MACA体系中循环浸出低品位氧化锌矿制备电解锌

以云南兰坪低品位氧化锌矿及其循环浸出渣的浮选精矿为原料,常温常压下在MACA(金属氨络合物)体系中进行循环浸出。浸出液先净化除砷和锑、再两段锌粉逆流置换深度净化,两次净化后液电积制取电解锌。考察工艺中循环浸出率、净化率、物质平衡以及电解锌质量和电耗等技术经济指标。结果表明:先用MACA法处理原矿粉,再浮选硫化锌的选冶结合流程是合适的兰坪低品位氧化锌矿的处理方案,原矿锌的平均浸出率为70.48%,其氨可溶锌浸出率达到89.14%,浮选精矿锌的浸出率为79.75%,杂质元素的净化率达到95%,电解锌纯度达到99.98%,电流效率可达97.02%。     

硫化锌精矿直接浸出动力学及其半经验模型（英文）

在中试规模管式反应器中直接浸出贫铁硫化锌精矿,研究其浸出动力学。为满足静压浸出条件,将含硫酸亚铁和硫酸的矿浆注入到垂直放置的直径8 m的管式反应器中,空气从反应器的底部吹入。考察了初始硫酸浓度、反应温度、矿粒直径、初始硫酸锌浓度、矿浆浓度和铁含量对反应动力学的影响。结果表明,硫化锌精矿的直接浸出遵循收缩核模型,过程受化学反应控制,其表观活化能为49.7 k J/mol。采用半经验模型描述该过程,得到铁含量、硫酸浓度、硫化锌精矿浓度和矿粒直径的反应级数分别为0.982、0.189、-0.097和-0.992。采用SEM-EDS对硫化锌精矿浸出反应前后的矿粒进行分析,发现当锌的浸出率低于60%时,由于反应器搅拌的问题,颗粒表面硫产物层脱落,会影响浸出溶液与颗粒表面的接触。     

富氧硫酸体系中硫化锌精矿的常压直接浸出动力学

采用常压浸出与加压浸出相结合的方法,研究硫化锌精矿在富氧硫酸体系中的常压直接浸出动力学。结果表明:锌浸出受界面化学反应控制,浸出反应表观活化能为(44.28±4.28)kJ/mol。浸出槽底部锌的浸出速率远高于浸出槽上部锌的浸出速率,且随槽底矿浆压力的增大,锌的浸出速率明显提高。进一步基于收缩核模型,通过二次回归方法,建立浸出槽底部锌的浸出动力学方程。在硫化锌精矿常压富氧直接浸出中,对于位于浸出槽底部的锌浸出,温度的影响明显大于矿浆压力的影响。     

湿法炼锌渣中锌和铟的还原浸出（英文）

铁酸锌是锌中性浸出渣中的主要物相,热酸浸出是处理中性浸出渣的主要方法之一。研究了一种采用硫化锌精矿作为还原剂对锌中性浸出渣进行还原浸出的方法。研究发现,采用硫化锌精矿作为还原剂不仅能高效浸出锌中性浸出渣中的有价金属,而且同时实现溶液中Fe~(3+)向Fe~(2+)的还原。采用两段逆流浸出工艺,98.1%锌和97.5%铟被浸出,浸出液中Fe~(2+)/Fe~(3+)的摩尔比达到9.6。同时发现,浸出过程中铁和铜几乎完全浸出,而锡只有部分浸出。     

云南冶金集团永昌铅锌公司“高铁硫化锌精矿加压浸出项目”投产

云南冶金集团“高铁硫化锌精矿加压浸出项目”10000吨电锌示范工程在云南永昌铅锌股份有限公司建成投产，实现了高铁锌精矿湿法冶炼的重大突破。     

硫化锌精矿直接还原蒸馏的热力学分析

对硫化锌精矿直接还原蒸馏涉及的一些重要热力学问题进行了分析、探讨。从热力学角度来看,添加石灰或石灰石的硫化锌精矿直接还原蒸馏在1300～1400 K 是易于实现的,锌的还原挥发率和硫的固定率均几乎达到100%。分别绘制了 ZnS-CaO-C 三组元系在1300K 和1400K 的热力学状态图,并根据状态图推荐了硫化锌精矿直接还原蒸馏的实际过程中易于控制的气氛条件。     

