Electrogenerative leaching of nickel sulfide concentrate with ferric chloride

In order to utilize the chemical energy in hydrometallurgical process of sulfide minerals reasonably and to simplify the purifying process, the electrogenerative process was applied and a dual cell system was introduced to investigate FeCl3 leaching of nickel sulfide concentrate. Some factors influencing the electrogenerative leaching, such as electrode structure, temperature and solution concentration were studied. The results show that a certain quantity of electrical energy accompanied with the leached products can be acquired in the electrogenerative leaching process.The output current and power increase with the addition of acetylene black to the electrode. Varying the components of electrode just affects the polarization degree of anode. Increasing FeCl3 concentration results in a sharp increase in the output of the leaching cell when c(FeCl3) is less than 0.1mol/L. The optimum value of NaCl concentration for electrogenerative leaching nickel sulfide concentrate with FeCl3 is 3.0 mol/L. Temperature influences electrogeneratire leaching by affecting anodic and cathodic polarization simultaneously. The apparent activation energy is determined to be 34.63 kJ/mol in the range of 298 K to 322 K. The leaching rate of Ni^2 is 29.3% after FeCl3 electrogenerative leaching of nickel sulfide concentrate for 620 min with a filter bag electrode.     

中等嗜热菌浸出高砷铜精矿研究

高砷铜精矿因含砷较高存在砷害问题,限制了其利用.论文针对云南某高砷硫化铜精矿,采用某中等嗜热菌S.P进行浸出,对比研究了精矿粒度、浸出方式、矿浆浓度、浸出时间和菌液初始Fe3+等因素对浸矿过程的影响.在最佳浸矿条件下中等嗜热菌S.P浸矿时Cu,As和Fe浸出率分别为82.39%,78.21%和40.38%.此外,试验表明高浓度的初始Fe3+显著促进铜精矿中铜、砷的浸出,在初始Fe3+浓度为0.08~0.32 mol/L时,铜浸出率为86.34%~97.06%,As浸出率为89.22%~94.13%.浸渣的X射线衍射结果表明中等嗜热菌S.P浸矿过程中生成单质硫和少量砷酸铜.研究为该类矿的生物冶金处理提供了一定的研究基础,对高砷硫化铜精矿资源的开发利用具有重要意义.     

Mechanism of influence of chloride ions on electrogenerative leaching of sulfide minerals

1 INTRODUCTIONElectrogenerative leaching process is a newtechnique in hydrometallurgy .ZHANG et al[1]in-troduced the principle and technique of electrogen-erative process to metallurgy field by the leachingof synthetic Ni3S2with FeCl3.In order to utilizethe chemical energy in leaching process reasonablyand si mplify the purifying process , Wang et al[2 8]studied the electrogenerative leaching of a series ofsulfide minerals through a dual cell system withFeCl3and acidic MnO2as oxidant…     

Precipitation of antimony from the solution of sodium thioantimonite by air oxidation in the presence of catalytic agents

The behavior of antimony oxidation in the solution of sodium thioantimonite was studied in the presence of catalytic agents. The catalytic effects of the respective addition of cupric sulfate, sodium tartrate, potassium permanganate, phenol, 1,2-dihydroxybenzene and their combination on the oxidation of sodium thioantimonite were investigated. A pilot test was carried out. The results show that the respective use of sodium tartrate, cupric sulfate, potassium permanganate, phenol and 1,2-dihydroxybenzene have little catalytic effect on the oxidation of sodium thioantimonite. However there exists obvious catalytic oxidation by the combination of 0.25 g/L 1,2-dihydroxybenzene, 0.5 g/L potassium permanganate and 1.0 g/L phenol. Moreover, high blast intensity, the increase of temperature and NaOH concentration favor the oxidation of antimony. The oxidation process of antimony has such advantages as quick reaction and low operation costs. The results of the pilot test are consistent with those of laboratory experiments.     

