A kinetic and electrochemical study of the ammonia cyanide process for leaching gold in solutions containing copper

One of the biggest challenges for the gold industry in the 21st century is the presence of copper in gold containing ore bodies. This is because copper consumes large quantities of cyanide. In addition, copper cyanide species are more stable than free cyanide, and hence are problematic in events of tailings spillage. One of the methods which has been suggested for treating copper containing ores is to leach with an ammonia cyanide solution. The effect of copper and ammonia addition on gold leaching kinetics was studied in the present paper. It will be shown that when the solutions do not contain copper, the addition of ammonia decreases the rate of gold leaching. When copper is added to solution, the leach rate does decrease due to the formation of the copper cyanide complexes. However it will be shown that under conditions of zero free cyanide, gold does leach readily via the Cu(CN)₃²⁻ complex. It was found that the addition of ammonia had little effect on the leaching of gold by Cu(CN)₃²⁻, but did increase the leaching kinetics when the major cyanide species present is Cu(CN)₂⁻. Under these conditions, leaching in the absence of ammonia is very slow. The effect of copper(II) addition was also studied, and it was found that in the absence of free cyanide the presence of copper(II) increases the leach rate, provided there is enough ammonia to stabilise it against reaction with the copper(I) cyanide complexes.     

Analysis of valley-fill dump leach 2: Modeling and observations

Abstract

Valley-Fill Dump Leach 2, located at a gold mine in the Great Basin region of the United States, began operation on My 3, 1986 and was operated for six years until July 3, 1995. Run-of-mine low grade gold ore, also termed “sub-ore,” was placed in valley-fill dump (heap) leaches and processed for gold values with a cyanide lixiviant. Valley-fill dump leach 2 covered approximately 21 acres, and was comprised of 15 twelve-foot lifts containing a total of 4.5 million short tons of run-of-mine ore at an average grade of 0.037 ounces Au/short ton. Recovery was estimated at 50%, which accounts for more than 80,000 ounces of gold produced.

Available precious metals leaching models were evaluated and a modified version of the simple diffusion leaching model was compared to column test data and operational data.

Gold extraction curves from both the production lifts and the column tests compared reasonably well with the predicted gold extraction curves generated from the modified diffusion model. The major difficulty with modeling, however, was obtaining representative physiochemical characteristics of the ore. This circumstance would indicate that a sustained test program to determine these characteristics during the operation of the heap leach would be required if modeling is used as a technique to predict gold production.     

A new collector for rare earth mineral flotation

A new collector for bastnaesite flotation - modified hydroxamic acid (MOHA) has been developed through several years laboratory research. The experimental results of flotation of pure bastnaesite mineral and rare earth ores, and the application experience in flotation plants showed that MOHA is an efficient collector for bastnaesite flotation. MOHA has stronger collector ability and higher selectivity compared to the other commonly used collectors. Through the measurements of zeta-potential, adsorption and infrared adsorption spectrum and in terms of the electronegativity theory of reagent groups, the flotation mechanism has been discussed. It was concluded that the adsorption of MOHA on bastnaesite surfaces is chemical adsorption in nature through three oxygen atoms in the polar group of MOHA chelating the surface Ce(III) of bastnaesite to form a pentagon-cycle chelate: O---C=N---O---Ce(III)---). Additionally, the chemisorption is accompanied with the multilayer and non-homogeneous physical adsorption of the MOHA molecules. The MOHA adsorption equation at the surfaces of bastnaesite can be expressed as: Γ = 6.76×10⁻²C^1/1.02, while the adsorption rate constant k is: k = 2.64×10⁵min⁻¹mol⁻¹m².     

