The leaching of copper from sulphur activated chalcopyrite with cupric sulphate in nitrile-water mixtures

If chalcopyrite is roasted with sulphur at 400–450°C pyrite and idaite or bornite are produced. Bornite plus pyrite are also prepared by roasting a 1:1 mixture of chalcopyrite and covellite. These copper-iron sulphides were leached with acidified aqueous cupric sulphate solutions containing acetonitrile or hydracrylonitrile and the results are compared with leaching with acidified cupric chloride in brine. The nitrile route has the advantage of a less corrosive sulphate medium for subsequent copper recovery processes.Bornite appears to be the most attractive product from the roasting of sulphur and chalcopyrite because much of its copper can be readily leached. Iron reports to the solution only in the latter stages of extraction. Up to 80% of the copper in this bornite is leached with CuSO₄/RCN/H₂O at 60°C. Copper is recovered from the resulting cuprous sulphate solution by electrowinning with inert anode. The products are copper cathodes and cupric sulphate, which is recycled. The leach residue may be used to reactivate further chalcopyrite or is leached of its copper by established routes.     

Use of X-ray diffraction in a study of ammonia leaching of multimetal sulfides

X-ray diffraction has been used to study the changes in mineralogy that occur during ammonia leaching of sulfide minerals and complex bulk sulfide concentrates. Leaching results in high extraction rates (>90 pct) of copper from chalcopyrite, zinc from sphalerite, and lead from galena. However, under experimental leaching conditions (temperature, 115 °C to 135 °C; par-tial pressure of oxygen, 1.5 kg/cm²; pH ∼ 10.0), the pyrite grains are practically inert. Ap-parently, the amount of pyrite in leach residue is constant in absolute terms. However, its relative percentage changes because the amount of copper and zinc minerals is reduced in the leach residue during progressive leaching. The products formed during the leaching reaction, such as goethite and lead sulfate, tend to increase the weight of the leach residue, and thus the relative weight of pyrite remains nearly unchanged. The ratios of selected line pair intensities of pyrite lines and characteristic (selected) lines of chalcopyrite, sphalerite, and galena are used to establish the oxidative ammonia leaching kinetics of Cu-Zn-Pb bulk concentrates. That is, the variation in the line pair intensity ratios, with time, correlates with the changes in the el-emental concentrations in the leach liquor.     

硫化铅精矿三氯化铁浸出新工艺研究

研究了三氯化铁浸出硫化铅精矿,并在浸出后期沉淀银、铜、铋。通过试验,得出新工艺的最佳工艺条件为：三氯化铁质量浓度为140 g/L,氯化钠浓度为6 mol/L,浸出时间2 h,浸出温度90℃,沉淀剂用量为理论量的1.6倍,沉淀反应温度85℃,沉淀反应时间10 min。在此条件下,铅的浸出率达到98.41%,浸出渣中硫、银、铜、铋的富集率分别达90%,99%,98.5%,97%以上。     

用湿法处理BK铅厂阳极泥的研究

研究了处理BK铅厂阳极泥的湿法冶金过程。试验过程包括三部分:氧化浸出阳极泥,对浸出渣进行了处理回收银和金,对浸出液进行处理回收铜、铋和锑。试验结果表明,银和金的回收率分别为96%和86%;银产品的纯度达99.6%;锑、铋和铜的回收率分别为93.2%、90.0%和86.3%。最后的含砷溶液加入石灰和铁盐加以处理,废水达到了排放标准。     

锌精矿多种金属分离新工艺研究

以内蒙较大型锌矿选矿精矿为研究对象,最终确立焙烧-硫酸浸锌、铜-氰化浸银工艺。Zn、Cu、Ag浸出率分别为95.01%、94.13%、89.10%,并具有试剂耗量低、技术简单、易实施等优势。     

A ferrous chloride-oxygen leach process for nickel-copper sulphide concentrations

This paper describes a ferrous chloride-oxygen leach process for recovery of nickel and copper values from sulphide concentrates available in India. The sulphide concentrates aareleached with a stirred solution of ferrous chloride in a glass-lined reactor operated at various temperatures and oxygen pressures. In this single step process, copper and nickel are converted into their water soluble chloride forms, whereas iron is rejected as hydrated iron oxide with simultaneous generation of sulphur in the non-polluting elemental form. The influence of various parameters such as (i) amount of ferrous chloride (60–100% stoichiometric), (ii) oxygen pressure (0.308–0.515 MPa), (iii) leaching temperature (90–120°C) and (iv) duration of leaching (2–10 h) on the leaching process has been examined. It has been possible to recover 94–99% nickel and 98% copper with low iron contamination by using optinum conditions, such as a stoichiometric amount of ferrous chloride, a temperature of 110°C, an oxygen pressure of 0.377 MPa and a duration of 8 h.     

