 1.90
MgO TiO2 MnO ZnO PbO Na2O L.O.I
0.73 0.47 0.12 0.43 0.9 b0.1 6.76
0.36 0.25 0.56 b 0.05 b 0.05 – 4.20
Table 3 Elemental analysis of the primary ores samples by ICP-OES technique. Elements (g/t)
First sample
Second sample
Au Ag As Pb Zn Cu
10.5 2.1 16676.5 9000 4000 3000
2.5 160 543.3 14 53 56
performed to define the main and the trace minerals and their interlocking, as well. The results of the XRD analysis and optical mineralogy of the primary two gold samples are presented in Table 1. The major phases of minerals were quartz, clay minerals (illite), gypsum, hematite, arsenopyrite, feldspar and calcite for the first sample and quartz, goethite, calcite and feldspar for the second one. Microscopic analysis confirmed the existence of the phases detected by XRD method. In addition, arsenopyrite was reported in the optical mineralogy studies for the first sample which due to low amount of arsenopyrite, was not detected in XRD analysis. The samples were analyzed qualitatively by using X-ray fluorescence (XRF) technique in which its result is presented in Table 2. Due to the presence of feldspar and clay minerals in first sample, the amount of K2O and Al2O3 was twice as much compared to the second one. Also CaO content in the first sample was ten times higher than the second one that could cause some difficulties in cyanidation process. The iron oxide amount was 5.47% and 14.06% respectively. The analysis of gold, silver, arsenic, lead, zinc and copper were carried out by using atomic adsorption, inductivity coupled plasma and optical emission spectrometry (ICP-OES) techniques in which results are given in Table 3. It is seen that in the first sample, gold and arsenic content are much higher than the second one (by 4 and 32 times respectively), whereas the second sample contains silver much higher than the first one (80 times).
Mineralogical studies were carried out by optical microscope in order to determine gold liberation size and evaluation of various mineral interlocking in different size fractions including + 2000, −2000 + 840, −840 + 590, −590 + 297, − 297 + 149, −149 + 74, −74 + 62, −62 + 53 and −53 μm by using polished and thin sections in which their results are previously presented in Table 1. 2.3.1. High grade gold ore (sample1) The major minerals were quartz, clay minerals, gypsum, hematite, arsenopyrite, feldspar and calcite whereas the minor ones were galena, sphalerite, chalcopyrite, pyrite, malachite, azurite and celestite. The relationships between gold and sulfide minerals e.g. pyrite, sphalerite and galena are shown in Fig. 1. For further identification of first sample characterization, scanning electron microscope (SEM) method was used and its results are presented as particles photos and its component histogram in Fig. 2 series. As it can be seen in Fig. 2(A) and (C), components such as Pb, Zn, As and Fe were detected as major parts of particles in the first sample. Titanium and manganese are also found in the first sample which is given in Fig. 2(B) and (E) respectively. Aluminosilicate minerals such as illite and also feldspar and calcite can be distinguished in Fig. 2(D), due to the presence of the high amounts of Si, O, Al, Na, Ca and K elements in this particle. In Fig. 2(F) and (G), antimony is identified as a major part of the particle and Sr element in the particle of Fig. 2(H). By comparison of mineralogical data for both samples, it can be resulted that the ore characteristics of the first sample are more sophisticated than the other one and high amount of sulfide minerals causes more cyanide consumption. In some cases, porous particles are observed in the sample which is shown in Fig. 2(F) and (G). Also all the qualitative data of the SEM analysis are presented in Table 4. After separation by heavy liquid, heavy minerals were analyzed by optical microscope and as it can be seen in Fig. 3 which liberated gold particle as nugget and associated gold with sulfide minerals were consequently distinguished. Also the presence of other heavy minerals such as galena and celestite were proved by this method. According to the mineralogical studies, it can be concluded that some parts of gold is disseminated in sulfide minerals such as arsenopyrite, galena, sphalerite, chalcopyrite, etc. 2.3.2. Low grade gold ore (sample2) The second sample is categorized as oxide gold ore, because the main part of its pyrite minerals was changed to iron oxide and hydroxide such as goethite and lepidocrosite which are presented in Fig. 4. According to thin and polished sections studies, the main gangue mineral was SiO2 which was in three different forms namely
Fig. 1. Interlocking of gold and sulfide minerals.
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Fig. 2. Scanning electron microscope (SEM) to characterize the first sample.
crystalline, microcrystalline and cryptocrystalline. Fig. 4 also shows that gold grains are free and interlocked with quartz matrix. In addition, the presence of a high amount of silver in free and sulfide forms and also the low content of other sulfide minerals such as base metal sulfides is a remarkable point of these studies. 3. Material and methods In order to prepare and grind samples for cyanidation process, a Denver ball mill was used. The cyanidation experiments were carried
out in a 5-l plastic container equipped with the Ika mechanical mixer (Ika-RW20, Germany) with a manual controller unit and stainless steel impeller with a 9.85 cm diameter. A 20-l filter press and vacuum filter were applied for pulp filtration. Also for determination of free cyanide in solution, titration was employed with standard silver nitrate (0.01 M) and potassium iodide (10%) as indicator. To control pH, a pH meter (744 Metrohm) was utilized and hydrated lime and hydrochloric acid (5%) were used to set the pulp pH. In cyanidation, sodium cyanide was the lixiviant with injected air to pulp by sparger and hydrogen peroxide as oxidizing reagents.
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Table 4 The qualitative data of the SEM. Figure
Qualitative data
A B C D E F G H
Ca, K, Mg, Fe, As, Pb, S, Au, Si, Zn Au, Si, Al, As, Ti, Fe, Mg, Ca, K Au, Sr, Si, Fe, Al, S, As, Ca, K Fe, Si, Al, Na, Au, K, Ca Mn, Fe, Si, Ca, K, Al, Au Ca, Sb, Fe, Si, As, Na Fe, Si, Sb, Ca, Na, Al, As Sr, Si, Fe, Ca, Al, As, K, Sb
2009). The primary pretreatment was performed on the 1 kg of first sample by using HCl in 4 h and at pH of 2.5, with 25% as solid percent. The second pretreatment was carried out by 200 ml of HNO3 (30%) with the same condition as to the first test. In order to decrease of detrimental effect of sulfide and metal ions, roasting test in various temperatures and times were studied. Samples with 5 gr weight were analyzed for arsenic, antimony, cadmium, bismuth and mercury after roasting (each test were duplicated). 5. Results 5.1. Effect of grinding time
Solution and solids were analyzed by atomic adsorptions, inductivity coupled plasma and optical emission spectrometry (ICP-OES). 4. Apparatus and experimental procedures Cyanidation experiments were conducted to determine the most suitable conditions for two samples of different origins. These parameters included grinding time, sodium cyanide concentration, pH and leaching time. In order to increase the process recovery and reduce cyanide consumption, diagnostic leaching techniques as “pretreatment procedure”, which included acidic leaching by hydrochloric, nitric acid and roasting, were implemented to eliminate the detrimental elements such as arsenic, antimony and carbonaceous matter for the first sample (Ubaldini et al., 1994; Sinadinovic et al., 1999; Celep et al.,
The effect of different grinding times, to reach the sought particle size, 75 μm were studied. The results obtained from these grinding times versus gold recoveries and passing percent are presented in Figs. 5 and 6. In these experiments, other parameters were kept constant at 4000 mg/L of cyanide concentration for the first sample, and 2500 mg/L for the second one, whereas pH and leaching time for both samples were identical at 10.5 and 24 h. Figs. 5 and 6 show that the most suitable grinding times for two samples were found as 35 and 55 min respectively in which 91.8% and 86.24% gold recoveries were achieved. 5.2. Cyanide concentration Different cyanidation tests with various cyanide concentrations were performed, to determine the best cyanide concentration. It is
Fig. 3. (A) Liberated gold nugget, (B) unliberated gold associated with sulfide minerals.
