Dissolution kinetics of zinc silicate (hemimorphite) in ammoniacal solution

The influence of ammonia-ammonium ratio, specific surface area of particles, associated anions, temperature and total ammonia concentration on the dissolution kinetics of zinc silicate (hemimorphite) has been investigated in ammoniacal solutions. The results show that the initial dissolution rate increases with the specific surface area, temperature and total ammonia concentration. The associated anions have a significant influence on the zinc extraction from hemimorphite due to their different complexing affinity with zinc ion. SEM analysis shows that cleavage of hemimorphite appears during dissolution. The silica formed in the dissolution process is absorbed back and covers some active sites of the particles. The Elovich equation typifies the dissolution behavior of hemimorphite in ammoniacal solution with an activation energy of 57.6 kJ/mol which is characteristic for a chemically controlled process.     

Effects of mechanical activation on physicochemical properties and alkaline leaching of hemimorphite

Effects of mechanical activation on physicochemical properties and alkaline leaching of hemimorphite from Lanping, Yunnan province were studied. The results showed that both dry milling and mechano-chemical milling could enhance zinc extraction rate and shorten the leaching time. Scanning electron microscope (SEM), particle size distribution (PSD), X-ray diffraction (XRD) analysis demonstrated that different milling times and modes could form different morphologies, particle size distributions and crystal structures, thus influence zinc recovery differently.     

Separation of zinc from spent zinc-carbon batteries by selective leaching with sodium hydroxide

Physical methods such as crushing and sieving followed by magnetic separation steps were applied to separate non-magnetic material (88.4 wt.%) from zinc-carbon spent batteries. The oversize material was processed by eddy current separation to recover zinc sheet, carbon rods, and plastics. The undersize fraction (− 2.36 mm) with a metal composition of 15.5% Zn, 17.5% Mn, and 1.4% Fe was used for the leaching experiments under different conditions such as concentration of sodium hydroxide, temperature, agitation speed, and pulp density. Selective leaching using 4 mol dm^− 3 NaOH at 100 g dm^− 3 solid/liquid ratio, 80 °C and 200 rpm gave a zinc extraction of 82% and a manganese extraction of less than 0.1%. An overall zinc recovery was about 88.5% by the processes described in this study.     

Recovery of manganese from zinc alkaline batteries by reductive acid leaching using carbohydrates as reductant

Acid leaching of manganese from the powders of exhausted zinc alkaline batteries has been studied using carbohydrates as reducing agents. A mono-saccharide (glucose) and a di-saccharide (lactose) have been tested, the second as a template of industrial wastes. Manganese leaching was complete in the presence of both carbohydrates, along with zinc. Data from the factorial test with lactose have been analysed with ANOVA. The most significant effects for manganese extraction were sulphuric acid concentration, temperature, and lactose concentration.     

Leaching of zinc sulfide in alkaline solution via chemical conversion with lead carbonate

A novel hydrometallurgical process for recovering Zn from ZnS in alkaline solution via chemical conversion with PbCO₃ was developed in this work. The S originally present in ZnS can be converted into PbS, while the Zn can be converted into Na₂Zn(OH)₄ in the alkaline solution in the presence of PbCO₃. And then, the Pb in PbS deposited in the leach residues can be converted into PbCO₃ again in the Na₂CO₃ solution. It was found that over 90% of Zn can be extracted from ZnS when the leaching process is operated in 6 mol/L NaOH solution at 90 °C with PbCO₃ as additive, and over 95% of PbS in the leach residues can be converted into PbCO₃ by stirring the leach residues in Na₂CO₃ solution with air bubble at a temperature of 80 °C. The leaching solution can be used to produce metallic zinc powder by electrowinning after chemical separation of impurities.     

含硅酸锌矿类锌资源利用现状

从硅酸锌物化性质的角度,总结分析了国内外含硅酸锌矿类锌资源利用现状,使用单一火法处理还原效率低,能耗高。强酸强碱浸出液浓度高,在浸出锌的过程中大量杂质也被同时浸出,对后续提锌除杂不利。氧化锌在氨浸提锌的过程中,杂质很少被浸出,根据氨对氧化锌这一选择性浸出特性,为提高锌资源中硅酸锌的利用率,建议采取焙烧矿相转化的方式先将硅酸锌矿转化为氧化锌物相,然后采用氨法浸出提取锌的技术路线,并对该技术路线需要开展的主要研究工作进行了分析。     

Kinetic study on bastnaesite concentrate mechanochemical decomposition in NaOH solution

