Leaching of lead from zinc leach residue in acidic calcium chloride aqueous solution

A process with potentially reduced environmental impacts and occupational hazards of lead-bearing zinc plant residue was studied to achieve a higher recovery of lead via a cost-effective and environmentally friendly process. This paper describes an optimization study on the leaching of lead from zinc leach residue using acidic calcium chloride aqueous solution. Six main process conditions, i.e., the solution pH value, stirring rate, concentration of CaCl₂ aqueous solution, liquid-to-solid (L/S) ratio, leaching temperature, and leaching time, were investigated. The microstructure and components of the residue and tailing were characterized using scanning electron microscopy (SEM) and X-ray diffraction (XRD). On the basis of experimental results, the optimum reaction conditions were determined to be a solution pH value of 1, a stirring rate of 500 r·min-1, a CaCl₂ aqueous solution concentration of 400 g·L-1, a liquid-to-solid mass ratio of 7:1, a leaching temperature of 80℃, and a leaching time of 45 min. The leaching rate of lead under these conditions reached 93.79%, with an iron dissolution rate of 19.28%. Silica did not take part in the chemical reaction during the leaching process and was accumulated in the residue.     

硫氧混合铅锌矿中锌的氧化氨浸动力学

常压低温条件下在NH3-(NH4)2 SO4体系中使用过硫酸铵作为氧化剂对硫氧混合铅锌矿中的锌进行浸出实验,系统研究了搅拌速度、浸出剂浓度、氧化剂浓度与温度对于锌浸出率的影响.结果表明,在最优条件下锌的浸出率可达93.2%,且浸出过程中几乎没有其他离子进入溶液,实现了锌的选择性高效浸出,从而简化了后续的浸出液净化与材料制备过程.动力学研究表明,硫氧混合铅锌矿中锌的氧化氨浸过程遵循固体产物层扩散控制的未反应核收缩模型,浸出反应的表观活化能为17.89 kJ·mol-1.     

从磷酸废砷渣中回收砷的碱浸工艺研究

根据目前广西区大量砷渣得不到有效利用的现状,以磷酸净化过程中产生的含砷废渣为原料,通过物相分析确定了碱浸出法回收砷的工艺,考察了浸出温度、摩尔比、液固比和反应时间等工艺条件对砷浸出率的影响。结果表明,n(NaOH)∶n(As2S3)是影响砷浸出率的主要因素,较适宜的碱浸工艺条件为:浸出温度为70℃,n(NaOH)∶n(As2S3)=6.0∶1,液固比=6.0∶1,反应时间为30 min,在此条件下砷浸出率可达97.1%。在现有基础上,该工艺为磷酸废砷渣的综合利用提供了一条简单高效的技术路线。     

Sulfuric acid leaching of high iron-bearing zinc calcine

Sulfuric acid leaching of high iron-bearing zinc calcine was investigated to assess the effects of sulfuric acid concentration, liquid-to-solid ratio, leaching time, leaching temperature, and the stirring speed on the leaching rates of zinc and iron. The results showed that the sulfuric acid concentration, liquid-to-solid ratio, leaching time, and leaching temperature strongly influenced the leaching of zinc and iron, whereas stirring speed had little influence. Zinc was mainly leached and the leaching rate of iron was low when the sulfuric acid concentration was less than 100 g/L. At sulfuric acid concentrations higher than 100 g/L, the leaching rate of iron increased quickly with increasing sulfuric acid concentration. This behavior is attributed to iron-bearing minerals such as zinc ferrite in zinc calcine dissolving at high temperatures and high sulfuric acid concentrations but not at low temperatures and low sulfuric acid concentrations.     

电解锰渣固结体中重金属浸出毒性及其在模拟酸雨下的淋溶特性分析

以一种硅铝基胶凝材料为固结剂,对某电解锰厂的锰渣进行固化处理,分析在不同养护周期下锰渣固结体中重金属的浸出毒性,通过动态淋溶试验,研究养护周期和模拟酸雨pH值对固结体淋溶特性的影响,并通过SEM和XRD手段,分析酸雨淋溶前后固结体的形貌和成分,探讨固结体淋溶过程中重金属的溶出机理。结果表明,固化处理可有效固定锰渣中的重金属,固结体的养护周期对其在酸雨淋溶过程中Mn的溶出影响较大,其中养护28天的锰渣固结体在不同pH值酸雨淋溶条件下,淋溶液中Mn和Zn的最高浓度分别为0.189、0.116mg·L~(-1),其它重金属含量也较低,均低于国家标准浓度限值;酸雨淋溶过程中,固结体的晶体结构没有受到破坏,大部分的重金属被包裹在晶体内部或以更稳定的化合态存在,不易被溶出。     

