A new extraction process of carbonaceous refractory gold concentrate

A new hydrometallurgical process for a carbonaceous refractory gold concentrate at ambient temperature and pressure was presented, including grinding-leaching, intensified alkaline leaching（IAL）, thiosulfate leaching and cementation by zinc powder. The experimental results show that the grinding-leaching and intensified alkaline leaching process result in the selective oxidation of arsenopyrite and pyrite. The oxidation ratio of As is 96.6%, and 46.7 % for S. The total consumption of NaOH in alkaline leaching is only 28 % of that theoretically calculated under the conditions of full oxidization for the same amount of arsenopyrite and pyrite transforming into arsenates and sulfates, and 83.6% of gold is synchro-dissoluted by thiosulfate self-generated during pretreatment. Since the carbonaceous matter in concentrate possesses a strong capability of preg robbing, the cyanidation process is not suitable for the extraction of gold after pretreatment. However, the gold leaching rate by thiosulfate leaching for 24 h is increased to 91.7% from 0 - 3.2% by ultra-fine grinding without the pretreatment. The recovery of gold by zinc cementation gets to 99.6%. Due to the thiosulfate self-generated during alkaline leaching, the reagent addition in thiosulfate leaching afterwards is lower than the normal one.     

Thermodynamic analysis and process optimization of zinc and lead recovery from copper smelting slag with chlorination roasting

An efficient chlorination roasting process for recovering zinc (Zn) and lead (Pb) from copper smelting slag was proposed. Thermodynamic models were established, illustrating that Zn and Pb in copper smelting slag can be efficiently recycled during the chlorination roasting process. By decreasing the partial pressure of the gaseous products, chlorination was promoted. The Box−Behnken design was applied to assessing the interactive effects of the process variables and optimizing the chlorination roasting process. CaCl₂ dosage and roasting temperature and time were used as variables, and metal recovery efficiencies were used as responses. When the roasting temperature was 1172 °C with a CaCl₂ addition amount of 30 wt.% and a roasting time of 100 min, the predicted optimal recovery efficiencies of Zn and Pb were 87.85% and 99.26%, respectively, and the results were validated by experiments under the same conditions. The residual Zn- and Pb-containing phases in the roasting slags were ZnFe₂O₄, Zn₂SiO₄, and PbS.     

Co-extraction of valuable metals and kinetics analysis in chlorination process of low-grade nickel-copper sulfide ore

To efficiently co-extract Ni and Cu from low-grade nickel-copper sulfide ore, chlorination roasting with NH₄Cl followed by a water leaching process was investigated. The results show that 98.4% Ni and 98.5% Cu can be synchronously extracted when the ore particle size is 75-80 μm, the roasting time is 2 h, the mass ratio of NH₄Cl to ore is 1.6:1 and the roasting temperature is 550 °C. The evolution behavior of various minerals was elucidated using X-ray diffraction (XRD) coupled with scanning electron microscopy (SEM). The kinetics of the chlorination process based on the differential thermal and thermogravimetric analysis (DTA-TG) data was analyzed by Kissinger method and Flynn-Wall-Ozawa (FWO) method. The chlorination process of low-grade nickel-copper sulfide ore mainly contains two stages: the decomposition of NH₄Cl and the chlorination of ore. The maximum apparent activation energies (E_a) at two stages are determined to be 114.8 and 144.6 kJ/mol, respectively. The condensed product of exhaust gas is determined to be ammonium chloride, which can be recycled as the reactant again, making the process economic and clean.     

响应曲面法优化NaClO浸出贵州难浸金矿中的金

采用响应曲面法(RSM)优化NaClO浸出贵州难浸金矿中的金,探究了Na ClO浓度、温度、液固比、p H值4个因素之间的交互作用及其对金浸金率的影响,并进行了参数优化。RSM分析表明,各因素对金浸出率的显著程度为：NaClO浓度>液固比>p H>温度,NaClO浓度和p H之间的交互作用较为显著;得到较优的浸金工艺条件为：反应时间3 h,Na ClO浓度0.9 mol/L,初始pH=13.4,液固比8.2,温度27℃,在此条件下金的浸出率为93.4%。     

