钴白合金湿法冶金工艺研究

本文研究了钴白合金湿法冶炼工艺流程，详细进行了氯气浸出－电溶脱铜－ＴＢＰ萃取除铁试验研究。试验条件下，钴、铜、铁浸出率均达到９９％，且分离效果良好。文章给出了钴白合金冶炼原则工艺流程。     

电化学—高压氧浸联合处理铜钴合金的工艺研究

以非洲某铜钴合金(红合金)为原料,采用电化学—高压氧浸联合处理工艺,研究了电解温度、电解液Cu浓度、电流密度、电解液杂质Fe含量、高压氧浸温度等工艺参数对工艺过程的影响。结果表明,电解温度55～60℃、Cu浓度30～35g/L、电流密度200～250A/m~2是较合适的电解处理铜钴合金工艺参数条件。采用高压氧浸出电解所产生的阳极泥,控制反应温度210℃较为合适。电化学—高压氧浸联合处理铜钴合金可实现Cu、Co有价金属回收率均达到99.9%以上。     

铅阳极泥选择性脱铜试验研究

采用选择性脱铜—混酸浸锑、铋—硝酸脱铅—火法熔炼回收贵金属工艺综合回收铅阳极泥中的有价金属。重点介绍了该工艺中选择性脱铜的试验研究。确定了最佳脱铜条件:浸出温度28℃,初始酸度H2SO420 g/L,鼓空气浸出3 h,液固比L/S=5/1(mL/g),添加剂Fe3+浓度1 g/L;在该条件下,铜的平均脱除率为91.30%,锑的平均浸出率仅为2.11%,Bi,Pb,Au,Ag等不被浸出。该研究取得了较好的选择性脱铜效果,有效解决了铅阳极泥传统湿法处理工艺中存在的金属分离不彻底、产品质量不高等问题。     

从铅冰铜中高效选择性提取铜的工艺研究

采用高温高压纯氧氧化法选择性提取铅冰铜中铜, 研究了硫酸用量、浸出温度、反应时间、液固比、氧气压力、搅拌速度以及分散剂木质素用量对铜浸出率的影响及对浸出液中铁含量的影响。铅冰铜经氧压浸出后进行液固分离, 铅冰铜中的铜进入液相中, 绝大部分铁以赤铁矿的形式与铅、银、金等有价金属一起进入渣相中; 浸出后的硫酸铜溶液经调酸后直接进行旋流电解可得到合格的阴极铜产品, 浸出渣返回铅冶炼系统综合回收铅、银、金等有价元素。高温氧压浸出铅冰铜, 铜浸出率可达93.5%, 阴极铜产品质量达到99.975%, 有效实现了铅冰铜中铜的选择性提取。     

高硅含量钴白合金的电化学溶解工艺研究

为了高效提取有价金属Co,在硫酸体系中研究了钴白合金电化学溶解过程的影响因素并进行了条件优化。用电化学工作站监测合金块的阳极电位,用扫描电镜SEM分析合金阳极表面的变化,用X射线衍射仪(XRD)分析阴极产物,用原子吸收光谱仪(AAS)分析电解液中各金属的含量。结果表明：添加NH4Cl的硫酸电解体系中,钴白合金阳极中有价金属有效溶出。研究了添加剂量、电解电流、温度、酸度、阴阳电极面积比等因素的影响后,在优化条件下可实现Co、Fe的溶出率分别为119%、128%,阳极电流效率为72%。最后对电化学溶解法处理钴白合金的工艺流程的设计提出建议。     

四川某铜多金属矿石选矿试验

四川某铜多金属矿石中除铜外,还伴生有钼、硫钴和铁。为了合理有效地利用该矿石,对其进行了选矿工艺研究。结果表明,采用铜钼混合浮选-铜钼分离浮选-混浮尾矿浮硫钴-浮选尾矿弱磁选回收铁的工艺流程,可在高效回收铜的同时较好地实现钼、硫钴和铁的综合回收,所获铜精矿铜品位为21.25%、铜回收率为93.38%,钼精矿钼品位为45.78%、钼回收率为45.72%,硫钴精矿硫品位为44.69%、钴品位为0.46%、硫回收率为41.53%、钴回收率为46.42%,铁精矿铁品位为63.73%、铁回收率38.29%。     