Mechanochemical reaction involves simultaneous chemical reaction and particle crushing;the latter increases the effective reaction area and improves the reactivity,thus enhancing its kinetics.The classical shrinking core model was used to model the kinetics of bastnaesite mechanochemical decomposition in NaOH solution,which shows a questionable result.Mechanochemical reaction is a dynamic process,where the particle shape and concentration in reaction interface undergo constant change.Thus,a physically consistent model was applied to describe the kinetics.The variations in OH~-concentration and particle shape were considered in the revision of model.Considering the variation in OH~-concentration in solution with time,the model with varying OH~-concentration agrees better with the experimental data,improving the regression coefficients to between 0.936 and 0.992.By introducing fractal geometry to deal with the irregular system,the model was further optimized,and the regression coefficients increase to between 0.940 and 0.997.All these models considere shrinking particle approaches and controlling mechanisms for the diffusion and chemical reaction.Finally,the fractal model with varying OH~-concentration was selected to describe the mechanochemical decomposition of bastnaesite,which indicates that the process is controlled by chemical reaction.     

氧化锌的碱法浸出

根据等温溶解度法所研究的25℃、75℃和100℃时Na₂O-ZnO-H₂O三元水盐体系的平衡相图,借鉴氧化铝湿法生产工艺,探讨了氧化锌在碱性体系中的溶解和析晶过程.结果表明,对100℃、Na₂O的质量分数w_Na2O<34%的平衡体系,可采用稀释分解实现氧化锌的分离,完成该碱浓度下的氧化锌循环;对100℃、Na₂O的质量分数w_Na2O>29%的平衡体系,可采用冷却降温实现锌酸钠的分离,锌酸钠稀释后易得到氧化锌,从而完成在该碱浓度范围的氧化锌循环.     

EDTA滴定法测定氧压浸出溶液中锌

在锌冶炼过程的氧压浸出溶液中由于铜、铁、锰等杂质元素的存在,在使用EDTA络合滴定锌时终点变色不明显,结果误差大,不符合生产要求。采用硝酸消解试样,加入氨水、氯化铵、氟化钾沉淀分离铁、铅等元素,加入过硫酸铵使锰以二氧化锰析出;在pH 5.5~6.0的乙酸-乙酸钠缓冲溶液中以二甲酚橙为指示剂,铜(II)用硫脲掩蔽、铝用氟化钠掩蔽,滴定前加入抗坏血酸解除铁(III)对指示剂的封闭,用EDTA标准溶液滴定锌,终点明显,从而建立了使用EDTA滴定法测定氧压浸出溶液中锌的方法。按照实验方法测定氧压浸出溶液中锌,结果的相对标准偏差(RSD,n=9)为0.43%~1.3%;加标回收率为99%~102%;实验方法用于测定6个氧压浸出溶液中锌,测定值与参考值相吻合。     

Leaching of silica from vanadium-bearing steel slag in sodium hydroxide solution

The work aims to selectively extract silica from vanadium-bearing steel slag by a leaching process. The effects of the particle size, the ratio of solid to liquid, the concentration of sodium hydroxide solution and the leaching temperature on the leaching behavior of silica from vanadium-bearing steel slag were investigated. The leaching kinetics of silica from vanadium-bearing steel slag in 30-50% w/w NaOH solutions was studied at 240 °C and the shrinking-core model was established to express the leaching kinetics of silica. The data showed that the leaching rate was controlled by the chemical reaction on the system interface and the activation energy for the process was found to be 36.4 kJ mol^− 1. By the leaching process, the majority of silica could be removed effectively from the vanadium-bearing steel slag and a residue with a low SiO₂ content of 4.28% and a high V₂O₅ content of 11.15% was obtained. Under these conditions there was partial dissolution of Al and slight dissolution of Cr, Mn and Ti.     

EDTA滴定法测定氧化锌浸出渣中铅含量

提出了一种采用EDTA滴定法测定含铁、硅较高的氧化锌浸出渣中铅的分析方法。将试样用盐酸和硝酸溶解, 加入氟化铵分解渣料中的硅酸盐, 使包裹在硅酸盐中铅全部溶解出来, 然后加入硫酸使铅形成硫酸铅沉淀, 过滤, 使铅与锌、铁等易溶于硫酸的共存元素分离。用乙酸-乙酸钠缓冲溶液溶解硫酸铅, 加入抗坏血酸掩蔽少量铁, 以二甲酚橙为指示剂, 在pH 5～6的条件下, 用EDTA标准溶液滴定溶液至亮黄色为终点。样品测定结果的相对标准偏差在0.23%～0.76 %之间, 平均加标回收率为100.2%。     

论堆浸工艺中溶浸液的渗透问题

本文在简单回顾了国内外堆浸技术发展状况的基础上，从理论和实践两方面讨论了目前堆橙工艺中溶浸液渗透问题，并指出了存在的问题和研究方向，强调了研究渗透机理的必要性。     

添加增浸剂提高金浸出率试验

通过添加少量的增浸剂,可提高氰化浸出速度和金的浸出率。试验结果表明加入少量增浸剂,金矿石的堆浸（池浸）金浸出率可提高5％。     

Preparation of an As(V) solution for scorodite synthesis and a proposal for an integrated As fixation process in a Zn refinery

