从难处理硫化银物料中提取银新工艺研究

研究了硫氰酸铵水溶液氧压浸出-铁粉置换法从湿法炼锌热酸浸出渣浮选银精矿、硫化锰银矿脱锰后渣中提银新工艺。试验结果表明,浮选银精矿与脱锰渣的银浸出率分别为大于95%与接近88%。     

Enrichment of platinum group metals (PGMs) by two-stage selective pressure leaching cementation from low-grade Pt-Pd sulfide concentrates

A hydrometallurgical process was developed for producing Pt-Pd enriched concentrates from low-grade sulfide concentrates on-site at Jinbaoshan mine in an isolated area in China. The developed process (pressure acidic leaching-pressure cyanidation leaching) includes the following two steps. (1) The flotation concentrates are treated by acidic pressure leaching to selectively dissolve all the base metals (Cu, Ni, Co) while leaving most (85 to 94 pct) of the precious metals in the iron residues. The leaching solution is then processed by copper cementation and solvent extraction (SX) to recover PGMs and CuNiCo, respectively. (2) The iron residues are treated by pressure cyanidation leaching to selectively dissolve precious metals. The cyanidation solution is then cemented by zinc power reduction to produce Pt-Pd concentrates. Testing results from both lab and pilot (5 kg/batch, 50 L autoclave) scale tests show that Pt+Pd content has been increased from 86 g/t (flotation concentrate) to 56 to 59 wt pct (zinc cementation residue) with an extraction recovery of 95.8 pct. The Pt-Pd enriched cementation residues can be sold as Platinum Group Metal (PGM) concentrates to refineries for further refining.     

某毒砂金矿硫氰酸盐氨性体系加压氧化提金探索试验

对某毒砂金矿进行了硫氰酸盐氨性体系氧压提取金的探索试验,考察了反应温度、Cu2+浓度、浸出时间、液固比、氨水浓度、氧分压和硫氰酸铵浓度等对金浸出率的影响。结果表明,在下述优化条件下金浸出率为61.7%,即硫氰酸铵浓度3mol/L,液固比5∶1,反应温度150℃,浸出时间6h,搅拌速度750r/min,氨水浓度4.64mo/L,铜加入量1.5g/L。而经400℃焙烧预处理后金浸出率达到86.2%。     

Leaching of a sphalerite concentrate with H2SO4–HNO3 solutions in the presence of C2Cl4

The enhanced leaching of sphalerite concentrates in H₂SO₄–HNO₃ solutions and the extraction of sulfur with tetrachloroethylene were studied. Variables of the process were investigated including leaching temperature, reaction time, liquid / solid ratio, and tetrachloroethylene concentration. The number of cycles that tetrachloroethylene could be recycled did not have a significant effect on zinc extraction. The results indicated that 99.6% zinc extraction was obtained after three hours of leaching at 85 °C and 0.1 MPa O₂, when 20 g of sphalerite concentrate were leached in a 200 ml solution containing 2.0 mol/L H₂SO₄ and 0.2 mol/L HNO₃, in the presence of 10 ml C₂Cl₄. Leaching rates were significantly improved under these conditions.     

Determination of the Optimum Conditions for Leaching of Zinc Cathode Melting Furnace Slag in Ammonium Chloride Media

