Extraction of lithium from lepidolite using chlorination roasting-water leaching process

Chlorination roasting followed by water leaching process was used to extract lithium from lepidolite.The microstructure of the lepidolite and roasted materials were characterized by X-ray diffraction（XRD）.Various parameters including chlorination roasting temperature,time,type and amount of chlorinating agents were optimized.The conditional experiments indicate that the best mass ratio of lepidolite to NaCl to CaCl2 is 1：0.6：0.4 during the roasting process.The extraction of lithium reaches peak value of 92.86% at 880 °C,potassium,rubidium,and cesium 88.49%,93.60% and 93.01%,respectively.The XRD result indicates that the major phases of the product after roasting lepidolite with mixture of chlorinating agents（CaCl2 and NaCl） are SiO2,CaF2,KCl,CaSiO3,CaAl2Si2O8,NaCl and NaAlSi3O8.

Extraction of metals from complex sulfide nickel concentrates by low-temperature chlorination roasting and water leaching

The recovery of valuable metals from complex sulfide concentrates was investigated via chlorination roasting followed by water leaching. A reaction process is proposed on the basis of previous studies and the results of our preliminary experiments. During the process, various process parameters were studied, including the roasting temperature, the addition of NH4Cl, the roasting time, the leaching time, and the liquid-to-solid ratio. The roasted products and leach residues were characterized by X-ray diffraction and vibrational spectroscopy. Under the optimum condition, 95% of Ni, 98% of Cu, and 88% of Co were recovered. In addition, the removal of iron was studied in the water leaching stage. The results demonstrate that this process provides an effective approach for extracting multiple metals from complex concentrates or ores. © 2017, University of Science and Technology Beijing and Springer-Verlag Berlin Heidelberg.

Kinetics of leaching refractory gold ores by ultrasonic-assisted electro-chlorination

The resources of refractory gold ores are abundant, and their effective treatment can bring good economic benefits. This paper investigated the kinetics of leaching gold from refractory gold ores by ultrasonic-assisted electro-chlorination. The effects of ultrasound time ratio, initial hydrochloric acid concentration and leaching temperature on the kinetic parameters were discussed. It is found that the leaching ratio goes up with all the factors increasing. The reaction kinetics is controlled by diffusion. When ultrasound improves the diffusion by reducing the diffusion resistance, the activation energy increases to 37.1 kJ/mol.

Behavior of silver and lead in selective chlorination leaching process of gold-antimony alloy

The behavior of silver and lead in the selective chlorination leaching process of gold-antimony alloy was analyzed in detail and appropriate recovery methods were developed.A reduction method by adding gold-antimony alloy powder was adopted to recover silver according to the thermodynamics calculation.The reducing rate of silver can exceed 99%at 80℃for 1.5 h when the dosage of gold-antimony alloy powder is 10%.The dissolution equilibrium curved surfaces of PbSO4 and PbCl2 under different conditions were drawn.The experimental results are well consistent with theoretical analysis that indicate lead may be precipitated in the form of lead chloride.The grade of gold in the residue can be further concentrated to 94.5%after being washed with hot water. These two methods have been applied successfully in the practice.

Chlorination roasting-coupled water leaching process for potash recovery from waste mica scrap using dry marble sludge powder and sodium chloride

The present paper reports the effective utilization of marble sludge powder(MSP)for the recovery of potash values from waste mica scrap using chlorination roasting-water leaching method.Characterization studies indicated the presence of dolomite as the major mineral phase in MSP,whereas muscovite and quartz were observed in the mica sample.The acid leaching studies suggest a maximum of 22%potash recovery under conditions:4 M H2SO4 acid,particle size of^100μm,stirring speed of 600 r/min,leaching temperature of 75℃,and leaching time of 90 min.The chlorination roasting-water leaching process was adopted to achieve the lowest level of 80%-90%potash recovery.The optimum conditions for the recovery of^93%potash from mica(~8.6wt%K2O)requires 900℃ roasting temperature,30 min roasting time,and 1:1:0.75 mass ratio of mica:MSP:NaCl.The roasting temperature and amount of NaCl are found to be the most important factors for the recovery process.The reaction mechanism suggests the formation of different mineral phases,including sylvite(KCl),wollastonite,kyanite,and enstatite,during roasting,which were confirmed by X-ray diffraction(XRD)analyses and scanning electron microscopy(SEM)morphologies.The MSP-blended NaCl additive is more effective for potash recovery compared with the other reported commercial roasting additives.

