Oxygen pressure acid leaching of Gacun complex Cu concentrates

The treatment of the Gacun complex Cu concentrate with high contents of Pb, Zn, Ag, etc by oxygen pressure acid leaching was studied. It is unusual that tetrahedrite, whose treatment was rarely studied, is the primary copper mineral of the concentrates. Most of silver also occurs in the mineral. The optimum operating parameters of oxygen pressure acid leaching were established by conditional tests. Pilot scale test was carried out under the parameters, and the leaching rates of copper and zinc are as high as 97.10% and 89.83% while lead and silver are transformed into sulfate and sulfide respectively and stay in leaching residue. The copper and zinc in lixivium were reclaimed by extraction-electrowinning and purification-electrowinning, respectively, and the lead and silver in the residue were reclaimed separately by chloride leaching and thiourea leaching. The extraction rate of copper achieves 96%, and the leaching rates of lead and silver reach 90% and 95%, respectively.     

Kinetics of nickel leaching from roasting-dissolving residue of spent catalyst with sulfuric acid

Sulfuric acid leaching process was applied to extract nickel from roasting-dissolving residue of a spent catalyst, the effect of different parameters on nickel extraction was investigated by leaching experiments, and the leaching kinetics of nickel was analyzed. The experimental results indicate that the effects of particle size and sulfuric acid concentration on the nickel extraction are remarkable; the effect of reaction temperature is mild; while the effect of stirring speed in the range of 400–1 200 r/min is negligible. Decreasing particle size or increasing sulfuric acid concentration and reaction temperature, the nickel extraction efficiency is improved. 93.5% of nickel in residue is extracted under suitable leaching conditions, including particle size (0.074–0.100) mm, sulfuric acid concentration 30% (mass fraction), temperature 80 °C, reaction time 180 min, mass ratio of liquid to solid 10 and stirring speed 800 r/min. The leaching kinetics analyses shows that the reaction rate of leaching process is controlled by diffusion through the product layer, and the calculated activation energy of 15.8 kJ/mol is characteristic for a diffusion controlled process. Foundation item: Project (50574101) supported by the National Natural Science Foundation of China; Project (2003UDBEA00C020) supported by the Collaborative Project of School and Province of Yunnan Province, China     

复杂铜铅混合精矿氧压浸出综合回收工艺

呷村铜铅混合精矿中铜、铅矿物主要为黝铜矿和方铅矿,还含有较高的锌、银、砷和锑.本试验针对该矿采用一段氧压浸出综合回收工艺进行处理,通过条件优化实验确定了氧压浸出的操作条件.扩大验证实验表明Cu、Zn的浸出率分别高达98.89%、94.92%,Pb、Ag转化为矾类和硫化物形式留在浸出渣中,铜锌与铅银分离彻底.浸出液中的铜、锌分别通过萃取、电积进行回收.浸出渣中的铅、银通过碳酸盐转化-硅氟酸浸铅-硫脲浸银进行回收.铜萃取率,铅、银浸出率分别为96%、94%和93%.     

Study on the method of recovering and separating indium from residue containing indium

A method of recovering indium from complex smelting residue containing indium was investigated. Indium was extracted by technique of low acid leaching and solvent extraction. The conditions of separating iron and indium were studied and the optimal conditions were determined. When the residue is twoclass-countercurrent leached with 2 mol/L H2SO4 and 30-40 g/L NaC1 at 100 ℃, the leaching rate of indium is 80%. The extraction rate of indium is over 98% and that of iron is less than 5% after it is third-class-countercurrent extracted by P204, and when sulfonated kerosene is used as solvent, acidity in aqueous phase remains the same as that of leaching liquid, and phase is for 1 : 3. Using 2 mol/L HC1 as back-extraction agent, with phase ratio being 5 : 1, by third-class-countercurrent back-extraction, the back-extraction rate of indium is over 99%, but that of iron is very low, which can meet the need of separating indium and iron.     