锌浸渣还原焙烧-磁选回收铁

在查明锌浸渣工艺矿物学的基础上,采用还原焙烧将铁酸锌分解为氧化锌和磁性氧化铁,再通过磁选的方法回收铁,达到锌、铁分离的目的。实验考查了焙烧温度、焙烧时间、还原剂用量对铁酸锌分解率、铁回收率和铁品位的影响。结果表明:在焙烧温度为950℃、焙烧时间为1 h及还原剂添加量为10%和5%的条件下,铁酸锌分解率达到72.05%,铁回收率可达到91.79%,精矿中铁的品位为50%左右。焙烧及磁选过程中颗粒的团聚包裹是铁精矿品位不高的主要原因。     

难处理金矿预处理及金回收技术进展

随着优质金矿不断被开发消耗,难处理金矿占比不断地提高,从难处理金矿中回收金是金产业未来发展的必然趋势。本文简要分析了难处理金矿浸出困难的原因,介绍了焙烧氧化法、热压氧化法、生物氧化法、机械活化法、微波法5种预处理技术和氰化法、硫脲法、硫代硫酸盐法、卤素法、火法5种金回收技术的研究进展,并比较了5种预处理技术和4种湿法金回收技术的优势与不足。在此基础上对难处理金矿预处理和金回收技术的前景进行了展望。     

难冶金精矿焙砂和烟尘氰化浸金的差异（英文）

采用对比方法研究难冶金精矿焙砂和烟尘氰化浸金的差异。结果表明,直接氰化时焙砂和烟尘中金的浸出率分别为85.31%和54.30%。砷、碳含量及其存在形式是导致两者金浸出率差异的主要原因。对于NaOH预处理后氰化浸金,焙砂和烟尘中金的最大浸出率分别为87.70%和58.60%。有害元素的脱除、碱浸预处理过程中金的损失及铁氧化物的阻碍共同决定碱浸预处理后焙砂和烟尘中金的浸出率。经H₂SO₄预处理后,焙砂和烟尘中金的最大浸出率分别达到94.96%和80.40%。碳质物的影响是焙砂和烟尘中金浸出率差异的主要原因。基于这些差异,提出两种适宜工艺,焙砂和烟尘中金的浸出率分别达到94.91%和91.90%。     

Bioleaching of refractory gold ore （ Ⅲ ）——Fluid leaching Jinya refractory gold concentrate by Thiobacillus ferrooxidans

A novel fluidized-bed reactor was designed and installed for bloleaching in a semi-continuous way, by which a process for biuleaching-cyanidation of Jinya refractory gold arsenical concentrate was studied. The arsenic extraction rate reaches 82.5 % after 4-day batch biooxidation of the concentrate under the optimized condition of pH 2.0, ferric ion concentration 6.Sg/L and pulp concentration 10%. And leached rate of gold in the following cyanida.tion is over 90%. The parameters of three series fluidized-bed reactors exhibit stability during the semi-continuous bioleaching of the concentrate.Armmic in the concentrate can be got rid of 91% after 6-day leaching. Even after 4 days, 82% of arsenic extraction rate was still obtained. The recovery rates of gold are 92 % and 87.5 % respectively in cyaniding the above bioleached residues.The results will provide a base for further commercial production of gold development.     

碳质金矿的碳质物及生物氧化预处理研究现状

碳质金矿是一种重要的难处理金矿。研究发现,其碳质物主要包括元素碳、有机酸和烃类物质。在氰化浸金过程中碳质物可通过类活性炭的吸附方式将已溶解的金劫走。目前,已有的预处理方法主要有高温焙烧法、生物氧化法、化学氧化法、竞争吸附法、覆盖抑制法、微波加热法。生物氧化法因具有条件温和、流程简单、能耗低、环境友好等优点得以迅速发展。与生物氧化预处理有关的微生物主要有氧化亚铁硫杆菌、氧化硫硫杆菌、氧化亚铁钩端螺旋菌等化能无机自养菌。有关黄孢原毛平革菌、假单胞菌、多毛链霉菌在碳质物降解和钝化方面的研究也已展开。最后,分析了该技术存在的问题,并对其应用前景进行了展望。     

氯化焙烧-水浸法从锂云母矿提锂(英文)

