赤泥酸浸的试验研究

以HCl和H2SO4为浸出剂,对赤泥中各有价金属的浸出条件进行了较系统的试验研究。结果表明,先采取低浓度HCl浸出,残渣再用高浓度H2SO4分解的两段浸出方式为佳。一段浸取盐酸浓度为6mol／L．液固比L／S=4：1．反应温度50℃,反应时间1h。Sc2O3的浸出率大于80％,TiO2的浸出率为1％左右。二段浸取硫酸浓度为92％．酸渣比为3：1．熟化温度为200℃,熟化时间为1．5h。TiO2的浸出率为96．57％。     

微波预处理对载醋酸锌废触媒锌浸出的影响

提出微波预处理废触媒-酸浸提锌新工艺,测定废触媒在微波场中的温升曲线,探索微波预处理温度和保温时间对浸出率的影响,并对微波预处理废触媒机理进行分析.结果表明:微波预处理可显著提高锌的浸出率,当微波预处理温度和时间分别为950 ℃和12 min时,锌的浸出率达到96.58%.微波预处理打开了堵塞的废触媒孔道,增大了浸出剂与锌的接触面积.     

煤在铁闪锌矿氧压酸浸中的应用

研究煤在铁闪锌矿氧压酸浸中的应用。实验发现,含碳量高于70％的煤无助于铁闪锌矿浸出。在温度为423K时,低碳煤（褐煤）可以用作铁闪锌矿氧压酸浸中优良的硫分散剂,其用量选定1．O％（相对于精矿质量）为宜,在98％以上精矿粒度小于43）am,液固比为4．5：1mL／g,浸出剂中初始硫酸浓度为1．43mol／L,锌浓度为0．62mol／L,浸出温度为423K,氧分压为0．5MPa,搅拌转速为800r／min,浸出1．5h条件下,添加1．0％褐煤后,锌浸出率达到96％,上述工艺运行高效、稳定。褐煤具有比木质素磺酸钠更强的分散单质硫的能力,并能基本消除单质硫对浸出残余硫化矿的沾染。     

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。     

中度嗜热微生物浸出复杂多金属铜精矿的优化(英文)

研究了中度嗜热微生物浸出复杂多金属铜矿过程中初始pH、温度、装液量、转速、原电池效应(黄铁矿比例)和矿浆浓度的影响。结果表明,初始pH为1.5的铜浸出率分别是初始pH为1.0和2.0的实验组浸出率的1.5倍和1.4倍。当温度为45°C时铜的浸出率比温度为50°C时的高出1236.8%。随着转速的升高和装液量的降低,铜的浸出率明显提高。当黄铁矿的比例增大时,铜的浸出率也逐渐提高,但是当黄铁矿比例高于20.0%后,铜浸出率不再增加。温度和pH及温度和黄铁矿比例之间存在显著的交互作用。     

从含银的废催化剂中回收白银

介绍了从含银的废催化剂中回收白银的一种新方法。采用氨浸，A还原剂还原，浸出率98．5％，还原率99．9％，直接可获高纯度的银粉，金属银纯度可达99．9％。     

Pb/Zn冶炼废渣中有价金属生物浸出条件优化

为提高生物浸出Pb／Zn冶炼废渣中有价金属的浸出率,利用正交设计,通过摇瓶实验,研究微生物浸出某Pb／Zn冶炼废渣过程中温度、pH值、废渣浓度及浸出时间等对废渣中Cu,Zn,In,Ga,Pb和Ag等有价金属浸出的影响。结果表明,在pH值为1．5、废渣浓度为5％、温度为65℃的优化浸出条件下生物浸出4d,Pb／Zn冶炼废渣中有价金属Cu,Zn,In和Ga的浸出率分别达到95．5％,93．5％,85．0％和80．2％,而Pb和Ag则主要以硫酸铅、黄钾铁矾类物质或硫化银形式富集在余渣中。     

德兴铜矿尾矿库重金属的浸出

德兴铜矿的4#尾矿库废水来源包括碱性浮选矿浆和附近露天矿山的酸性废水（AMD）。因此,酸性矿山废水的酸化作用极可能造成尾矿中重金属的浸出。通过实验室的批次实验对德兴铜矿尾矿中锌、铜、铁和锰的浸出行为进行了研究,对pH值、温度、颗粒大小和接触时间对重金属浸出的影响进行了讨论。结果表明,锌、铜、铁和锰是尾矿中的主要重金属,石英类脉石矿物是尾矿的主要成分。尾矿的溶解反应受酸度控制,其动力学可以根据非均相反应的模型来表达,通过表面化学反应作为速率决定步骤的核收缩模型来解释。这些重金属的浸出依赖于pH值和接触时间。批次实验研究得出,锌、铜、铁和锰在pH=2．0时的最大浸出率分别为5．4％、5．8％、11．1％和34．1％。重金属的溶出与温度呈正相关性。颗粒大小不会改变这些重金属的浸出趋势,但会略微导致重金属的最高浓度和平均水平的下降。     

