H2SO4分解铁-FeCl3浸出焙烧金精矿中金

采用硫酸分解焙烧金精矿,金从黄铁矿中解离的同时金得到了富集,可采用氯化铁溶液非氰浸出金。研究了硫酸浓度及过量系数、分解温度对铁分解率的影响,优化工艺条件为,焙烧温度180 ℃,反应时间90 min,硫酸过量系数1.2,在此条件下,铁分解率为92.14%,金含量从原来的51.7 g/t提高到106.1 g/t;研究了反应温度、液固比对氯化铁溶液浸出硫酸浸出渣中金的影响,优化浸出条件为,液固比1.5,80 ℃浸出90 min,在此条件下,金浸出率96.8%。     

A sulfuric-acid process with near-zero SO2 gas emissions

A sulfuric-acid process has been developed that is able to handle low and variable SO₂ concentrations with practically zero SO₂ emissions (less than 3 ppm). The plant comprises two stages—a single-bed converter contact plant and a modified tower plant. Acids of 95–98% and/or oleum with up to 32% free SO₃ can be produced in the first stage. Off-gas of the first stage is piped to the second stage, where the SO₂ is converted to near nontracability while producing 76% strong acid. It is then returned to the contact plant to produce stronger acid or oleum. This process does not generate any additional disposable waste. For more information, contact T. Keilpart, Process Engineering Associates, 707 Redway Lane, Houston, Texas, 77062; (281) 488-6633; fax (281) 488-2456; e-mail PEA@net1.net.     

高硅硫化锌精矿氧化焙烧中硅酸锌生成反应的动力学

研究了高硅硫化锌精矿氧化焙烧过程中硅酸锌生成反应的动力学。确定了温度、粒度对硅酸锌反应速率的影响,结果指出：硅酸锌生成反应的动力学符合收缩核西式,其过程为固膜扩散控制。测定了各反应条件下的反应速率常数并测得其活化能为406KJ/mol。提出了硅酸锌生成反应的总动力学方程。限制硅酸锌反应速率的有效方法是适当提高精矿粒度和降低焙烧温度至860℃左右。     

Transformation of chalcopyrite in the roasting process of copper concentrate in fluidized bed reactor

This work presents the results of investigation process of copper concentrate roasting in fluidized bed reactor with the aim of studying the transformations of copper concentrate minerals as well as to check the accordance with theoretical predictions. The roasted samples were examined using chemical analysis, x-ray diffraction, and mineral microscopy.     

Reaction mechanisms of low-grade molybdenum concentrate during calcification roasting process

The effects of Ca-based additives on roasting properties of low-grade molybdenum concentrate were studied. The results show that calcium-based additives can react with molybdenum concentrate to form CaSO₄ and CaMoO₄. The initial oxidation temperature of MoS₂ is 450 °C, while the formation of CaMoO₄ and CaSO₄ occurs above 500 °C. The whole calcification reactions are nearly completed between 600 and 650 °C. However, raising the temperature further helps for the formation of CaMoO₄ but is disadvantageous to sulfur fixing rate and molybdenum retention rate. Calcification efficiency of Ca-based additives follows the order: Ca(OH)₂>CaO>CaCO₃. With increasing the dosage of Ca(OH)₂, the molybdenum retention rate and sulfur-fixing rate rise, but excessive dosages would consume more acid during leaching process. The appropriate mass ratio of Ca(OH)₂ to molybdenum concentrate is 1:1. When roasted at 650 °C for 90 min, the molybdenum retention rate and the sulfur-fixing rate of low-grade molybdenum concentrate reach 100% and 92.92%, respectively, and the dissolution rate of molybdenum achieves 99.12% with calcines being leached by sulphuric acid.     