Bitumen-air aggregates flow to froth layer: II. Effect of ore grade and operating conditions on aggregate composition and bitumen recovery

Size, rise velocity, shape, and composition of bitumen-air aggregates floating to the froth layer during the processing of 11.1% bitumen Estuarine ore, and 7.2% bitumen Marine ore under various operating conditions (caustic dosage and feed density) were determined. The flux of the bitumen-air aggregates inside a Primary Separation Vessel (PSV) processing 2 tonnes/hour of the ore was in-situ monitored and video recorded. The mass of bitumen contained in the aggregates floating to the froth layer under different operating conditions was calculated on the basis of the measured rise velocities and dimensions of the aggregates. It was found that the type of ore processed was the parameter having the largest effect on size and composition of the bitumen-air aggregates floating to the froth layer inside the PSV during a 50°C warm slurry extraction process. In the case of the Estuarine ore the average mass of bitumen contained in an average aggregate was ca. 9^*10⁻⁴ g, and the average aggregate size was ca. 1.0 mm. During processing of the low grade (Marine) ore the aggregates were smaller (within 0.33–0.69 mm) and they carried much less (from 0.17–1.37^*10⁻⁴ g) bitumen, depending on other operating conditions (caustic dosage and feed density). A good correlation between size of the bitumen-air aggregates and bitumen recovery in the primary separation vessel was found. Thus, size of the aggregates can be used as an indication of better flotation recovery.     

Evaluation of leaching parameters for a refractory gold ore containing aurostibite and antimony minerals: Part I – Central zone

A cyanidation study was conducted on a mild refractory gold ore sample from the Central zone of Clarence Stream Property, owned by Freewest Resources Canada, to develop a leaching strategy to extract gold. Gold, at a grade of 8.00 g/t, is present as native gold, electrum and aurostibite. The ore also contains 2.8% pyrrhotite, together with several antimony minerals (0.8% berthierite and gudmundite, 0.18% native antimony and stibnite). It also exhibits weak preg-robbing properties with 0.16% organic carbon. Aurostibite, a gold antimony compound, is particularly known to be insoluble in cyanide solution. The antimony dissolves in cyanide solution to form antimonates, which retards gold dissolution. Industrial practice of extracting gold from aurostibite generally consists of producing a flotation concentrate, which is leached in a pipe reactor at low alkalinity and high oxygen pressure with about 20 g/L cyanide.The proposed new approach is efficient and allows the extraction of gold directly from an ore at atmospheric pressure and a low cyanide concentration at pH 10.5. The effects of grinding, pre-treatment, lead nitrate, kerosene and cyanide concentrations have been investigated. The maximum gold extraction obtained on the ore was 87.9% using 800 ppm NaCN, 500 g/t lead nitrate, 30 g/t kerosene, DO (dissolved oxygen) 10 ppm and pH 10.5 in 168 h. The associated cyanide consumption was 1.3 kg/t. The additions of lead nitrate and kerosene increased gold extraction. In comparison to a P₈₀ of 74 μm, a P₈₀ of 30 μm significantly increased gold extraction. Gold in solid solution in gudmundite and arsenopyrite was believed to be responsible for the un-leached fraction until mineralogical analysis of hydroseparation concentrates of leach residues showed that most of the un-leached gold occurs as aurostibite, either as locked grains in sulphides/sulpharsenides or as grains with passivation rims of an Au–Sb–O phase. Coarse gold was also found. Gold extraction was not sensitive to cyanide concentration from 250 to 1200 ppm NaCN and high pH was detrimental. Decreasing the cyanide concentration reduced the cyanide consumption from 1.39 to 0.85 kg/t. The removal of coarse gold using a Knelson concentrator and a Mosley table prior to leaching increased the gold extraction to 90.4% (leach residue at 0.77 g/t).     

含铜氧化金矿氨氰选择性提金试验研究及工业应用

针对含铜氧化金矿采用氨氰选择性浸出提金,考察了分段加药制度、硫酸铵用量、矿石粒度等对金浸出率及浸出液铜金比的影响。结果表明:当硫酸铵用量8.00 kg/t,氰化钠用量0.60 kg/t,石灰用量5.00 kg/t,矿浆浓度40.00%,磨矿细度-0.074 mm粒级含量不低于95.00%时,平均金、铜浸出率分别为86.66%和1.16%。工业试验连续运行70 d,氰化尾渣金品位约0.55 g/t,金吸附率99%,金解吸率99.2%,电积回收率99.5%,金精炼回收率99.5%,金锭纯度99.99%,产品金达到国标Au-1标准。     

