硫脲冻融循环法分离硫酸铅−黄钠铁矾共沉淀物中的铅

采用硫脲冻融循环法分离硫酸铅?黄钠铁矾共沉淀物中的铅.研究结果表明,在硫脲冻融循环过程中,共沉淀物中混杂的硫酸铅逐渐在黄钠铁矾表面聚集,同时通过冷冻浓缩作用促进硫酸铅与硫脲反应,形成硫脲铅(Pb-tu)新物相.随着冻融循环次数的增加,黄钠铁矾周围的硫酸铅不断转化为Pb-tu,直至反应完全.12次冻融循环后,黄钠铁矾和P...     

黄钠铁矾法除铁在阜冶的应用浅析

介绍了黄钠铁矾法除铁的操作和掌握及工艺控制条件。     

从黄钾铁矾渣中回收锌铟

由黄钾铁矾渣在焙解过程的化学变化，确定回收锌铟的适宜焙解温度为４２１．５～６７０℃。实验表明黄钾铁矾渣中铁酸锌转化为可溶硫酸锌的转化率随焙解温度升高而增加，可浸出的铟由焙解温度和时间决定。当温度为５６０～６２０℃、时间为３０～１０ｍｉｎ时，锌的浸出率为８０％、铟为９０％。     

NaOH分解含铟铁矾渣新工艺

提出NaOH分解含铟铁矾渣新工艺,考察NaOH用量、液固比、温度和时间对铁矾渣分解率的影响,并讨论铁矾渣中杂质金属,如Zn、In、Cu、Cd、Pb、As、Sb、Sn和Ag等在NaOH分解过程中的行为.结果表明:在m(NaOH)-m(铁矾渣)=0.381 4-1、温度60 ℃、液固比2-1、反应时间2 h的最优条件下,铁矾渣的分解率达到98.03%,而原料中的杂质金属,如Sn、Sb、Zn、In、Cu、Cd、Pb和Ag等绝大部分留在分解渣中,As则以AsO43-的形态大部分进入溶液,浸出率达到83.36%.DSC-TGA热分析和X射线衍射分析结果表明:在NaOH分解过程中,铁矾渣中的铁主要以Fe3O4形式沉淀入渣;分解渣中Fe、In和Zn的含量分别为38.81%、0.23%和12.89%;经稀盐酸选择性浸出铟和锌后,进一步磁选富集可作为炼铁原料.     

利用黄钾铁矾渣处理赤泥的研究

赤泥中含有大量的OH-和Na+,而黄钾铁矾中富含硫酸盐和H+。赤泥和黄钾铁矾在水中反应将生成含硫酸钠的液相和以氢氧化铁为主的固相。液相可以蒸发浓缩沉淀出硫酸盐而固相煅烧后得到的含铁固体可以作为赤铁矿砂。本文尝试利用赤泥和黄钾铁矾的反应制备成有价值的石膏、芒硝和赤铁矿砂产品。     

湿法炼锌危废铁矾渣水热分解及铁物相转化行为

湿法炼锌过程产出的铁矾渣含有大量的有价金属锌、铅以及伴生金属铁,在水热条件下,危废铁矾渣将发生高效分解与转化,有价金属转入溶液,伴生铁转化为赤铁矿。本文以湿法炼锌企业产出的铁矾渣为研究对象,研究了反应温度、反应时间、液固比、初始酸度、晶种浓度等宏观技术参数对铁矾渣分解与转化的影响规律。理论计算和实验结果均表明在高温水热体系中,铁矾渣中的黄钾铁矾、黄铵铁矾和铁酸锌物相均可有效转化为赤铁矿,而铅铁矾性质稳定不易转化。升高温度并延长反应时间有利于黄钾铁矾、黄铵铁矾和铁酸锌物相的水热分解与转化。在220℃下反应1 h后,铁矾物相转化基本完成,其转化率达94%;反应4 h后铁酸锌物相衍射峰完全消失,锌浸出率达87%,转化渣中赤铁矿含量达68%。适当提高初始酸度有利于铁酸锌的转化,但当体系初始酸度高于15 g/L时将抑制铁矾物相转化。在反应温度220℃、反应时间4 h、液固比(mL/g) 10:1、初始酸度0.01 g/L的条件下,锌浸出率为89%,铁矾物相的转化率可达95%,铁矾转化渣中主要物相为赤铁矿,其含量为68%。     

