Recovery of gallium from zinc concentrate by pressure oxygen leaching

Zinc concentrate with high gallium content is one of the main resources of gallium.The gallium presents in the form of isomorphism in tetrahedron coordination with sulfur in sphalerite.The research was to investigate the amenability of zinc concentrate with high gallium to pressure oxygen leaching.The particle size,sulfuric acid concentration,oxygen partial pressure,additive amount,and time of reaction were studied.The extraction yields of gallium and zinc are 86%and 98%,respectively.The optimal condition is 100 g of zinc concentrate with particle size smaller than 38 lm,sulfuric acid concentration150 g L-1,leaching temperature 150℃,leaching time120 min,oxygen partial pressure 0.7 MPa,additive amount of 0.2 wt%.

Leaching Gold Using Oxidation Products of Elemental Sulfur in Ca(OH)_2 Solution under Oxygen Pressure

A gold leaching process by using in situ oxidation products of added elemental sulfur in Ca(OH)2 solution was investigated. A gold concentrate containing 45 g/t Au was tested and 85%~87% of gold were leached. The leached gold depends mainly on the initial molar ratio of elemental sulfur to the hydroxyl ion, the consumption of oxygen and the reaction temperature. Adding some surfactants, such as lignosulfonic calcium, at lower concentration increased the leached Au but at higher concentration decreased it. Both of thermodynamic analysis and experimental results show that thiosulfate is the major complexing agent for gold in the process.

Gold leaching with elemental sulfur in alkaline solutions under oxygen pressure

A gold leaching process by using oxidation products of elemental sulfur in alkaline solutions was proposed and investigated. A gold concentrate and a residue from an arsenic refractory gold concentrate by acidic oxidation leaching were tested. The residue contains 16.3% elemental sulfur and no more elemental sulfur was added in tests. For the concentrate elemental sulfur was added before leaching tests. The leaching ratio of gold depends mainly on the initial equivalent ratio of elemental sulfur to hydroxyl ions, the consumption of oxygen and the reaction temperature in the process. Analysis of the experimental results shows that thiosulfate is the majority complexing reagent for gold in the process. Over 90% gold was leached from the residue and 82%87% from the concentrate by using this process.

Oxygen pressure acid leaching of Gacun complex Cu concentrates

The treatment of the Gacun complex Cu concentrate with high contents of Pb,Zn,Ag,etc by oxygen pressure acid leaching was studied.It is unusual that tetrahedrite,whose treatment was rarely studied,is the primary copper mineral of the concentrates.Most of silver also occurs in the mineral.The optimum operating parameters of oxygen pressure acid leaching were established by conditional tests.Pilot scale test was carried out under the parameters,and the leaching rates of copper and zinc are as high as 97.10% and 89.83% while lead and silver are transformed into sulfate and sulfide respectively and stay in leaching residue.The copper and zinc in lixivium were reclaimed by extraction-electrowinning and purification-electrowinning,respectively,and the lead and silver in the residue were reclaimed separately by chloride leaching and thiourea leaching.The extraction rate of copper achieves 96%,and the leaching rates of lead and silver reach 90% and 95%,respectively.

An investigation of oxygen pressure acid leaching of Gacun complex Cu-Pb bulk concentrate

The treatment of the Gacun complex Cu-Pb bulk concentrate with high Zn,Ag,etc.,by oxygen pressure acid leaching was studied.The pri-mary copper and leadminerals in the concentrate are tetrahedrite and galena.The treatment of tetrahedrite was rarely studied,and most of silver occurred in themineral too.The optimum operating parameters of oxygen pressure acid leaching were established by conditional tests.Under these parameters,the result of pilot scale test showed that the leaching percentages of copper and zinc were separately as high as 98.9 wt.% and 94.9 wt.%,while lead and silver were transformed into sulfate and sulfide precipitations,respectively.The copper and zinc in lixivium were reclaimed by extraction-electrowinning and purification-electrowinning,respectively,and the lead and silver in the residue were reclaimed separately by carbonate transformation-silicofluoric acid leaching and thiourea leaching.