High pressure acid leaching of a refractory lateritic nickel ore

This paper describes the experimental findings of the extraction of nickel and cobalt by high pressure acid leaching (HPAL) of a refractory limonitic nickel laterite ore from the Gördes region of Manisa in Turkey. By optimizing the basic HPAL process parameters: leaching at 255 °C with 0.30 sulfuric acid to ore weight ratio with a particle size of 100% −850μ for 1 h of leaching, it was found that 87.3% of nickel and 88.8% of cobalt present in the ore could be extracted into the pregnant leach solution (PLS). However, these extraction results were found to be relatively low compared with other similar studies. In order to understand the possible reasons for this relatively lower extraction, further investigations have shown that together with a problem related to the kinetics of the dissolution reactions, a persistent acid resistant refractory mineral present in this sample also limited the leaching process. Attempts were made with different additives to solve this problem. The effects of chemical additives such as HCl, Na₂SO₄, FeSO₄, Cu⁺ and sulfur were tested and the effect of each addition on the degree of extraction of nickel and cobalt was determined.     

Improvement of cyanidation practice at Goldcorp Red Lake Mine

Goldcorp Red Lake Mine processes a mildly refractory high grade gold ore with 77.8 g/t Au and 1.9% arsenopyrite, 1.7% pyrite and 0.2% pyrrhotite at a grind of 77% −37 microns. A cyanidation study was conducted to determine the necessary retention time of the leach circuit and the optimal gold extraction and cyanide consumption. Results have shown that the gold leaching kinetics could be significantly improved by the addition 100–200 g/t lead nitrate directly in the grinding circuit, followed by a 6 h cyanidation. A 1 month trial indicated a throughput of 810 tpd can be sustained without detrimental effect on gold extraction. However, the mill is currently limited by the underground operation to a throughput ranging from 650 to 700 tpd for a 55 h retention time. In the event of future throughput increase over 810 tpd value, lead nitrate could become an essential part of the leaching efficiency. The experimental work indicated a gold extraction at 87.1–87.4% (tailings ∼9.8 g/t Au) which is similar to plant performance. The associated cyanide consumption was 0.7 kg/t. Gold dissolution can efficiently be conducted at an average cyanide concentration of 400 ppm. Previous control of cyanide concentration in the plant was 700 ppm in the first leach tank to 350 ppm NaCN at the end. Testwork realised in the mill allowed a 32% cyanide reduction predicted by the on-site cyanidation study. The new cyanide concentration is ranging between 500 ppm in the first leach tank and 300 ppm NaCN at the end of the leaching circuit. The cyanide consumption of the plant was reduced from 1.0 to 0.68 kg/t with reduction of effluent treatment costs. Grinding finer (91% −37 microns) slightly increased the gold extraction (tailings lower by 0.4 g/t) but lowers the leaching kinetics, increases cyanide consumption (35%) and the arsenic dissolution.     

Initial loss of cyanide, thiocyanate, and thiosulfate adjuvants following amendment to an oxidic gold ore

Gold extraction and gold phytomining involve the use of adjuvants such as cyanide, thiocyanate, or thiosulfate to facilitate the dissolution of gold and its subsequent recovery. The adjuvants can be applied at high rates, in excess of 1 kg t⁻¹ for some ores. At these concentrations, concerns have been expressed about adjuvant persistence and associated detrimental repercussions for the environment. The incubation study here examined the initial disappearance of these adjuvants in the dark following their amendment at 1 g kg⁻¹ gold ore. Over a 24 d period, nearly 75% of the initial thiocyanate was lost within 24 d while >50% of thiosulfate and cyanide was lost over the same time frame. Cyanide and thiosulfate loss showed an exponential loss while thiocyanate was observed to follow a biphasic relationship. For the oxidic ore used, the adjuvants were not inherently persistent, suggesting that natural attenuation may rapidly promote adjuvant loss under specific conditions.     