Fig. 4. A: The fine gold grain associated with quartz about 12 μm, B: replace most of pyrite minerals by goethite and lepidocrosite.
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Fig. 5. The gold recovery and passing percent versus grinding times for sample 1 (pH: 10.5, cyanide concentration: 4000 mg/L, cyanidation time: 24 h).
noteworthy that within these tests the other parameters were constant at 35 and 55 min for grinding time, pH = 10.5 and 24 h for leaching time. According to the obtained results that are demonstrated in Fig. 7, by increasing cyanide concentration from 4000 to 8000 mg/L in first sample and 2500 to 6500 mg/L for second one, the quantity of free cyanide in pulp was increased and recovery continued its upward trend, whereas recovery values were reached from 92.45% to 96.82% and 88.31% to 89.92% respectively. Regarding the standard free
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Fig. 8. The trend of recovery versus pH for two samples (grinding time: 35 and 55 min for the first and second samples, cyanide concentration: 4000, 2000 mg/L for the first and second samples, cyanidation time: 24 h).
cyanide in pulp (500 to 1000 mg/L) and acceptable recoveries (higher than 88%), 4000 and 2500 mg/L for cyanide concentrations were selected as best values for the two samples. 5.3. pH Cyanidation tests were conducted at different incremental pHs from 9.5 to 12 (with approximately 0.5 interval step). Fig. 8 presents the results of these experiments, in which the highest recoveries were 94.91% in pH 11.1 and 92.5% in pH 10.5, for the two samples respectively. The leaching time was 24 h. 5.4. Leaching time To find out the best cyanidation time, a 48 h test was conducted for each sample. Samples (10 mL) from the clear solution obtained after vacuum filtration were used for free cyanide determination and a 30 mL sample was provided to analyze gold content. The results are presented in Fig. 9. It is seen that the trend of gold dissolution was almost identical for two samples, and there was a significant inclined rate in first 10 h whereas maximum recovery reached 92% and 90% respectively. 5.5. Kinetics of gold dissolution (sample 1)
Fig. 6. The gold recovery and passing percent versus grinding times for sample 2 (pH: 10.5, cyanide concentration: 2000 mg/L, cyanidation time: 24 h).
Fig. 7. The results of cyanide concentration for two gold samples (grinding time: 35 and 55 min for the first and second samples, pH: 10.5, cyanidation time: 24 h).
Various types of oxidizing reagent such as air, hydrogen peroxide and a combination of air and H2O2 were utilized to increase the gold dissolution rate for the first sample (Yannopoulos, 1991; Guzman
Fig. 9. The gold recovery versus cyanidation times for two samples (grinding time: 35 and 55 min for the first and second samples, cyanide concentration: 4000, 2000 mg/L for the first and second samples, pH: 11.1 and 10.5 for the first and second samples).
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Fig. 10. Gold recovery versus different leaching times without oxidizing reagent, k = 1.06, (grinding time: 35 min, cyanide concentration: 4000 mg/L, pH: 11.1).
Fig. 13. Gold recovery versus different leaching time by air, k = 1.75, (grinding time: 35 min, cyanide concentration: 4000 mg/L, pH: 11.1).
condition for cyanidation process. Cyanidation test was performed on the roasted sample in this condition and obtained results presented a considerable decrease of cyanide consumption by 1150 mg/L and increasing recovery by 5.2% (Roshan, 1990; Dunn and Chamberlain, 1997; Celep et al., 2009). It can be concluded that by increasing of roasting temperature and time, the amount of arsenic, bismuth and mercury decreased, but this trend for antimony and cadmium has a little difference which may be relevant to the boiling point of these elements in which their trends are illustrated in Figs. 14–18.
Fig. 11. Gold recovery versus different leaching time by H2O2, k = 1.35, (grinding time: 35 min, cyanide concentration: 4000 mg/L, pH: 11.1).
et al., 1999; Marsden and Lain House, 2006). The results are shown in Figs. 10–13. For each experiment, two curves are presented, the first one is related to actual data which are obtained from experiments, and the second curve is the modeled one, using first order equation as y = ymax(1 − exp(−kx)) which was generated by solver function in Excel, based on the least square criteria. 5.6. Pre-treatment experiments (of sample 1) According to the results, roasting in 1000 °C at 2 h had adequate effects on omitting of followed elements and warrants a good
Fig. 14. The content of arsenic after roasting in various temperatures and times (initial value: 16,650 ppm).
Fig. 12. Gold recovery versus different leaching time by H2O2 + air, k = 1.65, (grinding time: 35 min, cyanide concentration: 4000 mg/L, pH: 11.1).
Fig. 15. The content of antimony after roasting in various temperatures and times (initial value: 2214 ppm).
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Fig. 16. The content of cadmium after roasting in various temperatures and times (initial value: 1.8 ppm).
Fig. 17. The content of bismuth after roasting in various temperatures and times (initial value: 1.1 ppm).
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in 50 min for the first sample. This phenomenon was due to the fact that more liberated minerals in higher grinding times were achieved and also gold grains surface was more covered by carbonaceous matter, calcite, and clay minerals (illite). Increase of cyanide concentration in leaching pulp caused high cyanide consumption due to the presence of cyanicides and did not improve gold recovery. With increasing pH, the value of free cyanide was definitely increased, whereas the concentration of free cyanide was increased from 120 to 240 mg/L at pH = 9.5 and reached to 2250 and 1750 mg/L at pH = 12 for the first and second samples respectively. Therefore, the optimum recovery was in the pH of 10.5 and 11 for both samples so that it decreased outside of this range. This could be attributed to the formation of other complexes at low pH which caused the gold recovery to be decreased. In addition, at pH higher than 11, the majority of components were not able to make stable complexes with cyanide. Hence, the free cyanide intensively increased which is followed by a recovery decrease. Also at high pH of solution, the suspended particles were deposited immediately, because of the forming of Ca(OH)2 and Mg(OH)2 complexes, as the flocculants matters. Furthermore, gold passivation at higher pH has been reported and a dramatic decrease of gold extraction in sample 2 at higher pH may be related to this. For leaching time, the curve slope in Fig. 9, which shows the recovery rate, was high at the first 8 h and gradually decreased until 24 h (changes was up to 2%). For instance, gold recoveries were achieved 92% for the first 10 h for the first sample and 89.4% for the first 8 h using second sample. Meanwhile the recoveries finally reached to 96.56% and 93.5% after 48 h respectively. Gold dissolution rate followed by 1st order kinetics presented as y = ymax(1 − exp (−kx)), in which k (rate constant) values were obtained at 1.06 and 1.16 min− 1 for the first and second samples. The results of which are shown in Table 5. The introduction of air during leaching was more effective in kinetics rate increase (1.75 min− 1), whereas using mixture of air and H2O2 yielded similar results (1.65 min− 1). In addition, the gold recoveries reached 99.03, 98.07 and 97.56% for the 48 h respectively which showed the high effect of air as an oxidizing reagent. Despite the difference in rate, the final recovery value by using H2O2 as an oxidizer was higher than other reagents. By using of HCl and HNO3 as acidic pretreatments, gold recoveries reached to 95.11% and 98.87% respectively and 97.05% by roasting of sample in 1000 °C in 2 h. In addition, cyanide consumption decreased by 340, 210 and 1150 mg/L for acidic pretreatment and roasting. The amount of arsenic, antimony, cadmium, bismuth and mercury elements decreased considerably as 88, 34, 22, 75 and 82% on the roasted sample (1000 °C in 2 h). The results of gold recovery and cyanide consumption versus variation of time and temperature of roasting process are presented in Table 6.