Pressure acid leaching of zinc sulfide concentrate

Effects of particle size of the zinc sulfide concentrate,leaching temperature,solid-to-liquid ratio and additive amount on pressure acid leaching process of the zinc sulfide concentrate were studied.The results indicate that the additive can improve the reaction kinetics and the conversion rate.And sulfur can be successfully separated from the zinc sulfide concentrate as elemental sulfur.The reasonable experiment parameters are obtained as follows:the leaching temperature 150℃,oxygen partial pressure 1 MPa,additive amount 1%,solid-to-liquid ratio 1:4,leaching time 2 h,initial sulfuric acid concentration 15%,and particle size less than 44μm.Under the optimum conditions,the leaching rate of the zinc can reach 95%and the reduction rate of the sulfur can reach 90%.

Analysis of Oxygen Consumption in Lead and Zinc Metallurgy

This article analyzes the role of oxygen in lead zinc metallurgy,including shortening the metallurgical process,promoting energy conservation and environmental protection,improving metallurgical strength,enhancing raw material adaptability,and enhancing comprehensive recovery efficiency.This article introduces different lead zinc metallurgical processes and their oxygen consumption characteristics,including oxygen enriched side blowing lead smelting,oxygen bottom blowing lead smelting,oxygen enriched top blowing lead smelting,flash smelting lead,oxygen pressure leaching zinc smelting,and atmospheric pressure oxygen leaching zinc smelting.It is pointed out that oxygen enhanced metallurgy is the direction for the transformation and upgrading of lead zinc metallurgy.

熟化银精矿氧压强化酸浸行为研究

考虑到浮选银精矿极具回收价值及现有处理工艺存在不足,本文提出了浓硫酸熟化-氧压酸浸的工艺路线。首先,以湿法炼锌过程产出的浮选银精矿为对象,利用浓硫酸的强氧化性破坏浮选残留有机药剂矿物包裹体;其次,利用加压湿法冶金独特的技术优势,氧化溶解低价金属硫化物;最后,重点研究了工艺过程原理及氧压酸浸过程中Zn、Cu、In的浸出行为及浸出渣中物相演变规律。研究结果表明:银精矿颗粒表面残留浮选药剂的P—O、—OH极性基团被浓硫酸氧化破坏,避免产生水化膜及矿物包裹体,促进有价元素的高效浸出;在高温氧化性条件和Fe^(3+)/Fe^(2+)变价离子对的助浸作用下,银精矿物相组成由闪锌矿(Zn S)、锌铁尖晶石(Zn Fe_(2)O_(4))、雌黄铁矿(FeS)、氯化亚铜(Cu Cl)为主,转变为以单质硫(S8)、铁矾(MeFe_(3)(SO_(4))_(2)(OH)_(6),Me:1/2Pb、Na、K)为主,并残留少量的闪锌矿(Zn S);适当提高反应温度和初始质量浓度可促进Zn、Cu、In的浸出,但温度过高会加剧铟铁共沉淀问题,降低锌、铟的浸出率,而质量浓度过高会降低氧气对Zn S及Fe^(2+)的氧化能力,抑制银精矿的溶解;延长反应时间、提高液固比均有利于银精矿的解离浸出;增大氧分压可以加强Fe^(3+)的氧化能力,促进对Zn S及ZnxFe(1-x)S的氧化作用,有利于银精矿的溶解。在反应温度为155℃、初始质量浓度为160 g/L、反应时间为180 min、氧分压为0.8 MPa、液固比为7 m L/g的氧压酸浸条件下,银精矿中Zn、Cu、In的浸出率分别为94.84%、98.36%和91.13%,浸出渣中铅、银的富集率分别达2.88%和7754 g/t。

