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。     

某难选金矿加温化学预氧化浸出技术

以某实际含铜金矿为研究对象,在氯盐酸性加温体系下,分析浸出温度、时间、矿物粒度、NaCl浓度、H2SO4浓度、氧气流量等因素对化学预氧化浸出除铜和浸出渣氰化浸金的影响过程.结果表明:在90%矿样粒度小于37 μm、浸出温度95 ℃、初始H2SO4浓度0.75 mol/L、起始NaCl浓度0.7 mol/L、液固比5-1、浸出时间24 h、搅拌速度750 r/min的条件下,可使铜的浸出去除率达到80%以上,预氧化渣金的氰化浸出率达98.23%.     

二氧化硫还原分解铁酸锌及锌浸渣工艺

锌冶炼浸出渣中锌主要以铁酸锌的形式存在,针对锌浸渣中铁酸锌难于分解的问题,以铁酸锌作为研究对象,研究二氧化硫作用下铁酸锌中锌的溶出和Fe(Ⅲ)的还原行为。考察初始硫酸浓度、液固比、二氧化硫通入量、反应时间、反应温度对二氧化硫还原分解铁酸锌行为的影响。结果表明:最佳反应条件如下,初始硫酸浓度120 g/L、液固比11:1、二氧化硫通入量0.41×10~(-2)mol/g、反应时间120 min、反应温度105℃。在最佳反应条件下,对锌浸渣开展还原浸出实验,锌的浸出率能达到99%以上,Fe(Ⅲ)的还原率能达到98%。通过ICP-MS和XRD分析表明,锌浸渣中的铁酸锌完全分解,还原浸出渣的主要成分为锌和铅,分别以ZnS和PbSO_4的形式存在。     

锌冶炼中浸渣锌还原浸出机制与动力学

以锌冶炼中浸渣为研究对象,研究中浸渣的化学成分及锌的存在形态,锌主要以铁酸锌形式存在。采用SO2做还原剂,研究温度、初始硫酸浓度、二氧化硫分压对锌浸出效率的影响,并分析中浸渣中锌还原浸出反应机制及动力学。结果表明:H+在锌还原浸出过程中起关键作用,锌还原浸出反应活化能为31.67 k J/mol,为化学反应控制;SO2做还原剂时,反应时间、液固比及初始酸度均大幅降低。反应最佳工艺条件:初始硫酸浓度80 g/L、温度95℃、液固比(L/S)10 m L/g、二氧化硫分压200 k Pa、反应时间120 min。该工艺条件下,中浸渣中锌浸出率达99%以上。XRD和ICP分析表明:中浸渣中铁酸锌分解,硫化锌在该反应条件下未完全浸出,还原浸出渣中主要化学成分为铅和锌,主要物相为Pb SO4和Zn S。     

氰渣综合利用提取金银的试验研究

通过正交试验详细考察了矿浆浓度、硫酸过剩系数、反应温度和反应时间等因素对氰渣浸铁率的影响.结果表明矿浆浓度为35%、硫酸过剩系数为1.3、反应温度为100℃、反应时间为2.5 h的试验条件下,铁的浸出率最高,可达97.80%.对比氰渣和浸铁渣金、银的氰化浸出效果发现氰渣再氰化金、银的浸出率分别为5%和10%,而浸铁渣再氰化金、银的浸出率则分别高达87%和80%,因此氰渣浸铁再氰化是提高金、银回收率的有效途径之一.     

亚硫酸钠浸出-甲醛还原回收氯化渣中的银

为了改善氰化金泥湿法精炼过程中氯化渣铁粉置换工艺熔炼粗银时的工作环境,提高银的回收率,采用亚硫酸钠浸出-甲醛还原方法处理氯化渣。结果表明,在pH=8.5、液固质量比为20:1、35℃的条件下用浓度为250g/L的亚硫酸钠溶液浸出氯化渣3h后,银浸出率大于99%;浸银液在40%甲醛与银比例为5:2 (mL/g)、50℃的条件下还原1.5 h,银还原率达99%以上。浸出渣返回金泥氯化分金流程,浸银液还原后可再生循环使用,银综合回收率可达98%以上。     

湿法炼锌铜砷渣分离提铜及沉砷工艺研究

采用铁粉置换法处理湿法炼锌产生的锌浸渣还原浸出液,产出一种含砷铜渣,以该含砷铜渣为研究对象,利用氧压酸浸缓慢分解含砷铜渣,使其中的铜、锌等溶解进入溶液,同时,砷、铁以臭葱石的形式沉淀为浸出渣,从而将铜的浸出和砷、铁的沉淀在同一反应釜同一过程中完成,有效实现含砷铜渣中有价金属的浸出过程与杂质的沉淀过程在同一过程同步进行。结果表明：在反应温度为135℃、反应时间为4 h、液固体积质量比25 mL/g、硫酸浓度为50 g/L、氧分压500 kPa、铁砷摩尔比为1的条件下,浸出渣中铜含量仅为2.03%,浸出率达到97.72%,砷含量达到26.06%,沉淀率达到95.98%;浸出液中铜的浓度达到20.47 g/L,砷浓度小于0.63 g/L,实现了铜和砷的高效分离,提高了铜金属回收率和资源综合利用率。浸出渣中砷均以臭葱石(FeAsO₄·2H₂O)的形式存在,符合当前的环境友好型发展理念。     