Roasting and chlorine leaching of gold-bearing refractory concentrate:Experimental and process analysis

This paper deals with gold extraction from a refractory concentrate by chlorine leaching.The process considers a pre-treatment of refractory materials by low temperature oxidation carried out with low oxygen concentration.The oxidized material is treated by leaching with brine.After gold adsorption/reduction onto activated carbon,iron and base metals can be precipitated by NaOH.Roasting tests show the necessity to carry out a thermal pre-treatment at least at 550°C to obtain a reduction of sulfur and mercury in the material(50%and 90%,respectively).Highest gold extraction yield(around 93%)is obtained in the leaching test performed with material sample treated at 650°C.This result confrms the necessity to optimize the thermal pre-treatment to improve Au recovery and to reduce chlorine consumption.A comparison with conventional cyanidation confrms that chlorination could be an useful alternative:in fact,gold extraction yield is quite low:57%in non-pre-treated material and 25%in pre-treatment material.

Effect of Roasting Temperature for Copper Leaching of Sulfide Concentrate by Combined Methods

This study was undertaken to develop a cost-effective method of copper leaching in sulfide concentrate involving minerals as chalcopyrite and pyrite. The combined methods involving low temperature roasting of sulfide copper minerals with potassium chlorides without the formation of sulfurous gases and their evolution into the atmosphere is suggested. The effect of experi-mental parameters such as a roasting temperature and time has been pre-sented and discussed in detail. This study was carried out in the roasting temperature range 400&#176C - 600&#176C and roasting time ranges 1 - 4 h with potas-sium chloride. The obtained cake is leached by an aqueous sulfuric acid 60 g/l concentrated at pH = 2 - 2.5 and 4 h. The concentrate, a roasting cake and after leaching solid residue were analyzed by chemical analysis, XRD and SEM-EDX. At temperature ranges 450&#176C - 550&#176C and 3 h of roasting, the pro-cess is activated. The recovery of copper into the solution at optimal roasting condition (500&#176C, roasting time 3 h) is 93.2%. The advantage of the devel-oped method is that, it not requires high roasting temperature and the use of autoclave for leaching. The method gives possibility to leach the concentrate at room temperature and atmospheric pressure, which gives possibility to reduce cost of the process.

氯化焙烧-水浸法从富铷锂云母矿中提取Li、Rb

由于低铷锂云母中的Rb赋存形态差异较大,且含量低,提Rb工艺参数稳定性差,难以实现工业化生产。本文以澳大利亚富铷锂云母矿为原料,采用NaCl-CaCl_(2)焙烧-水浸法进行了提取Li、Rb的试验研究,考察焙烧温度、配料比、时间和浸出液固质量比、时间、温度对Li、Rb提取的影响。结果表明,在富铷锂云母矿:NaCl:CaCl_(2)的质量比为1:0.6:0.4、焙烧温度为800℃、焙烧时间为1.5 h、水浸温度为25℃、液固比4:1、浸出时间15 min的条件下,Li、Rb浸出效果最佳,Li、Rb浸出率分别达87.8%和94.9%。焙砂及浸出渣的X射线衍射(XRD)分析表明,氯化焙烧后,锂云母中的Li和Rb矿相转化为可溶性的LiCl和RbCl相,同时生成KCl和难溶物CaAl_(2)Si_(2)O_(8)、NaAlSi_(3)O_(8)、CaF_(2)。△GΘT-T和△HΘT-T关系图表明,NaCl-CaCl_(2)与碱金属氧化物的反应顺序为Rb>K>Li;矿石中的SiO_(2)与Al_(2)O_(3)参与对碱金属的氯化反应有促进作用。

钛冶金氯化收尘渣钒资源回收利用研究

对某钛冶金氯化收尘渣进行水洗无害化处理得到水洗渣;对水洗渣进行钙化焙烧,研究了焙烧温度和焙烧时间对水洗渣钒浸出率的影响。结果表明,氯化收尘渣中钒主要以低价钒形式存在;钒浸出率随焙烧温度升高而增加,焙烧温度850℃时,钒浸出率达88%;焙烧时间增加,不同产物进行二次反应,形成钒青铜等难溶化合物,出现不同价态钒酸盐相,导致在Na_(2)CO_(3)溶液中钒浸出率不同;适宜的焙烧温度为850℃,焙烧时间为120 min。

用氯化浸出法从铜阳极泥中回收单质砷

研究了采用氯化浸出—NaH_(2)PO_(2)选择性还原工艺回收铜阳极泥中的单质As。考察了液固体积质量比、初始酸度、浸出温度、初始Cl^(-)质量浓度、浸出时间对As浸出率的影响,以及还原温度、NaH_(2)PO_(2)与As的物质的量比、还原时间对As还原沉淀率的影响。结果表明:最佳氯化浸出条件为初始酸度220g/L,液固体积质量比6/1,浸出温度70℃,初始Cl^(-)质量浓度150g/L,浸出时间2.5h;最佳还原条件为还原温度80℃,NaH_(2)PO_(2)与As的物质的量比2/1,还原时间2h;在最佳条件下所得单质As产品中As平均质量分数为90.44%,产品质量符合单质As的精制要求。

