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.     

Mechanism of influence of ferric ion on electrogenerative leaching of sulfide minerals with FeCl3

1 Introduction Electrogenerative leaching process is a newly- developed technique in hydrometallurgy. Although its principle has developed since the late 1960s[1], this technique has been overlooked in metallurgy until ZHANG et al[2] introduced it to the …     

Nickel-cobalt separation at Sherritt Gordon

Treating a complex nickel-copper-cobalt sulfide concentrate in a leaching operation has required the development of special techniques for the separation of nickel and cobalt. The process described in this paper can be applied with only minor modifications to a wide range of nickel and cobalt bearing solutions.     

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。     

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 primary copper and lead minerals in the concentrate are tetrahedrite and galena.The treatment of tetrahedrite was rarely studied,and most of silver occurred in the mineral 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.     

Kinetics of zinc sulfide concentrate direct leaching in pilot plant scale and development of semi-empirical model

The direct leaching kinetics of an iron-poor zinc sulfide concentrate in the tubular reactor was examined. All tests were carried out in the pilot plant. To allow the execution of hydrostatic pressure condition, the slurry with ferrous sulfate and sulfuric acid solution was filled into a vertical tube (9 m in height) and air was blown from the bottom of the reactor. The effects of initial acid concentration, temperature, particle size, initial zinc sulfate concentration, pulp density and the concentration of Fe on the leaching kinetics were investigated. Results of the kinetic analysis indicate that direct leaching of zinc sulfide concentrate follows shrinking core model (SCM). This process was controlled by a chemical reaction with the apparent activation energy of 49.7 kJ/mol. Furthermore, a semi-empirical equation is obtained, showing that the order of the iron, sulfuric acid and zinc sulfate concentrations and particle radius are 0.982, 0.189, ?0.097 and ?0.992, respectively. Analysis of the unreacted and reacted sulfide particles by SEM–EDS shows that insensitive agitation in the reactor causes detachment of the sulfur layer from the particles surface in lower than 60% Zn conversion and lixiviant in the face with sphalerite particles.     

Bioleaching of low grade nickel sulfide mineral in column reactor

Jinchuan low grade nickel （0.4%-0.6% Ni, mass fraction） sulfide mineral ore contains a remarkably high content of magnesia （30%-35% MgO, mass fraction） present in the main gangue minerals. Bioleaching was performed to investigate the feasibility to process the mineral due to its relative simplicity, eco-friendly operation and low capital cost requirements. The mixed mesophiles were enriched from acid mine drainage samples collected from several acid mines in China. Considering that the magnesia is easily extracted by acid solution and the excessive Mg^2＋ will exceed the tolerance of the mixed mesophiles, three effective means were used to reduce the disadvantage of magnesia during the bioleaching operation. They were adaptation of the mixed mesophiles to improve the tolerance; pre-leaching to remove most leachable magnesia and periodic bleeds of a portion of the pregnant leaching solution to control the level of Mg^2＋ based on the tolerance of the mixed mesophiles. An extraction of nickel （90.3%） and cobalt （88.6%） was successfully achieved within a 300 d leaching process from the Jinchuan low grade nickel sulfide mineral ore using a column reactor at ambient temperature.     

高砷难处理硫精矿氰化浸出提银实验研究

某硫精矿中的银矿物嵌布粒度非常细小,包裹于硫化物中的银约占50.46%,属于难处理高砷含银硫精矿。采用细磨后化学预处理氰化浸出,银浸出率仍然低于80%;硫精矿经氧化焙烧后,As、S的脱除率都达到90%以上,但银浸出率却较低;对该含银硫精矿添加钠盐焙烧预处理,再采用常规氰化法浸出,银浸出率显著提高,达到85.15%,同时氰化钠耗量降低至2.0 kg/t。     

高压酸浸镍钼矿提取钼和镍

研究一项针对镍钼矿用高压酸浸的方法回收镍和钼的全湿法工艺。采用该工艺避免了传统上艺焙烧镍钼矿（15％～25％s）带来的人量S02和As2O3排放,减小了对环境的污染;与现有的湿法碱浸回收钼工艺相比,本工艺存酸浸过程中回收了儿乎全部的镍和人部分的钼。在氧压环境下,几乎全部的镍和大部分的钼都进入溶液,少部分的钼留在酸浸渣中,睃浸渣进一步用碱（NaOH）浸出。在最佳的实验条件下,97％的镍和96％的钼分别被浸出。     

