Recovery of valuable metals from zinc leaching residue by sulfate roasting and water leaching

Zinc leaching residue (ZLR), produced from traditional zinc hydrometallurgy process, is not only a hazardous waste but also a potential valuable solid. The combination of sulfate roasting and water leaching was employed to recover the valuable metals from ZLR. The ZLR was initially roasted with ferric sulfate at 640 °C for 1 h with ferric sulfate/zinc ferrite mole ratio of 1.2. In this process, the valuable metals were efficiently transformed into water soluble sulfate, while iron remains as ferric oxide. Thereafter, water leaching was conducted to extract the valuable metals sulfate for recovery. The recovery rates of zinc, manganese, copper, cadmium and iron were 92.4%, 93.3%, 99.3%, 91.4% and 1.1%, respectively. A leaching toxicity test for ZLR was performed after water leaching. The results indicated that the final residue was effectively detoxified and all of the heavy metal leaching concentrations were under the allowable limit.     

回转窑内渣成分优化实现锌铅元素的高效回收（英文）

在回转窑内从高硅锌浸出渣中实现高效回收锌铅金属。XRD、SEM、EDS及ICP表征结果表明,浸出渣含12.4%SiO₂、16.1%Zn和7.4%Pb(质量分数)。热力学分析表明,在回转窑内1150～1250℃的冶炼环境下,锌浸出渣中极易产生锌和铅的金属蒸气。通过分析13种冶金渣成分的黏度及熔点,发现3种渣成分(47%SiO₂-23%CaO-30%FeO、40%SiO₂-28%CaO-32%FeO、40%SiO₂-30%CaO-30%FeO,质量分数)具有合适的物理特性,即熔点和黏度分别为1150～1280℃和0.2～0.5 Pa·s。工业实验表明,采用优化的渣成分,锌和铅的回收率分别高达97.3%和94.5%,冶炼后渣内锌和铅的含量分别低至0.51%和0.45%(质量分数)。水淬渣的国标浸出实验表明,浸出液中含1.82 mg/L Zn、～0.01 mg/L Cu、0.0004 mg/L As、～0.01 mg/L Cd、0.08 mg/L Pb和～0.02 mg/L Hg,证实锌浸出渣同步实现无害化...     

湿法炼锌净化渣酸浸的动力学

对湿法炼锌净化渣的浸出动力学进行了研究,并探讨了硫酸浓度、反应温度、粒度等对钴、锌浸出率的影响规律。从动力学的角度分析了整个浸出过程,得到优化条件：液固比50：1(mL/g),硫酸浓度100 g/L,反应温度70°C,粒度75~80μm,反应时间20 min。在此优化条件下钴的浸出率为99.8%,锌的浸出率为91.97%。结果表明：在硫酸体系中钴的浸出符合不生成固体产物层的“未反应收缩核”模型。通过 Arrhenius 经验公式求得钴和锌表观反应活化能分别为11.693 kJ/mol和6.6894 kJ/mol,这表明浸出过程受边界层扩散控制。     

氧化锌贫矿提锌渣中铅和银的氯盐一步浸出

运用X射线衍射、扫描电镜和X射线能谱等分析手段,对山东某地深度氧化锌贫矿提锌后渣进行工艺矿物学特征分析可知,矿物中金属赋存状态复杂,属难选矿物。开发出氯盐一步法浸出铅和银的新工艺,考察反应温度、NaCl浓度、添加剂用量、液固比、HCl加入量和浸出时间对浸出过程的影响。结果表明:加入添加剂对铅的浸出率没有影响,但可以显著提高银的浸出率。条件试验研究得出最佳工艺条件如下:浸出温度90℃、NaCl浓度390 g/L、添加剂用量15 mL、液固比(质量比)7?1、HCl加入量3 mL、浸出时间3 h。在此最佳工艺条件下,铅的浸出率达到95%左右,银的浸出率达到90%左右。     

