Leaching performance of Al-bearing spent LiFePO4 cathode powder in H2SO4 aqueous solution

The leaching performance and leaching kinetics of LiFePO₄ (LFP) and Al in Al-bearing spent LFP cathode powder were systematically studied. The effects of temperature (273?368 K), stirring speed (200?950 r/min), reaction time (0?240 min), acid-to-material ratio (0.1:1?1:1 mL/g) and liquid-to-solid ratio (3:1?9:1 mL/g) on the leaching process were investigated. The results show that the concentration of reactants and the temperature have a greater impact on the leaching of Al. Under the optimal conditions, leaching efficiencies of LFP and Al are 91.53% and 15.98%, respectively. The kinetic study shows that the leaching of LFP is kinetically controlled by mixed surface reaction and diffusion, with an activation energy of 22.990 kJ/mol; whereas the leaching of Al is only controlled by surface chemical reaction, with an activation energy of 46.581 kJ/mol. A low leaching temperature can effectively suppress the dissolving of Al during the acid leaching of the spent LFP cathode material.     

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。     

Characterization and kinetic study on ammonia leaching of complex copper ore

Ammonia leaching kinetics of a complex Cu-ore assaying 8.8% Cu and 36.1% Fe was examined. Mineralogical characterization indicated that the major phase of the ore was siderite with chalcopyrite as the major sulfide mineral. The effects of parameters such as agitation, temperature, NH₃ concentration, particle size and oxygen partial pressure (p_O2) were investigated. Under the standard leaching conditions of 125–212 µm particle size, 120 °C, 1.29 mol/L NH₃ and 202 kPa of p_O2, about 83% Cu could be selectively extracted in 2.5 h. However, when using higher NH₃ concentration and lower particle size, more than 95% extraction was achieved. The leaching process was found to be surface reaction controlling. The estimated activation energy was (37.6±1.9) kJ/mol and empirical orders of reaction with respect to p_O2 and [NH₃] were about 0.2 and 1, respectively.     

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。     

有机溶剂在浆萃取对锌精矿直接浸出的影响

研究了锌精矿直接浸出时的在浆有机溶剂萃取，通过扫描电镜分析比较了浸出前矿粒以及两种不同浸出渣的表面形貌。结果表明，四氯乙烯是一种非常有效的萃硫有机溶剂，其加入时间控制为0．5－1．5h，当有机相与无机相（O／A）之比值小于1／50时，锌浸出率随O／A比值的增大而增大；但当O／A比值超过1／50时，锌浸出率并不受相比的影响。     

Treatment of cadmium dust with two-stage leaching process

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Role of oxygen and ammonium ions in silver leaching with thiosulfate–ammonia–cupric ions

The dissolution of silver with thiosulfate and copper ions was studied in a batch-stirred reactor at 25 °C in the absence and presence of various ammonium ions and oxygen concentrations in the aqueous solution. Leaching silver with 4 mg L-1oxygen in the aqueous solution requires high concentration of ammonium and thiosulfate ions. High silver dissolution is achieved at pH of high cupric tetraamine concentration, around pH 9.8, but a high degradation of thiosulfate ions is resulted from their oxidation by the presence of oxygen. When the oxygen concentration is below 1 mg L-1, dissolved silver thiosulfate species are fairly stable and the degradation of thiosulfate ions in the aqueous solution is low. Under these conditions,high silver dissolution is attained using low thiosulfate concentration. At the oxygen concentration of 4 mg L-1,dissolved silver thiosulfate species precipitate out of the aqueous solution significantly affecting silver dissolution.Precipitation of dissolved silver thiosulfate species occurs because of oxidation of the thiosulfate bonding to the silver.     

烧结电除尘灰中钾盐的回收及其浸出动力学

为了研究烧结电除尘灰中回收钾盐的强化浸出措施,使用ICP-AES、SEM-EDS和XRD分析技术对除尘灰的表面和内部形态,特别是钾盐的赋存形式进行分析。结果表明,该电除尘灰的主要成分是铁氧化合物,在其表面裸露吸附着一定含量的KCl晶体。水浸实验表明,该粉尘中的KCl可以通过水浸出、蒸发结晶的方式回收,其收率为18.56%。结晶产物的分析结果表明,KCl占61.21%,NaCl占13.40%,CaSO4占14.62%,K2SO4占10.86%。其水浸出动力学符合外扩散控制模型控制。强化浸出实验表明,提高浸出温度、加强搅拌、增加液固比等措施可以提高钾盐的浸出率和浸出速率。     

钒钛磁铁矿提钒尾渣浸取钒

采用硫酸氢氟酸次氯酸钠组合浸出体系浸取钒钛磁铁矿提钒尾渣中的钒,研究浸出过程中试剂浓度、浸出液固比、浸出温度、浸出时间、物料粒度对钒浸出率的影响。结果表明:钒的浸出率随试剂浓度、液固比、温度和时间的升高而增大;当矿物粒度小于0.20 mm时,钒浸出率有随矿物粒度变小而减小的趋势。在物料粒度0.15~0.25 mm、初始硫酸浓度150 g/L、初始氢氟酸浓度30 g/L、次氯酸钠加入量为矿量1.5%、矿浆液固比6:1、浸出温度90℃、浸出时间6 h、搅拌速度500 r/min的条件下,钒的浸出率可达85%以上。     