This research is part of a continuing effort to leach zinc from zinc cathode melting furnace slags (ZCMFSs) to produce zinc oxide. The slag with an assay of 68.05 pct Zn was used in ammonium chloride leaching for zinc extraction. In this paper, the effects of influential factors on extraction efficiency of Zn from a ZCMFS were investigated. The Taguchi’s method based on orthogonal array (OA) design has been used to arrange the experimental runs in order to maximize zinc extraction from a slag. The softwares named Excel and Design-Expert 7 have been used to design experiments and subsequent analysis. OA L ₂₅ (5⁵) consisting of five parameters, each with five levels, was employed to evaluate the effects of reaction time (t = 10, 30, 50, 70, 90 minutes), reaction temperature [T = 313, 323, 333, 343, 353 (40, 50, 60, 70, 80) K (°C)], pulp density (S/L = 20, 40, 60, 80, 100 g/L), stirring speed (R = 300, 400, 500, 600, 700 rpm), and ammonium chloride concentration (C = 10, 15, 20, 25, 30 pctwt), on zinc extraction percent. Statistical analysis, ANOVA, was also employed to determine the relationship between experimental conditions and yield levels. The results showed that the significant parameters affecting leaching of slag were ammonium chloride concentration and pulp density, and increasing pulp density reduced leaching efficiency of zinc. However, increasing ammonium chloride concentration promoted the extraction of zinc. The optimum conditions for this study were found to be t ₄: 70 minutes, T ₅: 353 K (80 °C), (S/L)₂: 40 g/L, R ₃: 500 rpm, and C ₄: 25 pctwt. Under these conditions, the dissolution percentage of Zn in ammonium chloride media was 94.61 pct.     

Leaching of a sphalerite concentrate with H2SO4–HNO3 solutions in the presence of C2Cl4

The enhanced leaching of sphalerite concentrates in H₂SO₄–HNO₃ solutions and the extraction of sulfur with tetrachloroethylene were studied. Variables of the process were investigated including leaching temperature, reaction time, liquid / solid ratio, and tetrachloroethylene concentration. The number of cycles that tetrachloroethylene could be recycled did not have a significant effect on zinc extraction. The results indicated that 99.6% zinc extraction was obtained after three hours of leaching at 85 °C and 0.1 MPa O₂, when 20 g of sphalerite concentrate were leached in a 200 ml solution containing 2.0 mol/L H₂SO₄ and 0.2 mol/L HNO₃, in the presence of 10 ml C₂Cl₄. Leaching rates were significantly improved under these conditions.     

Catalytic-Oxidative Leaching of Low-Grade Complex Zinc Ore by Cu (II) Ions Produced from Copper Ore in Ammonia-Ammonium Sulfate Solution

The catalytic-oxidative leaching of a mixed ore, which consists of low-grade oxide copper ore and oxide zinc ore containing ZnS, was investigated in ammonia-ammonium sulfate solution. The effect of the main parameters, such as mass ratio of copper ore to zinc ore, liquid-to-solid ratio, concentration of lixivant, leaching time, and temperature, was studied. The optimal leaching conditions with a maximum extraction of Cu 92.6?pct and Zn 85.5?pct were determined as follows: the mass ratio of copper ore to zinc ore 4/10?g/g, temperature 323.15?K (50?°C), leaching time 6?hours, stirring speed 500?r/min, liquid-to-solid ratio 3.6/1?cm³/g, concentration of lixivant including ammonia 2.0?mol/dm³, ammonium sulfate 1.0?mol/dm³, and ammonium persulfate 0.3?mol/dm³. It was found that ZnS in the oxide zinc ore could be extracted with Cu(II) ion, which was produced from copper ore and was used as the catalyst in the presence of ammonium persulfate.     

Mineralogical characterization of silver flotation concentrates made from zinc neutral leach residues

Silver flotation concentrates prepared from high-silver (1480 ppm Ag) and low-silver (300 ppm Ag) neutral leach residues have been examined mineralogically to determine the phases present and to elucidate the behavior of silver during zinc processing. The flotation concentrates consist principally of sphalerite although lesser amounts of zinc ferrite and PbSO₄, as well as traces of other phases, also are present. In the high-silver flotation concentrate, silver occurs mostly as Ag₂S or (Ag, Cu)₂S rims on sphalerite although (Ag, Cu)₂S inclusions within sphalerite also are present. Trace amounts of a Cu-Ag-S-Cl phase are present on rare copper oxide grains, and this silver-bearing phase may be a fine mixture of Ag₂S, AgCl, and Cu₂S. In the low-silver flotation concentrate, silver occurs mostly as Ag₂S although traces of silver-bearing CuS and Cu₂S also are present. The Ag₂S occurs as <1 μm particles disseminated in elemental sulfur-silica gel patches, as discontinuous rims or isolated patches on sphalerite grains, and as tiny free particles. Silver chloride was not detected. These studies suggest that silver dissolves during neutral leaching and subsequently reacts with sphalerite or other sulfides to form silver sulfide.     