Improvement of copper–molybdenum ore beneficiation using a combined flotation and biohydrometallurgy method

The authors present the results of analysis of material composition and experimental investigations of acid and biohydrometallurgical leaching of middlings on grain size, pH level, leaching process duration, temperature and slurry density. The rational parameters of flotation and acid-bacterial leaching of middlings providing an efficient release of valuable components from mineral complexes and recovery to flotation concentrate and leaching solution have been determined. A combined flowsheet and a beneficiation process for bulk flotation middlings of copper–molybdenum ore have been suggested, which include middlings grinding, sulfide minerals flotation, bacterial leaching of sulfide flotation tailings, liquid-phase extraction of dissolved copper and electrolysis of re-extraction eluates. The suggested combined method of cleaning of middlings of copper–molybdenum ores beneficiation provides the total copper recovery increase by 0.8% with a reduction of the cost price of saleable material by 0.5%.     

Molybdenum removal from copper ore concentrate by sodium hypochlorite leaching

The removal of molybdenum from a copper ore concentrate by sodium hypochlorite leaching was investigated. The results show that leaching time, liquid to solid ratio, leaching ternperature, agitation speed, and sodium hypochlorite and sodium hydroxide concentrations all have a significant effect on the removal of molybdenum. The optimum process operating parameters were found to be: time, 4 h: sodium hydroxide concentration, 10%; sodium hypochlorite concentration, 8%; liquid to solid ratio, 10:1; temperature, 50℃; and,agitation speed, 500 r/min. Under these conditions the extraction of molybdenum is greater than 99.9% and the extraction of copper is less than 0.01%. A shrinking particle model could be used to describe the leaching process. The apparent activation energy of the dissolution reaction was found to be approximately 8.8 kJ/mol.     

硫化锌精矿中镓锗在高压氧浸中的浸出行为研究

试验研究了硫化锌精矿中镓、锗在高压氧浸中的浸出行为,考察确定了酸锌的量比,浸出时间,浸出液初始Fe(II)离子浓度,木质素添加量,矿石粒度,温度等因素对各元素浸出率的影响。结果表明,锗与锌的浸出行为较为一致;而镓与铁的浸出行为密切相关,使铁充分浸出并保持浸出液具有足够的酸度,防止高温下Fe(II)离子的氧化水解,是提高金属镓浸出率的关键。对富含镓、锗等稀散金属的硫化锌精矿,采用两段逆流流程浸出,既可保证镓、锗浸出率在80%以上,又能使得一段浸出后液酸度(H2SO4)<20 g/L,从而有利于其后续的中和。     

Techniques of copper recovery from Mexican copper oxide ore

Mexican copper ore is a mixed ore containing mainly copper oxide and some copper sulfide that responds well to flotation. The joint techniques of flotation and leaching were studied. The results indicate that an ore containing 19.01% copper could be obtained at a recovery ratio of 35.02% by using sodium sulfide and butyl xanthate flotation. Over 83.33% of the copper oxide can be recovered from the tailings by leaching in suitable conditions, such as 1 h stirring at a temperature around 25 ℃ with a mixing speed of 500 r/min, an H2SO4 concentration of 1.0 mol/L and a mass ratio of the ore-slurry-liquid to solid (mL/mS) of 3. The overall yield of refined ore after flotation and leaching is over 89.18% of the copper, which is much better than sole flotation or leaching. A copper product containing more than 99.9% copper was obtained by using the process: flotation-agitation leaching-solvent extraction-electro-winning.     

高砷铜精矿浸出液制备砷酸铜及热力学研究

高砷铜精矿浸出液中因含砷较高存在砷害问题,砷回收利用具有重要的意义。研究针对某高砷硫化铜精矿除杂后的浸出液,探讨了以砷酸铜形式综合利用砷的热力学及工艺参数。绘制了Cu-As-H2O系电位-pH图,进行了砷酸铜制备工艺过程的热力学分析,对净化后的某高砷硫化铜精矿浸出液,以氨水为中和剂,当pH=5.0~8.0,温度80~90℃时,制得了CuAs2O4、C4H6As6Cu4O16、Cu5As2O10.5H2O、Cu5As4O15.9H2O、Cu2AsO4OH.3H2O及Cu2AsO4OH 6种不同结构的砷酸铜,沉砷后溶液中砷含量为9.11~35.82 mg/L。     

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.     