Ultrasound effects on zinc recovery from EAF dust by sulfuric acid leaching

In this work, an ultrasound-assisted leaching process was studied for the recovery of zinc from electric arc furnace (EAF) dust, in which zinc was mainly present in the form of franklinite (60%). Hydrometallurgy is emerging as a preferred process for the recovery of a variety of metals, and the use of ultrasound could offer advantages over the conventional leaching process, especially for the dissolution of franklinite. Franklinite is a refractory phase that is difficult to leach and represents the main obstacle in conventional hydrometallurgy processing. Atmospheric leaching with different sulfuric acid concentrations (0.2-2.0 M) at two temperatures (323 and 353 K) was performed. The tests were conducted using both conventional and ultrasound-assisted leaching. After the leaching tests, the solid residues were characterized by scanning electron microscopy (SEM) and X-ray diffraction (XRD) techniques, whereas the leach liquor was analyzed by inductively coupled plasma spectroscopy (ICP). The use of ultrasound facilitated the dissolution of franklinite at low acid concentrations and resulted in a greater zinc recovery under all of the investigated operating conditions.     

铁矾渣微波硫酸化焙烧水浸液的深度除铁

采用微波低温硫酸化焙烧-水浸和针铁矿除铁方法将Zn、Cu等富集到浸出液中,Pb和Ag富集到浸出渣中,使有价金属得到清洁的回收利用.研究了上述工艺中浸出液除铁的优化工艺条件,探究了反应体系的pH值、浸出液单次滴加量、浸出液的铁含量等因素对除铁效果的影响,并采用X射线衍射分析、扫描电子显微镜观察等手段对得到的沉淀渣进行了表征.研究获得的优化实验条件为：以200 mL的0.01 mol·L-1 ZnSO4溶液为底液,晶种添加量为20 g·L-1,除铁体系pH值控制在3左右,温度90℃,每隔5 min滴加3 mL水浸液(保持反应体系中铁的浓度<1 g·L-1).在此条件下,除铁后溶液残铁量仅为0.065 g·L-1,去除率可达99.3%,达到了深度除铁效果.除铁过程中,Zn的损失率仅为4.1%.     

锌浸出渣的还原焙烧和人工硫化矿物浮选过程中硫酸铅和硫酸锌的反应行为

To evaluate the feasibility of recovering Pb and Zn sulfides and Ag-containing minerals from Zn leaching residue by the process of reduction roasting followed by flotation, the reaction behaviors of Pb and Zn sulfates during this process were investigated. Chemical analysis showed that the transformation ratios of PbSO₄ and ZnSO₄ could reach 65.51% and 52.12%, respectively, after reduction roasting, and the introduction of a sulfidation agent could improve the transformation ratios of these sulfates. scanning electron microscopy–energy dispersive spectroscopy (SEM–EDS) revealed that temperature obviously affects the particle size, crystal growth, and morphology of the artificial Pb and Zn sulfide minerals. Particle size analysis demonstrated that the particle size of the materials increases after roasting. Flotation tests revealed that a flotation concentrate composed of 12.01wt% Pb, 27.78wt% Zn, and 6.975 × 10?2wt% Ag with recoveries of 60.54%, 29.24%, and 57.64%, respectively, could be obtained after roasting.     

氧化锌矿物浸出渣中铅、锶的提取及动力学分析

NH₄HCO₃ conversion followed by HCl leaching was performed and proven to be effective in extracting Pb and Sr from zinc extracted residual. The mechanism and operating conditions of NH₄HCO₃ conversion, including molar ratio of NH₄HCO₃ to zinc extracted residual, NH₄HCO₃ concentration, conversion temperature, conversion time, and stirring velocity, were discussed, and operating conditions were optimized by the orthogonal test. Experimental results indicate that NH₄HCO₃ conversion at temperatures ranging from 25 to 85°C follows the shrinking unreacted core model and is controlled by inner diffusion through the product layer. The extraction ratios of Pb and Sr under optimized conditions reached 85.15% and 87.08%, respectively. Moreover, the apparent activation energies of Pb and Sr were 13.85 and 13.67 kJ·mol?1, respectively.     