Fig. 18. The content of mercury after roasting in various temperatures and times (initial value: 1 ppm).
7. Summary and conclusions 6. Discussion By increasing of grinding time, cyanide consumption increased for both sample whereas 2910 mg/L in 20 min was changed to 3680 mg/L
In this study, two different types of gold samples were investigated in various cyanidation experiments. The first sample was a high grade ore and its gold and arsenic grades were 10.5 and 16676.5 g/t respectively and the second sample was a low grade ore with 2.5 g/t of
Table 5 Rate constant (k) of different oxidizing reagents for the first sample. Oxidant
Rate constant (min− 1)
Rate constant (min− 1)
Cyanide concentration (mg/L)
Cyanide consumption (mg/L)
Gold recovery (%)
Without oxidizer H2O2 H2O2 + air Air
1.06 1.35 1.65 1.75
10.5 10.5 10.5 10.5
4000 4000 4000 4000
3875 3540 3510 3650
96.56 99.03 98.07 97.56
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cyanidation condition, using hydrogen peroxide, air and mixture of H2O2 and air, were 1.06, 1.35, 1.75 and 1.65 min− 1 respectively.
Table 6 Au recovery versus roasting temperature and time. Condition Time (h)
Temperature (°C)
0.5
600 700 800 900 1000 600 700 800 900 1000 600 700 800 900 1000
1
2
Gold recovery (%)
Cyanide consumption (mg/L)
94.95 95.01 95.13 95.74 96.07 95.41 95.97 96.28 96.43 96.59 95.56 96.27 96.74 96.95 97.05
2950 2860 2740 2550 2405 2685 2480 2290 2210 2160 2605 2340 2085 2010 1940
Acknowledgments The authors would like to thank the Director of Mineral Processing Division in the Geological Survey of Iran for the support of this project; and Abedian and Movaghar Consultant, who helped in the collection of the samples and Beshkani for editing of the paper. References
Au, and low arsenic (543.3 g/t), but high silver content (160 g/t) which are presented in Table 3. The two samples were used in cyanidation experiments to determine their optimized parameters (Table 7), and the following conclusions were accordingly made. Type of detrimental minerals and their values caused to sophisticate the optimized conditions for the gold's processing circuit. Moreover detrimental matters in cyanidation process for the first sample were high amounts of arsenic, antimony, mercury, carbonaceous matters (calcite) and clay minerals (illite) which had negative effects on the process. The effects of toxic materials were caused by lower grinding time and a high amount of cyanide consumption. Ultimately the intensive effect of oxidizer on kinetic and final recovery was remarkable and optimized pH value for first sample was more than the second one, 10.5 against 11.1. Various pretreatment leaching under acidic condition, HNO3 and HCl, were studied on the first sample which its results are briefly given in Table 8. Cyanidation test was carried out using roasted sample, 2 h in 1000 °C, and its results showed that cyanide consumption was decreased considerably by 1150 mg/L and gold recovery was increased 5.2%. Different oxidizing reagents such as hydrogen peroxide (0.015 M), air (0.5 L/min) and mixture of H2O2 and air were employed to increase gold leaching kinetics for the first sample. The gold leaching followed 1st order kinetics, in which k, rate constant, value for normal
Table 7 The most suitable conditions for two different samples. Parameter
Sample 1
Sample 2
Grinding time (min) CN− concentration (mg/L) pH Cyanidation time (h) Max. gold recovery (%) Max. silver recovery (%)
35 4000 11.1 24 94.91 28.2
55 2500 10.5 24 92.5 93.5
Table 8 Results of acidic pretreatment on the sample 1. Acidic pretreatment
Decrease of cyanide consumption (mg/L)
Gold recovery (%)
HNO3 HCl
340 210
98.87 95.11
Adams, M.D., 2005. Advances in Gold Ore Processing, Developments in Mineral Processing. Elsevier, Amsterdam. Aghamirian, M.M., Yen, W.T., 2005. Mechanism of galvanic interactions between gold and sulfide minerals in cyanide solution. Minerals Engineering 18, 393–407. Angelidis, T.N., Kydros, K.A., 1995. Selective gold dissolution from a roasted auriferous pyrite–arsenopyrite concentrate. Hydrometallurgy 37 (1), 75–88. Brooy, S.R., Linge, H.G., Walker, G.S., 1994. Review of gold extraction from ores. Minerals Engineering 7 (10), 1213–1241. Celep, O., Alp, I., Deveci, H., Vicil, M., 2009. Characterization of refractory behaviour of complex gold/silver ore by diagnostic leaching. Transactions of Nonferrous Metals Society of China 19, 707–713. Dai, X., Jeffrey, M.I., 2006. The effect of sulfide minerals on the leaching of gold in aerated cyanide solutions. Hydrometallurgy 82, 118–125. Deng, T., Liao, M., 2002. Gold recovery enhancement from a refractory flotation concentrate by sequential bioleaching and thiourea leach. Hydrometallurgy 63 (3), 249–255. Deschenes, G., Prud home, P.J.H., 1997. Cyanidation of a copper–gold ore. Mineral Processing 50, 127–141. Dunn, J.G., Chamberlain, A.C., 1997. The recovery of gold from refractory arsenopyrite concentrates by pyrolysis–oxidation. Minerals Engineering 10 (9), 919–928. Ellis, S., Senanayake, G., 2004. The effects of dissolved oxygen and cyanide dosage on gold extraction from a pyrrhotite-rich ore. Hydrometallurgy 72, 39–50. Fink, C.G., Putnam, G.L., 1950. The action of sulphide ion and of metal salts on the dissolution of gold in cyanide solutions. Mining Engineering 187, 952–955. Guzman, L., Segarra, M., Chimenos, J.M., Fernandez, M.A., Espiell, F., 1999. Gold cyanidation using hydrogen peroxide. Hydrometallurgy 52, 21–35. Habashi, F., 1967. Kinetics and mechanism of gold and silver dissolution in cyanide solution. Montana Bureau of Mines and Geology 59, 1–42. Habashi, F., 1970. Principles of extractive metallurgy. Hydrometallurgy, vol. 2. Gordon and Breach, New York. Hedley, N., Tabachnick, H., 1968. Mineral Dressing Notes No. 23. Chemistry of Cyanidation. American Cyanamid Company, New Jersey, USA. Iglesias, N., Carranza, F., 1994. Refractory gold-bearing ore: a review of treatment methods and recent advances in biotechnological techniques. Hydrometallurgy 34, 383–395. Karimpour, M.H., 2006. Geology, Geochemistry, Petrology and Geophysics studies of Hirad Sn–Au mineralization. Eastern Mineral Research Center of Iran, pp. 16–113 (In Persian). Kondos, P.D., Deschenes, G., Morrison, R.M., 1995. Process optimization studies in gold cyanidation. Hydrometallurgy 39, 235–250. LaBrooy, S.R., Linge, H.G., Walker, G.S., 1994. Review of gold extraction from ores. Minerals Engineering 7 (10), 1213–1241. Lawrence, R.W., Bruynesteyn, A., 1983. Biological pre-oxidation to enhance gold and silver recovery from refractory ores and concentrates. CIM Bulletin 76 (857), 107–110. Lorenzen, L., van Deventer, J.S.J., 