Recovery of RE from Baotou rare earth concentrate with chlorination roasting

1　INTRODUCTIONBaotourareearthdepositisthelargestoneintheworld .Theconcentrateisthemixtureofbastnaesite(REFCO3)andmonazite (REPO4 )andtheratioofbastnaesitetomonaziteisgenerally 7∶3.ThemainprocessestorecoverREfromtheconcentrateincludetheroastingwithsulfuricacid ,alkaliheatingprocessandhightemperaturechlorinationetc[1] .Thecon ventionalprocessesproducewasteacid ,alkaliordis chargegastopollutetheenvironment ,thereforethemetallurgiststrytoseekthegreenchemistryprocesswithlowcost,simplepro…     

Vanadium recovery from Na2SO4-added V-Ti magnetite concentrate via grate-kiln process

Pilot experiments on Na₂SO₄-roasting of V-Ti magnetite (VTM) concentrate pellets were performed to activate vanadium with simulated grate-kiln machines. Combined with thermodynamic calculation, X-ray diffraction (XRD), scanning electron microscopy with energy-dispersive X-ray spectroscopy (SEM-EDS), and inductively coupled plasma-atomic emission spectrometry (ICP-AES) analyses, process parameters of pellet making were optimized, and the enhancement of vanadium oxidation and conversion was realized. The results show that, in balling stage, adding Na₂SO₄ can promote the thermal decrepitation resistance of green pellets; in the drying stage, higher wind temperature can improve the dispersion of sodium salt in pellets; in preheating stage, using O₂-rich hot gas can promote more thorough oxidation of vanadium; in roasting stage, the effect of vanadium conversion to sodium vanadate is better when the liquid amount is 15.5%-22.5%. Based on the above measures, the vanadium conversion rate achieved in the roasted pellets is 85.6%, proving the feasibility of vanadium recovery using an existing pellet production line.     

Simultaneous extraction of gold and zinc from refractory carbonaceous gold ore by chlorination roasting process

A novel process based on chlorination roasting was proposed to simultaneously recover gold and zinc from refractory carbonaceous gold ore by using NaCl as chlorination agent. The effects of roasting temperature, roasting time and NaCl content on the volatilization rates of gold and zinc were investigated. The reaction mechanism and the phase transition process were also analyzed by means of SEM, EDS and XRD. The results demonstrated that under the optimal conditions of NaCl content of 10%, roasting temperature of 800 °C, roasting time of 4 h and gas flow rate of 1 L/min, the rates of gold and zinc were 92% and 92.56%, respectively. During low-temperature chlorination roasting stage, a certain content of sulfur was beneficial to the chlorination reactions of gold and zinc; and during high-temperature chlorination roasting stage, the crystal structure of vanadium-bearing mica was destroyed, and the vanadium-containing oxides were beneficial to the chlorinating volatilization of gold and zinc. Eventually, the chlorinated volatiles of gold and zinc could be recovered by alkaline solution.     

Recovering gold from copper concentrate via the HydroCopper™ process

HydroCopper? technology comprises a chloride-leaching method for copper sulfide concentrates and copper production up to semi-products. As compared with the commonly used sulfate solutions, brine solutions offer aggressiveness and stability of the copper(I) ion and, consequently, a lower energy consumption in leaching. Copper(II) ions and oxygen are used as oxidants. Iron reports to the leaching residue as oxide and sulfur as elemental sulfur. Gold is dissolved and recovered in the third stage of the counter-current leaching when the redox potential reaches higher levels.     

Controlling limiting length of tunnels on Al foil electroetched in HCl-H2SO4 solution

The limiting length of tunnels, llim, of Al foil electro-etched in HCl-H2SO4 solution and the corresponding anodic polarization curves in the same solution were measured. It is found that there is a dependence of llim on the potential difference, △φ, between the pitting potential, φpit, and the corrosion potential, φcorr, of Al foil, when the temperature and H2SO4 concentration of HCl-H2SO4 electrolyte are changed. The dynamic equation on the tunnel growing and the linear equation between llim and △φ were deduced by analyzing the relationship among the over-potential on Al foil surface, the transport over-potential in tunnel solution and the over-potential at tunnel tip during the electro-etching. The results show that the growing velocity of tunnels decreases with their extending in length and the changing trend of llim can be judged by measuring △φ, which supplies a convenient access to explore new kinds of etchants.     

