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.     

Extraction of cobalt from laterite ores by citric acid in presence of ammonium bifluoride

Citric acid was used to selectively extract cobalt from limonite-type laterite ores in the presence of ammonium bifluoride. The results show that ammonium bifluoride enhances the leaching of cobalt by citric acid, and 84.5% cobalt is extracted from a laterite ore containing 0.13% Co when leached at ambient temperature for 2 h with 30 g/L citric acid and 10 g/L ammonium bifluoride. Pyrolusite is reduced by citric acid during leaching, cobalt intergrown with which is liberated and subsequently chelated by the citric acid. The extraction of cobalt is enhanced in the presence of ammonium bifluoride because the matrix of silicate minerals is destroyed by ammonium bifluoride and the adsorbed cobalt is subsequently liberated.     

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%。     

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。     

Leaching of complex sulphide concentrate in acidic cupric chloride solutions

The chemical analysis of a complex sulphide concentrate by emission spectrometry and X-ray diffraction shows that it contains essentially copper, lead, zinc and iron in the form of chalcopyrite, sphalerite and galena. A small amount of pyrite is also present in the ore but does not be detected with X-ray diffraction. The cupric chloride leaching of the sulphide concentrate at various durations and solid/liquid ratios at 100 ℃ shows that the rate of dissolution of the ore is the fastest in the first several hours, and after 12 h it does not evolve significantly. If oxygen is excluded from the aqueous cupric chloride solution during the leaching experiment at 100 ℃, the pyrite in the ore will not be leached. The determination of principal dissolved metals in the leaching liquor by flame atomic absorption spectrometry, and the chemical analysis of solid residues by emission spectrometry and X-ray diffraction allow to conclude that the rate of dissolution of the minerals contained in the complex sulphide concentrate are in the order of galena 〉 sphalerite 〉 chalcopyrite.     

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%。     

镍红土矿高压酸浸过程的金属元素浸出行为

以镍、钴的提取为目的,研究褐铁矿型镍红土矿高压酸浸过程中各金属元素的浸出行为,探讨硫酸加入量、浸出温度、浸出时间及液固比对各金属元素浸出率的影响.实验结果表明,在优化条件下Ni、Co、Mn和Mg的浸出率分别达到97%、96%、93%和95%以上,则Fe的浸出率小于1%.对高压浸出渣的分析表明,渣中的铁和硫主要分别以赤铁...     

机械活化对铟铁酸锌溶解动力学及物化性质的影响

以人工合成的高纯度铟铁酸锌为研究对象,采用搅拌球磨对其进行机械活化。以锌的浸出率为评价指标,研究不同活化试样在硫酸溶液中的浸出特性和溶解动力学,并利用X射线衍射仪、扫描电镜、激光粒度分析仪和比表面积分析仪分别考察机械活化对铟铁酸锌晶体结构、颗粒形貌、粒度和比表面积的影响。结果表明:机械活化使铟铁酸锌的物化性质发生明显改变,并由此提高了铟铁酸锌的反应活性。经机械活化30和60 min后,铟铁酸锌与硫酸反应的表观活化能由未活化时的76.4 kJ/mol分别降至58.6和51.8 kJ/mol,表观反应级数也由原来的0.79分别降至0.62和0.59。未活化铟铁酸锌的酸溶过程受化学反应控制,活化后则为混合控制。     

高压酸浸镍钼矿提取钼和镍

研究一项针对镍钼矿用高压酸浸的方法回收镍和钼的全湿法工艺。采用该工艺避免了传统上艺焙烧镍钼矿（15％～25％s）带来的人量S02和As2O3排放,减小了对环境的污染;与现有的湿法碱浸回收钼工艺相比,本工艺存酸浸过程中回收了儿乎全部的镍和人部分的钼。在氧压环境下,几乎全部的镍和大部分的钼都进入溶液,少部分的钼留在酸浸渣中,睃浸渣进一步用碱（NaOH）浸出。在最佳的实验条件下,97％的镍和96％的钼分别被浸出。     

氧压酸浸处理锌焙砂中浸渣的新工艺研究

锌焙砂一般采用中性-低酸-高酸三段浸出工序,该工艺在酸浸出中浸渣的过程中,铁也大量浸出进入到溶液中,加重了净化电积前除铁的负担。通过将传统锌湿法冶金工艺与氧压酸浸新工艺相结合,研究了氧压酸浸处理中浸渣的氧气压力、硫酸浓度、温度、浸出时间、粒度、液固比和分散剂等相关因素的影响。实验结果表明该工艺不仅提高了锌的浸出率（〉98％）,降低了铁的浸出率（〈50％）,缩短了生产周期,降低了生产成本,具有良好的经济效益;而且还具有环境友好和资源利用率高等优点,实现了简化工艺和节能减排的目的,为工业化生产提供了参考．     

Selective flotation of smithsonite,quartz and calcite using alkyl diamine ether as collector

The flotation separation of smithsonite from calcite and quartz using a alkyl diamine ether (GE-609) as the collector was investigated through micro-flotation experiments and the real ore flotation experiments. The results show that GE-609 exhibits good collecting capability to three minerals without selectivity. The presence of sodium sulfide enhances the flotation of smithsonite and calcite while inhibits quartz. Moreover, both sodium silicate and sodium hexametaphosphate exhibit good selective inhibition to calcite. The real ore test results show that a zinc concentrate containing 23.51% Zn with the recovery of 71.02% is obtained in the closed-circuit test. To understand the adsorption of GE-609 on smithsonite surface, zeta potential measurement and FTIR analysis were carried out, and the results indicate that the collector GE-609 can adsorb on smithsonite surface through both electrostatic adsorption and chemical adsorption, and the presence of sodium sulfide enhances the adsorption of GE-609.     