1992. Electrochemical interactions between gold and its associated minerals during cyanidation. Hydrometallurgy 30, 177–194. Luna, R.M., Lapidus, G.T., 2000. Cyanidation kinetics of silver sulfide. Hydrometallurgy 56, 171–188. Marsden, J.O., Lain House, C., 2006, “The chemistry Of Gold extraction”, Published by the SME Inc., 2nd Edition. Mason, P.G., 1990. Energy requirements for the pressure oxidation of gold-bearing sulphides. J. Met. 42 (9), 15–18. Murthy, D.S.R., Dey, M.L., Roychowdhury, S.K., Mathur, S.B., Akerkar, D.D., 1992. Studies on extraction of refractory gold silver through thiourea leaching. In: Misra, V.N., Halbe, D., Spottiswood, D.J. (Eds.), Int. Conf. Ext. Metall. Gold Base Met, pp. 329–332. Nicol, M.J., Fleming, C.A., Paul, R.L., 1987. The chemistry of the extraction of gold. In: Stanley, G.G. (Ed.), The Extractive Metallurgy of Gold, vol. 2. South African Institute of Mining and Metallurgy, Johannesburg, pp. 831–905. Qingcui, Li, Dengxin, Li, Fangjun, Qian, 2009. Pre-oxidation of high-sulfur and higharsenic refractory gold concentrate by ozone and ferric ions in acidic media. Hydrometallurgy 97, 61–66. Rees, K.L., van Deventer, J.S.J., 1999. The role of metal–cyanide species in leaching gold from a copper concentrates. Minerals Engineering 12, 877–892. Roshan, B.B., 1990. Hydrometallurgical processing of precious metal ores. Mineral Processing and Extractive Metallurgy Review 6, 67–80. Rubisov, D.H., Papengelakis, V.G., Kondos, P.D., 1996. Fundamental kinetic models for gold ore cyanide leaching. Canadian Metallurgical Quarterly 35 (4), 353–361. Senanayake, G., 2008. A review of effects of silver, lead, sulfide and carbonaceous matter on gold cyanidation and mechanistic interpretation. Hydrometallurgy 90, 46–73.
P. Karimi et al. / International Journal of Mineral Processing 95 (2010) 68–77 Sinadinovic, D., Kamberovic, Z., Vakanjac, B., 1999. Refractory gold ores, characteristics and methods of their procession. Proceedings of 8th Balkan Mineral Processing Conference, Belgrade, Yugoslavia, pp. 411–418. Stoychevski, M., Williams, L.R., 1993. Influence of oxygen, hydrogen peroxide and ozone on dissolution of gold from pyrite ore. Trans. Inst. Min. Metall. C 93–98. Tan, H., Feng, D., luckey, G.C., van Deventer, J.S.J., 2005. The behavior of carbonaceous matter in cyanide leaching of gold. Hydrometallurgy 78, 226–235. Ubaldini, S., Massidda, R., Abbruzzese, C., 1994. Factorial experiments in the study of a biooxidative pretreatment for a conventional cyanidation of a Turkey arsenical gold ore. Proceedings of the 5th Int Mineral Processing Symposium. Turkish Mining Development Foundation, Cappadocia, pp. 403–410. Weichselbaum, J., Tumilty, J.A., Schmidt, C.G., 1989. The effect of sulphide and lead on the rate of gold cyanidation. Aust. IMM Annu. Conf., Petth-Kalgoorlie, pp. 221–224. Yannopoulos, J.C., 1991. The Extractive Metallurgy of Gold. Van Norstand Rainhold, New York.
FURTHER READING Deschenes, G., Wallingford, G., 1995. Technical note effect of oxygen and lead nitrate on the cyanidation of a sulphide bearing gold ore. Minerals Engineering 8, 923–931.
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Fagan, P.A., Haddad, P.R., Mitchel, I., Dunne, R., 1998. Monitoring the cyanidation of gold–copper ores with ion-interaction chromatography. Determination of the CN: Cu(I) molar ratio. Journal of Chromatography 804, 249–263. Hausen, D.M., Bucknam, C.H., 1984. Study of preg-robbing in the cyanidation of carbonaceous gold ores from Carlin. Proceedings of the Second International Congress on Applied Mineralogy in the Minerals Industry. Metallurgical Society of AIME, Los Angeles, CA, pp. 833–856. Hewitt, D.M., Breuer, P.L, Jeffrey, M.I., Naim, F., 2009. Mechanisms of sulfide ion oxidation during cyanidation. Part II: Surface catalysis by pyrite. Minerals Engineering 22 (13), 1166–1172. Jeffrey, M.I., Breuer, P.L., 2000. The cyanide leaching of gold in solutions containing sulfide. Minerals Engineering 13, 1097–1106. Kudryk, V., Kellogg, H.H., 1954. Mechanism and rate controlling factors in the dissolution of gold in cyanide solutions. J. Met. 6, 541–548. Nguyen, H.H., Tran, T., Wong, P.L.M., “Copper interaction during the dissolution of gold”, Minerals Engineering 10, pp. 491–505.
Contents lists available at ScienceDirect
International Journal of Mineral Processing j o u r n a l h o m e p a g e : w w w. e l s ev i e r. c o m / l o c a t e / i j m i n p r o
Cyanidation of gold ores containing copper, silver, lead, arsenic and antimony P. Karimi a,⁎, H. Abdollahi a, A. Amini b, M. Noaparast a, S.Z. Shafaei a, F. Habashi c a b c
School of Mining Engineering, University of Tehran, Islamic Republic of Iran Geological Survey and Exploration of Iran, Islamic Republic of Iran Laval University, Quebec City, Canada G1V 0A6
a r t i c l e
i n f o
Article history: Received 11 December 2009 Received in revised form 27 February 2010 Accepted 13 March 2010 Available online 20 March 2010 Keywords: Gold ores Cyanidation Reaction kinetics Pretreatment
a b s t r a c t Cyanidation tests on two ore samples from two different gold deposits are reported. The first sample contained 10.5 ppm gold with high arsenic and antimony. The second sample had a low gold content (2.5 ppm) but a high silver content (160 ppm). The first series of test work focused on the determination of conditions for extracting gold from the samples ground to − 75 μm. The optimum parameters were 4000 mg/L for cyanide concentration, pH = 11.1 and 24-h cyanidation time for the first sample; and 2500 mg/L, pH = 10.5 and 24 h for the second sample. Silver and gold recoveries were 94.91% and 28.2% respectively for the first sample and 92.5% and 93.5% for the second sample. In the second series of test work, H2O2 (0.015 M), air (0.15 L/min) and a mixture of H2O2 and air were used as oxidizing agents to improve the gold extraction kinetics of the first sample. It was found that gold leaching recovery followed first order kinetics and that the injection of air had the maximum beneficial effect on the leaching kinetics. In the third series of test work, acid pretreatment (HNO3 and HCl) and roasting (600– 1000 °C) from 0.5 to 2 h were carried out before cyanidation. Acid pretreatment reduced cyanide consumption by 340 and 210 mg/L respectively and the corresponding gold recoveries increased to 98.87% and 95.11%. Cyanidation results on roasted samples showed that cyanide consumption was drastically reduced by 1150 mg/L and gold recovery increased by 5.2%. Furthermore, arsenic, antimony, cadmium and mercury contents were considerably reduced in the roasted sample (2 h at 1000 °C). © 2010 Elsevier B.V. All rights reserved.