Enhanced leachability of gold and silver in cyanide media: Effect of alkaline pre-treatment of jarosite minerals

Composite samples of tailings containing gold (1.35 g/t) and significant amounts of silver (155 g/t) were subjected to batchwise cyanide leaching to assess the feasibility of extracting gold and silver. The tailings are waste solids arising from flotation and leaching operations whereby the flotation product (sphalerite concentrate) is calcined and then solubilised into dilute sulphuric acid solution and eventually sequestered from the electrolyte by electrowinning. Silver and gold are part of the zinc refinery residue, flotation tailings and to a limited extent the calcine leach tailings. Mineralogical results showed that composite tailings are refractory in nature (44% quartz, 17% silico aluminates and 12% jarosites).The concept of enhancing gold and silver recovery from the tailings focused on firstly decomposing the jarosite minerals by alkaline pre-treatment and then secondly leaching with cyanide solution. These two steps ensured that free gold and silver found in the zinc refinery residue and in the jarosite minerals could be leached simultaneously. The composite tailings were treated with Ca(OH)₂ solutions and then heated to 90 °C for 2 h to decompose the silver-bearing mineral (Ag,PbFe₃(SO₄)₂(OH)₆). The alkaline pre-treated tailings were then subjected to cyanide leach tests at different NaCN dosages (2.5–10 kg/t) and particle size (96–200 μm). Without an alkaline pre-treatment stage, leach efficiencies achieved were 41% and 25% for gold and silver, respectively at 40 °C and 8 h mixing time. But, better leach efficiencies (55% for Au, 81% for Ag) were achieved after the feed was pre-treated with Ca(OH)₂. The leaching mechanism of gold was explained by the shrinking sphere model denoted by surface chemical reaction.     

Evaluation of equilibrium, kinetic and thermodynamic parameters for biosorption of nickel(II) ions onto bacteria strain, Rhodococcus opacus

The present study explores the ability of Rhodococcus opacus bacteria strain for the removal of nickel(II) under different experimental conditions. The effects of relevant parameters such as solution pH, biomass loading, ionic strength, and temperature on nickel(II) biosorption capacity were evaluated. Although the tests were carried out using synthetic solutions, the results are representative of typical waste effluents from chemical, dyestuffs, battery manufacture, porcelain enameling, metallurgical industries and electroplating circuits. The sorption data followed the Langmuir, Freundlich, Temkin, Dubinin–Radushkevich, and Sips isotherms. The maximum sorption capacity was found to be 7.63 mg g⁻¹ at pH 5. Thermodynamic parameters have also been evaluated and the results show that the sorption process was spontaneous and endothermic in nature. The dynamics of the sorption process were studied and the values of rate constants for different kinetic models were calculated. The activation energy of biosorption (E_a) was determined as 12.56 kJ mol⁻¹ using the Arrhenius equation.     

Investigations of accelerating parameters for the biooxidation of low-grade refractory gold ores

This research was designed to accelerate the biological oxidation rate of low-grade suphidic gold ores. The biological oxidation by tamed Thiobacillus ferrooxidans bacteria with a mixed low-grade refractory gold ore was investigated. These bacteria could well tolerate up to 18.0 g/L of arsenic (III). The effect on the oxidation rate due to the presence of a magnetic field on the culture medium was examined as well as the influence of surfactants (tween-80, tween-20 and emulsifier-OP) and selected ionic metals such as Ag⁺, Bi³⁺, Co²⁺ and Hg²⁺. The results showed that, based on a 60% oxidation rate for arsenic, the oxidation time was shortened by 83 hours, 90 hours, 98 hours and 100 hours under the optimal conditions of tamed bacteria, magnetized water, concentrations of tween-80 and Ag⁺ respectively. Under the combination of the above parameters, the biooxidation rate was shortened by 120 hours to achieve a 60% oxidation rate for arsenic. However it did not produce a simple additive effect on the biooxidation.     

