氧在硫代硫酸盐浸金过程中的作用

用纯金和金矿石对氧在硫代硫酸盐浸出金过程中的作用进行了研究。在硫代硫酸盐浸出纯金过程中,氧存在时金的溶解作用下降。空气或氧气使得硫代硫酸盐耗量增大,并且使金钝化。     

高铁硫化锌精矿直接浸出新工艺研究

在 10 0～ 110℃、0 4MPa及液固比 4 5～ 5 0条件下 ,硫酸氧化浸出高铁锌精矿 3～4h ,氧耗 90m3 /t矿 ,Zn、Cu、Cd和Fe的浸出率分别为 97%、80 %、95 %和 31% ,元素硫的产率为 2 10～ 2 32kg/t矿     

低品位含金硫精矿碱硫氧压浸金试验

对某低品位含金硫精矿开展氧压酸浸产元素硫、碱硫氧压浸取金银试验。酸浸试验最佳条件为:浸出温度130℃、浸出时间300min、浸出压力1.6 MPa、液固比4:1、始酸浓度20g/L、木质素2‰;碱硫氧压浸金试验最佳条件为:浸金剂(石灰+硫磺)为总干矿的40%、硫碱质量比0.35、催化剂Cu(NH_3)_4~(2+)0.06mol/L、活化剂木质素占干矿的0.2%、氧压0.4 MPa、温度105℃、浸金时间5h,在此条件下金、银的浸出率分别为88.23%和61.35%。     

氧在硫代硫酸盐浸出金过程中的作用

D．Feng和J．S．J．van Deventer研究了纯金和矿石体系中氧在硫代硫酸盐浸出金过程中的作用。纯金体系中有氧存在时金的溶解率下降。空气和氧气的存在使硫代硫酸盐耗量增大并使金钝化。氮气极大地提高了金的溶解率，在高氮气流速下，这种效果更明显。氮气泡可使硫代硫酸盐稳定并可防止金的钝化。氧气过饱和水的使用对金的溶解有较小影响。逆流浸出试验结果表明，金的钝化和硫代硫酸盐分解产物影响金的溶解。氮气氛下，浸出液可重复使用，对金的溶解没有影响。     

硫代硫酸盐对硫代硫酸铵浸出金的影响

D.Feng,et al.研究了在纯金和矿石体系中,用硫代硫酸钙、硫代硫酸钠和硫代硫酸铵浸出金。硫代硫酸盐的类型显著影响溶液和表面化学以及纯金和黄铁矿精矿在硫代硫酸铵溶液中的浸出。在纯金浸出体系中,硫代硫酸钙有较高的金浸出率,而耗量很低。浸出纯金时,硫代硫酸钠的耗量最少,金浸出     

硫代硫酸铵法从焙烧后的某含铜硫金精矿中回收金

研究了硫代硫酸盐法从焙烧后的含铜硫金精矿中回收金的工艺过程,研究了表明,焙烧前金的回收率很高,焙烧后,常温下硫代硫酸盐浸出,既可是较好的金浸出指标,又可降低硫代硫酸耗量。     

硫化铵提硫过程中汞、银、锌的浸出行为研究

硫化锌矿采用加压浸出技术处理后,得到的浸出渣经浮选和热过滤能获得纯度较高的硫磺,但硫的回收率低,且其中有价金属不易综合利用。利用硫化铵从热过滤渣中进一步回收硫,并对提取硫磺过程中汞、银、锌的浸出行为进行研究,分析了(NH4)2S浓度、液固比和浸出时间对浸出过程的影响。研究表明,在常温,(NH4)2S浓度为1.0 mol/L,液固比为6:1,浸出时间为60min的条件下,元素硫浸出率为95.36%,Hg、Ag、Zn的浸出率分别为4.71%、33.73%、0.32%。采用蒸馏热分解多硫化铵浸出母液,元素硫回收率为95%,获得的元素硫纯度达到99.5%以上。     

含砷难处理精金矿预氧化工艺对比试验

对某含砷高硫金精矿进行了两段焙烧与酸性氧压浸出两种预氧化工艺的对比试验。结果表明,精矿在两段焙烧试验条件下金回收率为88.59%,酸性氧压浸出条件下金回收率95.6%。经济效益评估结果表明该矿石宜采用氧压浸出工艺处理。     