硫酸铵焙烧红土镍矿的溶出液制备黄铵铁矾

为得到硫酸镍溶液除铁的合适工艺条件,以硫酸铵焙烧红土镍矿的熟料溶出液为原料,采用 NH4HCO3合成黄铵铁矾。考查了反应温度、反应时间、反应终点pH以及Fe3＋初始浓度对除铁率的影响。以上因素均对Fe3＋的去除率有显著影响,其中反应温度的影响最为显著。合适的反应条件为：Fe3＋初始浓度19.36 g/L、反应温度95℃、反应时间3.5 h、反应终点pH2.5。在此条件下所得到的黄铵铁矾为包含片状或棱形颗粒的花簇结构。     

铁矾、铅银混合渣提取有价成分研究

为了回收湿法炼锌“早熟”时所形成的铁矾、铅银混合渣中的有价成分,研究采用焙烧预处理——硫酸浸出工艺富集混合渣,可将渣中锌与铁浸出95％左右,渣率不到13％,铅银富集倍数为8倍以上,富集渣中铅含量达30％,银达3千克／吨,可以作为提铅原料。     

球铁中的奥氏体枝晶

综述了近年来国外及作者本人关于球铁中奥氏体枝晶的分类、形成的内在原因、影响因素以及枝晶对补缩、机械性能的影响的研究结果。     

过渡层红土镍矿中的镁质矿中和沉矾浸出

采用沉矾浸出法将铁质矿浸出液对镁质矿进行沉矾浸出。结果表明:镁质矿酸浸过程中,在镁质矿粒度为106～150μm、搅拌强度为150 r/min、终点pHe值为1.3、温度为95℃的条件下,浸出镁质矿3 h,镍、镁、铁的浸出率分别为93.34%、78.28%、26.4%;在沉矾浸出过程中,在反应温度为95℃、搅拌强度为150 r/min、硫酸钠中的钠与形成黄钠铁矾中的钠的摩尔比x为1.3、镁质矿粒度为106～150μm、反应终点pHe为1.3±0.2的条件下,沉矾浸出5 h,镍浸出率能达到92%,镁浸出率在74%以上,铁质矿浸出液除铁率达到87%以上,铁质矿浸出液中铁的浓度在15.87～42.16 g/L的范围内,对镁质矿的镍、镁浸出及铁质矿浸出液中Fe的浓度没有显著的不利影响,溶液中铁基本上控制在4 g/L以下。     

从硫酸溶液中还原制取金属碲粉

采用SO2作还原剂从含碲的硫酸溶液中制取碲粉,并确定了SO2从硫酸溶液中还原碲的最佳条件:反应温度80℃,NaCl浓度1.2 mol/L,SO2流量0.1 m3/h,反应时间40 min,碲的还原率达到99.63%。还原所得粗碲粉经亚硫酸钠脱硒,盐酸酸洗除杂处理,硒、砷、锡、铜的脱出率分别达到99%、93%、80%、87.5%,得到含碲99.669%的金属碲粉。XRD和SEM表征表明还原碲粉的形态为针状晶体。     

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.     

含铟硫酸浸出液中铟的富集

提出从含铟硫酸溶液中富集铟的试验方案:焙砂预中和-硫化锌精矿还原-石灰石中和沉铟.在一定的实验条件下确定各个工序的最佳条件:焙砂用量为理论量的1.3倍,预中和后溶液的酸度大约降至6 g/L;硫化锌精矿的用量为理论量的2.2～2.3倍,还原后溶液中Fe3+的浓度大约降至0.5 g/L;石灰石的用量为理论量的2.0倍,沉铟后溶液中In3+的浓度降至1 mg/L以下.在此最佳条件下,可使沉铟渣中铟的品位达到0.1%以上.     