铬渣柱浸生物解毒的研究

针对长沙铬盐厂铬渣污染的现状及铬渣的性质,进行了细菌柱浸解毒铬渣的实验。研究结果表明,实验前需将铬渣制成颗粒,解毒工艺的最佳粒径为4~8mm;最佳pH为10.0;最佳温度为30℃。在最佳工艺条件下,铬渣浸出液中Cr（Ⅵ）的浓度由250.6mg/L降为0mg/L,达到国家规定的废水排放标准;解毒后渣的浸出毒性为0.4mg/L,低于国家固体废物浸出毒性标准所规定的1.5mg/L。     

Recovery of valuable metals from waste diamond cutters through ammonia-ammonium sulfate leaching

Copper and zinc were recovered from waste diamond cutters through leaching with an ammonia-ammonium sulfate system and air as an oxidant. The effects of experimental parameters on the leaching process were investigated, and the potential-pH (E-pH) diagrams of Cu-NH₃-SO₄2--H₂O and Zn-NH₃-SO₄2--H₂O at 25℃ were drawn. Results showed that the optimal parameters for the leaching reaction are as follows:reaction temperature, 45℃; leaching duration, 3 h; liquid-to-solid ratio, 50:1 (mL/g); stirring speed, 200 r/min; ammonia concentration, 4.0 mol/L; ammonium sulfate concentration, 1.0 mol/L; and air flow rate, 0.2 L/min. The results of the kinetics study indicated that the leaching is controlled by the surface chemical reaction at temperatures below 35℃, and the leaching is controlled by diffusion through the product layer at temperatures above 35℃.     

高砷烟尘酸性氧化浸出砷和锌的试验研究

采用酸性氧化浸出工艺对某冶炼厂高砷烟尘进行湿法浸出砷、锌的试验研究。通过单因素试验确定最佳浸出工艺条件。结果表明,采用pH值为2的稀硫酸溶液,在浸出温度80℃、浸出时间105min、液固体积质量比10∶1、H2O2添加量1.75mL/g(烟灰)、搅拌速度705r/min的条件下,砷、锌浸出率分别达到78.25%和85.42%。     

湿法炼锌净化钴渣新处理工艺

提出了氨-硫酸铵体系处理湿法炼锌净化钴渣的新工艺.该工艺首先用铵-氨水溶液浸出烘烤后的钴渣,再用锌粉对浸出液净化除杂并进行锌与镉、钴、铜及铜与钴的分离.在最佳技术条件下,金属浸出率(质量分数)分别为Zn91.18%,Cu96.98%,Cd99.38%,Co89.35%;净化所得的富钴渣含Co3.79%,富集比达8.4,从这种钴渣中可直接提取钴或钴盐;而净化液可直接制取活性锌粉.氨-硫酸铵法具有原料适应性强,设备防腐要求低,能常温操作,能耗低,除杂容易等优点,是处理湿法炼锌废渣、综合利用有价元素的较好方法.     

氨氮和有机杂质对锌电积过程的影响

以硫酸锌-硫酸溶液模拟兰坪氧化锌矿反萃液,考察氨氮及有机杂质如260溶剂油和β-二酮对锌电积过程电流效率和电锌质量的影响。实验结果表明:氨氮对锌电积过程中电流效率和电锌质量基本无影响,而有机杂质260溶剂油和β-二酮影响较大。     

PTA氧化残渣中苯甲酸和Co2+、Mn2+的分离回收

为了控制精对苯二甲酸（PTA）产品质量,其生产过程排出大量氧化残渣,其主要成分为苯羧酸及微量的钴锰离子催化剂,不仅污染环境,而且造成资源浪费。采用混合溶剂双相浸取法处理PTA氧化残渣,从而实现残渣中苯甲酸和Co2+、Mn2+离子的有效分离回收。研究结果表明：在浸取温度为30℃,液料质量比乙酸异丁酯：正庚烷：0.2mol/L硫酸水溶液：干残渣=3:3:8:1,浸取时间为30min的条件下,有机相中苯甲酸浸取率达83%以上,纯度高于95%,水相中Co2+、Mn2+的浸取率分别为98.82%和99.31%。本文并提出一个从氧化残渣中回收苯甲酸和Co2+、Mn2+离子的工艺概念流程,实现氧化残渣资源化回收利用。     

Alkaline pressure oxidative leaching of bismuth-rich and arsenic-rich lead anode slime