提高某选矿厂金回收率的试验研究

针对某选矿厂金浸出率偏低的问题，对该厂原矿性质进行研究，确定了影响金浸出率的主要原因为原矿件铜、硫等耗氰元素含量偏高。通过该厂一段氰化浸出流程，分别对氯化钠用量、氰化时间、矿浆浓度、助浸剂种类和用量等影响因素进行了试验研究。研究结果表明，增加氯化钠用量或添加助浸剂可有效提高金浸出率，降低氰化尾渣中的金含量。在综合考虑现场具体情况的基础上，采用增加氯化钠用量的措施再次进行工业试验。试验结果表明，当氯化钠用量由原来的500 g/t增加至900 g/t时，金浸出率可达89%，比原指标提高了17%左右，金浸出率及尾渣金品位可连续、稳定达标。     

锌精矿氧压浸出渣熔硫工业生产改造实践

阐述了氧压浸出锌精矿后,浸出渣中元素硫的回收工艺及生产技术,由于在氧压浸出过程中,锌精矿中的硫主要以元素硫的形态进入到浸出渣中,浸出渣经过浮选、过滤得到可满足熔硫生产的硫精矿,在熔硫过程中,控制熔池的温度、液位及热滤时间,生产出硫磺产品品位在99.2%以上。     

Extraction of selenium from copper anode slimes in a sealed leaching system

A new method was proposed for extracting selenium from copper anode slimes with a low concentration of nitric acid in a sealed sulfuric acid leaching system. It is performed under an atmosphere of oxygen which allowed for a cyclic utilization of nitric acid. The effects of main parameters on selenium leaching were studied. The mineralogical characterizations of the typical samples were investigated by XRD and SEM. The results showed that the optimal conditions of the process are considered to be total gas pressure of 0.1 MPa, leaching temperature of 388 K, solid-liquid ratio of 0.20 g mL^–1, H₂SO₄ concentration of 2 mol L^–1, HNO₃ concentration of 0.07 mol L^–1 and leaching time of 2 h. The high selenium leaching efficiency of 99.23% was obtained under these conditions. According to the results of XRD and SEM-EDS, Cu–Ag selenide in the raw anode slimes is difficult to be leached with sulfuric acid alone; copper can be leached more easily from Cu-Ag selenide than silver; selenide is oxidized into the solution, undergoing the intermediate product of elemental selenium.     

铜阳极泥全湿法处理工艺研究

在 115℃及 0 4MPa操作压力下 ,氨浸新工艺具有流程短、金属直收率高、生产成本低和无环境污染等优点 ,适用于含金低的阳极泥。当其中硒碲具有明显回收价值时 ,可在氨浸金银铜之后增加一道稀硝酸浸出工序 ,以提取硒、碲和铅。对于含金及硒高的阳极泥 ,宜采用酸法工艺 ,采用本文中低浓度硝酸或亚硝酸盐的加压氧化酸浸工艺 ,金银铜硒的浸出率分别可达 99%、99%、99%及 98% ,且无污染 ,生产成本低     

Ferric ion leaching of chalcopyrites from different localities

Carefully sized and characterized chalcopyrite concentrates were produced from eleven different geographical localities, and were subsequently leached in both ferric chloride-hydrochloric acid and ferric sulfate-sulfuric acid media. When the leaching rates were corrected for the amount of CuFeS₂ in the particular sample, similar leaching rates (±50 pct) were observed for all eleven chalcopyrites under all experimental conditions. Furthermore, all the relatively pure chalcopyrites leached according to the same reaction and yielded >95 pct elemental sulfur as a sulfidic reaction product. In light of the current experimental observations, various reasons are offered for the order of magnitude variations in the chalcopyrite leaching rates often reported in the literature.     

Leaching of a sphalerite concentrate with H2SO4–HNO3 solutions in the presence of C2Cl4

The enhanced leaching of sphalerite concentrates in H₂SO₄–HNO₃ solutions and the extraction of sulfur with tetrachloroethylene were studied. Variables of the process were investigated including leaching temperature, reaction time, liquid / solid ratio, and tetrachloroethylene concentration. The number of cycles that tetrachloroethylene could be recycled did not have a significant effect on zinc extraction. The results indicated that 99.6% zinc extraction was obtained after three hours of leaching at 85 °C and 0.1 MPa O₂, when 20 g of sphalerite concentrate were leached in a 200 ml solution containing 2.0 mol/L H₂SO₄ and 0.2 mol/L HNO₃, in the presence of 10 ml C₂Cl₄. Leaching rates were significantly improved under these conditions.     