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%。     

Nonstationary processes that occur on nonpolarizable lead surface in sulfuric acid

The dynamics of variations in the currentless potential of a lead electrode in 2 M solution of sulfuric acid after cathode treatment by liberating hydrogen is studied. It is shown that, in the course of cathodic polarization, the liberation of hydrogen is accompanied by the formation of a film of lead sulfates due to corrosion that occurs on the metal surface. The major component of the measured potential is the voltage drop in the sulfate film. Two explanations for the simultaneous occurrence of the processes of hydrogen liberation and lead corrosion, which is impossible from the viewpoint of thermodynamics, are proposed. The first explanation is based on the electrical nonuniformity of the surface, which results from current localization at single active points (microzones), and on the absence of a protective cathodic potential at a distance from these points. The second explanation involves the voltage drop in the sulfate film, which is the component of the potential measured at the film–electrolyte interface. At the metal–film interface, the anodic polarization of the metal surface can occur, while nominally cathodic polarization takes place. Upon current interruption, the intricate processes of the growth and recrystallization of the sulfate film accompanied by the stepwise passivation of lead continue to occur. The limiting process for the corrosion system is the anodic reaction of the dissolution of lead.     

Leaching platinum-group metals in a sulfuric acid/chloride solution

A leaching process was established based on the ability of platinum-group metals to form stable chloro-complexes in acidic chloride solutions. Industrial catalyst losses were examined for the recovery of platinum, palladium, and rhodium by leaching with a mixture of sulfuric acid and sodium chloride to avoid using aqua regia or autoclave conditions. Extraction of platinum and rhodium in 60% H₂SO₄ at 135°C steadily increased with increasing NaCl concentrations reaching 95% and 85%, respectively, at 0.1 M NaCl after two hours. By comparison, palladium was dissolved more quickly but also reached 85% under the same conditions. Extraction of each metal increased with temperatures up to 125°C but plateaued at higher temperatures. Similar behavior was observed with increasing H₂SO₄ concentrations up to 60%. More than 99% extraction of each metal was obtained after ten hours using 0.1 M NaCl and 60% H₂SO₄ at 125°C. For more information contact Mohamed Hesham Hassan Mahmoud in the Extractive Metallurgy Department, Central Metallurgical R&D Institute, P.O. Box 87, Helwan, Cairo, Egypt; (202)5010642, Ext. 213; Fax (202)5010639; e-mail mheshamm@hotmail.com     

Leaching isomorphism rare earths from phosphorite ore by sulfuric acid and phosphoric acid

Most of the phosphorite deposits in the world contain isomorphism rare earths(RE) which are considerably difficult to be leached into solution in the wet phosphoric acid process. In this work, a systematic study of leaching RE using sulfuric acid, phosphoric acid, mixed acid and two-step leaching of phosphoric acid and sulfuric acid was performed. The aims are to illuminate the main factors that inhibit RE leaching and to provide insights into the further enrichment of RE in the wet phosphoric acid solution. The results indicate that H_2SO_4 is not an effective acid for leaching isomorphism RE from phosphorite ore.The low RE leaching efficiency attributes to the RE cocrystallized and encapsulated by phosphogypsum(PG) as well as the precipitation of RE by RE sulfates or phosphates. High concentration of H_3PO_4 can enhance the dissolution and diffusion of RE ions. Hence, the optimized leaching mode of improving RE leaching efficiency is to adequately dissolve phosphorite ores in high concentration of H_3PO_4 solution and then add H_2SO_4 to crystallize PG.The effect of co-crystallization or encapsulation of PG on RE can be decreased due to the crystallizing mode of PG in the bulk solution instead of on the interface of solid reactants. RE leaching efficiency can be high up to 65% by the optimized leaching mode.