A new alkaline pressure oxidative leaching process(with NaNO3 as the oxidant and NaOH as the alkaline reagent)is proposed herein to remove arsenic,antimony,and lead from bismuth-rich and arsenic-rich lead anode slime for bismuth,gold,and silver enrichment.The effects of the temperature,liquid-to-solid ratio,leaching time,and reagent concentration on the leaching ratios of arsenic,antimony,and lead were investigated to identify the optimum leaching conditions.The experimental results under optimized conditions indicate that the average leaching ratios of arsenic,antimony and lead are 95.36%,79.98%,63.08%,respectively.X-ray diffraction analysis indicated that the leaching residue is composed of Bi,Bi2O3,Pb2Sb2O7,and trace amounts of NaSb(OH)6.Arsenic,antimony,and lead are thus separated from lead anode slime as Na3AsO4·10H2O and Pb2Sb2O7.Scanning electron microscopy and energy-dispersive spectrometry imaging revealed that the samples undergo appreciable changes in their surface morphology during leaching and that the majority of arsenic,lead,and antimony is removed.X-ray photoelectron spectroscopy was used to demonstrate the variation in the valence states of the arsenic,lead,and antimony.The Pb(IV)and Sb(V)content was found to increase substantially with the addition of NaNO3.     

熔融还原法镍渣炼铁的热力学与动力学

利用熔融还原法进行了闪速炉水淬镍渣提铁的实验研究,探讨了熔渣二元碱度、反应温度和反应时间对提铁效果的影响.XRD测试结果表明水淬镍渣由正硅酸铁FeO.SiO2和玻璃态物质组成.镍渣中的氧化铁主要以FeO.SiO2的形式存在,通过常规的选矿方法很难实现铁氧化物的富集,故采用熔融还原方法进行镍渣提铁实验.实验结果表明增加配合料中CaO的加入量、提高反应温度以及延长熔制时间都能不同程度地提高镍渣中铁的还原率.通过比较1450～1600℃范围内各反应温度下不同类型还原反应的Gibbs自由能,镍渣熔融还原过程的主要反应形式为(FeO)+C(S)→[Fe]+CO.本实验确定的最佳配方组成为:镍渣100g、CaO34.7g、CaF24.04g和焦炭8.5g;最佳反应条件为1500℃熔制180min.以上条件下的渣铁分离效果较好,铁还原率达到96.32%.     

磁场对粉煤灰中重金属淋滤释放特性的影响

对循环流化床锅炉粉煤灰的性质进行了研究,根据磁场对水和矿物作用的一般规律,采用Tessier连续化学浸取方案和连续淋滤方法,研究了磁场对粉煤灰中重金属元素Pb,Cr的淋溶释放特性的影响及其规律．研究表明,外加磁场的大小对铅的溶出有着重要影响,但与溶出铅的浓度不是线性关系,而是存在一个最佳的点或范围．0．1T的磁场强度与pH=4．0的酸度条件下,最有利于铅的溶出．而外加磁场的大小对铬的溶出影响不明显．     

世界与我国铅锌产量和消费量的重构相空间混沌预测

通过对世界与我国铅锌产量和消费量的详细调查,采用混沌理论重构相空间预测方法,对世界与我国铅锌生产和消费趋势进行了预测.研究表明,世界与我国铅锌产量和消费量的演化均存在混沌现象,且锌金属产量和消费量演化的分形维数比铅金属高,说明锌市场变化比铅市场复杂.预测结果显示,2010年以后世界铅锌产量小于消费量,市场将出现短缺,且锌金属短缺程度比铅金属严重;我国铅金属将出现过剩,而且越来越严重,锌金属产消基本平衡.此外,探讨了我国铅锌原料供需矛盾.研究结论为:我国铅精矿将出现严重的短缺,锌精矿短缺程度比铅工业略好,但资源短缺将会长期存在.针对我国铅金属生产严重过剩,同时铅精矿严重不足的局面,提出了重点对我国铅工业进行产业结构调整的建议.     

Leaching of the residue from the dry off-gas de-dusting and desulfurization process of an iron ore sinter plant

The residue from a second-stage dry sinter plant off-gas cleaning process contains both the fine dust from the sinter plant and the sorbent used. Recycling of the material that is usually handled by landfills to the sinter plant feed is not possible because of its chloride content. Leaching of the chlorides allow the recycling of remaining solids. The saline leachate produced contains some heavy metals and must be treated before it is discharged into the sea. In laboratory experiments, leaching tests with the subsequent treatment of the leachate were conducted. After the process was optimized, all heavy-metal concentrations were below the permissible values. The optimum treatment conditions for heavy-metal precipitation were observed to be the filtration of the suspended solids followed by the dosing of liquid with lime milk (pH 10) and the subsequent precipitation using sodium sulfide.