堆浸法从含银氧化铁矿中提取银

云南个旧某矿氧化矿石中,除Sn以外还伴生有丰富的Ag。为了解决Ag的回收利用问题,作者等人对该矿石进行了柱浸、槽浸、堆浸和回收Ag的试验,取得较好的效果。用堆浸法进行较大规模的生产,Ag回收率为34．96％,达到综合回收利用资源的目的。     

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。     

失效汽车催化剂中铂族金属的加压氰化浸出

研究了加压氰化法从失效汽车催化剂中选择性浸出铂族金属, 考察了加压氰化过程影响金属浸出率的各种因素. 实验结果表明 对200 g处理批量的小实验铂族金属氰化浸出率分别达到Pt98%, Pd99%, Rh96%; 千克级放大实验可完全重现小实验结果, 且该工艺也适应于其他类型失效催化剂的处理.     

超声空化作用下高温合金废料中战略金属的浸出机理（英文）

为提高高温合金废料中战略金属的回收效率，提出一种超声辅助浸出工艺。在超声辅助浸出过程中，浸出时间为60 min时，高温合金废料中Re、Ni、Co、Al、Cr和W的浸出率分别为92.3%、95.2%、98.5%、98.7%、97.5%和27.2%，明显优于常规浸出工艺中浸出时间为120 min时的浸出效率。高温合金废料中战略金属浸出效率的提高可归因于超声空化作用下产生的物理和化学协同作用。超声空化的物理作用可以阻止高温合金废料表面溶解层的形成，促进浸出液与高温合金废料接触，从而提高高温合金废料中战略金属的浸出率；而超声空化的化学作用可以促进HCl-H₂O₂体系中羟基自由基的产生，从而增强浸出体系的电位，促进体系中Re向ReO₄^-的氧化反应过程。     

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...     

Mechanism of influence of ferric ion on electrogenerative leaching of sulfide minerals with FeCl3

1 Introduction Electrogenerative leaching process is a newly- developed technique in hydrometallurgy. Although its principle has developed since the late 1960s[1], this technique has been overlooked in metallurgy until ZHANG et al[2] introduced it to the …     

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.     

含金硫酸渣硫酸盐化焙烧-氯化浸金研究

针对难处理含金硫酸渣进行了硫酸盐化焙烧-氯化(氯酸钠-氯盐)浸出试验研究,考察了氯酸钠用量、氯化钠用量、液固比、浸出时间、浸出温度、焙烧预处理等因素对金浸出率的影响。研究结果表明,在优化的试验条件:氯化钠用量为80 kg/t,氯酸钠100 kg/t,反应温度80℃,液固比为3,反应时间为4 h下,处理该含金硫酸渣可以得到91.44%的金浸出率。     

Technological mineralogy and environmental activity of zinc leaching residue from zinc hydrometallurgical process

Chemical, physical, structural and morphological properties of zinc leaching residue were examined by the combination of various detection means such as AAS, XRF, XRD, Mössbauer spectrometry, SEM-EDS, TG-DSC, XPS and FTIR. The toxicity characteristic leaching procedure (TCLP) was used to investigate the environmental activity of zinc leaching residue for a short contact time. The phase composition analysis indicated that the zinc leaching residue mainly consists of super refined flocculent particles including zinc ferrite, sulfate and silicate. The physical structural analysis showed that it has a thermal instability and strong water absorption properties. The results of TCLP indicated that the amounts of Zn and Cd in the leaching solution exceed 40 and 90 times of limit, respectively, which demonstrate that this residue is unstable in weak acidic environment for a short contact time.     

石煤钒矿硫酸活化常压浸出提钒工艺

研究石煤钒矿的硫酸活化提钒方法。分别考察矿石粒度、硫酸浓度、活化剂用量、催化剂用量、反应温度、反应时间和浸出液固比等因素对钒浸出率的影响。结果表明:石煤提钒的优化条件为矿石粒度小于74μm的占80%、硫酸浓度150 g/L、活化剂CaF2用量(相对于矿石)60 kg/t、催化剂R用量20 g/L、反应温度90℃、反应时间6 h、液固比(体积/质量,mL/g)2:1,在此优化条件下,钒浸出率可达94%以上;在优化条件下,采用两段逆流浸出,可有效减少活化剂CaF2以及浸出剂硫酸的消耗量;经过两段逆流浸出萃取反萃氧化水解工艺,全流程钒资源总回收率可达86.9%;V2O5产品纯度高于99.5%。     

机械活化强化钒渣钙化提钒工艺

针对转炉钒渣钙化焙烧酸浸工艺中存在的钒转浸率低的问题,采用高能球磨对钒渣进行活化预处理,以期强化其提钒效果。采用激光粒度分析仪、BET比表面积测定仪和XRD对活化前后钒渣进行了粒度、比表面积及物相结构分析;采用浸出实验研究了机械活化对钙化焙烧和浸出的影响规律。结果表明:机械活化法增大了钒渣的比表面积,增加了晶格畸变与微观应力,使含钒物相充分解离,由此可改善钒渣钙化焙烧的动力学条件。在浸出20 min条件下,机械活化80 min可将钒浸出率提高10%,最佳焙烧温度降低100℃。