硫化锌精矿常压富氧直接浸出行为

借助工艺矿物学分析对常压富氧直接浸出条件下锌精矿中主要硫化物的浸出行为进行研究。结果表明,除黄铁矿外,其他硫化矿均会明显溶解。基于对浸出渣中单质硫与反应残余硫化物之间关系的分析,认为闪锌矿、黄铜矿、铜蓝、方铅矿的溶出可能遵循间接氧化方式,即硫化物首先酸溶,生成的H2S脱离矿物表面并迁移至溶液本体中进而氧化成单质硫。上述硫化矿的浸出过程可能受界面化学反应控制。对于磁黄铁矿的溶出,直接电化学氧化可能起主导作用,其浸出过程可能受产物层单质硫的扩散控制。     

锌浸出渣中重金属的环境活性和生态风险评价

采用矿物学分析、BCR三步连续浸提、动态淋洗实验以及Hakanson潜在生态风险评价4种方法对锌浸出渣重金属的环境活性以及生态风险进行评价.结果表明,锌浸出渣中重金属的环境活性大小依次为Cd＞Zn＞Cu＞As＞Pb.废渣中主要重金属的潜在生态风险评价表明,该种废渣对环境具有很高的生态风险,单个重金属的生态危害顺序为Cd＞Zn＞Cu＞As＞Pb.Cd有很严重的生态风险,是对生态环境造成毒性的主要原因.     

Pressure leaching of zinc silicate ore in sulfuric acid medium

Zinc silicate ore was characterized mineralogically and the results showed that zinc exists mainly as hemimorphite and smithsonite in the sample. Sulfuric acid pressure leaching of zinc silicate ore was carried out to assess the effect of particle size, sulfuric acid concentration, pressure, reaction time and temperature on the extraction of zinc and the dissolution of silica. Under the optimum conditions employed, up to 99.25% of zinc extraction and 0.20% silica dissolution are obtained. The main minerals in leaching residue are quartz and small amounts of undissolved oxide minerals of iron, lead and aluminum are associated with quartz.     

Leaching and recycling of zinc from liquid waste sediments

The selective leaching and recovery of zinc in a zinciferous sediment from a synthetic wastewater treatment was investigated. The main composition of the sediment includes 6% zinc and other metal elements such as Ca, Fe, Cu, Mg. The effects of sulfuric acid concentration, temperature, leaching time and the liquid-to-solid ratio on the leaching rate of zinc were studied by single factor and orthogonal experiments. The maximum difference of leaching rate between zinc and iron, 89.85%, was obtained by leaching under 170 g/L H2SO4 in liquid-to-solid ratio 4.2 mL/g at 65 ℃ for 1 h, and the leaching rates of zinc and iron were 91.20% and 1.35%, respectively.     

铟的还原浸出动力学及分步沉淀选择性分离（英文）

开展硫化锌精矿还原浸出高铁锌浸出渣高效浸铟及浸出液中铟选择性分离的研究。结果表明：在固体物料粒度74～105μm、反应温度90℃、浸出时间300 min、硫酸浓度1.4 mol/L的条件下,铟的浸出率达95%以上。采用收缩核模型对还原浸出动力学进行分析,不同条件下的浸出实验结果表明反应受穿过固体产物层的扩散控制,活化能为17.96 k J/mol,相对于硫酸浓度的反应级数为2.41。铁粉置换沉铜过程铜和砷的沉淀率均达99%以上。98%以上的铟从含高亚铁离子浓度的硫酸锌溶液中选择性分离,获得铟含量约为2.4%的富铟渣,经酸浸-萃取-电积工艺流程进一步处理后可得到纯铟。     

湿法炼锌废渣中硫脲浸出银的动力学

探讨了从湿法炼锌废渣中用硫脲浸出回收银的浸出反应动力学,从这种难浸的含银炼锌废渣中用硫脲浸出回收银,浸出反应是一种典型的氧化还原反应并可充分进行,同时,通过动力学推导得出,从湿法炼锌废渣中用硫脲浸出银的反应动力学模型为收缩核动力学模型,同时计算出浸出活化能为13．26kJ/mol。该模型反映了浸出过程中控制整个反应速率的决定步骤是固膜扩散速率,并较好地说明了浸出机理。     

Recovery of zinc and lead from Yahyali non-sulphide flotation tailing by sequential acidic and sodium hydroxide leaching in the presence of potassium sodium tartrate