Reflotation of Mt. Isa Zinc Concentrate after SO2 Leach Treatment

Laboratory flotation tests were carried out on the SO₂ leaching residues of zinc concentrates of Mt. Isa lead–zinc concentrator. In the evaluation of these tests grades, recoveries and separation index values of the metallic elements are used. Separation index value of a metallic element is a new concept proposed in this paper and is useful in the evaluation of any concentration process. In the bulk concentrates obtained after flotation tests, the grades, recoveries, and separation index values of silver were around 900–1000 ppm, 18–20%, and 17–19, respectively.     

锌加压酸浸渣中硫磺的回收方法研究

对比分析了浮选法、热过滤法和硫化铵法回收锌加压酸浸渣中硫磺的优缺点。考察了硫化铵溶液浸出浮选硫精矿、硫化物滤饼和多硫化铵母液热分解过程的影响因素。结果表明,液固比和硫化铵浓度对硫磺浸出效果影响较为明显,在最佳试验条件下硫化物滤饼中硫的浸出率约为95%,浮选硫精矿中硫的浸出率和回收率均达到98%,多硫化铵母液热分解后获得的硫磺产品纯度高达99.57%。硫化铵浸出渣中有价金属富集倍数较高,有利于锌加压酸浸渣的综合利用。     

Sulphuric acid leaching of polymetallic nodules using paper as a reductant

Sulphuric acid leaching of manganese nodule for extraction of Cu, Ni, Co and Mn in presence of a novel cellulosic reductant, waste newspaper, has been reported. The various parameters chosen for the study were: time, temperature, H₂SO₄ concentration and amount of paper. To quantify the linear and interaction coefficients a 2³ full factorial design of experiments was followed. The regression equations were determined and the adequacy of these equations was tested by Fischer’s test. The linear coefficients for time, acid concentration and amount of paper were found to be significant for extraction of Cu, Ni, Co and Mn. While acid concentration and amount of paper showed positive interactions for Cu, Ni and Co extractions it had negative significance for Mn dissolution. Under the conditions for >97% extraction of metal values iron extraction was ∼40%. In order to reduce the iron contamination by discarding iron as jarosite effect of addition of ammonium sulphate during leaching was also studied. Iron extraction could be brought down to 14% from ∼40% with the addition of 30g/L (NH₄)₂SO₄ without affecting the recoveries of other metals.     

Selective and Efficient Extraction of Zinc from Mixed Sulfide–oxide Zinc and Lead Ore

Mixed sulfide–oxide lead and zinc ores are generally composed of both sulfides and oxides. The dissolution of sulfides is more difficult than oxides thus the addition of oxidant is necessary. In this paper, oxidative leaching of mixed ore in NH₃-(NH₄)₂SO₄ solution using ammonium persulfate as oxidant under atmospheric pressure and relatively low temperature was investigated for the first time. The effects of factors on the leaching of pure ZnS were studied and the optimal conditions with zinc 98.7% were determined. Selective and efficient extractions of 93.9% and 94.9% zinc from zinc sulfide ore and mixed ore were also achieved, respectively.     