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.     

Effects of SO2 and SO3 on the Na2SO4 induced corrosion of nickel

The effects of SO₂ and SO₃ in the environment on the hot corrosion behavior of Ni in the temperature range 750–950°C has been studied. Below the melting point of Na₂SO₄ (884°C), rapid corrosion takes place by formation of a Na₂SO₄-NiSO₄ melt, which can penetrate the porous oxide scale and give rise to sulfide information by coming in contact with the metal. The distribution of the sulfides depends on the SO₂ level in the ambient gas. Continued corrosion occurs by a sulfidation-oxidation mechanism. At temperatures above the melting point of Na₂SO₄, accelerated degradation occurs via dissolution of the surface scale, followed by reprecipitation of the oxide in a nonprotective form.Deceased     

Dissolution theory of gold in alkaline thiourea solution(Ⅲ) ——Thermodynamics on dissolution of gold in alkaline thiourea solution containing Na _2SO_3

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Microbial oxidation of refractory gold sulfide concentrate by a native consortium

A defined mesophilic consortium including an iron oxidizing bacterium and a sulfur oxidizing bacterium was constructed to evaluate its ability for bioleaching a flotation concentrate from Andacollo mine in Neuquén, Argentina. Experiments were performed in shake flasks with a pulp density of 10% (w/v), using a basal salt medium containing ferrous iron at pH 1.8. The leaching solutions were analyzed for pH, redox potential (using specifics electrodes), ferrous iron (by UV-Vis spectrophotometry) and metal concentrations (by atomic absorption spectroscopy). The results showed that the consortium was able to reduce the refractory behavior of the concentrate, allowing 91.6% of gold recovery; at the same time, high dissolution of copper and zinc was reached. These dissolutions followed a shrinking core kinetic model. According to this model, the copper solubilization was controlled by diffusion through a product layer (mainly jarosite), while zinc dissolution did not show a defined control step. This designed consortium, composed of bacterial strains with specific physiological abilities, could be useful not only to optimize gold recovery but also to decrease the leachates metallic charge, which would be an environmental advantage.     

锌粉置换回收锑精矿碱浸液中的金

锑精矿碱浸液是湿法处理锑精矿产生的中间产物含有少量的金,必须回收。以山东某公司锑精矿碱浸液为原料,经多种提取剂对比,采用锌粉置换回收其中的金。考察了锌粉用量、料液pH、浸出温度、时间的影响,得到的最佳工艺条件为:锌粉用量4.0 kg/m³,溶液pH=11,反应时间3 h,反应温度80 ℃,多批次试验金回收率88%。工艺过程简单,经济性可行,有效解决了资源浪费问题,为锑精矿碱浸液回收金提供了一条新的工艺。     

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.     

细菌胞外聚合层在氧化含砷金精矿生物氧化过程中的作用

通过研究含砷金精矿细菌氧化过程中各项工艺指标的变化,测定pH值、电压、砷离子浓度、细菌数量和胞外聚合层中多糖含量及结构,探讨胞外聚合层中多糖在细菌氧化含砷金精矿过程中的作用。结果表明:当电压由500mV升到650mV、As(Ⅲ)快速转化为As(Ⅴ)时,单个细菌可通过消耗胞外聚合层中的多糖来抵御砷离子的毒害;当电压稳定在650mV、砷离子价态转化减缓时,细菌胞外聚合层中的多糖参与矿物的氧化反应;在砷离子胁迫下,多糖含量降低幅度可超过20%,即细菌可通过改变单个菌体胞外聚合层中多糖的产量来防止细胞受到侵害。     

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.