A QEM*SEM study of a suite of pressure leach products from a gold circuit

High lime consumption during the neutralisation of oxidised solids at the Lihir gold operation can be related to both ore mineralogy and conditions within the autoclaves. During the pressure oxidation process, basic iron sulphate and potassium jarosite precipitated and silicate clay minerals were dissolved. Illite in the plant feed was identified as a source of potassium for potassium jarosite precipitation — a greater extraction of which was obtained during a period of ‘low’ lime consumption. By contrast potassium feldspar was found to be largely un-reactive.The dissolution of basic iron sulphate under alkali conditions resulted in the consumption of lime. The difference between the conversion of 4% and 8% of the total solid, by weight, from basic iron sulphate to ferric hydroxide can account for either ‘low’ or ‘high’ lime consumption. Potassium jarosite was largely non-reactive during neutralisation.The quantity of basic iron sulphate present in the autoclave discharge is likely to be a function of both ore mineralogy and autoclave conditions. These mineralogical changes occur mostly in the ultrafine, −10 μm size fraction.In order to control the consumption of lime in the plant, it is clear that during pressure oxidation the formation of basic iron sulphates needs to be minimised and that of potassium jarosite promoted.     

Enhanced hydrochloric acid leaching of nickel

Dissolution rates of nickel in hydrochloric acid with and without the presence of thiourea have been studied by means of rotating disc methods with solution analysis. In the absence of thiourea an apparent activation energy value of 61.1 ± 12.1 kJ. mol⁻¹ was determined. Such a high value is inconsistent with a diffusion controlled process and this was supported by the lack of dependency of the dissolution rate on rotational speed. A reaction controlled process is indicated. In 8.4 mol. dm⁻³ hydrochloric acid the thiourea concentration range producing optimum dissolution rates was 10⁻³ to 10⁻² mol. dm⁻³. At 70° C the rate of leaching was enhanced by some 105 %. Tests using radiochemical labelling of the sulphur and of the carbon separately in the thiourea molecule confirmed previous proposals, for cobalt dissolution, that hydrogen sulphide is produced as a by-product of the cathodic depolarisation reaction involving the thiourea and is adsorbed on the metal surface in the molecular state, forming activated anodic sites which enhance anodic dissolution. At the 10⁻¹ mol. dm⁻³ thiourea concentration level stimulation of the nickel dissolution process gave way to its partial inhibition with a maximum diminution at 30°C of 50 %.     

Leaching of manganese ores using sawdust as a reductant

In this study sawdust was used as reductant for sulphuric acid leaching of manganese ore. Effects of pulp density, amount of acid, temperature, particle size of ore and amount of sawdust were studied. Manganese extraction of ∼98% was achieved under the conditions: leaching time 8 h, 5% H₂SO₄ (v/v), 10% pulp density, 90 °C and 5% sawdust (w/w), i.e. 0.5 g/g ore. Other Mn containing materials like low grade manganese ore, manganese nodule and Mn-nodule leach residues were tested and all these materials responded well giving more than 98% Mn extraction.     

响应面法优化微生物浸出低品位钼矿工艺条件

从钼矿区矿坑水中分离得到自然混合菌种,经驯化培养,用于浸出陕西洛南低品位钼矿。以浸出液中钼的含量作为评判指标,在单因素试验的基础上,通过响应面法优化得出了最佳浸出条件为:初始pH值2.0、Fe²⁺浓度5.00 g/L、Fe³⁺浓度2.90 g/L,此条件下钼浸出率为73.87%,而相应的无菌化学浸出实验的钼浸出率仅为12.65%。拟合得到二次回归方程模型的相关系数为R²=0.991 1,P<0.000 1,说明回归效果很好,失拟项P=0.671 0>0.05,失拟不显著,误差小,可用该模型对钼的微生物浸出率进行分析和预测。     

Solvent extraction of gold(I) from alkaline cyanide solutions by the cetylpyridinium bromide/tributylphosphate system

This paper explores the solvent extraction of KAu(CN)₂ from alkaline cyanide solutions using quaternary ammonium cetylpyridinium bromide (CPB) as an extractant with the addition of tributylphosphate (TBP) as a modifier. It also investigates the influence of several variables on gold extraction, including the molar ratio (β) of CPB to Au(I), the volume percentage of TBP (φ_TBP), NaCl concentration, phase ratio (AQ/ORG), and gold concentration in the aqueous phase. The results indicate that nearly all of the Au(I) (>98%) was transferred from the aqueous phase into the organic phase when β = 1 and φ_TBP = 30 vol%. We also carried out experiments for treating 20 L synthetic aurocyanide solution containing 10 mg/L Au(I) with column-shaped extraction equipment. The results demonstrated the recovery of more than 94.5% of Au(I) after two successive stages of extraction, and the Au(I) concentration in the raffinate was less than 0.5 mg/L. KSCN solution was used to strip the gold-loaded organic phase, and about 90% of Au(I) could be reverse extracted into the aqueous phase when the KSCN concentration reached 3.0 mol/L. The results obtained in this paper establish that the CPB/TBP extraction system has potential for practical application in the extraction and separation of gold from alkaline aurocyanide solutions.     