从刚果(金)铜萃余液中回收钴和铜及萃余液的循环利用

为了从刚果（金）某湿法炼铜厂铜萃余液中回收钴和铜,并实现萃余液的循环利用,采用除铁—沉铜—第一段沉淀钴—第二段沉淀钴工艺流程处理含钴铜萃余液。结果表明,采用该工艺分别得到铁渣、铜渣、粗氢氧化钴和第二段钴渣,铁渣中钙含量26.79%、铜含量0.05%、钴含量0.12%,除铁过程中铜和钴损失率分别为4.64%和3.14%;铜渣含铜3.74%、含钴2.06%;粗氢氧化钴中钴含量32.83%;第二段钴渣含钴7.14%;钴总回收率大于95%。铁渣用于建筑材料,铜渣返回浸出系统回收铜和钴,粗氢氧化钴可直接外售,第二段钴渣经萃余液调浆返回除铁工序,第二段沉钴后溶液回用作磨矿补加水,有价金属和处理后的废液都得到了资源化利用。经生产实践检验,该处理工艺具有良好的经济、环境和社会效益。      

Cobalt and zinc recovery from copper sulphate solution by solvent extraction

It has been demonstrated in earlier works that zinc as an impurity can be effectively removed from cobalt sulphate solutions (Zn/Co < 1) by solvent extraction with D2EHPA. Some process residues from copper plants contain both cobalt and zinc as valuable metals, which have to be separately extracted for their recovery. Leaching of such residues leads to solutions with higher Zn/Co ratios (Zn/Co > 10). Again, solvent extraction with D2EHPA has been successfully used to separate cobalt and zinc into their respective solutions, which could further be treated by appropriate techniques for the production of these metals.The method mainly consists of selective copper extraction with LIX 984, iron removal by precipitation with CaCO₃, simultaneous cobalt and zinc extraction with D2EHPA followed by their separation by selective stripping with sulphuric acid of different concentrations. The use of a specific cobalt extractant is not necessary. More than 95% copper has been recovered from the pregnant solution typically containing 1.0 g/l Co²⁺, 2.0 g/l Cu²⁺, 12.60 g/l Zn²⁺ and 8.4 g/l Fe³⁺. The cobalt and zinc recoveries were on an average of 90% each in their respective individual solutions.     

Textural,mineralogical and chemical characteristics of copper reverb furnace smelter slag of the Okiep Copper District,South Africa

The Okiep Copper District in South Africa has produced more than 110 million tons at a grade of 1.71% Cu from several small mafic ore bodies. The ore was smelted on site and generated ∼5 mt of slag. During the life of mine attempts to recover copper from the slag by flotation had limited success. After mine closure the challenge of environmental rehabilitation and the possible disposal of the slag, triggered a reinvestigation into the viability of slag as a copper resource. Characterisation of the slag as a contribution to the potential copper recovery is the objective of this study.The slags are hard, vitreous with a matrix of Si–Fe–Al–Mg–Ca glass and laths of Mg–Fe–olivine, Fe–Mg–orthopyroxene and minor Cr-spinel. Copper grade varies between 0.11% and 0.42% with minor nickel, cobalt, molybdenum, zinc and tungsten. All economic elements are hosted by disseminated spheroidal prills which consist mainly of the copper sulphides bornite, chalcocite, covellite and chalcopyrite with exsolved sulphide phases of the minor base metals as well as rhenium and silver. Prills consisting of metallic copper and alloys are minor constituents. Prill diameter is highly variable with most in the 40–60 μm range and the historically poor copper recovery is attributed to the small prill size. Crushing of slag to −45 μm as opposed to the previous −75 μm should significantly increase sulphide liberation and recovery of copper and minor base metal sulphides by conventional flotation.Provided the operation is economically viable, redistribution of the processed slag to environmentally acceptable sites will resolve the present pollution and rehabilitation challenge related to the dumps in the Okiep Copper District. The operation will also have a positive socio-economic impact on this poverty-stricken part of South Africa.     