采用氯化焙烧-水浸法处理锂云母矿,并对氯化处理温度、时间、氯化剂的类型及用量进行研究。条件优化实验表明,在锂云母、氯化钠、氯化钠的质量比为 1:0.6:0.4,氯化处理温度为 880 °C,氯化处理时间为 30 min时,锂的提取率可达 92.86%,钾、铷、铯的提取率分别为 88.49%、93.60%和 93.01%,。采用 XRD 对锂云母原矿及焙烧后物料的物相进行分析。XRD 结果分析表明,当将锂云母和混合氯化剂一起焙烧(氯化钙及氯化钠)时,所得物相为 SiO2、CaF2、KCl、CaSiO3、CaAl2Si2O8、NaCl 和 NaAlSi3O8。     

从镍钼矿中提取镍钼的工艺

针对现行镍钼矿处理工艺存在的钼镍需要分别提取的缺陷,提出镍钼矿加钙氧化焙烧-低温硫酸化焙烧-水浸提取镍钼的新工艺。以贵州遵义镍钼矿为原料,对CaO加入量、氧化焙烧温度、氧化焙烧时间、硫酸加入量、硫酸化焙烧温度、硫酸化焙烧时间以及焙砂水浸工艺参数对镍钼浸出率的影响进行研究。结果表明:在最佳工艺条件下,钼的浸出率为97.33%,镍的浸出率为93.16%,且最佳工艺参数为100 g镍钼矿加入35 g CaO,700℃氧化焙烧2 h,得到的焙砂加入70 mL浓硫酸,再经250℃硫酸化焙烧2 h;硫酸化焙烧得到的焙砂按液固比2:1加水搅拌,经98℃浸出2 h。加入CaO不仅能有效减少镍钼矿氧化焙烧烟气对环境造成的污染,而且能显著提高镍的浸出率。     

Recovery of valuable metals from zinc leaching residue by sulfate roasting and water leaching

Zinc leaching residue (ZLR), produced from traditional zinc hydrometallurgy process, is not only a hazardous waste but also a potential valuable solid. The combination of sulfate roasting and water leaching was employed to recover the valuable metals from ZLR. The ZLR was initially roasted with ferric sulfate at 640 °C for 1 h with ferric sulfate/zinc ferrite mole ratio of 1.2. In this process, the valuable metals were efficiently transformed into water soluble sulfate, while iron remains as ferric oxide. Thereafter, water leaching was conducted to extract the valuable metals sulfate for recovery. The recovery rates of zinc, manganese, copper, cadmium and iron were 92.4%, 93.3%, 99.3%, 91.4% and 1.1%, respectively. A leaching toxicity test for ZLR was performed after water leaching. The results indicated that the final residue was effectively detoxified and all of the heavy metal leaching concentrations were under the allowable limit.     

高砷难处理金精矿细菌氧化-氰化提金

通过在高砷金精矿中配入不同比例的低砷碳酸盐型金精矿,使其所含硫、砷及铁等主要矿物成分含量发生变化,研究给矿中铁砷摩尔比对难处理高砷金精矿细菌氧化一氰化浸出效果的影响.结果表明:含砷金精矿中铁砷摩尔比直接影响细菌预氧化的效果,同时也影响细菌的活性和溶液中铁砷摩尔比的变化,给矿中铁砷摩尔比越高,溶液中的铁砷摩尔比也越高,且随着给矿中铁砷摩尔比的增加,溶液中铁砷摩尔比的变化幅度加大,给矿中铁砷摩尔比介于4.6~2之间,有利于细菌预氧化和氰化浸出,铁、砷氧化率分别由6.14%和7.38%提高到89.90%和93.60%,金、银浸出率分别由64.18%和35.93%提高到97.78%和88.83%,较好地改善细菌氧化效果,稳定和优化细菌预氧化过程.     

Mineralogical characterization and pretreatment for antimony extraction by ozone of antimony-bearing refractory gold concentrates

The mineralogical characterization of antimony-bearing refractory gold concentrates and the antimony extraction by ozone in HCl solution were investigated. The mineralogical study shows that there exist stibnite(Sb₂S₃), arsenopyrite(FeAsS), pyrite(FeS₂) and quartz in the concentrates, and the gold is mainly (67.42%) encapsulated in sulfides. The antimony extraction by ozone in hydrochloric acid was employed and the influences of temperature, liquid/solid ratio, HCl concentration and stirring speed on the extraction of antimony were investigated. High antimony extraction (93.75%) is achieved under the optimized conditions. After the pretreatment by ozone, the antimony is recovered efficiently and the gold is enriched in the leaching residue.     