Improving zinc recovery by thermoacidophilic archaeon Acidianus copahuensis using tetrathionate

The attachment and bioleaching experiments were conducted to evaluate the zinc recovery from Hualilan ore by the thermoacidophilic archaeon Acidianus copahuensis. Cells of this species pregrown on tetrathionate showed higher capability of attachment to the ore than cells pregrown on other energy sources and such attachment seemed to be mediated by the product of extracellular polymeric substances. A. copahuensis achieved a successful bioleaching of the ore reaching 100% of zinc recovery when tetrathionate was added. Simultaneous addition of yeast extract and tetrathionate maintained the zinc extraction at higher rate. Zinc dissolution kinetics was controlled by chemical reaction in cultures with the external addition of tetrathionate but by the diffusion through a product layer of jarosite in the other cultures.     

Pressure leaching of converter vanadium slag with waste titanium dioxide

The process of pressure leaching the converter vanadium slag with waste titanium dioxide without roasting was studied. Mineralogy analysis indicates that the converter vanadium slag contains mainly three mineral phases: magnetite, titanium magnetite, and silicate phase. Vanadium is in combination with iron, titanium, manganese, aluminum, and silicon. The impacts of leaching temperature, leaching time, stirring speed, liquid-to-solid ratio, and initial leaching agent concentration were investigated on the waste titanium dioxide leaching process. The results indicate that under the optimal conditions, the vanadium and the iron leaching rates are 96.85 % and 93.50 %, respectively, and the content of titanium is 12.6 % in the residue. The main mineral phases for the residues under the optimal operation conditions are quartz, ilmenite, anatase, and silicate phase, and the residues can be reused as the extraction of titanium raw materials for titanium dioxide production technology by the sulfate method.     

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。     

锌浸出渣中银的浮选回收综述

锌浸出渣是湿法炼锌过程中产生的固体废弃物,其含有的银具有巨大的经济价值,浮选回收银是重要的途径。锌浸出渣中浮选回收银,面临着含银矿物种类多而分散、矿物粒度细、酸性强及难免离子含量高等综合叠加影响。部分锌浸出渣可采用直接浮选法回收银,但对银赋存状态异常复杂的锌浸出渣,直接浮选往往难以实现银的高效选别回收,多采用进行一定的预处理后再浮选的间接浮选法。常用的预处理技术主要有浆洗、磨矿、外加载体、焙烧及热酸浸出等,预处理可针对性的调控矿浆环境、改变矿物的赋存状态及矿物表面性质,以提高选择性和捕收能力。各种预处理都有其独特的优缺点,合适的预处理技术结合湿法炼锌工艺或预处理技术的有机联合应用,对浮选回收锌浸出渣中的银将更为有效。     

Technology and mechanism of desilication from roasted diasporic bauxite in atmosphere

The leaching desilication technology of roasted diasporic bauxite in atmosphere by caustic soda solution was investigated. The optimum parameters were: the grinding fineness of the roasted bauxite -0. 076 mm and 80 % -85 %,leaching time 2h, Na2Ok100-150g/L, L/S 4-5, leaching temperature 90-95℃. The desilication rate 55.20% and concentrate A/S (mass ratio of A12O3 to SiO2) 9.90, as good as those obtained at pressure, were obtained respectivdy.Investigation of two-stage leaching shows that it can both improve desilication rate of roasted ore and reduce leaching time.When time of the first stage and the second stage is 30 min and 30 min respectively, desilication rate can reach 59.65 %.X-ray diffraction analysis of the concentrate has proved that desilication procedure is accompanied with the formation of sodium aluminosilicate hydrate. X-ray spectra also show that silica removed during leaching is amorphous silica. SiO2 occurrina as ouartz in raw ore or mullite formed during roasting can not dissolve in alkali solution.     

An investigation of oxygen pressure acid leaching of Gacun complex Cu-Pb bulk concentrate

The treatment of the Gacun complex Cu-Pb bulk concentrate with high Zn,Ag,etc.,by oxygen pressure acid leaching was studied.The primary copper and lead minerals in the concentrate are tetrahedrite and galena.The treatment of tetrahedrite was rarely studied,and most of silver occurred in the mineral too.The optimum operating parameters of oxygen pressure acid leaching were established by conditional tests.Under these parameters,the result of pilot scale test showed that the leaching percentages of copper and zinc were separately as high as 98.9 wt.％ and 94.9 wt.％,while lead and silver were transformed into sulfate and sulfide precipitations,respectively.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 carbonate transformation-silicofluoric acid leaching and thiourea leaching.     

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。     

三水型铝土矿盐酸浸出过程的热力学研究

针对低品位三水型铝土矿使用拜耳法提取氧化铝存在溶出率低的情况,采用盐酸法浸出,利用X衍射分析和半定量分析手段研究铝土矿的矿物组成和化学组成,并对矿石中主要有价金属的浸出反应进行热力学分析.结果表明:铝土矿的主要成分为三水铝石,其次是针铁矿和石英,以及锐钛矿等,其主要矿物三水铝石和针铁矿的酸浸反应很容易进行,在配矿液固比100∶7、浸出时间2h、矿石粒度小于55 μm、盐酸浓度10％、浸出温度373～383K的条件下矿物中铁和铝的浸出率均在95％以上,浸出渣主要成分为SiO2,易于综合利用.     

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.