铟的还原浸出动力学及分步沉淀选择性分离（英文）

开展硫化锌精矿还原浸出高铁锌浸出渣高效浸铟及浸出液中铟选择性分离的研究。结果表明：在固体物料粒度74～105μm、反应温度90℃、浸出时间300 min、硫酸浓度1.4 mol/L的条件下,铟的浸出率达95%以上。采用收缩核模型对还原浸出动力学进行分析,不同条件下的浸出实验结果表明反应受穿过固体产物层的扩散控制,活化能为17.96 k J/mol,相对于硫酸浓度的反应级数为2.41。铁粉置换沉铜过程铜和砷的沉淀率均达99%以上。98%以上的铟从含高亚铁离子浓度的硫酸锌溶液中选择性分离,获得铟含量约为2.4%的富铟渣,经酸浸-萃取-电积工艺流程进一步处理后可得到纯铟。     

硫酸介质中氯化物参与下氧化浸出铜渣过程

研究了镍铳选择性氧化浸出产生的铜渣在硫酸介质中氯离子参与下的氧气氧化浸出,考察和比较了氯化物用量、氧气流量、硫酸浓度、温度等因素对铜和镍浸出的影响,并讨论了其原因,确认了在硫酸介质中加入少量氯化物即可在常压下用氧气有效地浸出主要由辉铜矿或类似的硫化铜矿物组成的铜渣。     

云南某铜尾矿中铜和金的选矿回收试验

云南某铜尾矿主要金属矿物有黄铜矿、黄铁矿、磁黄铁矿等,黄铜矿以原生硫化铜为主,金以裸露金和黄铜矿包裹金为主。为综合回收其中有价铜、金,进行了选矿试验。试样在磨矿细度为-200目占85%的情况下采用两次粗选、第二次粗选后扫选、两次精选、第二次精选后扫选、混精矿再磨至-325目占85%、粗选后扫选精矿再磨至-325目占85%、中矿循序返回流程处理。最终获得铜品位15.51%、回收率68.34%、产率1.41%的铜金精矿,其中的金品位19.93 g/t、回收率54.04%,银品位231.72 g/t、回收率41.89%。     

Technology and mechanism of desilication from roasted diasporic bauxite in atmosphere

The leaching desilication technology of roasted diasporic bauxite in atmosphere by caustic soda solution was investigated. The optimum parameters were: the grinding fineness of the roasted bauxite -0. 076 mm and 80 % -85 %,leaching time 2h, Na2Ok100-150g/L, L/S 4-5, leaching temperature 90-95℃. The desilication rate 55.20% and concentrate A/S (mass ratio of A12O3 to SiO2) 9.90, as good as those obtained at pressure, were obtained respectivdy.Investigation of two-stage leaching shows that it can both improve desilication rate of roasted ore and reduce leaching time.When time of the first stage and the second stage is 30 min and 30 min respectively, desilication rate can reach 59.65 %.X-ray diffraction analysis of the concentrate has proved that desilication procedure is accompanied with the formation of sodium aluminosilicate hydrate. X-ray spectra also show that silica removed during leaching is amorphous silica. SiO2 occurrina as ouartz in raw ore or mullite formed during roasting can not dissolve in alkali solution.     

氧压酸浸低品位富银硫化矿富集提取银和锌

由于含有大量的黄铁矿和白铁矿(它们约占原矿的70%,质量分数),以及闪铅矿一定程度上的氧化,云南澜沧铅矿股份有限公司所产的富银硫化矿难以富集.本文通过对该矿在90-170℃下氧压酸浸,以期连同后面的氰化能提取精矿中的银.通过进行2L高压釜的小型试验,考察了温度、酸度、碘化钠用量、氧分压、氧气流速对银和锌回收率的影响.结果表明,银的回收率取决于银是否进入黄钾铁矾渣,或者与碘化钠反应生成碘化银沉淀.在优化的条件下,银和锌的回收率分别达到71.5%和41.29%.     

富银低品位锰矿和闪锌矿的联合氧压酸浸

进行了硫化锌和含银低品位锰矿联合氧压酸浸的小型试验,以考察影响浸出的各种因素,诸如浸出温度、浸出时间、氧分压和搅拌速度.实验在2L的高压釜内进行.实验结果表明,硫化锌和含银低品位锰矿能相互促进浸出,但这种耦合作用须在一定条件下才会起作用.作者对一些浸出做了探讨.为下一步扩大试验的需要,本文给出了合理的浸出条件:浸出温度, 110℃; 浸出时间,2h; 氧分压, 0.6MPa; 硫酸用量, 1.2倍理论用量.     