微波焙烧预处理难浸含金硫精矿

对难浸含金硫精矿进行微波焙烧,考察微波功率、矿量(即样品质量)和焙烧时间对样品质量损失率和浸出率的影响。结果表明:当微波功率为16 k W、焙烧时间为50 min、矿量为900 g时,样品质量损失率可达28.12%,浸出率可达71.56%,较原矿直接碘化浸出率(9.82%)有了大幅提高。利用XRD和SEM技术分析焙烧前后样品的成分和表面形貌,微波焙烧后的样品较原矿更为松散、多孔,更利于浸出。     

Fundamental study on utilization of tin and zinc-bearing iron concentrate by selective chlorination

The feasibility and technologies of comprehensive recovery of tin, zinc, arsenic and iron from the complex iron ores by selective chlorination roasting were studied by thermodynamic analysis and roasting experiments. Investigation shows that the product pellets with the compression strength of 2 625 N/P, the tumble index of 97. 26%, the abrasion index of 1. 35%, tin, arsenic and zinc residue of 0. 043%, 0. 046% and 0.058% respectively can be achieved if bailing a concentrate containing 0.39 % tin, 0.40% arsenic and 0.28% with addition of 8% coke breeze and 0. 5% CaCl2 and roasting the pellets at 1 060 - 1 080℃ for 40 min. The volatilization of tin, arsenic and zinc is 91.75 %, 93.42 % and 81.12 % respectively. The performances of the product pellets are able to meet the requirements of blast furnace ironmaking.     

Selective extraction of zinc from sulfate leach solution of zinc ore

1　INTRODUCTIONSolventextractionwasregardedasahighlyeffi cienttechniqueofseparationandpurification .Ithasbeenwidelyappliedinmetalextractionandrecoveryduringhydrometallurgicalprocessandenvironmentalprotection[14 ] .Solventextractionhasbeencommer ciallyusedintheextractionofuranium ,rarenoblemetals ,copper ,cobaltandnickel,whichwastypical lypresentedbythetechniqueofleach solventextrac tion electrowinning .Presently ,thecopperproductionfromsolventextractionhasbeenaccountedformorethan 2 0 %oft…     

Extraction of manganese and zinc from their compound ore by reductive acid leaching

Comprehensive utilization of low grade manganese–zinc compound ore containing lead and silver with a method of reductive acid leaching was studied. According to the ?–pH diagram of Mn–Zn–H₂O system, Mn and Zn can be leached simultaneously in the pH range of –2 to 5.61. The results showed that both hydrogen peroxide and sucrose were effective reductants which could intensify the simultaneous leaching of Mn and Zn into leachate as well as enrich Pb and Ag in the residue. 95.88% of Mn and 99.23% of Zn were extracted when the compound ore was leached with hydrogen peroxide in sulfuric acid media, meanwhile the contents of Pb and Ag in the residue were enriched to 13.21% and 489.36 g/t, respectively. When sucrose was used as the reductant, the leaching efficiencies of Mn and Zn separately achieved 98.26% and 99.62%, and contents of Pb and Ag in the residue were as high as 13.92% and 517.87 g/t, respectively.     

循环流态化焙烧加压浸出从极难浸石煤中提取钒

以广西某极难浸石煤钒矿为研究对象,研究循环流态化焙烧试样在加压浸出条件下的钒浸出率。结果表明:在相同酸浸条件下,循环流态化空白焙烧试样的钒浸出率高于钠化焙烧的钒浸出率。系统的焙烧浸出工艺对比研究表明:该石煤钒矿只有在循环流态化焙烧并加压高浓度酸浸作用下才能获得最高的钒浸出率,应属于极难浸石煤钒矿。在V(H2SO4):V(HF)=1:1和MnO2添加量(质量分数)为3%的条件下,循环流态化空白焙烧矿的最佳酸浸条件为液固比1:1、浸出温度150℃、浸出时间6 h,钒浸出率可达98.11%。同时,研究循环流态化空白焙烧矿加压浸出的动力学模型、浸出控制步骤及表观活化能。循环流态化空白焙烧能避免钠化焙烧产生的Cl2及HCl等有害气体的排放问题。从焙烧反应设备的创新应用着手,探索试验工艺条件,为极难浸石煤钒矿的工业化利用提供参考和依据。