株冶集团将投资6．58亿元建设搭配锌浸出渣Kivcet直接炼铅项目

株洲冶炼集团股份有限公司3月8日召开2008年度董事会,审议通过了关于建设搭配锌浸出渣Kivcet（基夫赛特）直接炼铅项目的提案。项目设计生产能力为12万吨／年粗铅、13．6万吨／年硫酸、1．93万吨／年次氧化锌,总建筑面积29251平方米;搭配处理锌直接浸出尾矿渣9．6万吨／年、硫化物热滤渣2万吨／年。     

采用碱浸预脱硅-氰化工艺从氰化尾渣中回收金

采用高浓度碱浸对氰化尾渣进行预脱硅处理,考察搅拌速度、固液比、Na OH浓度及温度对硅浸出速率的影响,研究脱硅过程的反应动力学,得到相应的动力学方程。结果表明:当搅拌速度为400 r/min、固液比为1:5、Na OH浓度为80%、反应温度为280℃时,二氧化硅的浸出率为91.8%;碱浸过程受产物层内扩散控制,表观反应活化能为37.375 k J/mol。通过正交实验对氰化浸金的条件进行了优化,在Si O2浸出率为91.8%,Na CN浓度为1.5 g/L,固液比为1:3,浸出时间为48 h的条件下,金的浸出率为87.83%。     

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.     

Coupling leaching of sphalerite concentrate

Coupling process of sphalerite concentrate leaching in H2SO4-HNO3 and tetrachloroethylene extracting of sulfur was investigated. Effects of leaching temperature, leaching time, mass ratio of liquid to solid and tetrachloroethylene addition on zinc leaching processes were examined separately. SEM images of sphalerite concentrate and residues were performed by using JEM-6700F field emission scanning electron microscope. The relationship between the number of recycling and extraction ratio of zinc was studied. The results indicate that 99.6% zinc is obtained after leaching for 3 h at 85℃ and pressure of 0.1MPaO2, with 20g sphalerite concentrate in 200 mL leaching solution containing 2.0mol/L H2SO4 and 0.2mol/L HNO3, in the presence of 10 mL C2Cl4. The leaching time of zinc is 50% shorter than that in the common leaching. The coupling effect is distinct. The recycled C2Cl4 exerts little influence on extraction ratio of zinc.     

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.     

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%。     

次氧化锌中砷的物相及砷的浸出(英文)

研究烟化炉次氧化锌中砷的物相类型。结果表明:按砷的物相可将次氧化锌分为3种类型。在一型次氧化锌中砷以As2O3形态存在,而在二型和三型次氧化锌中砷分别以亚砷酸锌(Zn(AsO2)2)和砷酸铅(Pb(As2O6),Pb4As2O9)形态存在。在热力学分析基础上,对二型次氧化锌进行浸出脱砷。结果表明:采用30g/LNaOH溶液,在液固比3、温度20°C的条件下,砷的浸出率在1h内可达到65%～70%,而铅、锌的损失均小于1%。     

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。     

Ida~（2-）-H_2O体系浸出低品位氧化锌矿

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%。     

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。     

A new extraction process of carbonaceous refractory gold concentrate

A new hydrometallurgical process for a carbonaceous refractory gold concentrate at ambient temperature and pressure was presented, including grinding-leaching, intensified alkaline leaching（IAL）, thiosulfate leaching and cementation by zinc powder. The experimental results show that the grinding-leaching and intensified alkaline leaching process result in the selective oxidation of arsenopyrite and pyrite. The oxidation ratio of As is 96.6%, and 46.7 % for S. The total consumption of NaOH in alkaline leaching is only 28 % of that theoretically calculated under the conditions of full oxidization for the same amount of arsenopyrite and pyrite transforming into arsenates and sulfates, and 83.6% of gold is synchro-dissoluted by thiosulfate self-generated during pretreatment. Since the carbonaceous matter in concentrate possesses a strong capability of preg robbing, the cyanidation process is not suitable for the extraction of gold after pretreatment. However, the gold leaching rate by thiosulfate leaching for 24 h is increased to 91.7% from 0 - 3.2% by ultra-fine grinding without the pretreatment. The recovery of gold by zinc cementation gets to 99.6%. Due to the thiosulfate self-generated during alkaline leaching, the reagent addition in thiosulfate leaching afterwards is lower than the normal one.