含锗银精矿浓硫酸熟化—氧压酸浸行为研究

针对银精矿极具回收价值及现有处理工艺不足,以湿法炼锌过程产出的含锗银精矿为研究对象,提出了浓硫酸熟化—氧压酸浸的工艺路线,重点研究了工艺过程原理、含锗银精矿熟化—氧压酸浸过程中Zn、Cu、Ge的浸出行为及浸出渣中物相演变规律。结果表明:浓硫酸熟化预处理可氧化分解含锗银精矿中低价金属硫化物及精矿表面残留的有机药剂;在高温氧化性条件和Fe^(3+)/Fe^(2+)变价离子对的助浸作用下,物相组成由含锗银精矿中的铁闪锌矿、闪锌矿、雌黄铁矿、氯化亚铜为主演变为浸出渣中单质硫、硫酸铅、铁矾(MeFe_(3)(SO_(4))_(2)(OH)_(6),Me:1/2Pb、Na、K)为主。最佳氧压酸浸条件为:反应温度150℃、初始酸度150 g/L、反应时间180 min、氧分压0.6 MPa、液固比7 mL/g,Zn、Cu、Ge的浸出率分别为95.34%、98.47%和84.35%,铅、银分别富集至3.19%和7074 g/t,实现了Zn、Cu、Ge的深度浸出和铅、银的高效富集。

硫化铜精矿氧压浸出工艺研究

针对硫化铜精矿火法冶炼工艺副产硫酸,在硫酸滞销地区难以大规模应用的难题,研究了硫化铜精矿氧压浸出工艺。对某硫化铜精矿原料进行了成分、物相及有价元素赋存状态进行了分析,从理论上分析了硫化铜精矿氧压浸出的可行性,研究了一段氧压浸出、两段氧压浸出工艺的浸出效果。结果表明,采用两段逆流氧压浸出,一段控制温度135~140℃、总压1.3~1.4 MPa、时间2~2.5 h、分散剂加入量4‰~5‰、沉矾剂加入量2.5%~3%,二段控制温度165~170℃、总压1.4~1.5 MPa、时间2.5~3 h、分散剂加入量6‰~7‰、沉矾剂加入量3%~3.5%,铜浸出率为90.49%~92.25%。一段浸出液含铜90.78~92.34 g/L、铁10.81~13.84 g/L、硫酸28~30 g/L,可送后续除铁净化—电积生产电铜。

硫化锌精矿氧压浸出过程中硫回收率的影响机制

锌氧压浸出特点是硫化锌精矿直接湿法浸出,硫不进入烟气,而是直接转化成单质硫回收。该工艺生产流程更简单,环境友好,综合能耗低,硫磺产品相对更容易储存、运输及销售。因此该技术的推广应用对锌冶炼行业节能减排乃至碳达峰碳中和目标的实现意义重大。从精矿矿物特征及金属的浸出率、浸出过程中硫元素的酸化率、硫磺结晶过程控制、硫浮选效率以及硫精矿热过滤温度等方面着手,研究其对硫磺回收率的影响机制。

锌氧压酸浸硫精矿和硫尾矿的矿物学特征与环境活性

采用电感耦合(ICP-OES)、X射线衍射(XRD)、扫面电镜-能谱(SEM-EDS)、红外光谱(FTIR)及差热差重(TG-DTA)等多种检测方法对锌氧压酸浸硫精矿和硫尾矿的化学组成、粒度分布、物相结构等进行研究,并采用毒性浸出程序(TCLP)法考察其短期环境活性。研究结果表明:硫精矿1和硫精矿2主要以球状的单质硫存在,其中,少量硫精矿1以小颗粒的闪锌矿、黄铁矿等附着于块状的铁矾颗粒上,平均粒径为22.1μm;少量硫精矿2以块状的闪锌矿、黄铁矿和小颗粒存在,平均粒径为59.2μm。2种硫精矿在185~340℃时会剧烈燃烧,热稳定性较差。2种硫尾矿主要以块状的铅黄铁矾和黄钠铁矾存在,其中,硫尾矿1平均粒径为21.1μm;硫尾矿2含一定的赤铁矿(16.86%),平均粒径为5.3μm。硫精矿和硫尾矿中锌和镉元素在弱酸条件下不稳定,硫尾矿1中锌和镉的浸出量分别超标1.93倍和2.50倍;硫精矿2的分别超标2.03倍和6.56倍,硫尾矿1的分别超标3.46倍和5.94倍,硫尾矿2的分别超标11.10倍和8.28倍。