Ag recovery from copper anode slime by acid leaching at atmospheric pressure to synthesize silver nanoparticles简

In this paper, recovery of silver from anode slime of Sarcheshmeh copper complex in Iran and subsequent synthesis of silver nanoparticles from leaching solution is investigated. Sarcheshmeh anode slime is mainly consisted of Cu, Ag, Pb and Se. Amount of Ag in the considered anode slime was 5.4% (by weight). The goal was to recover as much as possible Ag from anode slime at atmospheric pressure to synthesize Ag nanoparticles. Therefore, acid leaching was used for this purpose. The anode slime was leached with sulfuric and nitric acid from room to 90 °C at different acid concentrations and the run which yielded the most recovery of Ag was selected for Ag nanoparticles synthesis. At this condition, Cu, Pb and Se are leached as well as Ag. To separate Ag from leach solution HCl was added and silver was precipitated as AgCl which were then dissolved by ammonia solution. The Ag nanoparticles are synthesized from this solution by chemical reduction method by aid of sodium borohydride in the presence of PVP and PEG as stabilizers. The synthesized Ag nanoparticles showed a peak of 394 nm in UV–vis spectrum and TEM images showed a rather uniform Ag nanoparticles of 12 nm.     

Process optimization of reaction of acid leaching residue of asbestos tailing and sodium hydroxide aqueous solution

Silica is the major component of the acid leaching residue of asbestos tailing. The waterglass solution can be prepared by the reaction of the residue with sodium hydroxide aqueous solution. Compared to the high temperature reaction method, this process is environmental friendly and low cost. In this paper, the reaction process of the residue and the sodium hydroxide aqueous solution is optimized. The optimum reaction process parameters are as follows: the usage of sodium hydroxide is 26.4 g/100 g acid leaching residue, the reaction temperature is 90℃, the reaction time is 1 h, and the ratio of the liquid/solid is 2.0. The significance sequence of the process parameters to the alkali leaching reaction effect is the usage of sodium hydroxide > the ratio of the liquid/solid > the reaction time > the reaction temperature. The significance sequence to the leaching ratio of SiO2 is the ratio of the liquid/solid > the usage of sodium hydroxide > the reaction time > the reaction temperature. The significance sequence to the modulus of the sodium silicate is the ratio of the liquid/solid > the usage of sodium hydroxide > the reaction time > the reaction temperature. Under the optimum conditions, the leaching ratio of the SiO2 is 77.5%, and the modulus of the sodium silicate is 3.15. The XRD analysis result indicates that the major components of the alkali leaching residue are serpentine, talc, quartz and some albite.     

Simultaneous electrogenerative leaching of chaicopyrite concentrate and MnO2

In order to enhance the electrogenerative leaching rate of chalcopyrite concentrate reasonably, the principle of generative process was applied to simultaneous leaching of chalcopyrite concentrate and MnO2. The results show that Cu^2＋ and Mn^2＋ in addition to electrical energy could be acquired in the simultaneous electrogenerative leaching process. The leaching cell has the open circuit potential of about 1.0 V and gains quantity of electricity of about 700 C. The optimum leaching rates of Cu^2＋ and Mn^2＋ are 23.10% and 22.1%, respectively after electrogenera- tive leaching for about 10 h under the present conditions.     

Recovery of molybdenum from residues by simultaneous ultrafine milling and alkali leaching

A new technology of treating molybdenum residues by simultaneous ultmfme milling and alkali leaching was put forward to recover molybdenum from metallurgical residues. The effects of residue size, milling time, solid content, n(Na2CO3)/n(Mo) and slurry pH value on molybdenum leaching rate were investigated. The results indicate that a simpler process, lower slurry temperature, 50% shorter treating time, 60% decrease of Na2CO3 content and 15% increase of molybdenum leaching rate can be obtained by the new technology compared with the traditional process. The leaching kinetic equation was determined, and calculation of active energy (E = 56.2 kJ/mol) shows that the leaching process of molybdenum residues by simultaneous ultmfine milling and alkali leaching is controlled by chemical reaction. Potential exists for the new process to form the basis for an economically viable, environmentally friendly process to recover valuable elements from residues.     

Extraction of nickel from molybdenum leaching residue of metalliferous black shale by segregation roasting and acid leaching

The recovery of nickel from molybdenum leach residue by the process of segregation roasting-sulfuric acid leaching-solvent extraction was investigated. The residue was characterized by microscopic investigations, using X-ray fluorescence spectrometry (XRF) and X-ray diffractometry (XRD) techniques and the residue after segregation roasting was characterized by chemical phase analysis method. A series of experiments were conducted to examine the mass ratio of activated carbon (AC) to the residue, segregation roasting time and temperature, sulfuric acid concentration, liquid-to-solid ratio, leaching time, leaching temperature, addition amount of 30% H₂O₂, stirring speed (a constant) on the leaching efficiency of nickel. A maximum nickel leaching efficiency of 90.5% is achieved with the mass ratio of AC to the residue of 1:2.5, segregation roasting time of 2 h, segregation roasting temperature of 850 °C, sulfuric acid concentration of 4.5 mol/L, liquid-to-solid ratio of 6:1, leaching time of 5 h, leaching temperature of 80 °C, addition of 30% H₂O₂ of 0.6 mL for 1 g dry residue. Under these optimized conditions, the average leaching efficiency of nickel is 89.3%. The nickel extraction efficiency in the examined conditions is about 99.6%, and the nickel stripping efficiency in the examined conditions is about 99.2%.     

