含炭难处理金矿石碘法浸出

用非氰试剂——碘从矿石中浸出黄金，目的是研究碘化浸金取代剧毒氰化物提金的可行性；采用了热力学方法对碘-碘化物溶金的可能性进行了简单分析，用碘-碘化物溶液对含炭难处理金矿石进行了浸出工艺条件实验及与氰化浸出的对比实验，取得了较理想的实验结果：浸出4h金浸出率达到95％，氰化法直接浸出12h金浸出率80％；碘化法浸金速度快，金浸出率高。     

用碘--碘化物溶液从含金矿石中提取金

考虑不同电解液中影响金溶解的热力学因素,由热力学计算结果绘制25℃时的金--碘--水体系的Eh-pH图,分析得到金-碘系统在水环境中的稳定区域：pH=0～14;结合对碘化浸金热力学、动力学的分析,在浸金实验的基础上,分析碘浸出金的机理;以碘-碘化物为浸出剂,实验研究含碳矿石中金的溶解,讨论金萃取的反应时间、碘-碘化物质量分数、pH对浸出的影响,并与氰化物的浸出结果进行对比;探讨从浸出液中回收金及碘的再生方法.所得金浸出的合理条件为：碘的质量分数在0.8%～1.0%,n（12）：n（I^-）=1：8～1：10,助氧化剂双氧水质量分数在1%～2%,浸出时间为4h,液固比在3：1～5：1,浸出温度为常温,矿浆为中性或酸性.金的浸出率可达93%以上,浸出液中的金和碘均可用电解方法回收.分析证明,在达到相同浸出率的情况下,与氰化法相比,碘化法浸出时间短、综合成本低.     

含铜难处理金矿选择性浸出试验研究

针对某含铜难处理金矿进行了碘化法和石硫合剂(lime sulfur synthetic solution,LSSS)法的选择性浸金的研究。结果表明,在碘单质质量浓度为8g/L,浸出时间为2h的条件下,碘化法浸出金的浸出率为88.1%,而且铜的浸出率不足1%。在石硫合剂质量分数为25%,浸出时间为6h的条件下,LSSS法浸出金的浸出率仅为73.5%。对比碘化浸出和石硫合剂浸出效果可知,碘化法对该含铜难处理金矿不仅浸出速度快、浸出率高而且铜几乎不被浸出,具有很强的选择性浸金作用。     

某难处理多金属矿石综合浸出铀、铜、银试验研究

针对某难浸铀矿石,采用“氯化焙烧-硫酸浸出”工艺进行处理提取铀、铜、银。研究结果表明,最佳氯化焙烧实验条件为氯化钠用量6%,氯化焙烧温度 460 ℃,氯化焙烧时间2 h,焙烧液固比0.2∶1。对氯化焙烧后的矿样进行硫酸浸出,浸出条件为:硫酸浓度30 g/L、浸出时间30 min、浸出温度70 ℃、液固比2∶1,此时金属离子铀、铜、银的浸出率分别为铀85.08%、铜95.82%、银91.80%。     

砷硫铁的脱除对高砷金精矿氰化浸金的影响

通过焙烧脱砷和硫,并采用硫酸浸出金焙砂脱铁,研究不同条件下砷、硫和铁的脱除对高砷金精矿氰化浸金的影响。研究结果表明:金精矿中,砷、硫和铁的质量分数分别为3.20%,27.35%和23.50%;在焙烧温度为500℃,焙烧时间为4 h和空气流量为0.2 m~3/h条件下,砷和硫脱除率分别达到51.53%和79.16%;所得金焙砂经过质量分数为30%硫酸浸出,铁浸出率高达98.12%,酸浸渣中砷、硫和铁质量分数分别为0.10%,0.55%和0.44%;采用质量分数为6‰的氰化钠溶液浸出酸浸渣,金浸出率达98.05%;经过对砷、硫和铁进行脱除,金品位从32.98 g/t增加到68.22 g/t;金焙砂通过酸浸,单体金和裸露金总质量分数从93.87%增加到96.66%;低温焙烧和酸浸适合高砷金精矿氰化浸金。     