辉钼精矿氧压水浸工艺及动力学研究

采用正交试验确定辉钼精矿氧压水浸的最佳工艺条件,并在最佳工艺条件下进行氧压水浸过程MoS_(2)氧化分解反应的动力学研究。结果表明:最佳工艺条件为矿石粒度D_(90)=46μm、氧压1.1 MPa、温度230℃、液固比7∶1和反应时间1 h;在此条件下,钼转化率最高可达99.77%,铜浸出率为97.56%,铁浸出率为88.71%。在170~230℃浸出温度范围内(其他工艺条件均为上述最佳值),辉钼精矿中MoS_(2)氧化分解反应遵循未反应核收缩模型,该过程主要受化学反应控制,表观活化能为45.71 kJ/mol。

钼精矿氧压水浸渣碱性浸出工艺研究

氧压水浸工艺处理辉钼矿,钼回收率达95%以上,此工艺中,采用氨水浸出氧压水浸渣会导致氨浸渣中存在部分难溶钼酸盐,降低了钼回收率。本文尝试采用NaOH或Na_(2)CO_(3)为浸出介质对氧压水浸渣进行浸出实验,考察浸出剂用量、温度、液固比及浸出时间等参数对钼浸出率的影响,并结合浸出渣的物相组成和热力学分析阐明浸出过程反应机理,得到以下主要结论。辉钼矿氧压水浸渣中含有CaMoO_(4)这一物相,采用NaOH浸出时,无法有效将CaMoO_(4)中的Mo浸出,但Na_(2)CO_(3)浸出时,在温度20℃以上,即可自发实现CaMoO_(4)的分解,且温度越高,分解越容易;采用Na_(2)CO_(3)作为浸出剂时,最佳工艺条件为浸出温度90℃、Na_(2)CO_(3)用量系数1.2、液固比4∶1、浸出时间90 min、搅拌转速400 r/min,此条件下钼浸出率为97.15%;采用NaOH作为浸出剂,最佳工艺条件下的浸出率为93.98%。试验结果表明,采用Na_(2)CO_(3)作为浸出剂可有效提升钼的浸出率。

锌精矿氧压浸出技术研究进展

传统湿法炼锌工艺存在烟气处理成本高、综合回收率低等问题,近年来,氧压浸出技术因其原料适应性强、锌回收率高、环境污染小等优点被应用于锌冶炼工业。介绍了锌精矿氧压浸出的生产工艺,论述了浸出过程反应的基本原理,以及离子催化、温度、酸度、分散剂等条件对浸出过程离子行为的影响与机理。讨论了锌精矿中伴生元素铁、镓、锗、铟、银等的浸出行为及在不同浸出条件下浸出行为的变化。

锌精矿与锌浸出渣加压协同助浸过程优化及动力学研究

针对现有锌精矿加压浸出过程氧气利用率低、闪锌矿溶解缓慢等问题,利用锌浸出渣中Fe^(3+)的强氧化性以及Fe^(2+)/Fe^(3+)氧化还原离子对输氧效能,提高闪锌矿中S2-的高效氧化和O2利用率,实现锌精矿与锌浸出渣协同强化浸出。研究结果表明,Fe^(3+)可以提高氧气溶解度,同时Fe^(2+)/Fe^(3+)的往返变价行为加速了浸出体系中闪锌矿(ZnS)的氧化溶解,锌的浸出速率和浸出率得到显著提升。以锌浸出渣形式向体系添加6.1 g/L Fe^(3+)、在初始硫酸浓度95 g/L、氧分压0.8 MPa、反应温度160℃、反应时间120 min的优化条件下,锌浸出率可以达到98.60%,同时92.47%的铁沉淀入渣。锌精矿与锌浸出渣协同助浸受外扩散控制,表观活化能为15.61 kJ/mol,协同浸出的动力学方程为:1-(1-α)^(1/3)=4.78×10^(-3)×{[exp(-15610/8.314T)]×c(H_(2)SO_(4))1.21×P(O_(2))0.33×c(Fe^(3+))0.43×t}。