Mineralogical characterization of silver flotation concentrates made from zinc neutral leach residues

Silver flotation concentrates prepared from high-silver (1480 ppm Ag) and low-silver (300 ppm Ag) neutral leach residues have been examined mineralogically to determine the phases present and to elucidate the behavior of silver during zinc processing. The flotation concentrates consist principally of sphalerite although lesser amounts of zinc ferrite and PbSO₄, as well as traces of other phases, also are present. In the high-silver flotation concentrate, silver occurs mostly as Ag₂S or (Ag, Cu)₂S rims on sphalerite although (Ag, Cu)₂S inclusions within sphalerite also are present. Trace amounts of a Cu-Ag-S-Cl phase are present on rare copper oxide grains, and this silver-bearing phase may be a fine mixture of Ag₂S, AgCl, and Cu₂S. In the low-silver flotation concentrate, silver occurs mostly as Ag₂S although traces of silver-bearing CuS and Cu₂S also are present. The Ag₂S occurs as <1 μm particles disseminated in elemental sulfur-silica gel patches, as discontinuous rims or isolated patches on sphalerite grains, and as tiny free particles. Silver chloride was not detected. These studies suggest that silver dissolves during neutral leaching and subsequently reacts with sphalerite or other sulfides to form silver sulfide.     

Processing of Tungsten Preconcentrate from Low Grade Ore to Recover Metallic Values

Abstract

For the exploitation of a low grade tungsten deposit of Tapaskonda, A.P., India containing 0.1-0.16% WO₃, preconcentrate with 13.06% WO₃ was produced by physical beneficiation. This concentrate containing ferberite along with small amount of wolframite minerals was treated following two routes viz. soda ash roast-leach and alkali pressure leach processes. The parameters such as time, temperature, concentration of alkali, etc., have been studied to optimise the recovery of tungsten. In the soda roast -leach process about 89% tungsten was recovered by roasting the concentrate at 1073 K for 4h under the oxidising conditions. In the alkali pressure leaching process, tungsten recovery was 94% at 463K for 20bar pressure, 100g/L sodium hydroxide and 60 minutes time. The leaching kinetics followed diffusion control model with lixiviant reacting the mineral phase through the porous product layer. An activation energy of 31.2kJ/mole was acquired in the temperature range 428-473 K. The leach liquor was purified with respect to different impurities by a two-stage precipitation process. Tungsten from the purified leach solution was extracted by 10% Alamine-336 and 10% isodecanol in kerosene. The loaded metal when stripped with NH₄OH produced ammonium paratungstate (APT) solution which was crystallised to get the crystal of APT. The alkali pressure leach-solvent extraction process was thus found attractive for treating such concentrates.     

Dissolution kinetics of malachite in ammonia/ammonium carbonate leaching

The leaching of oxide copper ore containing malachite, which is the unique copper mineral in the ore, by aqueous ammonia solution has been studied. The effect of leaching time, ammonium hydroxide, and ammonium carbonate concentration, pH, [NH₃]/[NH₄⁺] ratio, stirring speed, solid/liquid ratio, particle size, and temperature were investigated. The main important parameters in ammonia leaching of malachite ore are determined as leaching time, ammonia/ammonium concentration ratio, pH, solid/liquid ratio, leaching temperature, and particle size. Optimum leaching conditions from malachite ore by ammonia/ammonium carbonate solution are found as ammonia/ammonium carbonate concentrations: 5 M NH₄OH+0.3 M (NH₄)₂CO₃; solid/liquid ratio: 1:10 g/mL; leaching times: 120 min; stirring speed: 300 rpm; leaching temperature: 25 °C; particle size finer than 450 μm. More than 98% of copper was effectively recovered. During the leaching, copper dissolves as in the form of Cu(NH₃)₄⁺² complex ion, whereas gangue minerals do not react with ammonia. It was determined that interface transfer and diffusion across the product layer control the leaching process. The activation energy for dissolution was found to be 15 kJ mol⁻¹.     