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。     

氰化尾渣还原焙烧酸浸提铁及氰化浸金新工艺

以氰化尾渣为原料,采用还原焙烧酸浸工艺对其进行处理。当还原温度为850℃、加入煤粉质量为氰化尾渣质量的13%、还原时间为100 min时,对氰化尾渣进行还原,氰化尾渣中Fe2O3转化为Fe3O4或FeO。还原后采用硫酸浸出,当硫酸浓度为50%、硫酸用量系数为1.2、反应温度为105℃、反应时间为3 h时,铁的浸出率达到93.66%。还原焙烧渣在600℃氧化焙烧2 h经过脱碳后氰化浸金,当氰化钠用量为4 kg/t、反应时间为28 h、液固比为2:1时,金的浸出率达到92.4%。经过还原焙烧、硫酸浸出、氧化焙烧及氰化浸金,氰化尾渣渣量减少了38.8%。     

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%。     

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.     

Kinetics of zinc sulfide concentrate direct leaching in pilot plant scale and development of semi-empirical model

The direct leaching kinetics of an iron-poor zinc sulfide concentrate in the tubular reactor was examined. All tests were carried out in the pilot plant. To allow the execution of hydrostatic pressure condition, the slurry with ferrous sulfate and sulfuric acid solution was filled into a vertical tube (9 m in height) and air was blown from the bottom of the reactor. The effects of initial acid concentration, temperature, particle size, initial zinc sulfate concentration, pulp density and the concentration of Fe on the leaching kinetics were investigated. Results of the kinetic analysis indicate that direct leaching of zinc sulfide concentrate follows shrinking core model (SCM). This process was controlled by a chemical reaction with the apparent activation energy of 49.7 kJ/mol. Furthermore, a semi-empirical equation is obtained, showing that the order of the iron, sulfuric acid and zinc sulfate concentrations and particle radius are 0.982, 0.189, ?0.097 and ?0.992, respectively. Analysis of the unreacted and reacted sulfide particles by SEM–EDS shows that insensitive agitation in the reactor causes detachment of the sulfur layer from the particles surface in lower than 60% Zn conversion and lixiviant in the face with sphalerite particles.     

硫化锌银精矿富集银的工艺研究

基于国内外硫化锌矿处理的火、湿法研究进展,对含锌银精矿采用硫酸化焙烧、稀硫酸浸出工艺脱除锌、富集银,考察了焙烧和浸出过程中的主要影响因素。结果表明,硫酸配比为150%,在300℃焙烧90 min,以5%稀硫酸为浸出剂,液固比8:1,搅拌转速200~300 r/min,85℃浸出120 min,最终锌的浸出率可达到98%以上,浸出渣中银含量为7.24%,银被富集7倍。     

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.     

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。     

铜阳极泥处理过程中中和渣中碲的提取与制备

采用硫酸浸出二氧化硫还原方法从中和渣中制取单质碲。研究表明:采用硫酸浸出中和渣,当反应温度为30℃、反应时间为0.5 h、硫酸浓度为53.9 g/L、硫酸用量为理论用量的1.5倍时,碲浸出率为99.99%;采用亚硫酸钠还原酸浸液中碲时,碲(Ⅳ)发生水解生成二氧化碲;采用二氧化硫还原酸浸液中碲时,当反应温度为75℃、反应时间为2 h、盐酸浓度为3.2 mol/L、二氧化硫流量为0.4 L/min时,碲回收率达到99.84%。X射线衍射(XRD)分析表明二氧化硫还原得到的产物为单质碲,电感耦合等离子体发射光谱(ICP)分析表明,碲粉中碲含量为98.27%。扫描电子显微系统(SEM)分析表明,碲粉的形态为针形。     

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．     

Kinetics of indium dissolution from marmatite with high indium content in pressure acid leaching

The kinetics of oxygen pressure acid leaching marmatite with high indium content was studied. The effects of particle size, agitation rate, temperature, H_2SO_4 concentration, and oxygen partial pressure on leaching rate of indium were investigated. The results show that when the agitation rate is above 600 r-min~(-1), its influence on indium leaching rate is insignificant. It is determined that the leaching rates increase with the increase in sulfuric acid concentration, temperature, partial oxygen pressure, and the decrease in particle size. Moreover, the results demonstrate that the process of indium leaching is controlled by interface chemical reaction, with apparent activation energy of 65.7 k J-mol~(-1). The apparent reaction orders of sulfuric acid and oxygen partial pressure are determined to be 0.749 and 1.260, respectively. The leaching reaction process follows shrinking unreacted core model. And finally, the kinetics model equation is established for indium.