Potential use of carbon felt in gold hydrometallurgy

The use of carbon felt as a three-dimensional electrode appears to be very promising for the recovery of heavy metals, and toxic compounds removal from dilute solutions, considering its favourable physico-chemical properties: high specific surface area, good fluid permeability and compressibility, chemical inertness and good electrical conductivity.This work presents the contribution of the carbon felt electrode in two different steps of the gold cyanidation recovery process

• &#x02022;- firstly as a cathode for the electrowinning of gold solution obtained after the elution of loaded carbon;
• &#x02022;- secondly as an anode for the electro-oxidation of cyanide ions present at low concentration (200–300mg/l) in waste streams.

In the first case, more than 10 kg of gold per m² of felt (2000 kg of gold per m³ of felt) can be loaded at 400 A/m² from dilute gold solutions (30 mg/l Au) with classic Faradic yields (6–12 %) and high overall extraction efficiency (>90 %). The felt homogeneously loaded with adhesive gold deposit can be smelted in an electric furnace without addition of fluxes since the carbon felt decomposes in gaseous products without formation of ashes.In the second case, the cyanide ion concentration can be lowered to as low as 10 mg/l by electro-oxidation to cyanate form (CNO⁻) at 400 A/m² in the anodic compartment of a divided cell. The presence of copper ion significantly improves the current efficiency of this electro-oxidation and sharply reduces the oxidation of the carbon felt. Copper ions accelerate the oxidation rate of free cyanide ions through the formation of easily oxidizable complexes owing to a mechanism in which the Cull Cu^ll redox couple acts as an electron transfer mediator.     

膜技术处理金矿氰化浸出溶液的应用

综述了处理金矿氰化浸出溶液的技术研究现状，分析现有技术的优缺点，介绍了电渗析法和液膜分离法在处理金矿氰化浸出溶液的应用情况，对该领域的发展现状及目前面临的主要问题作了简要的概述，并对未来的研究趋势进行了展望。     

The mechanism of leaching of gold from refractory ores

Refractory gold ores commonly contain free gold, submicroscopic gold, base metal sulphides, pyrite, pyrrhorite, labile sulphides and carbonaceous material. The analysis of gold deportment with the diagnostic leaching technique has opened a research field to gain more insight into the fundamentals of the kinetic behaviour of gold dissolution from refractory ores. Hence, more detailed studies on the kinetics of the reactions that occur during leaching with cyanide are now possible.Previous research work has indicated that the dissolution rate of gold during cyanidation becomes depressed after a certain time with no further increase in the rate. However, a second noticeable increase in the gold dissolution rate has been observed after a leaching time of 5 to 12 hours during some of the cyanidations, following the destruction of a mineral with the diagnostic leaching technique. If such a second increase in dissolution rate does not occur, the re-leaching of the filtered solids with a fresh cyanide solution yields an additional gold extraction of 4 to 8 %.Various experimental results have indicated that a passivative film forms on the surface of some of the liberated gold. The selective destruction of the various minerals with oxidative acid leaches destroys and/or decomposes certain minerals which may form films on the gold surface by precipitation. The chemical composition of these films and precipitates depends on the mineralogy of the sample. These films may be oxides, sulphides, carbonates and cyanide complexes.The complexes can be destroyed, depending on the nature of the film, by interstage dilute acid and/or cyanide washes in an agitated vessel. The destruction of the films exposes the gold surface for cyanidation. Whereas most of the previous studies on the leaching of gold have focused on the whole ore, the emphasis in this study was on the leaching behaviour of gold from various ore constituents. This approach of studying the leaching behaviour of different gold bearing minerals has provided reasons why some ores leach better than others.