采用加压浸出工艺优化传统湿法炼锌流程研究

用传统湿法炼锌厂的热酸浸出液在高压釜中浸出锌精矿。结果表明,在温度130℃,液固比14∶1,精矿粒度-50μm占96%,浸出时间3h,氧分压600kPa,添加木质素磺酸钙0.4%的条件下,锌浸出率达97%以上,浸出液中的铁含量低于2g/L,加压浸出液可直接返到传统湿法炼锌流程的中性浸出,同时精矿中的硫以元素硫形式进入渣相。该工艺流程易与传统湿法炼锌厂现有流程结合,具有同时浸锌除铁、工艺流程简单、对环境友好等优点。     

Leaching of sulfidized chalcopyrite with H<Subscript>2</Subscript>SO<Subscript>4</Subscript>-NaCl-O<Subscript>2</Subscript>

The recovery of copper from chalcopyrite by leaching is complex not only due to the slow dissolution kinetics of this mineral in most aqueous media but also due to the production of solutions that are heavily contaminated with iron. On the contrary, the leaching of sulfidized chalcopyrite is very attractive because of a faster and more selective dissolution of copper compared to the leaching of the untreated chalcopyrite. In this work, the results of leaching in H₂SO₄-NaCl-O₂ solutions of sulfidized chalcopyrite concentrate are discussed. Experiments were carried out with chalcopyrite concentrates previously reacted with elemental sulfur at 375 °C for 60 minutes. The results showed that the concentration of chloride ions below 0.5 M, temperature, and leaching time are important variables for the extraction of Cu. On the other hand, Fe extraction was little affected by the same variables, remaining below 6 pct for all the experimental conditions tested. Microscopic observations of the leached particles showed that the elemental sulfur produced by the reaction does not form a coherent layer surrounding the particle, but rather concentrates in certain locations as large clusters. The leaching kinetics can be accurately described by a nonreactive core-shrinking rim topochemical expression for spherical particles 1 − (1 − 0.45X)^1/3=kt. The activation energy found was 76 kJ/mol for the range 85 °C to 100 °C.     

Sulfuric acid oxygen-pressure leaching of Ni3S2 prepared by a wet process

Ni₃S₂ prepared by a wet process was easily leached as nickel sulfate at 383 K, p_o2 1 MPa, and sulfuric acid concentration of 0.1–0.15 mol L⁻¹. The leaching reaction proceeds through the intermediate formation of NiS prior to complete dissolution. A constant leaching rate was observed for most of the duration of the reaction, and this has been attributed to an increase in the specific surface area of the sulfide particles. A thin sulfur layer was formed on the sulfide; the diffusion of oxygen through the sulfur layer was found to be rate-determining.     

高砷含锑难浸金精矿提金工艺的研究

研究了高砷含锑难浸金精矿的中温低压氧化酸浸- 石灰液或氨浸氧化的联合预处理技术的工艺和优化操作条件,讨论了催化氧化酸浸、石灰液氧化浸出及氧化氨浸中主要影响因素的作用规律。预处理后, 金的氰化浸出率由未预处理时的几乎为零提高到90%以上。同时提出了一种酸浸渣直接氧压浸金的新过程,论述了酸浸渣中的元素硫在石灰液中氧压浸出的过程原理及主要条件因子的影响规律。结果表明, 金的浸出率基本上与联合预处理后金的氰化率相同。     

氰化尾渣提金预处理试验研究

针对某选矿厂高硫高砷难处理金矿石经两段焙烧—氰化浸金后产生的尾渣,根据其性质分析,采用硫酸浸铁—硫脲浸金工艺进行提金试验研究。其结果表明:在最优浸铁条件下,即浓硫酸用量45 m L、温度80℃、时间4 h、液固比4∶1,产出的浸铁渣再进行硫脲浸金,铁和金的浸出率分别达到60.37%和63.44%;这表明采用该工艺可有效地从该氰化尾渣中浸出回收金。     

LEACHING OF MASSIVE RICH COPPER ORE WITH ACIDIFIED FERRIC CHLORIDE

Massive rich copper ore, taken from the Küre region of Turkey and containing 7.95% Cu, 2.13% Pb, 0.341% Zn, 0.087% Co, 0.039% Ni, 0.9 (g/t) Au, and 17 (g/t) Ag, was subjected to the experimental study. Effects of leaching time, ferric ion concentration, solid/liquid ratio, acid concentration and temperature on the metal dissolution efficiencies were investigated in order to determine optimum leaching conditions and related leaching kinetics. As a result of these experiments carried out under the optimum leaching conditions, 76% of Cu, 55% of Co, 96% of Ni, 100% of Pb, and 91% of Zn were extracted. The value of apparent activation energy for copper dissolution was found as 43.8?kJ/mole, which is in agreement with the published data. The rate-controlling step was found as the diffusion of ferric ions into the sulfur layer formed on the surface of partially leached ore particles.