废旧电路板中金的碘化法回收工艺

针对电路板中金的回收问题,提出粉碎—浮选—碘化浸出工艺。采用浮选法对经粉碎分级的电路板粉末进行分选实验,分选出的金属粉末用硝酸去除贱金属,过滤,由碘化法浸出滤渣中的金。实验结果表明:电路板中金属和非金属完全解离粒度为0.450 mm;浮选分离小于0.450 mm电路板粉末,沉物金属质量分数为87.32%,金属回收率为90.11%;碘和碘化物溶液可以在3 h内有效浸出电路板粉末中的金,金浸出率为95.53%。     

载金硫化物焙烧--自浸出过程研究

针对传统氧化焙烧-氰化浸金工艺环境污染严重的现状,采用焙烧-自浸出工艺提取载金硫化物中的金.研究焙烧温度、焙烧时间和试样量对单质硫转化率和金浸出率的影响,通过X射线衍射分析、扫描电镜观察、能谱分析等手段分析焙烧过程中载金硫化物中硫的物相转变规律.载金硫化物中黄铁矿发生热分解反应生成单质硫和磁黄铁矿,随焙烧温度的升高和焙烧时间的延长,黄铁矿的特征衍射峰强度逐渐减小直到消失,磁黄铁矿的特征衍射峰逐渐生成并增强,原本致密状的黄铁矿颗粒变得疏松多孔.50 g试样在氮气流量1 L·min-1、焙烧温度800℃、焙烧时间60 min的条件下,单质硫的转化率达到42.53%,金浸出率达到88.70%,实现载金硫化物的高效非氰浸出.     

双氧水还原浸出非洲氧化铜钴矿的试验研究

采用双氧水还原浸出非洲氧化铜钴矿,研究了还原剂用量、初始酸浓度、液固比、浸出温度和浸出时间等参数对浸出过程的影响。结果表明：使用双氧水与铜钴矿计量比为0.2 mL/g、浓硫酸与铜钴矿质量比为0.46、液固比为5：1（mL/g）,在温度75℃下浸出2 h,钴、铜的浸出率分别达到了99.50%,99.42%。     

碘化浸金过程的基本理论

碘化浸金被证明是一种有效环保的非氰浸金方法,但金溶解过程中的基本理论目前研究很少。结合热力学数据,计算了实际的浸金体系中主要反应平衡电位,建立了可能存在的化学反应及电位和pH之间的关系式,完整地做出了实际浸金体系的V—pH图,并对碘化浸金机理进行分析。该研究为碘化浸金的工业化提供了理论指导。     

硫脲法浸金研究

硫脲法浸金是近年来引人注目且比较活跃的研究课题。本文研究了硫脲浓度、氧化剂用量、浸取温度和浸取时间对金的浸出率的影响,探求了硫脲法浸金的适宜条件。     

贵州高硫高品位铝土矿硫溶出性能研究

对不同硫含量的贵州高硫铝土矿硫的溶出率进行了研究。在铝土矿硫含量一定的条件下,考查石灰添加量、溶出温度、苛性碱浓度和溶出时间等对硫的溶出率的影响,得出对硫溶出率影响较大的影响因素。实验结果表明,中等硫含量的高硫铝土矿能得到较低硫溶出率,较适宜的溶出条件为溶出温度260℃,碱浓度195 g/l,石灰添加量CaO12%,时间50 min.     

Recovery of gold from hydrometallurgical leaching solution of electronic waste via spontaneous reduction by polyaniline

The present study is primarily designed to develop an environmentally-benign approach for the recovery of precious metals, especially gold, from the ever increasingly-discarded electronic wastes (e-waste). By coupling the metal reduction process with an increase in the intrinsic oxidation state of the aniline polymers, and the subsequent re-protonation and reduction of the intrinsically oxidized polymer to the protonated emeraldine (EM) salt, polyaniline (PANi) films and polyaniline coated cotton fibers are able to recover metallic gold from acid/halide leaching solutions of electronic wastes spontaneously and sustainably. The current technique, which does not require the use of extensive extracting reagents or external energy input, can recover as much as 90% of gold from the leaching acidic solutions. The regeneration of polyaniline after gold recovery, as confirmed by the X-ray photoelectron spectroscopy measurements, promises the continuous operation using the current approach. The as-recovered elemental gold can be further concentrated and purified by incineration in air.     