Bioleaching of chalcocite by mixed microorganisms subjected to mutation

Mixed microorganisms with elevated activity of chalcocite-leaching were screened by mutation methods. The original microorganisms collected from acid mine drainage of different sites were mixed and then treated with mutagens NO₂⁻, diethyl sulfate (DES), UV and their combinations, respectively. Five groups of mixed microorganisms with much stronger ore-leaching ability were obtained by screening on the leaching media. Among them, group E of mixed microorganisms (treated with 1% DES for 60 min) with the best performance on chalcocite-leaching, increases the content of Cu²⁺ by 101.4% in 20 d of leaching compared with the control culture. In addition, group E is more tolerant to Cu²⁺ in media than the control without mutation treatment. Analysis for the diversity of microbial clones indicates that half of operational taxonomic units (OTUs) in group E are Acidithiobacillus ferrooxidans. These observations suggest that group E might have potentials for industrial application. Foundation item: Project(50321402) supported by the National Natural Science Foundation of China; Project(2004CB619201) supported by the Major State Basic Research and Development Program of China     

Effects of Ag on microstructure and mechanical properties of 2519 aluminum alloy

The effects of Ag on the microstructure and mechanical properties of 2519 aluminum alloy were investigated by means of tensile test, micro-hardness test, transmission electron microscope and scanning electron microscope. The results show that the addition of 0.3 % （mass fraction） Ag accelerates 2519 aluminum alloy＇s age-hardening, increases its peak hardness and reduces 4 h of peak aged time at 180 ℃. The addition of 0. 3% （mass fraction） Ag increses the tensile strength at room temperature and elevated temperature. This increment at room temperature and 200 ℃ is 24 MPa and 78 MPa, respectively. In contrast, the elongation of 2519 aluminum alloy is decreased with Ag addition. The increase of tensile strength of 2519 aluminum alloy with Ag addition is attributed to the high volume fraction of Ω phase.     

真空蒸馏硬锌综合回收有价金属

硬锌是火法炼锌过程中的副产物，含有多种金属元素，采用真空蒸馏的方法综合处理硬锌，使锗铟银富集在蒸馏残渣中，锗的富集倍数为１０倍，直收率大于９６％，铟的富集倍数为４倍，直收率大于９０％．     

金泥处理新工艺的研究

针对某金矿厂金泥生产过程中存在的问题：坩埚使用寿命低，熔炼产品合金中金银品位低及渣含金高，金收率低等，提出了金泥酸浸除锌、浸渣再熔炼，一步获得金银合金的新工艺。同时研究了硫酸加入量、浸出时间对锌浸出率的影响。试验结果表明，采用该新工艺不仅能消除锌对熔炼过程的危害，延长坩埚使用寿命，而且可将锌回收，当硫酸加入量为理论量１５％￣１７％，浸出时间不少于８０ｍｉｎ，固液比１：３时，锌浸出率≮９５％。含金浸     

New technique of comprehensive utilization of spent Al2O3 -based catalyst

A new technology was developed to recover multiple valuable elements from the spent Al2O3-based catalyst by X-ray phase analysis and exploratory experiments. The experimental results show that in the condition of roasting temperature of 750 ℃ and roasting time of 30 min, molar ratio of Na2O to Al2O3 of 1.2, the leaching rates of alumina, vanadium and molybdenum in the spent catalyst are 97.2%, 95.8% and 98.9%, respectively. Vanadium and molybdenum in sodium aluminate solution can be recovered by precipitators A and B, and the precipitation rates of vanadium and molybdenum are 94. 8% and 92. 6%. Al（OH）3 was prepared from sodium aluminate solution in the carbonation decomposition process, and the purity of Al2O3 is 99. 9% after calcination, the recovery of alumina reaches 90. 6% in the whole process; the Ni-Co concentrate was leached by sulfuric acid, a nickel recovery of 98. 2% and cobalt recovery over 98.5% can be obtained under the experimental condition of 30% H2SO4, 80 ℃, reaction time 4 h, mass ratio of liquid to solid 8, stirring rate 800 r/min.     