The recovery of zinc and lead from Yahyali non-sulphide flotation tailing using sulfuric acid followed by sodium hydroxide leaching in the presence of potassium sodium tartrate was experimentally investigated. In the acidic leaching stage, the effects of pH, solid-to-liquid ratio and temperature on the dissolution of zinc from the tailing were explored. 82.3% Zn dissolution was achieved at a pH of 2, a temperature of 40 °C, a solid-to-liquid ratio of 20% and a leaching time of 2 h, whereas the iron and lead dissolutions were determined to be less than 0.5%. The sulfuric acid consumption was found to be 110.6 kg/t (dry tailing). The leaching temperature had no beneficial effect on the dissolution of zinc from the tailing. The acidic leach solution was subjected to an electrowinning test. The cathode product consisted of 99.8% Zn and 0.15% Fe. In the alkaline leaching stage, the Pb dissolution increased slightly in the presence of potassium sodium tartrate. More than 60% of Pb was taken into the leach solution when the leaching temperature increased from 40 to 80 °C. The final leach residue was analyzed by XRD and XRF. The XRD results indicated that the major peaks originated from the goethite and quartz while minor peaks stem from smithsonite and cerussite. The XRF analysis demonstrated that the residue contained 70.3% iron oxide. Based on the sequential leaching experiments, the zinc and lead were excellently depleted from the flotation tailing, leaving a considerable amount of iron in the final residue.     

Selective recovery of zinc from zinc oxide dust using choline chloride based deep eutectic solvents

Deep eutectic solvents (DESs) are a kind of potential lixiviant for selective metal processing due to their versatile complexation properties. In this study, we investigated the recovery of zinc from zinc oxide dust using choline chloride−urea− ethylene glycol (ChCl−urea−EG) DESs. The zinc extraction efficiency can be up to 85.2% when the slurry concentration is 50 g/L, leaching temperature is 80 °C and stirring speed is 600 r/min. The leaching process is controlled by the diffusion and the corresponding activation energy is 32.1 kJ/mol. The resultant solution was directly used for the electrodeposition of zinc. The pure zinc deposit is obtained with a current efficiency of 82.6%. Furthermore, the ChCl−urea−EG DESs can be recycled. This approach is shown to be promising for the recycling of zinc from the zinc-containing dust.     

Physicochemical and environmental characteristics of alkali leaching residue of wolframite and process for valuable metals recovery

Physicochemical and mineralogical characteristics of an alkali leaching residue of wolframite were studied by XRD, SEM?EDS, chemical phase analysis, mineral liberation analyzer (MLA), and TG?DSC methods. Batch leaching tests, toxicity characteristic leaching procedure (TCLP) tests and Chinese standard leaching tests (CSLT) were conducted to determine the environmental mobility of toxic elements. The results show that, due to the high contents of W, Fe, Mn, Sn, and Nb, the residue is with high resource value, but the content of a toxic element, As, is also high. The existing minerals of the investigated elements mainly occur as monomer particles, but it is difficult to extract these valuable metals by conventional acid leaching due to their mineral properties. The release of As increases over time in acidic environment. The leaching concentration of all investigated harmful elements through TCLP is within the limiting value, while the leaching concentration of As through CSLT exceeds the limiting value by more than 4 times, so the residue is classified as hazardous solid waste based on the Chinese standard. A process for valuable metals recovery from this residue was proposed. Preliminary experimental results indicated that the main valuable metals could be extracted effectively.     

Hydrometallurgically recovering zinc from electric arc furnace dusts

The increasing use of zinc-containing scrap for steel production has lead to a high zinc content in the electric arc furnace and converter flue dusts. The cost of disposing of this residue is high due to environmental restrictions. Various recycling processes have been developed for these dusts, but most never reached the pilot plant stage and many investigations were stopped because of metallurgical and economical inefficiencies. While pyrometallurgical methods have to deal with high energy consumption, low zinc yield, and valueless residues, hydrometallurgical processes could offer an economical recycling alternative. This paper describes hydrometallurgical methods for recovering zinc from steel industry dust. These methods can be integrated in the primary zinc-winning process or in galvanization. Investigations of sulfuric-acid leaching show high zinc solubility but also a high iron content in the final liquor. As a result, steps for purification are required that cannot be conducted economically. Alternatively, a NaOH leaching gives a satisfying zinc yield and a very low solubility for the iron that remains in the residue. For more information, contact J. Antrekowitsch, University of Leoben, Nonferrous Metallurgy Department, Franz-Josef-Str. 18, Leoben, Styria A-8700, Austria; +43-3842-402322; fax +43-3842-402627; e-mail antreko@unileoben.ac.at.     