Recovery of Mn^2＋, Co^2＋ and Ni^2＋ from manganese nodules by redox leaching and solvent extraction

The recovery of Mn, Co and Ni from deep-sea manganese nodules was conducted by acid oxidative leaching and solvent extraction. The results indicate that pyrrhotite used during leaching can effectively facilitate the leaching out of manganese, cobalt and nickel. The leaching behaviors of Mn, Ni and Co were determined and the influences of temperature, leaching time and sulphuric acid concentration on leaching rate were also investigated. Co and Ni are precipitated from the leaching liquor by adding sodium sulfide into solution with agitation for 2 h at 50 ℃, and the manganese sulphate is obtained by concentrating the resulting solution. By re-dissolving the precipitates of cobalt and nickel, the separation of cobalt and nickel is performed using di（2-ethylhexyl） phosphoric acid （D2EHPA） for impurities elimination with 8 stages at organic-to-aqueous（O/A） volume ratio of 3：5, and 2- ethylhexyl phosphonic acid mono-2-ethylhexyl ester （known as PC88A or P507） for cobalt extraction with 3 stages counter-current operations at O/A volume ratio of 2：3 followed by their scrubbings and strippings, respectively. The final maximum recovery rates for manganese, cobalt and nickel are 85%, 75% and 78%, respectively.     

Vanadium recovery from stone coal through roasting and flotation

A new process for vanadium recovery from stone coal by roasting–flotation was investigated based on the mineralogy. The process comprised four key steps: decarburization, preferential grinding, desliming and flotation. In the decarburization stage, roasting at 550 °C effectively avoided the negative effect of the carbonaceous materials in raw ore and generation of free CaO from calcite decomposition during roasting. Through preferential grinding, the high acid-consuming minerals were enriched in the middle fractions, while mica was enriched in the fine and coarse fractions. Through flotation, the final concentrate can be obtained with V₂O₅ grade of 1.07% and recovery of 83.30%. Moreover, the vanadium leaching rate of the final concentrate increased 13.53% compared to that of the feed. The results reveal that the decarburization by roasting at 550 °C is feasible and has little negative impact on mica flotation, and vanadium recovery from stone coal is conducive to reducing handling quantity, acid consumption and production cost.     

Pressure nitric acid leaching of alkali-pretreated low-grade limonitic laterite

The pressure nitric acid leaching of alkali-pretreated low-grade limonitic laterite, as well as removing impurity Al(III) and preparing intermediate product of nickel/cobalt sulphide from leaching liquor were investigated. After pretreatment, iron exists in the form of amorphous iron oxides, while nickel is adsorbed on the surface of iron oxides in the form of nickel oxide. The preferable pressure leaching conditions are determined as follows: leaching temperature of 458 K, leaching duration of 60 min, initial acidity of nitric acid of 1.90 mol L-1and liquid to solid ratio of 3:1(volume to mass ratio). Under these conditions, the leaching efficiencies of Ni, Co and Al are 95 %, 88 % and 55 %, respectively, and that of Fe is less than 1 %. The loss rates of Ni and Co are 1.8 % and1.5 %, respectively, during the step of removing impurity Al(III). The sulphide precipitation process produces the interim production of nickel/cobalt sulphides, recovering greater than 99 % of Ni and Co in the purified solution.The iron-rich([60 %) pressure leaching residue with low Cr, S can be further reclaimed as the raw materials for iron making.     

EFFECT OF FINE GRINDING ACTIVATION ON LEACHING BEHAVIOUR OF SCHEELITE CONCENTRATE

The influence of mechanical activation on the leaching behaviour of scheelite was studied bymeans of fine grinding in an attritor and subsequently HCl leaching in presence of PO_4~(3-).Results showed that after fine grinding in the attritor,the reaction rate of scheelitewith HCl-Na_3PO_4 solution was remarkably increased,the extraction of W increased fromabout 8 to 99%.The IR spectra and X-ray diffraction analysis indicated that in addition toan enlargement of surface area the fine grinding action had made also changes of fine struc-ture and reactivity of solid surface,hence the leaching process of scheelite can be carried outunder mild leaching conditions.     

Kinetic study on pressure leaching of high iron sphalerite concentrate

The kinetics of pressure leaching high iron sphalerite concentrate was studied.The effects of agitation rate,temperature, oxygen partial pressure,initial acid concentration,particle size,iron content in the concentrate and concentration of Fe2 added into the solution on the leaching rate of zinc were examined.The experiment results indicate that if the agitation rate is greater than 600 r/min,its influence on Zn leaching rate is not substantial.A suitable rise in temperature can facilitate the leaching reaction,and the temperature should be controlled at 140-150℃.The increase trend of Zn leaching rate becomes slow when pressure is greater than 1.2 MPa,so the pressure is controlled at 1.2-1.4 MPa.Under the conditions of this study,Zn leaching rate decreases with a rise in the initial sulfuric acid concentration;and Zn leaching rate increases with a rise of iron content in the concentrate and Fe 2 concentration in the solution.Moreover,the experiment demonstrates that the leaching process follows the surface chemical reaction control kinetic law of“shrinking of unreacted core”.The activation energy for pressure leaching high iron sphalerite concentrate is calculated,and a mathematical model for this pressure leaching is obtained.The model is promising to guide the practical operation of pressure leaching high iron sphalerite concentrate.