1. Introduction
reactions and the preferential formation of Au (CN)− 2 on the surface instead of AuSx (Senanayake, 2008).
1.1. General Early studies on the dissolution of gold in cyanide solution in the presence of sulfide minerals have shown that heavy metal components, such as Cu, Fe and Zn, significantly increase the consumption of both cyanide and oxygen (Habashi, 1967; Dai and Jeffrey, 2006). In addition, the sulfide component has been shown to have a strong impact on the gold leaching kinetics (Fink and Putnam, 1950; Hedley and Tabachnick, 1968; Dai and Jeffrey, 2006). The leaching behavior of gold in the presence of sulfide minerals depended strongly on both the solubility of the sulfides and the oxygen concentration in the solution (Dai and Jeffrey, 2006). Based on the experiments observation, it is postulated that sulfide ions (formed by the decomposition of sulfide minerals) show detrimental effect on the cyanidation kinetics of gold and silver. However, the detrimental effect of dissolved sulfide on gold cyanidation according to reactions R1 and R2 which diminishes at higher cyanide concentrations, and favor the reverse
⁎ Corresponding author. Tel./fax: +98 21 64592257. E-mail address: [email protected] (P. Karimi). 0301-7516/$ – see front matter © 2010 Elsevier B.V. All rights reserved. doi:10.1016/j.minpro.2010.03.002
−2 R1 M2S + 4CN− ⇆ 2M(CN)− 2 +S − − R2 M2S + 4CN− + H2O ⇆ 2M(CN)− 2 + OH + HS
The presence of galena, as well as both sulfide and added lead(II), also improved the rate of gold dissolution. The dissolution rate was found to be affected by ions such as lead(II), sulfide, and iron(III) released by the sulfide mineral. The significant decrease in the rate of electrochemical and chemical dissolution of gold caused by chalcopyrite, pyrite, and pyrrhotite has been related to the various films formed on the gold surface as well as the galvanic interaction (Lorenzen and van Deventer, 1992; Aghamirian and Yen, 2005; Dai and Jeffrey, 2006; Senanayake, 2008). The dissolution of sulfides results in high cyanide consumption with the formation of cyano complexes (e.g., of the ions Fe+ 2, Zn+ 2, Cu+ 2, Ni+ 2, Mn+ 2) and SCN−. When sulfide minerals are present in gold ores, gold dissolution can be affected in various ways. In one hypothesis (Weichselbaum et al., 1989; Kondos et al., 1995), soluble sulfide (S2− or HS−) generated from mineral dissolution reacts with gold and forms a passive film, which decreases the rate and extent of leaching. In another theory (Lorenzen and van Deventer, 1992; Kondos et al., 1995), a dynamic coupling of reduction forms at the sulfide mineral surface which
P. Karimi et al. / International Journal of Mineral Processing 95 (2010) 68–77
results in oxidation on the gold grains (Kondos et al., 1995). In the leaching of gold using cyanide, copper presents difficulties due to the lack of selectivity of cyanide for gold over copper (Rees and van Deventer, 1999). Cyanide also forms complexes with a number of 2− 2− and Ag(CN)− metal irons: Fe(CN)4− 6 , Ni(CN)4 , Zn(CN)4 2 which decrease the free cyanide concentration in slurry and retard gold leaching(Rees and van Deventer, 1999). The refractoriness of gold ores can result primarily from the inherent mineralogical features with particular reference to the mode of presence and association of gold, and to the presence of carbonaceous matter (Rubisov et al., 1996; Celep et al., 2009). A suitable pretreatment process is often required to overcome the refractoriness and render the gold accessible to the lixiviant action of cyanide and oxygen (Ubaldini et al., 1994; Celep et al., 2009). Roasting, pressure oxidation, biooxidation, ultrafine grinding and modified cyanidation are the pretreatment methods currently practiced for refractory gold ores and concentrates (Sinadinovic et al., 1999; Iglesias and Carranza, 1994; Celep et al., 2009). Roasting, a simple stage process using a fluidized bed roaster at around 650 °C, is one of the most common methods for the treatment of gold-bearing pyrite/arsenopyrite and pyrrhotite concentrates to produce porous calcine with the increased amenability to cyanidation (Roshan, 1990; Dunn and Chamberlain, 1997; Celep et al., 2009). Diagnostic leaching has been also proved to be very useful as an analytical tool for establishing the distribution of gold within different mineral phases of an ore (Celep et al., 2009). It was reported that pre-leaching successfully overcame the effect of pyrite and pyrrhotite on cyanide consumption and the gold leaching kinetics, but it had no beneficial effect when chalcopyrite was present (Dai and Jeffrey, 2006).
69
effect by presence of preg-robber, it considerably diminishes in solutions of high concentration of cyanide (Tan et al., 2005). In such refractory gold ores, gold particles may sometimes be occluded or included in the sulfide minerals and pretreatment is necessary to decompose the mineral structure to liberate gold for subsequent recovery. These ores are usually pretreated by some oxidizing process after which gold and silver can be recovered by standard cyanidation process. The main available oxidizing processes for refractory gold ores are roasting (LaBrooy et al., 1994; Angelidis and Kydros, 1995; Qingcui et al., 2009), biological oxidation (Lawrence and Bruynesteyn, 1983; Deng and Liao, 2002) and pressure oxidation (Mason, 1990; Murthy et al., 1992; Qingcui et al., 2009). It has also been demonstrated that the addition of other oxidizing reagents increases the kinetics of gold dissolution (Stoychevski and Williams, 1993; Deschenes and Prud home, 1997). In this study, main cyanidation parameters for two different gold samples (grade and mineralogy) were compared. The studied samples were from two different gold deposits in Iran. The first sample (high gold grade) was taken from “Hirad gold deposit”, situated in Southern Khorasan province, and the second one (low gold grade) was provided from “Latala gold deposit”, located in Kerman province (Karimpour, 2006). After optimization of parameters, and due to the presence of arsenic, antimony, carbonaceous matters and clay minerals, pretreatment processes such as acidic leaching and roasting for the first sample were investigated. Due to the complexity of mineralogy of first sample, gold dissolution kinetic studies were also evaluated by using different oxidizing reagents which was not applied for the second one. 2. Ore preparation and characterization study