炼铜尾渣提取重介质产品的特性及矿物学分析

资源绿色开发和冶炼废渣的高效利用成为战略技术需求。本文结合炼铜尾渣的矿物学性质和重介质选矿的现实需求,采用分级-磁选-浓缩脱泥流程获得炼铜尾渣重介质产品,采用化学分析、XRD、SEM和EDS等手段,考察了其化学成分、物相组成及残余铅锌杂质的矿物学特征,探讨了炼铜尾渣重介质产品应用的环境影响。研究表明：重介质产品密度为4.42 t/m3,含铁(TFe)56.57%、SiO2含量为23.49%,少量Pb、Zn、Cu等金属杂质;主要矿物为磁铁矿、铁橄榄石、铅铁硅质玻璃体和石英,含量达99.41%。因磁铁矿和铁橄榄石的嵌布粒度较细,解离度低,磁性物含量可达95.4%,便于回收使用;残余铅锌铜重金属元素溶出率很低、环境影响风险较小,为炼铜尾渣的资源化应用开辟了新的途径。     

从炼铜厂炉渣中回收铜铁的研究

针对铜转炉渣中铜铁硅矿物紧密共生、呈细粒不均匀嵌布及渣硬度高、难磨的特点，进行了多种磨矿与选别流程组合的对比试验，最后选用磨矿(-0.043mm 79．6％)-浮选-磁选-浮选中矿与磁性矿合并再磨(-0．040mm99．32％)-再浮-再磁的阶段磨矿阶段选别的流程，其中第一段磁选精矿再磨是铁硅单体分离获得合格铁精矿的关键．在转炉渣含铜1．58％(硫化铜和金属铜占78．68％)、含铁53．54％(磁性氧化铁占28．53％)的情况下，获得铜精矿品位19．82％，回收率85．48％的选铜指标，同时综合回收了渣中磁性氧化铁，得到铁品位62．525％、回收率35．02％、含SiO2 9．94％的合格铁精矿．     

四川某多金属硫化铜矿的综合回收

针对四川某多金属硫化铜矿矿石的性质,对该矿中的铜、钴、铁进行了综合回收试验。结果表明,采用混合浮选—铜钴分离浮选工艺,能获得铜品位22.41%、回收率91.32%的铜精矿和品位0.53%、回收率56.22%的钴精矿。浮选尾矿再用磁选回收铁,可以获得品位64.49%、回收率38.04%的铁精矿。     

Pyrometallurgical upgrading of PGM-rich leach residues from the Western Platinum base metals refinery through roasting

The production of Platinum Group Metals (PGMs) normally entails the smelting of PGM flotation concentrates, converting of the furnace matte and removal of the bulk of the Ni, Cu, Co, S and Fe through atmospheric and pressure leaching in a base metals refinery to produce a PGM-rich concentrate. A number of impurities, mostly Se, Te, As, Bi, Os and Pb, are not removed significantly during the oxidising leach process in sulphuric acid media. In addition slag inclusions in matte leads to contamination of the PGM residues with silica, fayalite, magnetite and trevorite phases. Furthermore some Cu, Ni, Fe and S also remain. For this reason a typical Precious Metal Refinery (PMR) feed material contains less than 65% PGMs. The PMR is based on a chloride process and requires contaminants to be within narrow specification limits to prevent the formation of PGM residues that must be reprocessed or tolled, leading to poor first pass metal efficiencies and extending the duration of the production pipeline for efficient recovery.A process has been developed to significantly upgrade the BMR leach residues through pyrometallurgical processing, which include a multistep process of roasting under oxidising atmospheres, a two-step smelting process of the roasted calcine (with engineered slag chemistry and slag-refractory interactions) and subsequent atomisation of the molten alloy which can be fed as a slurry into the HCl/Cl₂ dissolution reactors in the precious metals refinery. These pyrometallurgical steps upgrade the BMR residue from a 45–50% grade up to an alloy grade of ca. 90% PGMs, whilst removing the most deleterious elements with major process impacts on the PMR.This paper will focus primarily on the roasting step and it will investigate the thermochemical and mineralogical changes occurring during roasting. These changes were evaluated through a combination of thermochemical modelling and experimental investigation. The roasting step needs to be in an oxidative environment in order to achieve the vapourisation of Se, Te, As, Os and S. The speciation of PGMs and their vapourisation behaviour are presented, as well as the sensitivity of precious metals deportment to changes in roast conditions.     

