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。     

Treatment of cadmium dust with two-stage leaching process

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湿法炼锌废渣中硫脲浸出银的动力学

探讨了从湿法炼锌废渣中用硫脲浸出回收银的浸出反应动力学,从这种难浸的含银炼锌废渣中用硫脲浸出回收银,浸出反应是一种典型的氧化还原反应并可充分进行,同时,通过动力学推导得出,从湿法炼锌废渣中用硫脲浸出银的反应动力学模型为收缩核动力学模型,同时计算出浸出活化能为13．26kJ/mol。该模型反映了浸出过程中控制整个反应速率的决定步骤是固膜扩散速率,并较好地说明了浸出机理。     

铜渣综合利用中铟的回收

介绍了从铜渣中回收铟的试验情况，试验表明：该过程铟直收率可达619％，所选工艺合理、简单，易于实现工业化。     

氧化锌贫矿提锌渣中铅和银的氯盐一步浸出

运用X射线衍射、扫描电镜和X射线能谱等分析手段,对山东某地深度氧化锌贫矿提锌后渣进行工艺矿物学特征分析可知,矿物中金属赋存状态复杂,属难选矿物。开发出氯盐一步法浸出铅和银的新工艺,考察反应温度、NaCl浓度、添加剂用量、液固比、HCl加入量和浸出时间对浸出过程的影响。结果表明:加入添加剂对铅的浸出率没有影响,但可以显著提高银的浸出率。条件试验研究得出最佳工艺条件如下:浸出温度90℃、NaCl浓度390 g/L、添加剂用量15 mL、液固比(质量比)7?1、HCl加入量3 mL、浸出时间3 h。在此最佳工艺条件下,铅的浸出率达到95%左右,银的浸出率达到90%左右。     

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．     

锌浸出渣中银的回收方法综述

湿法炼锌是当今世界炼锌主要方法,生产中所得锌浸渣中含有较多的锌及其他有价金属(如银、铅、铜、镉、铟、锗等),作为尾矿堆存不仅会占用大量的土地资源,而且还会造成资源的严重浪费,加之锌浸出渣的长期堆存环境危害性高,因此对锌浸出渣中的有价金属进行综合回收具有十分重要的意义。本文结合国内外锌浸出渣的研究状况,重点针对锌浸出渣中银的回收现状进行论述,全面阐述锌浸出渣中银的工艺矿物学性质、提取工艺流程及发展趋势,对该类资源的综合回收利用具有十分重要的借鉴作用。     

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.     

神经网络优化ITO废靶酸浸回收铟(英文)

以提高铟浸出率为目标,通过工艺实验结合神经网络研究ITO废靶酸浸的优化工艺。首先,固定其他工艺因素,进行单因素如残酸度、氧化剂加入量、酸浸温度及时间对铟浸出率影响的实验研究。结果表明,增大残酸度可提高铟浸出率;氧化剂的加入可明显提高铟浸出率,但增加到一定程度后浸出率提高不明显;升高温度可明显提高浸出率,但继续升高则会降低铟浸出率;延长浸出时间也可提高铟浸出率。通过反向传播算法的人工神经网络(BPNN)研究多因素的综合作用对铟浸出率的影响规律,预测值与实验值相差很小,表明所建立的BP模型铟浸出率能比较准确地预测。最终,通过BPNN预测以及实验验证,获得高达99.5%浸出率的工艺参数:残酸度50~60 g/L、氧化剂含量10%、浸出温度70°C和浸出时间2 h。     

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.     

Development of hydrometallurgical process for indium recovery from waste liquid crystal display using Cyphos IL 101

The main goal of this study was to recover indium from the waste liquid crystal display (LCD) panel. In this context, an ionic liquid Cyphos IL 101 was explored. The extraction parameters such as equilibration period, acid concentration, chloride ion concentration, hydrogen ion concentration were examined on the extraction efficiency of Cyphos IL 101 towards In(III). Quantitative extraction of indium was found at 2.0 mol/L HCl using 0.005 mol/L Cyphos IL 101 and quantitative stripping with 1.0 mol/L H₂SO₄. Job's method was used to determine the extracted species and R₃R‘PInCl₄ (R=C₆H₁₃; R′=C₁₄H₂₉) was proposed. Based on the observations on multi-metal studies, Cyphos IL 101 was further employed for the removal of indium, tin and copper from the leach liquors of waste LCDs. Optimized conditions were generated for the recovery of indium from waste LCDs. McCabe−Thiele diagram analysis, counter-current extraction and selective stripping were carried out to separate the metal ions, i.e., indium, tin and copper. Two stages at O/A ratio of 1:3 were required for complete removal of tin from the feed and selective stripping of In and Sn was achieved using 0.1 mol/L H₂SO₄. A scheme for separating indium from the waste LCDs was proposed.     