Morphology of chalcopyrite leaching in acidic ferric sulfate media

Two-dimensional computer simulations based on percolation theory were used to explain the morphology associated with atmospheric chalcopyrite leaching in acidic ferric sulfate solution. The aim of this study was to understand the differences in observed morphology between chalcopyrite residues leached with and without pyrite in the leach environment. The study of chalcopyrite morphology is of interest because there are no records of similar investigations available. Simulations showed high copper extractions from chalcopyrite when surface atoms were mobile leading to agglomeration of like atoms and the formation of highly porous mineral structures. High degrees of surface mobility are associated with active anodic behavior. The simulated morphology was consistent with previously observed residue morphology from chalcopyrite leach experiments in the presence of pyrite. Thus it was found that the enhanced recoveries and peculiar morphology observed during pyrite catalyzed leaching are attributable to active anodic behavior. Conversely, the simulations also showed that the recovery of copper was low when surface atoms were effectively locked in place resulting in mineral passivation. The simulation morphology obtained in this case was also consistent with experimental results of chalcopyrite leached without the presence of pyrite which have shown non-porous film like product layers.     

Leaching of chalcopyrite with ferric ion. Part III: Effect of redox potential on the silver-catalyzed process

This study examines the effect of redox potential on silver-catalyzed chalcopyrite leaching. Leaching tests were carried out in stirred Erlenmeyer flasks with 0.5 g chalcopyrite mineral, 1 g Ag/kg Cu and 100 mL of a sulphate solution of Fe³⁺/Fe²⁺ (with redox potential ranging between 300 and 600 mV Ag/AgCl) at pH 1.8, 180 rpm and 35°C or 68 °C. Unlike uncatalyzed leaching, an increase of the redox potential increased copper dissolution in the presence of silver ions, as the regeneration of Ag⁺ requires a high concentration of oxidizing agent, Fe³⁺. Additionally, the high reactivity of the mineral surface when silver was present could have been responsible for inhibiting the nucleation of hydrolysis products of Fe³⁺ on it. Excessive addition of silver transformed the chalcopyrite surface into copper-rich sulphides such as covellite, CuS, and geerite, Cu₈S₅, preventing the formation of CuFeS₂/Ag₂S galvanic couple and the recycling of silver ions.     

从高硫多金属金精矿中提取金银铜试验研究

采用焙烧—酸浸—氰化工艺从高硫多金属金精矿中提取金、银、铜。其试验结果表明:在最佳条件下,金、银、铜的平均浸出率分别可达到96.56%、79.12%、91.33%。通过对比金精矿、焙砂、氰化渣中金、银的化学物相可知,硅酸盐包裹金、银不易被氰化浸出,而加入复合添加剂焙烧,硅酸盐包裹的金、银品位大幅度下降,由直接焙烧的2.05 g/t、163.35 g/t分别降到0.81 g/t、25.24 g/t。     

Microbiological leaching of a chalcopyrite concentrate by Thiobacillus ferrooxidans

The microbiological leaching of a chalcopyrite concentrate has been investigated using a pure strain of Thiobacillus ferrooxidans. The optimum leaching conditions regarding pH, temperature, and pulp density were found to be 2.3, 35 degrees C, and 22% respectively. The energy of activation was calculated to be 16.7 kcal/mol. During these experiments the maximum rate of copper dissolution was about 215 mg/liters/hr and the final copper concentration was as high as 55 g/liter. This latter value is in the range of copper concentrations which may be used for direct electrorecovery of copper. Jarosite formation was observed during the leaching of the chalcopyrite concentrate. When the leach residue was reground to expose new substrate surface, subsequent leaching resulted in copper extractions up to about 80%. On the basis of this experimental work, a flow sheet has been proposed for commercial scale biohydrometallurgical treatment of high-grade chalcopyrite materials.     