Towards understanding nickel converter matte solidification

Nickel converter mattes are complex metallurgical solutions of Ni, Cu, S, Fe and O along with low concentrations of many other elements including Co, Pb and PGEs. Studies on how such complex mixed solutions evolve upon cooling may contribute towards an improved understanding of matte solidification. Liquidus and primary phase equilibria were calculated for Cu–Ni–S ternaries including fixed iron and cobalt concentrations. True liquid matte starting compositions and calculated assays were subsequently superimposed on relevant Cu–Ni–S_FeCo ternary systems. Multiphase cooling equilibria were also calculated for variable Cu–Ni–S–Fe–Co–O matte systems. In addition, actual industrial mattes were analysed using automated mineralogy, electron probe microanalysis and field emission scanning electron microscopy.The effect of the end composition of the ternary systems at fixed iron and cobalt concentrations on the liquidus temperature range has been quantified. The liquidus temperature range is lowered when the fixed iron and cobalt concentration decreases. The solidification pathway of oxygen-free liquid matte has been estimated. Moreover, it has been shown that variations in the starting composition of oxygen-free matte alter the path of solidification towards the eutectic. The examination of multiphase cooling equilibria for variable Cu–Ni–S–Fe–Co–O liquid phase systems provided a quantitative understanding of solidification processes to within ±2.5 °C. The analysed nickel and copper-sulphide phase structures have shown to exhibit chemical non-equilibrium within high and low iron matte. It can be concluded that the present study has provided a coherent insight into nickel converter matte solidification processes.     

底吹熔炼渣工艺矿物学研究

为了回收铜渣中的有价金属,采用XRF、XRD、SEM、EDS和BPMA等分析手段对底吹熔炼铜渣进行了工艺矿物学研究,查明了熔炼渣的主要成分、主要矿物成分、铜物相赋存状态,并对渣中重要矿物相的嵌布 特征、嵌布粒度和主要矿物解离度进行了深入研究,结果表明：①熔炼渣中主要有价金属为Cu、Fe、Pb、Zn等,杂质成分主要为SiO2。②熔炼渣中主要矿物为冰铜、铁橄榄石、铁酸盐和玻璃相;主要含铜矿物为冰铜 、金属铜、黄铜矿和氧化铜等,以冰铜含量最高,分布率为92.69%。③熔炼渣中冰铜粒度分布不均匀,主要呈粗细不等的粒状或圆点状分布于渣中,与硫化铅、铁橄榄石、玻璃相、铁酸盐等矿物嵌布关系密切。④金 属铜主要呈长粒状和圆粒状,产出的多数金属铜被铁酸盐、铁橄榄石、玻璃相等矿物包裹或连生。⑤铁酸盐在放大后呈叶状雏晶,与金属铜和冰铜关系密切,易与铁橄榄石和其他硫化矿紧密共生。⑥铁橄榄石与金属 铜和冰铜关系密切,与铁酸盐相互包裹、夹杂、连生组成熔渣的基底物相。⑦玻璃相充填于铁酸盐、铁橄榄石、金属铜、冰铜粒间起胶黏作用。⑧主要含铜矿物金属铜与冰铜的单体解离度较低,分别为46.13%和 33.81%,主要分布于-0.038+0.020 mm粒级内,因此对粗粒冰铜和金属铜进行回收的同时,也应注重细粒冰铜和金属铜的回收。     