硫化锌银精矿富集银的工艺研究

基于国内外硫化锌矿处理的火、湿法研究进展,对含锌银精矿采用硫酸化焙烧、稀硫酸浸出工艺脱除锌、富集银,考察了焙烧和浸出过程中的主要影响因素。结果表明,硫酸配比为150%,在300℃焙烧90 min,以5%稀硫酸为浸出剂,液固比8:1,搅拌转速200~300 r/min,85℃浸出120 min,最终锌的浸出率可达到98%以上,浸出渣中银含量为7.24%,银被富集7倍。     

硫化锌精矿加压浸出过程的热平衡

笔者对硫化锌精矿加压浸出过程的热平衡进行了计算。计算结果表明,反应热不足以平衡物料升温至150℃温度时所需的热量,需从外部补充热量。针对这一问题,提出一种为保持设定的浸出温度返回浸出的废电解液需要达到的预热温度的计算方法。     

原料对湿法炼锌的影响与浸出工艺选择

阐述了在硫化锌精矿中,关键元素在湿法炼锌过程中的行为,及其对流程选择和生产操作产生的影响;对多种浸出及渣处理方案进行了分析、比较,并结合工程实际及原料成分,对浸出工艺流程进行了选择。     

浮选方铅矿精矿中铊的脱除工艺

针对一种铊(Tl)品位为48.75 g/t的含Tl方铅矿精矿,分别采用直接酸浸、次氯酸钠氧化-酸浸、碱式氧化-酸浸的方法,进行精矿脱Tl工艺研究。考查了矿浆浓度、次氯酸钠氧化时间、碱式氧化时间、充气速率、硫酸用量、浸出温度、浸出时间等因素对Tl去除率的影响。结果表明:采用直接酸浸和次氯酸钠氧化-酸浸工艺,脱Tl效果不佳,Tl去除率分别为17.9%和33.1%。采用碱性氧化-酸浸脱Tl工艺,脱Tl效果良好,当工艺条件为:充气速率为1.25 L/min,矿浆浓度20%,氧化温度为50℃,氧化时间32 h;浸出体系H_2SO_4用量12%,浸出温度40℃,浸出时间3.5 h,Tl去除率可达到83.4%,脱Tl铅精矿Pb品位为61.9%,Pb回收率为96.9%。脱Tl后的铅精矿中Tl的品位为8.05 g/t,符合铅冶炼企业的要求。     

Iron control in zinc pressure leach processes

The occurrence of zinc in sulfide ore deposits is generally accompanied by various iron minerals. Hence, even the most efficient concentrators generally produce a zinc concentrate with significant iron content. The efficient recovery of zinc metal from zinc concentrates requires the rejection of iron residue in a form that minimizes the zinc entrainment. Careful control of the iron precipitation step is important, so that the iron residue produced is amenable to efficient liquid-solid separation in order to obtain high zinc recoveries. In hydrometallurgical zinc processes, the coprecipitation of minor impurities along with iron precipitation is also important in producing zinc-sulfate solution from which high-purity zinc cathode can be electrowon. The integration of Dynatec’s zinc pressure leach process with existing roast-leach-electrowin plants employing various methods of iron rejection is briefly described in this article, along with the application of two-stage pressure leaching in stand-alone processes. For more information, contact W.D. Vardill, Dynatec Corporation, 8301-113 Street, Fort Saskatchewan, Alberta, Canada T8L 4K7; (780) 992-8190; fax (780) 992-8100; e-mail wvardill@mettech.dynatec.ca.     

Processing of Zinc Oxide Fume at Flin Flon,Manitoba

The hydrometallurgical processing of an impure zinc oxide fume is described. Flowsheet includes roasting for fluorine elimination and countercurrent leaching to produce a neutral sulphate solution. This solution is combined with zinc sulphide leach solution for subsequent purification and electrolysis. Novel to the neutral leach step is the use of automatic pH control.     

The activox® process: Growing significance in the nickel industry

Metal-market analysts project a global nickel supply gap which will be filled by further high-pressure acid leach treatment of laterite ores and the commercialization of hydrometallurgical refineries to recover nickel from sulfide ores. The Activox® process is one such hydrometallurgical technology developed to recover a range of base and precious metals from sulfide ores and concentrates. The combination of ultrafine grinding and oxidative leaching extracted and enabled the recovery of 96.2% nickel, 88.3% cobalt, and 82.9% copper in the 310 kg/h Tati Hydrometallurgical Demonstration Plant.     

Aqueous lixiviants: Principle,types, and applications

Aqueous lixiviant is a leach liquour capable of dissolving all or part of ore or concentrate. It plays a critical role in the hydrometallurgical process. Although cyanide leaching of gold and silver has demonstrated success in the industry for 100 years, searching for innovative lixiviants to leach gold, copper, and other heavy metals effectively, economically, and environmentally has never stopped. This paper reviews the aqueous lixiviant’s principle, summarizes the types used in numerous sites, and presents its new applications.     