The effect of microwave irradiations on the leaching of zinc from bulk sulphide concentrates produced from Rampura–Agucha tailings

This work reports the effect of microwave irradiation on extraction of zinc from a bulk sphalerite (ZnS)/pyrrhotite (FeS) concentrate produced by flotation of tailings from the Rampura Agucha Lead–Zinc mines in India. This material could not be treated economically by conventional hydrometallurgy or pyrometallurgical methods due to low zinc concentration. Consequently, microwave assisted leaching was tested. Zinc leaching in sulphuric acid was rapid with > 90% dissolution after 6 min, further irradiation increased the zinc to 96% after 16 min. Iron dissolution increased up to 12 min irradiation and then decreased due to precipitation. Power consumption for > 90% Zn recovery was 0.36 kWh/kg concentrate. The leach solution contained 50–55 g/L zinc and 18–20 g/L iron which could be reduced to 0.54 g/L by jarosite precipitation. The ratio of Zn/Fe leached was 3.4 compared with the ratio of 0.6 for the concentrate showing significant selectivity for zinc over iron using this method.     

Copper recovery from chalcopyrite concentrates by the BRISA process

The technical viability of the BRISA process (Biolixiviación Rápida Indirecta con Separación de Acciones: Fast Indirect Bioleaching with Actions Separation) for the copper recovery from chalcopyrite concentrates has been proved. Two copper concentrates (with a copper content of 8.9 and 9.9 wt.%) with chalcopyrite as the dominant copper mineral have been leached with ferric sulphate at 12 g/L of ferric iron and pH 1.25 in agitated reactors using silver as a catalyst. Effects of temperature, amount of catalyst and catalyst addition time have been investigated. Small amounts of catalyst (from 0.5 to 2 mg Ag/g concentrate) were required to achieve high copper extractions (>95%) from concentrates at 70 °C and 8–10 h leaching. Liquors generated in the chemical leaching were biooxidized for ferrous iron oxidation and ferric regeneration with a mixed culture of ferrooxidant bacteria. No inhibition effect inherent in the liquor composition was detected. The silver added as a catalyst remained in the solid residue, and it was never detected in solution. The recovery of silver may be achieved by leaching the leach residue in an acid-brine medium with 200 g/L of NaCl and either hydrochloric or sulphuric acid, provided that elemental sulphur has been previously removed by steam hot filtration. The effect of variables such as temperature, NaCl concentration, type of acid and acidity–pulp density relationship on the silver extraction from an elemental sulphur-free residue has been examined. It is possible to obtain total recovery of the silver added as a catalyst plus 75% of the silver originally present in concentrate B (44 mg/kg) by leaching a leach residue with a 200 g/L NaCl–0.5 M H₂SO₄ medium at 90 °C and 10 wt.% of pulp density in two stages of 2 h each. The incorporation of silver catalysis to the BRISA process allows a technology based on bioleaching capable of processing chalcopyrite concentrates with rapid kinetics.     

Leaching of sulfidized chalcopyrite with H<Subscript>2</Subscript>SO<Subscript>4</Subscript>-NaCl-O<Subscript>2</Subscript>

The recovery of copper from chalcopyrite by leaching is complex not only due to the slow dissolution kinetics of this mineral in most aqueous media but also due to the production of solutions that are heavily contaminated with iron. On the contrary, the leaching of sulfidized chalcopyrite is very attractive because of a faster and more selective dissolution of copper compared to the leaching of the untreated chalcopyrite. In this work, the results of leaching in H₂SO₄-NaCl-O₂ solutions of sulfidized chalcopyrite concentrate are discussed. Experiments were carried out with chalcopyrite concentrates previously reacted with elemental sulfur at 375 °C for 60 minutes. The results showed that the concentration of chloride ions below 0.5 M, temperature, and leaching time are important variables for the extraction of Cu. On the other hand, Fe extraction was little affected by the same variables, remaining below 6 pct for all the experimental conditions tested. Microscopic observations of the leached particles showed that the elemental sulfur produced by the reaction does not form a coherent layer surrounding the particle, but rather concentrates in certain locations as large clusters. The leaching kinetics can be accurately described by a nonreactive core-shrinking rim topochemical expression for spherical particles 1 − (1 − 0.45X)^1/3=kt. The activation energy found was 76 kJ/mol for the range 85 °C to 100 °C.