加压湿法炼锌过程中镓的富集

针对冶炼厂在两段逆流氧压浸出湿法炼锌工艺流程中镓回收率不高的问题,采用XRD、SEM-EDS与ICP-AES对工艺流程中各阶段溶液及渣物相进行分析,对预中和、锌粉置换过程影响镓回收率低的机理进行探究。结果表明,一段浸出液预中和过程中由于酸度过低导致镓与黄钾铁矾或铅黄钾铁矾共沉淀的现象是导致镓损失的原因。在锌粉置换过程中,单纯依靠锌粉提高pH无法使镓完全沉淀。据此提出在预中和过程中通过将预中和终点酸度控制为6 g/L,反应时间1 h,反应温度85℃,以避免镓损失;在置换阶段采用先中和再置换的方法:先使用优质氧化锌7 g/L进行中和反应1 h,然后加入锌粉3 g/L反应1 h,反应温度85℃。经优化后,镓沉淀率可达100%。

极低品位硫钴精矿钴回收工艺试验

随着新能源汽车、储能系统等产业的快速发展,对钴的需求持续增长,钴回收尤为重要。硫钴精矿是将钒钛磁铁矿原矿经过破碎、磨矿、选铁、浮硫等工序加工而得的粉状矿物,主要成分是硫和铁,同时含有贵重金属钴、铜、镍等。硫钴精矿是生产钴产品的主要原料之一,但其中的钴含量通常较低,需要通过进一步的处理和回收来提高其利用价值。本研究针对硫钴精矿的钴品位极低、元素赋存状态复杂的特点,采用氧压浸出工艺。结果表明:控制浸出温度为160±5℃、浸出氧分压1.2 MPa、反应时间2 h、液固比4∶1、始酸浓度16%的条件下,钴、铜、铁、硫的浸出率分别达到78.5%、75.4%、32.7%、31.4%。硫钴精矿处置直接经济效益为30.36美元/t,按年生产加工80万t硫钴精矿测算,则年毛利润2 428.80万美元,具有良好的经济回收前景。

草食性动物排泄物制备氧压酸浸炼锌中分散剂的研究与应用

对木质素在牧草中的分布进行了详细论述,对牧草的体外消化情况进行了分析,归纳了木质素的制备方法,创新地提出了草食性动物排泄物富集—提纯木质素—制备氧压浸出分散剂的技术新思路,利用草食性动物消化—分解—吸收植物中蛋白质、脂肪、淀粉和粗纤维等营养物质,使得木质素残留在排泄物中,实现了木质素的天然富集,并以草食性动物排泄物为原料制备了木质素磺酸盐,提高了木质素磺酸盐单位产出量,降低了处理成本。该技术结合了冶炼企业与草原生态的特点,充分利用当地草原牧草、草食性动物排泄物为锌氧压浸出提供制备木质素的原材料,形成了植物-动物-锌冶炼多领域相结合的特色工艺。

电解废液与锌冶炼烟尘协同处理试验研究

为综合回收锌冶炼烟尘及电解废液中的有价金属,本研究进行了利用锌电解废液浸出锌冶炼烟尘试验,通过一段浸出、碳酸钠中和沉锌及氢氧化钠中和沉镁等工序获得了沉锌产品和沉镁产品,通过两段浸出、萃取反萃、中和沉铟等工序获得了富铟渣及副产物铅银渣。试验最佳工艺条件:一段浸出为液固比4:1,浸出温度80℃,浸出时间75 min,在此条件下,锌浸出率可达78.69%,铟浸出率仅为8.4%;锌镁分离最佳终点pH值区间为6.86~7.80;二段浸出最佳工艺条件为终点pH值1.08,液固比3:1,浸出温度75℃,浸出时间10 h,在此条件下,铟浸出率可达86.84%。该研究可为炼锌厂开路除杂及综合回收有价金属提供新思路。