铅阳极泥H2SO4 NaCl选择性浸出研究

采用选择性脱铜—H2SO4+NaCl选择性浸锑、铋—硝酸脱铅—火法熔炼回收贵金属工艺综合回收铅阳极泥中的有价金属。重点介绍了该工艺中H2SO4+NaCl选择性浸锑、铋试验研究。确定了最佳浸出条件:初始硫酸浓度2.5～3 mol/L,NaCl浓度为75～100 g/L,浸出温度80℃,液固比L/S=8/1(mL/g),浸出时间2 h;在该条件下锑、铋、铜的平均浸出率均大于99%,铅的平均浸出率仅1.68%,金银不被浸出,锑、铋、铜得以有效选择性浸出,铅、金、银在渣中得到了有效富集,为后续工艺中硝酸脱铅和贵金属火法综合回收工艺创造了有利条件,解决了传统铅阳极泥湿法综合回收出现的金属分离不彻底,贵金属直收率不高等问题。     

碳热还原法从铜渣中回收铁钼合金

湖北某铜冶炼厂电炉渣浮选铜后的尾渣，Fe品位为35.37%，Mo品位为0.30%，其中铁主要以磁铁矿和铁橄榄石形式存在，钼存在形式复杂，以氧化物为主，同时与铜渣中Si、Fe等之间形成化学键。若采用 直接磁选回收铁，常规浮选回收钼，铁与钼均不能被有效回收。为使铜渣中的铁与钼资源可最大化回收再利用，以煤粉作还原剂，氧化钙与氧化铝作造渣剂，采用熔融直接还原工艺制备铁钼合金，从而一并回收铜渣 中的铁和钼。探讨了还原温度、还原时间、煤粉用量、氧化钙用量、氧化铝用量等因素对Fe、Mo在合金中的回收率及品位的影响。结果表明在还原温度1 400 ℃、还原时间60 min、煤粉用量、氧化钙用量、氧化铝用 量分别是铜渣量的20%、20%、10%等优化条件下，Fe、Mo在合金中回收率分别为89.03%、98.44%，品位分别为91.70%、0.86%。     

用直流电弧炉还原熔炼大洋多金属结核

根据大洋多金属结核中各主要元素的还原性差异,用碳作还原剂,在直流电弧炉中选择性还原熔炼大洋多金属结核,获得了合格的富锰渣产品和熔炼合金。富锰渣中锰收率达91％,铜、钴、镍、铁进入熔炼合金的比率分别为：96．73％,98．13％,98．53％,92．90％。     

赞比亚某复杂铜钴矿选矿工艺研究

本文针对赞比亚某铜钴矿进行了工艺矿物学研究,在其基础上进行了详细的选矿试验研究。根据矿石中铜、钴矿物的可浮性差异,通过使用高效选择性捕收剂BK404,采用优先选铜-尾矿选钴的工艺流程,最终获得实验室小型闭路试验结果为：铜精矿含铜32.48%,回收率89.75%;钴精矿含钴1.58%,回收率46.75%。     

Extractive separation of copper and nickel from copper bleed stream by solvent extraction route

Bleed stream from electro refining step of copper smelter was processed to recover the metals as high value products such as copper and nickel powders or salts. The process consists of partial decopperisation of the bleed stream followed by crystallization of a mixed salt of copper and nickel sulphate, leaching of the mixed salt, removal of iron, solvent extraction for the separation of copper and nickel and winning the solution to produce metal powders. After partial decopperisation of copper bleed stream, the resultant solution was subjected to crystallization which produces composite crystals with the chemical composition of 8.4–12.5% Cu, 13.7–14.38% Ni and 1–2 ppm of Fe as impurity. This mixed salt was leached with water and was treated for iron precipitation. The purified solution was subjected to solvent extraction using two solvents namely LIX 84 or Cyanex 272. A 20% LIX 84 in kerosene extracted 99.9% copper and 0.059% nickel at a pH of 2.5, similarly a 5% Cyanex 272 in kerosene extracted 98.06% copper and 0.51% nickel at a pH of 4.85. LIX 84 was used for metal separation in the mixer-settler unit and a flow sheet was developed using this solvent. The pure solutions of copper after stripping it from the loaded organic and nickel left in the raffinate were further electrolysed to produce pure copper (99.9%) and nickel (99.8%) powders, alternatively pure sulphate salts could also be crystallized. Since it is well known that Cyanex 272 is used for the separation of cobalt and nickel, a few experiments were performed on the extraction of copper by using Cyanex 272. A complete study is yet to be done to develop a flow sheet by using this solvent.