Prediction of pre-oxidation efficiency of refractory gold concentrate by ozone in ferric sulfate solution using artificial neural networks

An artificial neural network model was developed to predict the oxidation of refractory gold concentrate （RGC） by ozone and ferric ions. The concentration of ozone and ferric ions, pulp density, oxygen amount, leaching time and temperature were employed as inputs to the network; the output of the network was the percentage of the ferric extraction iron from RGC. The multilayered feed-forward networks were trained by 33 sets of input-output patterns using a back propagation algorithm; a three-layer network with 8 neurons in the hidden layer gave optimal results. The model gave good predictions of high correlation coefficient （R2=0.966）. The predictions by ANN are more accurate when compared with conventional multivariate regression analysis （MVRA）. In addition, calculation with ANN model indicates that temperature is the predominant parameter and ozone concentration is the lesser influential parameter in the pre-oxidation process of refractory gold ore. The ANN neural network model accurately estimates the ferric extraction during pretreatment process of RGC in gold smelter plants and can be used to optimize the process parameters.

Pretreatment and thiosulfate leaching of refractory gold-bearing arsenosulfide concentrates

A hydrometallurgical process for refractory gold-bearing arsenosulfide concentrates at ambient temperature and pressure was presented, including fine grinding with intensified alkali-leaching （FGIAL）, enhanced agitation alkali-leaching （EAAL）, thiosulfate leaching and displacement. Experimental results on a refractory gold concentrate showed that the total consumption of NaOH in alkaline leaching is only 41% of those theoretically calculated under the conditions of full oxidization for the same amount of arse- nides and sulfides transformed into arsenates and sulfates, and 72.3% of gold is synchro-dissoluted by thiosulfate self-generated during alkaline leaching. After alkaline leaching, thiosulfate leaching was carried out for 24 h. The dissolution of gold is increased to 91.9% from 4.6% by cyanide without the pretreatment. The displacement of gold by zinc powder in the solution gets to 99.2%. Due to an amount of thiosulfate self-generated during alkaline leaching, the reagent addition in thiosulfate leaching afterwards is lower than the normal.

Biooxidation-thiosulfate leaching of refractory gold concentrate

A process of biooxidation followed by thiosulfate leaching of gold from refractory gold concentrate was investigated.Mineralogical studies on the concentrate showed that very fine gold grains(<10μm)were encapsulated in pyrite and arsenopyrite,while the proportion of monomer gold was only 21%.The gold-bearing sample was identified as a high-sulfur fine-sized wrapped-type refractory gold concentrate.The gold leaching efficiency obtained by direct cyanidation was only 59.86%.After biooxidation pretreatment,the sulfide minerals were almost completely decomposed,92 wt%of the mineral particles of the biooxidation residue were decreased to<38μm,and the proportion of monomer gold in the biooxidation residue was over 86%.Meanwhile,the gold content in the biooxidation residue was enriched to 55.60 g/t,and the S,Fe,and As contents were reduced to approximately 19.8 wt%,6.97 wt%,and 0.13 wt%,respectively.Ammoniacal thiosulfate was used for gold extraction from the biooxidation residue of the refractory gold concentrate.The results showed that the optimal reagent conditions were 0.18 M thiosulfate,0.02 M copper(II),1.0 M ammonia,and 0.24 M sulfite.Under these conditions,a maximum gold leaching efficiency of 85.05%was obtained.

Differences of cyanide leaching between calcine and dust from refractory gold concentrates

Differences of cyanide leaching between the calcine and the dust from a refractory gold concentrate were investigated by comparative method. Results showed that gold leaching efficiencies of the calcine and the dust were 85.31% and 54.30%, respectively, with direct cyanidation. Contents and existing forms of arsenic and carbon were the main reasons for those differences. The maximum gold leaching efficiencies of the calcine and the dust were 87.70% and 58.60%, respectively, with cyanidation after NaOH pre-leaching. Harmful elements removal, gold loss in NaOH pre-leaching and iron oxides hindrance codetermined gold leaching efficiencies of the calcine and the dust. After H2SO4 pre-leaching, the maximum gold leaching efficiencies of the calcine and the dust achieved 94.96% and 80.40%, respectively. The effect of carbonaceous matter was the main reason for differences for leaching efficiencies of the calcine and the dust. Based on those differences, two proper gold extraction processes were put forward, and gold leaching efficiencies for the calcine and the dust achieved 94.91% and 91.90%, respectively.

