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.     

焙烧氟碳铈矿硫酸浸出稀土的动力学(英文)

研究了硫酸浸出德昌稀土与天青石共伴生矿的焙烧矿过程。考查粒度、搅拌速度、硫酸浓度和温度对稀土浸出率的影响,并对稀土的浸出动力学进行分析。在选定的浸出条件下:粒径0.074～0.100mm、硫酸浓度1.5mol/L、液固比8:1、搅拌速度500r/min,稀土浸出反应受内扩散控制,表观活化能为9.977kJ/mol。     

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。     

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。     

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%。     

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...     

Leaching of vanadium from stone coal with sulfuric acid

The effects of roasting, mass ratio of H2SO4 to stone coal, leaching temperature, liquid-to-solid ratio, grinding fineness of stone coal, and two-stage counter-current leaching on the vanadium leaching ratio were studied. The results show that the vanadium leaching ratio of roasted stone coal through two-stage counter-current leaching can reach 65.1% at the mass ratio of H2SO4 to stone coal of 20%, leaching temperature for the production of vanadium from stone coal.     

高硅氧化锌矿加压酸浸处理

通过对广西某地高硅氧化锌矿矿物分析可知,矿物中的锌主要以硅锌矿和异极矿的形式存在.采用加压酸浸技术对该高硅氧化锌矿进行处理,条件试验研究得出最佳工艺条件为:矿物粒度0.104 mm,硫酸浓度120 g/L,釜内压力1.0 MPa,浸出时间90 min,反应温度120 ℃,液固比3-1.综合性试验研究得出:在最佳工艺条件下,锌浸出率可达97%以上,SiO2截留率大于99.2%,浸出矿浆过滤速率大于880 L/(m2-h),浸出矿浆具有良好的过滤性能.     

硫酸浸取法提取粉煤灰中氧化铝

以粉煤灰和硫酸为原料,经活化、磁选、硫酸浸出、硫酸铝溶出、结晶制备出Al2(SO4)3.18H2O。硫酸铝经煅烧、γ-Al2O3碱溶和晶种分解、氢氧化铝煅烧等过程制备出冶金级氧化铝。本文给出了完整的工艺流程,对此工艺进行了详细阐述。在最佳条件下,粉煤灰中氧化铝提取率可达92.3%。     

以废茶叶为还原剂在硫酸溶液中还原浸出氧化锰矿（英文）

采用废茶叶在硫酸溶液中还原浸出加蓬和湘西氧化锰矿石,探索废茶叶用量、硫酸浓度、固液比、浸出温度和反应时间对浸出过程的影响。对加蓬氧化锰矿,优化的浸出条件为:氧化锰矿与废茶叶的质量比10:4、硫酸浓度2.5 mol/L、固液比7.5:1、浸出温度368 K、浸出时间8 h;在此条件下,加蓬氧化锰矿的浸出率几乎达100%。对于湘西氧化锰矿,优化浸出条件为:氧化锰矿与废茶叶的质量比10:1、硫酸浓度1.7 mol/L、液固比7.5:1、温度368 K、浸出时间8 h;在此条件下,锰的浸出率达到99.8%。氧化锰矿的还原浸出过程符合内扩散控制模型,加蓬和湘西氧化锰矿石的还原浸出反应表观活化能分别为38.2 kJ/mol和20.4 kJ/mol。采用X-射线衍射(XRD)和扫描电子显微镜(SEM)对浸出前、后的锰渣进行表征。     

浸出高硅锌焙砂中硅胶聚沉的规律

考察了pH值,静置时间,CaO快速中和以及添加Fe^3 离子对浸出高硅锌焙砂时硅胶聚沉的影响,发现pH=2.0-4.8是硅胶的稳定区,此时浸出液过滤十分困难,加速硅胶聚沉以提高过滤速率的有效方法是：用CaO或锌焙砂快速中和越过pH=2.0-4.8范围,以达到pH=5.0-5.5;延长静置时间,在浸出锌焙砂时加入带有相反电荷的胶体,均有利于硅胶的聚沉,实验条件下可使Zn浸出液和硅胶聚沉物分离的过滤速率提高36－60倍,同时使锌的回收率达到88％－90％。     

Transformation and leaching kinetics of silicon from low-grade nickel laterite ore by pre-roasting and alkaline leaching process

The mineralogical phase transformation of a low-grade nickel laterite ore during pre-roasting process and the extraction of silicon during alkaline leaching process were investigated. The results indicate that the reaction activity of nickel ores is effectively improved by pre-roasting at 650 °C for 2 h, because of the transformation of lizardite into magnesium olivine and protoenstatite. When finely ground ore samples (44–61 μm) pre-roasted firstly react with sodium hydroxide solution (60 g/L) with a solid/liquid ratio of 1:5 at 140 °C for 120 min, the extraction of silicon can reach 89.89%, and the other valuable elements of magnesium, iron and nickel are accumulated in the solid residues. The leaching kinetics of nickel laterite ore can be described successfully by the diffusion through the product layer control model. The activation energy is calculated to be 11.63 kJ/mol and the kinetics equation can be expressed as 1–3(1–x)^2/3+2(1–x)=13.53×10^?2exp[–11.63/(RT)]t.     

Solvent extraction of vanadium from sulfuric acid solution

The behaviour of vanadium(V) extracted from sulfuric acid solution was investigated using Cyanex 923 as an cxtractant. The effects of the concentration of Cyanex 923 and the pH of the solution were studied. The extraction of vanadium(V) increases with the increase of Cyanex 923 concentration and shaking time. Cyanex 923 can extract vanadium(V) fi'om sulfuric acid solution at low pH conditions, and the best pH conditions for exuaction of vanadium(V) are at pH 1.0-2.0. The species extracted into the organic phase is VO2HSO4 with one molecule of Cyanex 923. Equilibrium studies were used to assess the extraction efficiency of vanadium(V) recovery from the sulfuric acid solution.     

Effect of mechanical activation on alkali leaching of chromite ore

Mechanical activation was used to improve the extraction of chromium in molten NaOH. It is observed that the extraction ratio reaches 97% after leaching for 200 min when chromite ore is mechanically activated for 10 min, but only 34% if not activated. Mechanical activation can decrease the particle size, increase the surface area, and enhance the lattice distortion. Further, the mechanisms for mechanical activation were exposed. The results show that the mechanical activation mainly focuses on chromite ore particle size decrease and the lattice distortion. The formation of aggregation weakens the strengthening effect of mechanical activation for releasing high surface energy.     

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。     

铁酸锌选择性还原及其在锌浸出渣处理中的应用

传统湿法炼锌过程产生大量富含有价金属资源的铁酸锌废渣,铁的分离是实现铁酸锌废渣中有价金属资源回收的关键。提出含大量铁酸锌的锌浸出渣选择性还原焙烧?浸出分离铁和锌的新方法。通过热力学分析确定铁酸锌分解过程中Fe3O4和ZnO产物的优势区域,并发现V(CO)/V(CO+CO2)比是控制铁酸锌还原焙烧产物物相的关键因素,在V(CO)/V(CO+CO2)比在2.68%?36.18%范围内,铁酸锌优先分解生成在Fe3O4和ZnO。通过TG分析,确定铁酸锌还原焙烧的最佳条件为焙烧温度700?750°C,CO体积分数6%,V(CO)/V(CO+CO2)30%。基于上述研究结果,对富含铁酸锌的锌浸渣进行还原焙烧处理,焙烧产物经酸浸后,锌的浸出率达70%,铁的浸出率仅为18.4%,实现锌浸渣中锌和铁的有效分离。     

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.     

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.