Isolation of Leptospirillum ferriphilum by single-layered solid medium

According to physiological and biochemical characteristics of Leptospirillum ferriphilum,a strain of object bacteria was isolated successfully.Bacteria were enriched by selective liquid medium and plated on designed single-layered agar solid medium.Colony was cultured and bacteria were collected.The morphologies of the object bacteria were observed using crystal violet staining,scanning electron microscope(SEM)and transmission electron microscope(TEM).The result of 16S rDNA identification shows that this bacterium belongs to Leptospirillum ferriphilum and it is named as Leptospirillum ferriphilum strain D1.These results indicate that this new single-layered agar solid medium is efficient and simple for isolation of Leptospirillum ferriphilum.Additionally,physiological-biochemical characteristics show that the optimum initial pH value and its growth temperature are 1.68 and 40 ℃,respectively.The culture of it is used to leach a complex concentrate chalcopyrite,the leaching efficiencies of copper and iron are 1.93 % and 13.74 %,respectively,and it is more effective than the A.ferrooxidans culture in the leaching of the complex concentrate chalcopyrite.     

A new process of extracting vanadium from stone coal

A new process of extracting vanadium from the stone coal vanadium ore in Fangshankou, Dtmhuang area of Gansu Province, China was introduced. Various leaching experiments were carried out, and the results show that the vanadium ore in Fangshankou is difficult to process due to its high consumption of acid and the high leaching rate of impurities. However, the leaching rate can be up to 80% and the content of V2O5 in the residue can be between 0.22%-0.25% in the process of ore fine grinding→oxidation roasting→mixing and ripen-ing→aqueous leaching→P2O4 solvent extraction→sulfiuie acid stripping→oxidation and precipitation→decomposition by heat. Also, the quality of flaky V2O5 produced by this process can meet the requirements of GB3283-87. The total leaching rate of vanadium is 70%. Also, three types of wastes are easy to treat. The vanadium extraction process is better in relation to the aspect of environmental protection than the sodium method.     

含铟锌渣浸出和萃取铟的研究

湿法炼锌产出的锌渣含铟达700~850 g/t,采用两段酸浸,铟的浸出率可达90%以上,并提出了酸浸的工艺流程和最佳浸出条件.分别对酸液和有机相采用三级逆流萃取和反萃,铟的萃取率和反萃率分别达98.5%和99%以上,并提出了最佳萃取和反萃条件.     

Extracting gold from pyrite roster cinder by ultra-fine-grinding and resin-in-pulp

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盐酸对冶金污泥中铜锌镉铅的浸出工艺优化

为了能够充分回收利用冶金污泥中的有价金属,采用盐酸作为浸取剂浸出污泥中的重金属,并进行浸出工艺的优化.首先对污泥性质进行分析,分别采用烘干法测定冶金污泥的含水率,用X射线荧光光谱仪测定试样中金属成分及质量分数,用X射线衍射仪对试样中各元素的物相特征进行定性和定量分析.结果显示:污泥含水率为75.88%;干污泥含铜和锌的质量分数分别是1.51%和1.71%;污泥矿物相中铜主要以单质形式存在,锌主要以闪锌矿形式存在.然后采用盐酸作浸取剂,在单因素条件下进行浸出反应.研究了冶金污泥中铜、锌、镉、铅等重金属的浸出规律,并考察了浸出温度、浸出时间、盐酸浓度、液固比、粒径等因素对浸出率的影响.盐酸浸出污泥中重金属最佳工艺参数:浸出温度为25℃,浸出时间为10 min,盐酸浓度为1mol/L,液固比为25∶1(mL/g),干污泥粒度小于150μm.在此条件下,铜、锌、镉、铅的浸出率可分别达到84.4%、88.1%、98.8%、85.4%.盐酸浸出最佳工艺条件的确定,对工业应用有一定实用价值.     

氨浸置换法资源化废杂铜的研究

采用碳铵浸取-置换沉积新方法资源化、无害化回收废杂铜.考察浸取时间、氨水浓度、碳酸铵用量、催化剂用量对浸取效果的影响,并讨论还原剂用量、置换时间及搅拌速度对沉积过程的影响.结果表明:在(NH4)2CO3:6.5 g,NH3.H2O:6 mol/L、浸取催化剂:12 mL、浸取时间3 h的最优条件下,铜、锌的浸取率分别达到93.12%、95.03%,而伴生元素留在滤渣中.二段置换沉积过程中,在最佳工艺条件下,制备出产品纯度与附加值高的铜粉和七水硫酸锌.铜的回收率达到98.3%.该工艺具有操作简单、生产效率高、成本低、无污染等特点.