稀酸浸出还原焙烧红土矿时铁还原度对浸出的影响

对含镍的红土矿进行了微波加热还原焙烧-稀硫酸浸出的实验研究,考察了活性炭粉加入量、微波功率和加热时间对铁还原程度的影响,并分析了铁还原度对浸出过程中镍和铁浸出率的影响.结果表明红土矿中铁的还原程度随碳粉加入量、微波功率和加热时间的增加而增大,在800 W的微波功率下加热约12.5 min即可完成还原反应.镍的浸出率与铁的还原度近似呈线性关系,但铁的浸出率在铁还原度超过60%后增长迅速.因此铁还原度控制在60%为宜,相应的镍浸出率约为85%,而铁的浸出率不超过30%.     

难处理金矿中伴生矿物对氰化浸出的影响

对金矿石中常见的几种伴生矿物在氰化浸出中的影响进行了分析,同时采用化学试剂配制标准液的方式,考察了Fe2+,Cu2+,As3+对氰化物消耗的影响.试验结果表明:铁矿物中,磁黄铁矿对氰化浸出的影响较大,使溶金速率下降28.1%,氰化物耗量增加4倍,而黄铁矿与赤铁矿对氰化浸出的影响较小;铜矿物中,黄铜矿与辉铜矿对氰化浸出都具有很大影响,其中辉铜矿可使溶金速率下降36.81%,氰化物耗量增加10倍;砷矿物中,雄黄与雌黄对氰化浸出极其有害,使溶金速率分别下降41.95%和49.90%,氰化物耗量分别增加13.8倍和15.0倍,相反毒砂在氰化体系中比较稳定,对氰化浸出的影响较小.离子耗氰试验中,Fe2+...     

从难处理金矿中回收金:黄铁矿对硫脲浸金体系稳定性的影响

The extraction of gold from refractory gold ores (RGOs) without side reactions is an extremely promising endeavor. However, most RGOs contain large amounts of sulfide, such as pyrite. Thus, investigation of the influence of sulfide on the gold leaching process is important to maximize the utilization of RGOs. In this work, the effects of pyrite on the stability of the thiourea system were systematically investigated under different conditions. Results showed that the decomposition rate of thiourea was accelerated sharply in the presence of pyrite. The effect of pyrite on gold recovery in thiourea leaching systems was then confirmed via a series of experiments. The decomposition efficiency of thiourea decreased by 40% and the recovery efficiency of gold increased by 56% after the removal of sulfide by roasting. Under optimal conditions, the efficiency of the gold recovery system increased to 83.69% and only 57.92% of thiourea decomposition was observed. The high consumption of thiourea by the leaching system may be attributed to not only adsorption by mineral particles but also catalytic decomposition by some impurities in the ores, such as pyrite and soluble ferric oxide.     

用高品位高硫铝土矿生产氧化铝时的除硫新方法研究

本文研究了贵州高硫型高品位铝土矿溶出性能,考查了石灰添加量、溶出温度、苛性碱浓度和溶出时间对氧化铝溶出率的影响以及铝土矿的硫向溶液中的转化率(即硫的溶出率)的影响;还研究了用铝酸钡净化拜耳赤泥洗液的除硫条件和效果。     

Study on gold concentrate leaching by iodine-iodide

Gold extraction by iodine-iodide solution is an effective and environment-friendly method. In this study, the method using iodine-iodide for gold leaching is proved feasible through thermodynamic calculation. At the same time, experiments on flotation gold concentrates were carried out and encouraging results were obtained. Through optimizing the technological conditions, the attained high gold leaching rate is more than 85%. The optimum process conditions at 25℃ are shown as follows: the initial iodine concentration is 1.0%, the iodine-to-iodide mole ratio is 1:8, the solution pH value is 7, the liquid-to-solid mass ratio is 4:1, the leaching time is 4 h, the stirring intensity is 200 r/mim, and the hydrogen peroxide consumption is 1%.