从含铜金黄铁矿中综合回收金属铜与金的研究

针对某含有金、银、铜等多种有价元素的黄铁矿,在对其原矿物化性质分析的基础上,通过低温氧化焙烧,烟气制酸,焙砂硫酸浸铜,浸铜渣氰化浸金的工艺对该黄铁矿实现了综合利用.使用上述工艺对含硫45.85%(质量分数)、含铜1.92%(质量分数)、含金1.60 g/t的黄铁矿进行处理,得到铜的浸出率为90.09%,金的浸出率可达70%,氰化渣中铁的含量为63.46%,可作为铁精矿外售.金、铜、铁等有价组分实现了综合回收.      

黄铜矿湿法冶金研究进展

阐述了黄铜矿湿法冶金的现状,各种工艺的原理、流程、优缺点及现行发展规模。当前,浸出／溶剂萃取／电积技术在铜湿法冶金中得到普遍应用,是其未来发展趋势。加压酸浸的实际应用很少;卤化盐类浸出工艺比较复杂,多数还处于试验或示范发展阶段;堆浸／萃取／电积主要处理低品位铜矿,其作业周期长,但已经取得很好的工业实践效果。介绍了绿色溶剂-离子液体浸出黄铜矿的研究。     

铜湿法冶金现状及未来发展方向

目前氧化铜矿堆浸／溶剂萃取／电积生产阴极铜，已被证实是一种低成本的铜冶炼方法。堆浸／萃取／电积法的成功也带来了开发湿法冶金工艺从黄铜矿和其他铜精矿中提铜的复兴，本文论述了湿法炼铜工艺的现状并考察了商业上最具吸引力的潜在应用，对湿法处理黄铜矿精矿的优缺点及其初步的经济指标与现有最好的铜熔炼和精炼生产进行了比较。     

铜湿法冶金现状及未来发展方向

目前氧化铜矿堆浸/溶剂萃取/电积生产阴极铜,已被证实是一种低成本的铜冶炼方法。堆浸/萃取/电积法的成功也带来了开发湿法冶金工艺从黄铜矿和其他铜精矿中提铜的复兴,本文论述了湿法炼铜工艺的现状并考察了商业上最具吸引力的潜在应用,对湿法处理黄铜矿精矿的优缺点及其初步的经济指标与现有最好的铜熔炼和精炼生产进行了比较。     

Arsenic removal from copper ores and concentrates through alkaline leaching in NaHS media

Removal of arsenic impurity in ores and concentrates containing copper (Cu) through alkaline leaching in NaHS media was investigated in this work. Samples containing Cu from 10 to 40 wt.% and arsenic from 0.8 to 14 wt.% with enargite (Cu₃AsS₄) as main arsenic bearing mineral were used as starting materials and all leaching tests were conducted at 80 °C under normal atmospheric pressure. Solution and/or slurry potential and pH were maintained consistently below − 500 mV (SHE) and above 12.5 respectively with the addition of NaHS and NaOH, creating a reducing environment for arsenic dissolution and conversion of Cu₃AsS₄ to Cu₂S. Pulp density ranged from 100 to 1000 g/L, NaHS and NaOH reagents were added at 50–200 g/L each and leaching time varied from 10 min to 10 h.Characterization of solid samples (original and leach residue) by XRD and XRF analyses and chemical analysis of both solid and solution samples by ICP analysis showed that Cu₃AsS₄ in the starting material was completely decomposed or transformed to Cu₂S and arsenic released into solution as As (III)/As³⁺ ions (Na₃AsS₃). Over 90% of arsenic in the starting materials was removed within 1–3 h for materials with arsenic content from 1 to 4 wt.% and within 3–6 h for materials with arsenic content over 4–10 wt.%. Dissolution and analysis of leach residues obtained after leaching by ICP indicated that arsenic in the starting materials has been reduced in all cases to below 0.5 wt.%. In all test conditions dissolution of Cu and Fe into solution was not detected, indicating selective leaching of arsenic. NaHS application for removal of arsenic in Cu-ores and/or concentrates was demonstrated in this work and further research is in progress to develop a process to include treatment of arsenic leached into solution.     