A method for producing an As(V) solution from an As-bearing material obtained from a nonferrous hydrometallurgical process was investigated. Preparation of the As(V) solution included oxidative leaching of As with a NaOH solution, elimination of As as a calcium arsenate precipitate (johnbaumite: Ca₅(AsO₄)₃(OH)) with a Ca(OH)₂ secondary salt, washing the Ca–As compounds, and reaction of the Ca–As precipitate with H₂SO₄. This process was shown to be industrially applicable. The As ions and Cu ions were effectively separated by oxidative leaching with O₂ gas injection under strongly basic conditions. In this system, As dissolved in the NaOH solution and Cu precipitated with the residue. The dissolved As in this highly concentrated NaOH solution was then effectively precipitated from the solution by addition of a surplus amount of CaO, which allowed recycling of the NaOH solution. The addition of surplus Ca precipitated Ca₅(AsO₄)₃(OH) and Ca(OH)₂, which inhibited the leaching of As but did leach Ca and Na. When the Ca–As compounds were dissolved with H₂SO₄, Ca ions precipitated in the form of gypsum from the As-bearing solution. The gypsum produced by this process is likely to give rise to a number of As-related issues and the As level, therefore, needs to be reduced. This process is advantageous for the treatment of As since it is stabilized as scorodite. The production of an As(V) solution could be applied to hydrometallurgical operations as it necessary for the removal of As. This process is shown to be practically useful to As removal in Zn refining and a closed flow circuit is proposed for integration of this process into a Zn hydrometallurgical operation.     

Kinetics of reductive leaching of manganese oxide ore with molasses in nitric acid solution

The kinetics of reductive leaching of manganese from a low grade manganese ore in dilute nitric acid in the presence of molasses is examined. The rate is controlled by diffusion through the “product” layer composed of the associated minerals. The leaching process follows the kinetic model 1 − 2/3X − (1 − X)^2/3 = kt with an apparent activation energy of 25.7 kJ/mole. It is concluded that the concentration of HNO₃ and molasses as well as temperature are the main factors influencing the leaching rate. The results indicate a reaction order of 1.2 for HNO₃ concentration and 1.9 for molasses concentration.     

高硅氧化铅锌矿加压酸浸中锌的浸出动力学

探讨了高硅氧化铅锌矿加压酸浸中锌的浸出反应动力学。考察了温度、硫酸浓度、搅拌速度、压力对锌浸出的影响。结果表明:氧化铅锌矿加压酸浸过程遵循未反应核缩减模型,浸出过程属于固体膜层扩散控制,表观反应级数为1.09,在试验选取的温度范围内,浸出的扩散活化能14.3kJ/mol。     

二乙基三胺五乙酸滴定法测定赤泥酸浸液中三氧化二钪

吸取10mL赤泥酸浸液于烧杯中,用氨水溶液调节溶液pH值为3.0~3.5并搅拌20min,经固液分离去除沉淀物,将滤液倒入500mL锥形瓶中,再用盐酸和氨水调节溶液pH值为1.6~6.0,加入5滴二甲酚橙指示剂,用二乙基三胺五乙酸(DTPA)标准溶液滴定至由酒红色变为亮黄色即为滴定终点。据此,建立了DTPA滴定法测定赤泥酸浸液中三氧化二钪的方法。考察了滴定pH值、三氧化二钪质量浓度、滴定温度、放置时间对测定结果的影响。结果表明:三氧化二钪质量浓度的检测范围为20~500mg/L,合适的滴定pH值范围为1.6~6.0,滴定温度为20~60℃;滴定前将溶液放置20min对测定无影响。共存元素干扰试验表明,将样品预先用氨水搅拌沉淀分离后再测定,可消除样品中铁和铝对测定的干扰;样品中其他共存元素不干扰测定。采用实验方法测定赤泥酸浸液中三氧化二钪的结果与电感耦合等离子体原子发射光谱法测得结果基本一致,相对标准偏差RSD(n=5)小于0.2%。     

难处理金矿石难浸的原因及预处理方法

本文分析了难处理金矿石难浸的原因，指出该类金矿石浸出前必须进行预先氧化，才能取得理想的浸出效果，为此，介绍目前已在国内外获得工业应用的几种预氧化方法的原理，应用概况及优缺点，最后这几种方法进行技术经济比较，供金矿开发者建厂时选择参考。     

氨浸法在湿法炼锌中的优点及展望

介绍了氨浸法在湿法炼锌中的优点及展望,分析了氨浸法具有原料适应性广,工艺流程短,净化负担轻,环境污染小,产品品种多,投资少等特点,并展望了该法在21世纪的应用前景,对传统湿法炼锌工艺提出新的挑战。