Flotation technology of refractory low-grade molybdenum ore

Because of the low grade, high oxidation rate and the accumulation of little associated metal sulfide ore in the molybdenum concentrate during flotation, the Qingyang molybdenum ore is difficult to beneficiate. The experimental studies of grinding fineness, the amount of roughing modifier, depressant and collector were completed. In the cleaning process, the contrast experiments of one regrinding, the regrinding and scrubbing, two-stage regrinding was carried. The result shows that the grade of molybdenum ore concentrate is 45.31%, the recovery is 65.98% and the rich ore ratio reaches 20.59% by the regrinding and scrubbing seven cleaning, the regrinding of concentrations from middling of molybdenum-sulfur separation. The regularly-concentrated material from the apparatus was as the middling products. Hence, ideal beneficiation index can be obtained with a rational mineral processing, which offers new beneficiating technology for the refractory low-grade molybdenum ore in China.     

低品位铜锌混合矿加压浸出研究

研究了低品位多金属铜锌混合矿的浸出条件．探讨了氧分压、酸度、温度、反应时间、添加剂等因素对铜锌浸出率的影响．结果表明，铜和锌的浸出率可达98％和99％．提高了矿产资源的利用率，利用直接加压浸出的方法取代传统的焙烧——浸出工艺。从生产源头上消涂了烟气污染．     

Molybdenite alkali fusion and leaching: reactions and mechanism

The production of MoO_3 from Sarcheshmeh molybdenite concentrate via a pyro-hydrometallurgical process was studied.The molybdenite concentrate and sodium carbonate were premixed and fused under air atmosphere.Then the fused products were leached in water and the dissolved molybdenum was recovered as ammonium molybdate.The ammonium molybdate was then calcined to produce mo-lybdic oxide.At the fusion stage,the effect of the mass ratio of carbonate to sulfide on the reaction products and the solubility of t...     

难处理高结合率氧化铜矿选冶联合工艺研究

采用选冶联合工艺对含铜1.53％、氧化率47.06％、结合率21.57％的高结合率氧化铜矿进行回收.原矿的砂光片分析结果表明,矿石中大部分铜矿物嵌布粒度极细,多呈星点状和不均匀浸染状分布,与硅、钙、镁、铝等脉石共生严重,导致浮游性较差.针对该矿石的特点,研究了工艺参数及流程结构对指标的影响,确定了“三次粗选—粗精矿再磨—三次精选”的硫化浮选工艺流程,获得了含铜品位为23.43％、回收率为53.72％的铜精矿.对尾矿的形貌及矿物组成表征发现：铜矿物呈细粒浸染状或被硅酸盐矿物包裹,导致这部分铜损失在尾矿中.在最佳的酸浸工艺条件下,对浮选尾矿进行酸浸试验,获得了相对原矿的浸出率为33.21％的试验指标;铜综合回收率为86.93％.     

Simultaneous electrogenerative leaching of chaicopyrite concentrate and MnO2

In order to enhance the electrogenerative leaching rate of chalcopyrite concentrate reasonably, the principle of generative process was applied to simultaneous leaching of chalcopyrite concentrate and MnO2. The results show that Cu^2＋ and Mn^2＋ in addition to electrical energy could be acquired in the simultaneous electrogenerative leaching process. The leaching cell has the open circuit potential of about 1.0 V and gains quantity of electricity of about 700 C. The optimum leaching rates of Cu^2＋ and Mn^2＋ are 23.10% and 22.1%, respectively after electrogenera- tive leaching for about 10 h under the present conditions.     