1.2. Factors affecting gold leaching 2.1. Sample preparation Cyanidation has been used for over 100 years to extract precious metals from sulfide ores. Despite this fact, the reactions involved are not fully understood. The chemical reactions that take place during the cyanidation of concentrate for ores can be very complex (Luna and Lapidus, 2000). Since a large proportion of gold ores contain sulfide minerals, the effects of these minerals on gold dissolution in cyanide solution have interested many researchers. The initial rate of gold dissolution is largely controlled by the factors such as cyanide and oxygen concentrations, pH, solid–liquid interface, Eh of the slurry, alkalies, particle size and temperature. The presence of other catalytic ions in solution and salinity of water also affect the rate of leaching (Habashi, 1970; Nicol et al., 1987; Ellis and Senanayake, 2004). From the electrochemical studies, the rate of gold leaching in air-saturated solutions increases with increasing concentration of cyanide but becomes independent of cyanide concentration when it exceeds 0.075% KCN. Clearly the high concentrations of oxygen and cyanide improve initial gold leaching kinetics but high cyanide concentrations also increase cyanide consumption (Ellis and Senanayake, 2004). The use of oxygen or an oxidizing reagent is essential for the dissolution of gold under normal conditions. Such oxidizing reagents as sodium peroxide, potassium permanganate, bromine, and chlorine have been used with more or less success; but they are not used now, due to their cost and complications involved in their handling (Yannopoulos, 1991; Guzman et al., 1999; Marsden and Lain House, 2006). Generally, gold ores can be classified as “free milling” and “refractory” depending on their response to cyanide leaching (Brooy et al., 1994; Celep et al., 2009). While high gold recoveries (N90%) from free milling ores can be readily achieved, refractory gold ores are often characterized by the low gold extractions (50–80%) within a conventional cyanide leaching process (Rubisov et al., 1996; Adams, 2005; Celep et al., 2009). The presence of preg-robber components such as carbonaceous matter, sulfide minerals and aluminosilicate have more influence in decreasing of gold and silver dissolution rate and recovery. In despite of negative
Two representative samples were supplied by using a systematic sampling procedure and about 400 kg from each sample were used for experiments. After four crushing stages (two jaw crushers, cone crusher and roll crusher), primary sample size reached to −2830 μm. In the next step, 2 kg samples were prepared by riffling and cone and quartering method. 2.2. XRD, XRF and chemical analysis For ore characterization, optical mineralogy, using the prepared polished and thin sections, X-ray diffraction (XRD) studies were Table 1 Mineralogical compositions of two gold samples. Gold ore type
Mineral name
Common mineral formula
Composition percentage
Sample 1
Quartz Illite
SiO2 (K,H3O)(Al,Mg,Fe)2(Si,Al)4O10[(OH)2, (H2O)] CaSO4.2H2O Fe2O3 FeAsS KAlSi3O8 CaCO3 PbS ZnS CuFeS2 FeS2 Cu2(CO3)(OH)2 Cu3(CO3)2(OH)2 SrSO4 SiO2 FeO(OH) CaCO3 KAlSi3O8
61.6 11.1
Sample2
Gypsum Hematite Arsenopyrite Feldspar Calcite Galena Sphalerite Chalcopyrite Pyrite Malachite Azurite Celestite Quartz Goethite Calcite Feldspar
9.5 5.2 3.6 3 3 1 0.6 0.5 0.2 0.2 0.2 0.2 73 21.6 2.7 2.4
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2.3. Microscopic mineralogy
Table 2 Qualitative analysis by XRF method. Components (wt.%)
First sample
Second sample
Components (wt.%)
First sample
Second sample
SiO2 Al2O3 Fe2O3 P2O5 CaO SO3 K2O
59.05 17.20 5.47 0.10 4.12 0.60 3.91
68.91 8.34 14.06 0.10 0.44 0.86 1.90
MgO TiO2 MnO ZnO PbO Na2O L.O.I
0.73 0.47 0.12 0.43 0.9 b0.1 6.76
0.36 0.25 0.56 b 0.05 b 0.05 – 4.20
Table 3 Elemental analysis of the primary ores samples by ICP-OES technique. Elements (g/t)
First sample
Second sample
Au Ag As Pb Zn Cu
10.5 2.1 16676.5 9000 4000 3000
2.5 160 543.3 14 53 56
performed to define the main and the trace minerals and their interlocking, as well. The results of the XRD analysis and optical mineralogy of the primary two gold samples are presented in Table 1. The major phases of minerals were quartz, clay minerals (illite), gypsum, hematite, arsenopyrite, feldspar and calcite for the first sample and quartz, goethite, calcite and feldspar for the second one. Microscopic analysis confirmed the existence of the phases detected by XRD method. In addition, arsenopyrite was reported in the optical mineralogy studies for the first sample which due to low amount of arsenopyrite, was not detected in XRD analysis. The samples were analyzed qualitatively by using X-ray fluorescence (XRF) technique in which its result is presented in Table 2. Due to the presence of feldspar and clay minerals in first sample, the amount of K2O and Al2O3 was twice as much compared to the second one. Also CaO content in the first sample was ten times higher than the second one that could cause some difficulties in cyanidation process. The iron oxide amount was 5.47% and 14.06% respectively. The analysis of gold, silver, arsenic, lead, zinc and copper were carried out by using atomic adsorption, inductivity coupled plasma and optical emission spectrometry (ICP-OES) techniques in which results are given in Table 3. It is seen that in the first sample, gold and arsenic content are much higher than the second one (by 4 and 32 times respectively), whereas the second sample contains silver much higher than the first one (80 times).
Mineralogical studies were carried out by optical microscope in order to determine gold liberation size and evaluation of various mineral interlocking in different size fractions including + 2000, −2000 + 840, −840 + 590, −590 + 297, − 297 + 149, −149 + 74, −74 + 62, −62 + 53 and −53 μm by using polished and thin sections in which their results are previously presented in Table 1. 2.3.1. High grade gold ore (sample1) The major minerals were quartz, clay minerals, gypsum, hematite, arsenopyrite, feldspar and calcite whereas the minor ones were galena, sphalerite, chalcopyrite, pyrite, malachite, azurite and celestite. The relationships between gold and sulfide minerals e.g. pyrite, sphalerite and galena are shown in Fig. 1. For further identification of first sample characterization, scanning electron microscope (SEM) method was used and its results are presented as particles photos and its component histogram in Fig. 2 series. As it can be seen in Fig. 2(A) and (C), components such as Pb, Zn, As and Fe were detected as major parts of particles in the first sample. Titanium and manganese are also found in the first sample which is given in Fig. 2(B) and (E) respectively. Aluminosilicate minerals such as illite and also feldspar and calcite can be distinguished in Fig. 2(D), due to the presence of the high amounts of Si, O, Al, Na, Ca and K elements in this particle. In Fig. 2(F) and (G), antimony is identified as a major part of the particle and Sr element in the particle of Fig. 2(H). By comparison of mineralogical data for both samples, it can be resulted that the ore characteristics of the first sample are more sophisticated than the other one and high amount of sulfide minerals causes more cyanide consumption. In some cases, porous particles are observed in the sample which is shown in Fig. 2(F) and (G). Also all the qualitative data of the SEM analysis are presented in Table 4. After separation by heavy liquid, heavy minerals were analyzed by optical microscope and as it can be seen in Fig. 3 which liberated gold particle as nugget and associated gold with sulfide minerals were consequently distinguished. Also the presence of other heavy minerals such as galena and celestite were proved by this method. According to the mineralogical studies, it can be concluded that some parts of gold is disseminated in sulfide minerals such as arsenopyrite, galena, sphalerite, chalcopyrite, etc. 2.3.2. Low grade gold ore (sample2) The second sample is categorized as oxide gold ore, because the main part of its pyrite minerals was changed to iron oxide and hydroxide such as goethite and lepidocrosite which are presented in Fig. 4. According to thin and polished sections studies, the main gangue mineral was SiO2 which was in three different forms namely
Fig. 1. Interlocking of gold and sulfide minerals.