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 ozonein 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 inhydrochloric acid was employed and the influences of temperature,liquid/solid ratio,HCl concentration and stirring speed on theextraction of antimony were investigated.High antimony extraction(93.75%)is achieved under the optimized conditions.After thepretreatment by ozone,the antimony is recovered efficiently and the gold is enriched in the leaching residue.

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 of10%(w/v),using a basal salt medium containing ferrous iron at pH1.8.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,allowing91.6%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.

Well-controlled stirring tank leaching to improve bio-oxidation efficiency of a high sulfur refractory gold concentrate

For the high sulfur refractory gold concentrate with 41.82%sulfur and 15.12 g/t gold,of which 82.11%was wrapped in sulfide,a well-controlled stirring tank leaching was carried out to improve the bio-oxidation efficiency.Results show that bio-oxidation pretreatment can greatly improve the gold recovery rate of high-sulfur refractory gold concentrate,and at the optimum pH 1.3 in this study,compared with the process without pH control,the oxidation rate of sulfur increased from 79.31%to 83.29%,while the recovery rate of gold increased from 76.54%to 83.23%;under this condition the activity of mixed culture could be sustained and the formation of jarosite could diminish.The results also displayed that for the high sulfur refractory gold concentrate,the recovery of gold is positively correlated with the oxidation rate of sulfur,and the recovery rate of gold increases with the increase of sulfur oxidation rate within a certain range.

A new extraction process of refractory gold arsenosulfide concentrates

A new hydrometallurgical process for the refractory gold arsenosulfideconcentrates under normal temperature and pressure was presented. The experimental results of arefractory gold concentrate show that the total consumption of NaOH in alkaline leaching is only 40%of those theoretically calculated under the conditions of full oxidization at the same oxidation ofarsenic to ar-senate and sulfur to sulfate. After alkaline leaching, cyanidation and adsorptionwere carried out for 24 h. The dissolution of gold by NaCN is increased to 95.3% from 12.8% beforepretreatment, and meanwhile 99.3% of the adsorption of gold by activated charcoal. The consumptionof NaCN for one ton ore is 10 kg, which is 1.2 times less than that before pretreatment. As it iscarried out under normal temperature and pressure, the investment of installations is alsodecreased.

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 03.2% 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.

Intensified alkaline leaching pretreatment of refractory gold concentrates at common temperature and pressure

A new process for the hydrometallurgy of refractory gold concentrates was presented. The process comprises grinding leaching, intensified alkaline leaching (IAL), cyanidation and adsorption. In a stirring type pulverizing leaching tower mill, the concentrate is ground to <35.6 μm of 95.5 % while simultaneously leached by NaOH of 12 kg/t, then carried out intensified alkaline leaching for 48 h by NaOH of 108 kg/t in enhanced agitation tanks with the pulp concentration of 40% solids at the environmental temperature of 9.5 ～ 13.5 ℃ and the environmental pressure of 10 5 Pa. The oxidation rate of As is 94.9%, and 47.6% for S. The total consumption of NaOH is only 20% of that theoretically calculated under the conditions of full oxidation at the same oxidation rates of arsenic to arsenate and sulfur to sulfate. The gold leaching rate by NaCN in 24 h is increased from 9.2% before pretreatment to 94.2%. The consumption of NaCN is 7.5 kg/t, which is one times less than that before pretreatment. The extraction cost of gold is about 422 Yuan/t.

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 bioleaching in a semi continuous way, by which a process for bioleaching 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 6.5?g/L and pulp concentration 10%. And leached rate of gold in the following cyanidation is over 90%. The parameters of three series fluidized bed reactors exhibit stability during the semi continuous bioleaching of the concentrate. Arsenic 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.