玉龙铜矿硫化矿选矿产品方案的优化选择

通过对玉龙铜矿湿法冶金工艺中的硫化铜矿石选矿方案进行比较,阐述了生产铜硫混合精矿比生产单一铜精矿和硫精矿,具有以下特点:不仅简化了选矿工艺流程,减少了选矿药剂的消耗,降低了精矿和浸渣的脱水成本,同时,还简化了精矿焙烧的生产工艺和管理,改善了洗涤效果,保护了环境,提高了铜回收率,其经济效益十分显著。     

Utilizing Iron Residues from Zinc Production in the U.S.S.R.

When zinc calcine leach residues are subjected to conventional hydro-metallurgical treatment, iron is removed from the production circuit in the form of jarosite or goethite. A combined hydrometallurgical treatment of zinc calcine and zinc oxide fume leach residues applied at a zinc plant in the U.S.S.R. produces potassium jarosite containing undesirable impurities of 1.5–2.0 wt.% Zn, 0.2–0.3 wt.% Cu, 0.2–0.6 wt.% Pb, 0.005–0.01 wt. % Cd and 27–29 wt. % Fe. After some study, it was found that low-contaminant jarosite can be used in iron-oxide pigments and in cement clinker production. Methods for manufacturing such products have been developed and tested on a pilot-plant scale, and commercial tests are in progress. The investigations carried out for low-contaminantjarosite utilization resulted not only in the development of a wasteless and environmentally acceptable technology for zinc calcine treatment, but made it possible to recover one more valuable component—iron—from zinc raw materials.     

APPLICATIN OF ATTRITION GRINDING IN ACID LEACHING OF NICKEL SULFIDE CONCENTRATE

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Copper and elemental sulphur from chalcopyrite by pressure leaching

Processes employing direct oxidation under an over-pressure of air or oxygen in an aqueous sulphuric acid medium have been developed in the Sherritt Gordon Laboratories for iron, nickel, cobalt, zinc and lead sulphide concentrates. This study has recently also been extended to chalcocite, Cu₂S, concentrates. The rising interest in processes employing direct aqueous oxidation is stimulated by the fact that elemental sulphur can be produced as a by-product rather than sulphur dioxide or sulphuric acid.The present paper outlines a process which features the direct pressure oxidation of the most abundant copper sulphide mineral, chalcopyrite, CuFeS₂. The optimum conditions for a practical pressure leaching step have now been developed in the laboratory which results in the production of copper sulphate solution suitable for copper winninq by electrolysis, hydrogen reduction, solvent extraction combined with electrolysis, or other means. The leach residue yields pure elemental sulphur by-product. Copper and elemental sulphur recoveries of 98 and 85% respectively have been recorded. The fastest oxidation rate, corresponding to a leach retention time of 2.5 hr, was obtained when the copper concentrate was ground to 99.5% — 325 mesh, when a 50% stoichiometric excess of concentrate over the amount of available sulphuric acid for copper was used and when the oxygen partial pressure and temperature were maintained at 500 psi and 240°F, respectively. In an idealized form, the pressure leaching reaction can be expressed as follows:—CuFeS₂ + H₂SO₄ + 1 1/40₂ + 1/2H₂O → CuSO₄ + Fe(OH)₃ + 2S°After separation of the copper sulphate solution by filtration, elemental sulphur and excess concentrate ore recovered from the iron oxide tailing by flotation. The tailing, containing iron oxide and insolubles, is rejected. The elemental sulphur is separated from the concentrate by hot filtration, solvent extraction, distillation, or other means, and the unleached chalcopyrite is recycled to the leaching step.     

高铁锌焙砂在高梯度磁场中的元素分布行为(英文)

采用高梯度磁选将难溶铁酸锌从锌焙砂中分离,并利用ICP、XRD、穆斯堡尔、SEM及激光粒度仪分析锌焙砂中的元素组成及物相结构。考察磁感应强度对铁酸锌及杂质元素如钙、硫和铅在磁选过程中的分布行为的影响。结果表明,85%以上的铁酸锌在0.70T的磁感应强度下能分布到精矿中,60%的钙和 40%的硫主要分布在非磁性物相中,并在磁选过程中富集于尾矿中,大部分的铅以超细颗粒均匀分散在锌焙砂中。