Treatment of cadmium dust with two-stage leaching process

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Iron control in zinc pressure leach processes

The occurrence of zinc in sulfide ore deposits is generally accompanied by various iron minerals. Hence, even the most efficient concentrators generally produce a zinc concentrate with significant iron content. The efficient recovery of zinc metal from zinc concentrates requires the rejection of iron residue in a form that minimizes the zinc entrainment. Careful control of the iron precipitation step is important, so that the iron residue produced is amenable to efficient liquid-solid separation in order to obtain high zinc recoveries. In hydrometallurgical zinc processes, the coprecipitation of minor impurities along with iron precipitation is also important in producing zinc-sulfate solution from which high-purity zinc cathode can be electrowon. The integration of Dynatec’s zinc pressure leach process with existing roast-leach-electrowin plants employing various methods of iron rejection is briefly described in this article, along with the application of two-stage pressure leaching in stand-alone processes. For more information, contact W.D. Vardill, Dynatec Corporation, 8301-113 Street, Fort Saskatchewan, Alberta, Canada T8L 4K7; (780) 992-8190; fax (780) 992-8100; e-mail wvardill@mettech.dynatec.ca.     

Increasing silver leaching rate from leaching-resistant zinc residues by thiourea leaching method with pressurized preoxidation process

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Extraction of manganese and zinc from their compound ore by reductive acid leaching

Comprehensive utilization of low grade manganese–zinc compound ore containing lead and silver with a method of reductive acid leaching was studied. According to the ?–pH diagram of Mn–Zn–H₂O system, Mn and Zn can be leached simultaneously in the pH range of –2 to 5.61. The results showed that both hydrogen peroxide and sucrose were effective reductants which could intensify the simultaneous leaching of Mn and Zn into leachate as well as enrich Pb and Ag in the residue. 95.88% of Mn and 99.23% of Zn were extracted when the compound ore was leached with hydrogen peroxide in sulfuric acid media, meanwhile the contents of Pb and Ag in the residue were enriched to 13.21% and 489.36 g/t, respectively. When sucrose was used as the reductant, the leaching efficiencies of Mn and Zn separately achieved 98.26% and 99.62%, and contents of Pb and Ag in the residue were as high as 13.92% and 517.87 g/t, respectively.     

用碱性Na_2EDTA溶液从次氧化锌烟灰中选择性回收铅(英文)

采用碱性Na2EDTA溶液从次氧化锌烟灰中回收铅。探讨温度、浸出时间、Na2EDTA浓度和起始NaOH浓度对铅、锌浸出率的影响。得到最优实验条件如下:液固比5:1 mL/g、搅拌速度650 r/min、Na2EDTA浓度0.12mol/L、NaOH初始浓度0.5 mol/L、温度70°C、浸出时间120 min。在最优实验条件下,铅、锌、氟和氯的平均浸出率分别为89.92%、0.94%、62.84%和90.02%。浸出液用于电沉积铅粉。在温度为60°C、电流密度为200A/m2、H3PO4浓度为1.5 g/L、铅离子浓度不低于5 g/L时,电沉积铅粉平均电流效率大约为93%,阴极铅纯度高于98%。电沉积1 kg铅粉大约消耗0.218 kg Na2EDTA和0.958 kW·h电能。     

Ida~(2-)-H_2O体系浸出低品位氧化锌矿(英文)

采用Ida2--H2O体系(亚氨二乙酸盐水溶液)处理高碱性脉石型低品位氧化锌矿,考察浸出时间、液固比、配体总浓度、温度及pH值对矿物中主金属Zn及杂质元素Ca、Mg、Cu、Ni、Fe、Pb、Cd的溶出影响。结果表明:在弱碱性Ida2--H2O体系中,Ca、Mg、Fe不会被大量溶出,有价金属Cu、Ni、Pb、Cd可部分随主金属Zn溶出而进入浸出液;在浸出时间4h、液固比5:1、配体总浓度0.9mol/L、温度70℃、pH8的优化条件下,锌浸出率为76.6%。