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Fig. 2. Scanning electron microscope (SEM) to characterize the first sample.
crystalline, microcrystalline and cryptocrystalline. Fig. 4 also shows that gold grains are free and interlocked with quartz matrix. In addition, the presence of a high amount of silver in free and sulfide forms and also the low content of other sulfide minerals such as base metal sulfides is a remarkable point of these studies. 3. Material and methods In order to prepare and grind samples for cyanidation process, a Denver ball mill was used. The cyanidation experiments were carried
out in a 5-l plastic container equipped with the Ika mechanical mixer (Ika-RW20, Germany) with a manual controller unit and stainless steel impeller with a 9.85 cm diameter. A 20-l filter press and vacuum filter were applied for pulp filtration. Also for determination of free cyanide in solution, titration was employed with standard silver nitrate (0.01 M) and potassium iodide (10%) as indicator. To control pH, a pH meter (744 Metrohm) was utilized and hydrated lime and hydrochloric acid (5%) were used to set the pulp pH. In cyanidation, sodium cyanide was the lixiviant with injected air to pulp by sparger and hydrogen peroxide as oxidizing reagents.
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Table 4 The qualitative data of the SEM. Figure
Qualitative data
A B C D E F G H
Ca, K, Mg, Fe, As, Pb, S, Au, Si, Zn Au, Si, Al, As, Ti, Fe, Mg, Ca, K Au, Sr, Si, Fe, Al, S, As, Ca, K Fe, Si, Al, Na, Au, K, Ca Mn, Fe, Si, Ca, K, Al, Au Ca, Sb, Fe, Si, As, Na Fe, Si, Sb, Ca, Na, Al, As Sr, Si, Fe, Ca, Al, As, K, Sb
2009). The primary pretreatment was performed on the 1 kg of first sample by using HCl in 4 h and at pH of 2.5, with 25% as solid percent. The second pretreatment was carried out by 200 ml of HNO3 (30%) with the same condition as to the first test. In order to decrease of detrimental effect of sulfide and metal ions, roasting test in various temperatures and times were studied. Samples with 5 gr weight were analyzed for arsenic, antimony, cadmium, bismuth and mercury after roasting (each test were duplicated). 5. Results 5.1. Effect of grinding time
Solution and solids were analyzed by atomic adsorptions, inductivity coupled plasma and optical emission spectrometry (ICP-OES). 4. Apparatus and experimental procedures Cyanidation experiments were conducted to determine the most suitable conditions for two samples of different origins. These parameters included grinding time, sodium cyanide concentration, pH and leaching time. In order to increase the process recovery and reduce cyanide consumption, diagnostic leaching techniques as “pretreatment procedure”, which included acidic leaching by hydrochloric, nitric acid and roasting, were implemented to eliminate the detrimental elements such as arsenic, antimony and carbonaceous matter for the first sample (Ubaldini et al., 1994; Sinadinovic et al., 1999; Celep et al.,
The effect of different grinding times, to reach the sought particle size, 75 μm were studied. The results obtained from these grinding times versus gold recoveries and passing percent are presented in Figs. 5 and 6. In these experiments, other parameters were kept constant at 4000 mg/L of cyanide concentration for the first sample, and 2500 mg/L for the second one, whereas pH and leaching time for both samples were identical at 10.5 and 24 h. Figs. 5 and 6 show that the most suitable grinding times for two samples were found as 35 and 55 min respectively in which 91.8% and 86.24% gold recoveries were achieved. 5.2. Cyanide concentration Different cyanidation tests with various cyanide concentrations were performed, to determine the best cyanide concentration. It is
Fig. 3. (A) Liberated gold nugget, (B) unliberated gold associated with sulfide minerals.
Fig. 4. A: The fine gold grain associated with quartz about 12 μm, B: replace most of pyrite minerals by goethite and lepidocrosite.
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Fig. 5. The gold recovery and passing percent versus grinding times for sample 1 (pH: 10.5, cyanide concentration: 4000 mg/L, cyanidation time: 24 h).
noteworthy that within these tests the other parameters were constant at 35 and 55 min for grinding time, pH = 10.5 and 24 h for leaching time. According to the obtained results that are demonstrated in Fig. 7, by increasing cyanide concentration from 4000 to 8000 mg/L in first sample and 2500 to 6500 mg/L for second one, the quantity of free cyanide in pulp was increased and recovery continued its upward trend, whereas recovery values were reached from 92.45% to 96.82% and 88.31% to 89.92% respectively. Regarding the standard free
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Fig. 8. The trend of recovery versus pH for two samples (grinding time: 35 and 55 min for the first and second samples, cyanide concentration: 4000, 2000 mg/L for the first and second samples, cyanidation time: 24 h).
cyanide in pulp (500 to 1000 mg/L) and acceptable recoveries (higher than 88%), 4000 and 2500 mg/L for cyanide concentrations were selected as best values for the two samples. 5.3. pH Cyanidation tests were conducted at different incremental pHs from 9.5 to 12 (with approximately 0.5 interval step). Fig. 8 presents the results of these experiments, in which the highest recoveries were 94.91% in pH 11.1 and 92.5% in pH 10.5, for the two samples respectively. The leaching time was 24 h. 5.4. Leaching time To find out the best cyanidation time, a 48 h test was conducted for each sample. Samples (10 mL) from the clear solution obtained after vacuum filtration were used for free cyanide determination and a 30 mL sample was provided to analyze gold content. The results are presented in Fig. 9. It is seen that the trend of gold dissolution was almost identical for two samples, and there was a significant inclined rate in first 10 h whereas maximum recovery reached 92% and 90% respectively. 5.5. Kinetics of gold dissolution (sample 1)
Fig. 6. The gold recovery and passing percent versus grinding times for sample 2 (pH: 10.5, cyanide concentration: 2000 mg/L, cyanidation time: 24 h).
Fig. 7. The results of cyanide concentration for two gold samples (grinding time: 35 and 55 min for the first and second samples, pH: 10.5, cyanidation time: 24 h).
Various types of oxidizing reagent such as air, hydrogen peroxide and a combination of air and H2O2 were utilized to increase the gold dissolution rate for the first sample (Yannopoulos, 1991; Guzman
Fig. 9. The gold recovery versus cyanidation times for two samples (grinding time: 35 and 55 min for the first and second samples, cyanide concentration: 4000, 2000 mg/L for the first and second samples, pH: 11.1 and 10.5 for the first and second samples).
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Fig. 10. Gold recovery versus different leaching times without oxidizing reagent, k = 1.06, (grinding time: 35 min, cyanide concentration: 4000 mg/L, pH: 11.1).
Fig. 13. Gold recovery versus different leaching time by air, k = 1.75, (grinding time: 35 min, cyanide concentration: 4000 mg/L, pH: 11.1).
condition for cyanidation process. Cyanidation test was performed on the roasted sample in this condition and obtained results presented a considerable decrease of cyanide consumption by 1150 mg/L and increasing recovery by 5.2% (Roshan, 1990; Dunn and Chamberlain, 1997; Celep et al., 2009). It can be concluded that by increasing of roasting temperature and time, the amount of arsenic, bismuth and mercury decreased, but this trend for antimony and cadmium has a little difference which may be relevant to the boiling point of these elements in which their trends are illustrated in Figs. 14–18.