Studies on the leaching of an arsenic–uranium ore

Preliminary leaching studies were carried out to develop a suitable method for the recovery of uranium and the elimination of arsenic from a low grade carbonate/silicate ore containing 64 ppm U and 2446 ppm As, as well as some Cu, Pb, Ni and Zn. An examination of the mineralogy found mostly uranium(VI) minerals, such as uraninite, and various base metal sulfides and arsenates in veins and fissures. Roasting the ore at 500–800 °C to volatilize arsenic proved to be unsuitable. Therefore, the ground ore was subjected to direct leaching with sulfuric acid, sodium sulfide and ferric chloride at 80–90 °C with a liquid to solid ratio of 1:1. With sulfuric acid at a concentration of 180 kg/t ore, complete recovery of both uranium and arsenic was achieved giving undesirable arsenic in the leach liquor. The maximum recovery of uranium and arsenic by leaching with sodium sulfide was only 20% and 18%, respectively. However, 3 M ferric chloride leached approximately 92% U(VI) and precipitated arsenic as ferric arsenate. Therefore, maximum uranium can be extracted and arsenic eliminated as impurity by selective leaching with ferric chloride.     

Hydrochloric acid-hydrogen peroxide leaching and metal recovery from a Greek zinc-lead bulk sulphide concentrate

A zinc-lead bulk sulphide concentrate from Kirki (Greece) was leached in aqueous solution with HClH₂O₂at atmospheric pressure and 95°C to extract up to 97% zinc, 40% lead, 80% silver and less than 12% iron after 6 h. Highly pure PbCl₂ crystallized from the leach filtrate on cooling. Sulphur was oxidized to the elemental form; its loss as sulphate ion in solution and residue was 7.5%. During leaching no emission of H₂S or SO₂ was detected. Conditions were determined for producing a pregnant solution (130 g/L Zn) low in iron and lead, to facilitate zinc extraction by electrolysis. Leaching experiments were conducted at 40% solids, 0.55 g HCl and 0.26 g H₂O₂ per gramme concentrate in a 1-L reactor. After a second leaching of the residue in aqueous solution with HCl (1 M) for 1 h at 90°C, the residual lead sulphate was extracted, so that total lead recovery was over 98%. Other values, such as silver and sulphur, could be recovered by additional treatment of either the leach solution or leach residue.     

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.     

超细磨技术在铜金精矿湿法综合回收铜金硫中的应用研究

对于某矿山浮选产出的铜金精矿,由于铜的影响及黄铜矿和黄铁矿等物质对金的包裹作用的影响,直接氰化浸出的金、铜回收率均很低,虽然采用火法工艺处理该类精矿可较好地解决金、铜回收率低的问题,但是存在环境污染等问题。重点对铜金精矿超细磨-热压浸出-萃取-电积提取铜、浸出渣硫代硫酸钠法提取金和煤油溶解回收单质硫的综合回收工艺进行了研究,结果表明应用该综合回收工艺之后,铜、金和单质硫的回收率分别达到95％、98％和99％,具有有价金属综合回收率高且对环境污染小的优点。     

Chloride leaching of chalcopyrite

Two new process flowsheets have been developed which combine chloride leaching of copper from chalcopyrite with solvent extraction, to selectively transfer copper to a conventional sulfate electrowinning circuit. Chloride leaching with copper(II) as oxidant offers significant advantages for copper including increased solubility and increased rates of leaching. Both process flowsheets were similarly designed with a two stage counter-current leach but differ with respect to iron deportment. The goethite model flowsheet includes sparging of air or oxygen to the second leach stage to aid precipitation of iron as goethite (FeOOH). The hematite model flowsheet precipitates iron as hematite (Fe₂O₃) downstream from the leach in a dedicated autoclave. A mass balance has been completed for both process flowsheets and this determined the concentrations of copper and iron species in feed liquor returning to the leach following copper solvent extraction.The optimum leach extraction conditions were determined by varying grind size, temperature and residence time for both leach model scenarios. Leach tests were conducted using a chalcopyrite concentrate from Antamina in northern Peru, which contains a low to moderate amount of gangue material. The hematite model was also examined using a Rosario concentrate from Chile which contained chalcocite in addition to chalcopyrite and significant pyrite. Leach tests based on the hematite model were successful in achieving copper extractions > 95% in 4–6 h at 95 °C after fine grinding the concentrate (P₉₀ = 41 μm). However, copper extraction exceeded 99% from the finely ground Rosario concentrate (P₉₀ = 37 μm). In the goethite model leach tests, 89% copper extraction was achieved under optimum conditions in the atmospheric conditions tested.