Effect of Pretreatment of Sulfide Refractory Concentrate with Sodium Hypochlorite, Followed by Extraction of Gold by Pressure Cyanidation, on Gold Removal

The majority of the refractory gold and silver occurs in occlusion in sulphides, then precious metal dissolution is possible by first oxidizing auriferous sulfide concentrate using sodium hypochlorite-sodium-hydroxide solution followed of pressure cyanidation of the oxidized concentrate, for the extraction of precious metals. This process was conducted and evaluated under cyanide and oxygen pressure. This versatile approach offers many advantages, including low temperatures, low pressure and less costly materials of construction than conventional pressure oxidation. In this study, the effect of oxygen pressure, concentration of sodium hypochlorite, temperature, and initial pH, in precious metals recovery and As removal were evaluated using a 24 factorial design. Characterization of the ores showed that pyrite and arsenopyrite were the main minerals present on the concentrate. The best results for gold extraction were obtained with oxygen pressure of 80 psi, 10% (w/w) sodium hypochlorite, temperature of 80℃, at pH = 13, and a constant stirring speed of 600 rpm. These conditions allowed an approximated 60% of gold and 90% of silver extractions in 1 hr.

塔吉克斯坦某难处理金精矿低温二段焙烧预处理试验研究

高砷高硫金精矿中的金矿物被毒砂(As_(2)O_(3))和黄铁矿(FeS_(2))包裹,直接氰化浸出浸出率较低,其预处理工艺一直是研究关注的热点。本文以塔吉克斯坦高砷高硫金精矿为研究对象,在热重分析和热力学分析的基础上采用二段焙烧法进行预处理,脱除大部分砷、硫、有机碳等有害杂质,并对工艺参数进行优化。试验考察了添加剂及焙烧温度、时间对金精矿中碳、硫、砷脱除率的影响,并在最终焙烧产物物相分析基础上进行了焙烧-氰化浸出验证试验,得到以下主要结论。一段焙烧采用弱氧焙烧气氛焙烧脱砷,不需使用添加剂,较优工艺参数为焙烧温度500~550℃、焙烧时间2 h;二段焙烧采用富氧气氛焙烧脱硫,较优工艺参数为焙烧温度650~700℃、时间2 h;在上述较优工艺参数下的焙烧-氰化浸出的碳脱除率、硫脱除率和砷脱除率分别为91.16%、96.93%和92.52%,金浸出率达到95.30%。该结果相对于直接氰化浸出,金浸出率提高了20%,一定程度上实现了资源的高值化利用,为难处理金精矿高效提金提供了技术依据。

某高硫含砷低品位金精矿选矿工艺研究

为提高某高硫含砷低品位难处理金精矿选别指标,分别采用尼尔森离心重选和高效旋流器分级工艺进行金精矿分选试验研究。试验结果表明:在保障金精矿冶炼净返系数等级的基础上,采用试验条件为重力倍数60G、反冲洗水量4 L/min、给矿速度400 g/min的尼尔森离心重选工艺,可获得金品位为30.97 g/t,金回收率为19.97%的重选精矿;采用试验条件为矿浆浓度20%、沉砂口直径22 mm的高效旋流器分级工艺,可获得金品位为26.02 g/t,金回收率为12.50%的溢流。尼尔森离心重选、高效旋流器分级工艺可使最终精矿产品的金回收率分别提高1.6%~1.9%和0.7%~1.0%。考虑到尼尔森离心重选工艺成熟工艺少、后期投资高,故推荐高效旋流器分级工艺进行该金精矿分选试验,以提高该金精矿利用价值。

难处理金精矿加压氧化预处理

研究了某含硫21.2%、含金40.3 g/t难处理金精矿的加压氧化预处理工艺。通过单因素试验研究加压氧化过程中磨矿细度、反应压力、反应温度、反应时间、搅拌转速及矿浆浓度等对氰化效果的影响,得到易于氰化的加压氧化渣。再通过氰化提金过程中矿浆浓度、氰化钠用量、pH、浸出时间等对金浸出率的影响研究,获得最佳加压氧化预处理条件为:磨矿细度-0.045 mm占95%、反应压力1.6 MPa、矿浆浓度20%、反应温度160℃、反应时间240 min、搅拌转速350 r/min。在此条件下,金氰化平均浸出率达到97.3%。