Fig. 11. Gold recovery versus different leaching time by H2O2, k = 1.35, (grinding time: 35 min, cyanide concentration: 4000 mg/L, pH: 11.1).
et al., 1999; Marsden and Lain House, 2006). The results are shown in Figs. 10–13. For each experiment, two curves are presented, the first one is related to actual data which are obtained from experiments, and the second curve is the modeled one, using first order equation as y = ymax(1 − exp(−kx)) which was generated by solver function in Excel, based on the least square criteria. 5.6. Pre-treatment experiments (of sample 1) According to the results, roasting in 1000 °C at 2 h had adequate effects on omitting of followed elements and warrants a good
Fig. 14. The content of arsenic after roasting in various temperatures and times (initial value: 16,650 ppm).
Fig. 12. Gold recovery versus different leaching time by H2O2 + air, k = 1.65, (grinding time: 35 min, cyanide concentration: 4000 mg/L, pH: 11.1).
Fig. 15. The content of antimony after roasting in various temperatures and times (initial value: 2214 ppm).
P. Karimi et al. / International Journal of Mineral Processing 95 (2010) 68–77
Fig. 16. The content of cadmium after roasting in various temperatures and times (initial value: 1.8 ppm).
Fig. 17. The content of bismuth after roasting in various temperatures and times (initial value: 1.1 ppm).
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in 50 min for the first sample. This phenomenon was due to the fact that more liberated minerals in higher grinding times were achieved and also gold grains surface was more covered by carbonaceous matter, calcite, and clay minerals (illite). Increase of cyanide concentration in leaching pulp caused high cyanide consumption due to the presence of cyanicides and did not improve gold recovery. With increasing pH, the value of free cyanide was definitely increased, whereas the concentration of free cyanide was increased from 120 to 240 mg/L at pH = 9.5 and reached to 2250 and 1750 mg/L at pH = 12 for the first and second samples respectively. Therefore, the optimum recovery was in the pH of 10.5 and 11 for both samples so that it decreased outside of this range. This could be attributed to the formation of other complexes at low pH which caused the gold recovery to be decreased. In addition, at pH higher than 11, the majority of components were not able to make stable complexes with cyanide. Hence, the free cyanide intensively increased which is followed by a recovery decrease. Also at high pH of solution, the suspended particles were deposited immediately, because of the forming of Ca(OH)2 and Mg(OH)2 complexes, as the flocculants matters. Furthermore, gold passivation at higher pH has been reported and a dramatic decrease of gold extraction in sample 2 at higher pH may be related to this. For leaching time, the curve slope in Fig. 9, which shows the recovery rate, was high at the first 8 h and gradually decreased until 24 h (changes was up to 2%). For instance, gold recoveries were achieved 92% for the first 10 h for the first sample and 89.4% for the first 8 h using second sample. Meanwhile the recoveries finally reached to 96.56% and 93.5% after 48 h respectively. Gold dissolution rate followed by 1st order kinetics presented as y = ymax(1 − exp (−kx)), in which k (rate constant) values were obtained at 1.06 and 1.16 min− 1 for the first and second samples. The results of which are shown in Table 5. The introduction of air during leaching was more effective in kinetics rate increase (1.75 min− 1), whereas using mixture of air and H2O2 yielded similar results (1.65 min− 1). In addition, the gold recoveries reached 99.03, 98.07 and 97.56% for the 48 h respectively which showed the high effect of air as an oxidizing reagent. Despite the difference in rate, the final recovery value by using H2O2 as an oxidizer was higher than other reagents. By using of HCl and HNO3 as acidic pretreatments, gold recoveries reached to 95.11% and 98.87% respectively and 97.05% by roasting of sample in 1000 °C in 2 h. In addition, cyanide consumption decreased by 340, 210 and 1150 mg/L for acidic pretreatment and roasting. The amount of arsenic, antimony, cadmium, bismuth and mercury elements decreased considerably as 88, 34, 22, 75 and 82% on the roasted sample (1000 °C in 2 h). The results of gold recovery and cyanide consumption versus variation of time and temperature of roasting process are presented in Table 6.
Fig. 18. The content of mercury after roasting in various temperatures and times (initial value: 1 ppm).
7. Summary and conclusions 6. Discussion By increasing of grinding time, cyanide consumption increased for both sample whereas 2910 mg/L in 20 min was changed to 3680 mg/L
In this study, two different types of gold samples were investigated in various cyanidation experiments. The first sample was a high grade ore and its gold and arsenic grades were 10.5 and 16676.5 g/t respectively and the second sample was a low grade ore with 2.5 g/t of
Table 5 Rate constant (k) of different oxidizing reagents for the first sample. Oxidant
Rate constant (min− 1)
Rate constant (min− 1)
Cyanide concentration (mg/L)
Cyanide consumption (mg/L)
Gold recovery (%)
Without oxidizer H2O2 H2O2 + air Air
1.06 1.35 1.65 1.75
10.5 10.5 10.5 10.5
4000 4000 4000 4000
3875 3540 3510 3650
96.56 99.03 98.07 97.56
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cyanidation condition, using hydrogen peroxide, air and mixture of H2O2 and air, were 1.06, 1.35, 1.75 and 1.65 min− 1 respectively.
Table 6 Au recovery versus roasting temperature and time. Condition Time (h)
Temperature (°C)
0.5
600 700 800 900 1000 600 700 800 900 1000 600 700 800 900 1000
1
2
Gold recovery (%)
Cyanide consumption (mg/L)
94.95 95.01 95.13 95.74 96.07 95.41 95.97 96.28 96.43 96.59 95.56 96.27 96.74 96.95 97.05
2950 2860 2740 2550 2405 2685 2480 2290 2210 2160 2605 2340 2085 2010 1940
Acknowledgments The authors would like to thank the Director of Mineral Processing Division in the Geological Survey of Iran for the support of this project; and Abedian and Movaghar Consultant, who helped in the collection of the samples and Beshkani for editing of the paper. References
Au, and low arsenic (543.3 g/t), but high silver content (160 g/t) which are presented in Table 3. The two samples were used in cyanidation experiments to determine their optimized parameters (Table 7), and the following conclusions were accordingly made. Type of detrimental minerals and their values caused to sophisticate the optimized conditions for the gold's processing circuit. Moreover detrimental matters in cyanidation process for the first sample were high amounts of arsenic, antimony, mercury, carbonaceous matters (calcite) and clay minerals (illite) which had negative effects on the process. The effects of toxic materials were caused by lower grinding time and a high amount of cyanide consumption. Ultimately the intensive effect of oxidizer on kinetic and final recovery was remarkable and optimized pH value for first sample was more than the second one, 10.5 against 11.1. Various pretreatment leaching under acidic condition, HNO3 and HCl, were studied on the first sample which its results are briefly given in Table 8. Cyanidation test was carried out using roasted sample, 2 h in 1000 °C, and its results showed that cyanide consumption was decreased considerably by 1150 mg/L and gold recovery was increased 5.2%. Different oxidizing reagents such as hydrogen peroxide (0.015 M), air (0.5 L/min) and mixture of H2O2 and air were employed to increase gold leaching kinetics for the first sample. The gold leaching followed 1st order kinetics, in which k, rate constant, value for normal
Table 7 The most suitable conditions for two different samples. Parameter
Sample 1
Sample 2
Grinding time (min) CN− concentration (mg/L) pH Cyanidation time (h) Max. gold recovery (%) Max. silver recovery (%)
35 4000 11.1 24 94.91 28.2
55 2500 10.5 24 92.5 93.5
Table 8 Results of acidic pretreatment on the sample 1. Acidic pretreatment
Decrease of cyanide consumption (mg/L)
Gold recovery (%)
HNO3 HCl
340 210
98.87 95.11
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