锑矿废渣模拟酸雨淋溶重金属Sb、As、Hg淋溶释放规律

为了全面系统地研究锑矿废渣中重金属的淋溶释放规律,以湖南省冷水江锡矿山锑矿的锑矿废渣为研究对象,通过对锑矿废渣的消解试验与半动态淋溶试验,测定锑矿废渣与淋溶液中重金属的含量,分析锑矿废渣淋溶液中Sb、As、Hg淋溶释放规律。结果表明:锑矿废渣中重金属含量较高,除Pb以外,其余重金属均超《土壤环境质量农用地土壤风险管控标准》(GB 15618—2018)的风险筛选值(pH≤5.5),重金属含量均超湖南省土壤背景值;锑矿冶炼渣在模拟酸雨淋溶下重金属析出浓度在第2～3天达到峰值,锑矿废石与锑矿尾砂淋溶一天后就达到峰值,锑矿废渣重金属累计析出量均为锑矿冶炼渣>锑矿尾砂>锑矿废石;锑矿废渣重金属的累计淋出量与模拟酸雨累计淋出量之间呈二次函数关系。     

Investigation into the flotation response of refractory molybdenum ore to depressant mixtures:A case study

In this paper, bulk flotation followed by separation was investigated to concentrate purified molybdenite product from Jinduicheng molybdenum ores(Shanxi province, China). The bench scale tests mainly focussed on separation of molybdenite from other sulfide minerals using the new type of depressants.The effect of each single depressant, including organic depressant-modified dextrin(MD), P-Nokes reagent(PN) and sodium trithiocarbonate(ST), and their mixtures on galena, chalcopyrite and other sulfide ores, was examined in turn by changing the concentrations used in cleaner flotation tests. Closed circuit experiments were carried out under the optimal condition and satisfying recovery and grade of molybdenite concentrate could be achieved(86.294% and 53.157%, respectively). A potential reagent regime was developed, with more environmental friendly and more economical advantages due to the introduction of modified dextrin.     

Extraction of vanadium from stone coal by modified salt-roasting process

In order to reduce the pollution of Cl₂ and HCl released during extracting vanadium from stone coal by sodium chloride roasting, a modified salt-roasting process was proposed by adding calcined lime in roasting process followed by H₂SO₄ leaching. The effects of parameters including roasting temperature, roasting time, addition mass ratio of NaCl, calcined lime upon leaching rate of vanadium and curing rate of chlorine were investigated, and the effects of leaching time and leaching temperature on leaching rate of vanadium were also studied. The results show that the vanadium leaching rate and the curing rate of chlorine are 67.3% and 51.5% (mass fraction), respectively, at roasting temperature of 750 °C, roasting time of 4 h, 15% sodium chloride and 8% (mass fraction) calcined lime, leaching temperature of room temperature, and leaching time of 4 h.     

Conversion of coal gangue into alumina,tobermorite and TiO_2-rich material

A large amount of coal gangue from coal mining and processing is regarded as waste and usually stockpiled directly. In order to recycle the valuable elements from the coal gangue, an integrated process is proposed. The process consists of three steps: 1concentrating alumina from the coal gangue via activation roasting followed by alkali leaching of Si O2 which produces alumina concentrate for alumina extraction by the Bayer process; 2) synthesizing tobermorite whiskers from the filtrated alkali liquo containing silicate via a hydrothermal method and reusing excess caustic liquor; and 3) enriching titanium component from the Baye process residue by sulfuric acid leaching. Alumina concentrate with 69.5% Al_2O_3 and mass ratio of alumina to silica(A/S) of 5.9pure 1.1 nm tobermorite whisker and TiO_2-rich material containing 33% TiO_2 are produced, respectively, with the optimal parameters Besides, the actual alumina digestion ratio of alumina concentrate reaches 80.4% at 270 oC for 40 min in the Bayer process.     

复杂铜铅混合精矿氧压浸出综合回收工艺

呷村铜铅混合精矿中铜、铅矿物主要为黝铜矿和方铅矿,还含有较高的锌、银、砷和锑.本试验针对该矿采用一段氧压浸出综合回收工艺进行处理,通过条件优化实验确定了氧压浸出的操作条件.扩大验证实验表明Cu、Zn的浸出率分别高达98.89%、94.92%,Pb、Ag转化为矾类和硫化物形式留在浸出渣中,铜锌与铅银分离彻底.浸出液中的铜、锌分别通过萃取、电积进行回收.浸出渣中的铅、银通过碳酸盐转化-硅氟酸浸铅-硫脲浸银进行回收.铜萃取率,铅、银浸出率分别为96%、94%和93%.