Thermodynamics of Solvent Extraction of Indium with EHEHPA

Molalities of In³⁺ were measured at ionic strength from 0.1 to 2.0 mol·kg⁻¹ in aqueous phase containing Na₂SO₄ as supporting electrolyte and at constant molality extractant at temperatures from 278.15 to 303.15 K in organic phase. The standard extraction constants K ⁰ at various temperatures were obtained by methods of extrapolation and polynomial approximation. Thermodynamic quantities for the extraction process were calculated.     

锌浸渣还原焙烧-磁选回收铁

在查明锌浸渣工艺矿物学的基础上,采用还原焙烧将铁酸锌分解为氧化锌和磁性氧化铁,再通过磁选的方法回收铁,达到锌、铁分离的目的。实验考查了焙烧温度、焙烧时间、还原剂用量对铁酸锌分解率、铁回收率和铁品位的影响。结果表明:在焙烧温度为950℃、焙烧时间为1 h及还原剂添加量为10%和5%的条件下,铁酸锌分解率达到72.05%,铁回收率可达到91.79%,精矿中铁的品位为50%左右。焙烧及磁选过程中颗粒的团聚包裹是铁精矿品位不高的主要原因。     

Kinetics of simultaneous leaching of Ag and Pb from hydrometallurgical zinc residues by chloride

The leaching kinetics of silver and lead simultaneously from zinc residue by chloride was investigated.The effects of stirring speed,temperature,sodium chloride concentration,particle size and liquid/solid ratio on Ag and Pb dissolution in sodium chloride were studied.It was determined that the dissolution rates increased with increasing sodium chloride concentration,temperature and decreasing particle size.The dissolution kinetics followed a shrinking core model,with inter-diffusion through gangue layer as the rate determining step.This finding is in accordance with the apparent activation energy(Ea)of 26.8 kJ.mol-1(Ag)and 26.5 kJ.mo1-1(Pb),and a linear relationship between the rate constant and the reciprocal of squared particle size.The orders of reaction with respect to sodium chloride concentration,temperature and particle size were also achieved.The rate of reaction based on diffusion-controlled process can be expressed by semi-empirical equations.     

铟纳米颗粒及纳米线的制备

采用化学方法使有机金属前驱体[In（η^5-C5H5）]在不同条件下分解制备金属铟纳米颗粒和纳米线。前驱体在甲苯或四氢呋喃溶剂中分解时，配位体的种类与相对量、溶剂中的水含量和紫外线照射对分解产物的成分、形貌和尺寸有很大的影响。利用十六烷基胺（HDA）或氧化三辛基膦（TOPO）配位体的作用，前驱体均可分解得到铟纳米颗粒；而在HDA相对含量高、溶剂中不含水和在紫外线照射作用下，前驱体分解反应产物为铟纳米线。     

Recovering indium with sulfating roasting from copper-smelting ash

A technology for recovering indium from Jinchuan copper-smelting ash was developed. Indium in the ash was first enriched to the leaching-slag in leaching process,and then recovered by sulfating roasting. The method included mixing the leaching-slag with sulfuric acid,making them into particles,roasting the mixture,and then leaching the calcine with hot water. Above 90% of indium in calcine could be dissolved into the leaching solution. The optimized conditions were determined as follows： the mass ratio of sulfuric acid to leaching slag was 0.1,the roasting time was about 1 to 1.5 h in the temperature range of 200-250℃,and the calcine was leached for 1 h with 5：1 of liquid/solid ratio at 60℃. Over 99% of indium in leaching solution was finally enriched by Zn substitution or sulfide precipitation.     

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。     

Thermodynamics of Solvent Extraction of Indium with P507

At high acidity and high ionic concentration, molalities of In³⁺ were measured in aqueous phase containing Na₂SO₄ as supporting electrolyte at ionic strength from 0.1 to 2.0 mol/kg and at constant molality extractant at temperatures from 278.15 to 303.15 K in organic phase. The standard extraction constants K ⁰ at various temperatures were obtained by methods of extrapolation and polynomial approximation. Thermodynamic quantities for the extraction process were calculated.     

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.     

Increasing silver leaching rate from leaching-resistant zinc residues by thiourea leaching method with pressurized preoxidation process

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