某复杂难处理硫精矿综合回收的试验研究

以昆明某复杂难选硫精矿为原料,先在马弗炉中进行了焙烧脱硫砷的条件试验,得出了焙烧的最佳操作条件后,以沸腾炉(工业生产炉床面积74 m^(2))中产出的烧渣为原料进行了氰化提金银的条件试验。该硫精矿综合处理的最佳条件为:焙烧温度840~870℃,焙烧时间2~2.5 h;烧渣磨矿细度-0.038 mm≥90%,氰化钠加入量2 kg/t,浸出液固比2∶1等,此时硫脱除率达99%以上,金浸出率达84%;经分析氰化尾渣可作为铁精矿出售。该项目已在云南某企业建成处理量300 t/d复杂含金硫精矿提金的工业生产线。

氯离子对难处理金精矿生物预氧化-炭浸氰化的影响

高盐矿区往往缺乏淡水资源供生产使用。针对矿区盐水现状,进行高砷难处理金精矿生物氧化-氰化提金试验,考查不同氯离子浓度对硫化物氧化率、金浸出率等技术指标的影响。结果表明,氯离子浓度对金精矿生物氧化的不利影响随着浓度的升高而增强,氯离子浓度1.5 g L以上时,硫化物的氧化率显著降低,10 g L氯离子盐水生物氧化延长至18 d,砷浸出率88.4%,硫氧化率仅为35.3%。相同生物氧化渣在相同氰化条件下采用清水和10 g L氯离子盐水进行氰化浸出时,盐水将降低金的浸出率。1.5 g L、5 g L氯离子盐水生物氧化10 d后的氧化渣采用10 g L氯离子盐水氰化,金的浸出率分别比清水氰化时的分别低2.8、3.4个百分点。难处理金精矿生物氧化砷浸出率均高于硫氧化率,表明毒砂氧化率高于黄铁矿,金浸出率与毒砂氧化率的相关性较黄铁矿氧化率更加密切。

难处理含金硫精矿的焙烧氧化-硫代硫酸盐浸出

为了提高难处理含金硫精矿中金的浸出率,采用同步热分析仪研究在马弗炉中焙烧氧化难处理含金硫精矿的最佳条件,通过优化实验确定硫代硫酸盐浸出的最佳工艺参数。结果表明:在马弗炉中焙烧氧化难处理含金硫精矿最佳条件为在700℃温度下焙烧2 h,难处理含金硫精矿硫的去除率可达94.7%。焙烧后,金的浸出率大幅度提高。使用组成为0.03 mol/L CuSO4、1.0 mol/L NH3·H2O、0.3 mol/L Na2S2O3、0.1 mol/L(NH4)2SO4和0.3 mol/L Na2SO3的硫代硫酸盐溶液作为浸出剂,最佳浸出工艺参数如下:浸出时间18 h、液固比2:1、振荡速度250 r/min、浸出温度50℃。在此浸出工艺参数下,金浸出率达71.2%。

高砷硫金精矿提金研究

广西贵港金精矿含15%～22%As,并含碳、铅等不利于氰化的元素,金直接氰化率8%～36%。采用催化氧化酸浸法预处理后,金氰化率可达92%～98%,氰渣浮选后精矿的金总收率达97%～99%。预氧化工艺在中温自热、低压、低酸操作条件下进行,废渣、废液符合环保要求。多批次小型试验及扩大试验结果表明该工艺技术指标稳定,经济可行。

湖南某高砷难处理金精矿的细菌氧化-氰化提金实验研究

湖南某高砷金精矿属于难处理矿石,含砷11.28%,含金66.18g/t,金的直接浸出率仅为21.91%。通过该样品5%、10%、15%、20%矿浆浓度的细菌氧化试验,发现金精矿砷的氧化率达到93%以上。细菌氧化渣的金浸出率随着矿浆浓度的增大而降低,5%矿浆浓度下细菌氧化渣的金浸出率为93.15%;10%矿浆浓度下细菌氧化渣的金浸出率为92.46%;15%矿浆浓度下细菌氧化渣的金浸出率为90.50%;20%矿浆浓度下细菌氧化渣的金浸出率为87.58%,比未经处理时金的直接氰化浸出率21.91%有了很大的提高,预处理效果很好。