基于电解-氯化法的废电路板中金回收试验

随着科技的进步,电子设备的更新换代速度加快,电子废弃物产生量与日俱增。电子废弃物中含有各种贵金属,具有巨大的回收价值。目前常用的提取方法是湿法冶金技术,其不仅程序复杂,且需使用酸、碱等有毒有害物质,故亟需开发一种环境友好且操作简单的回收技术。本文提出了一种将电解法和氯化法相结合,在单个反应器中,无需使用任何酸或碱,且其电解质成分只有NaCl的一步回收金的方法,并建立了阳极浸金、阴极还原金的电解氯化反应体系。以废电路板为研究对象,开展了金回收试验,结果表明:当NaCl浓度为3.8 mol/L、固液比为5 g/L、电流为600 A、温度为40℃、浸出时间为4 h时,金的浸出率为93.62%,金的回收率为66.74%。研究成果可为废弃电子设备中的贵金属高效安全回收提供借鉴。

金浸出尾渣微波氯化焙烧试验研究

金浸出尾渣因含有多种有价元素如金、铜等,其处理工作一直面临着挑战。传统氯化焙烧可有效提取金浸出尾渣中有价元素,但该过程需高温且消耗煤炭、天然气,释放大量二氧化碳和烟气。为此,提出采用微波氯化焙烧的新方法,该方法具有低温、清洁的优势。研究了微波氯化焙烧对Au挥发率和Cl脱除率等指标的影响,探讨了微波功率、氯化剂种类、焙烧温度和CaCl_(2)用量等因素的作用。试验结果不仅阐明了微波氯化焙烧的工作原理,还评估了焙烧渣的毒性。与传统方法相比,微波氯化焙烧显著降低了有毒元素含量。

钴锍氯化浸出工艺试验研究

钴锍是冶炼钴的常用原料之一,其富含钴、铜、铁等成分,化学性质稳定,常规浸出方法无法有效溶出有价金属,而且其硬度大,磨料处理难度大。本文以进口的钴锍颗粒为研究对象,从磨矿设备选择、氯化浸出试验、验证试验和工业试验等方面系统考察了钴锍氯气酸浸处理工艺的可行性,得到以下主要结论:振动磨料能有效处理硬度大的钴锍颗粒,经磨料后钴锍的排矿粒度90%在150μm以下,能耗为1.2 kg/kW·h;在15%盐酸与钴锍料液固质量比6∶1、拌速度60 r/min、反应温度95℃、反应时间8 h,并同时通入氯气的条件下,钴锍中铜、钴浸出率均达到90%以上,铁浸出率达到70%以上;工业试验中对磨料及浸出工序进行了优化,磨料能耗耗降至4 kg/kW·h,铜、钴、铁浸出率均达到了98%以上,效果提升显著。浸出液经针铁矿法除铁、萃取除铜后可以得到富钴溶液,而且浸出渣量小,能较好应用于工业化生产。

探究生活垃圾焚烧飞灰中氯对铅、镉浸出毒性影响

通过对18个不同区域飞灰进行半定量分析、定量分析、浸出毒性分析,探究飞灰中氯对铅、镉浸出毒性的影响。结果表明:飞灰中的主要元素为Ca、Al、Si、Na、Cl,重金属含量较多的为Pb、Zn、Cu,我国生活垃圾焚烧飞灰中氯含量为6.55%~16.8%,不同地区飞灰中氯含量差异性比较明显;原灰中氯含量和浸出液中铅、镉含量之间无明显关系,浸出液中氯含量和浸出液中铅、镉含量之间无明显关系。

氯化分金过程中的优化研究

本研究以某冶炼厂水浸分铜后产生的分铜渣为对象,探究了氯化分金过程中多个因素对分金渣中Au含量和Au浸出率的影响,包括NaClO_(3)的质量浓度、NaClO_(3)的反应温度、NaClO_(3)的添加时长、酸度、酸比(HCl∶H_(2)SO_(4))、反应固液比、反应温度、反应时间和分铜渣粒度。分铜渣粒度对Au浸出率影响较小,仅为1.99%。在最佳参数下,分金渣中Au含量下降了51.28 g/t,Au浸出率提高了4.06%。

铅冶炼次氧化锌烟灰浸出脱氯试验研究

次氧化锌烟灰主要成分为氧化锌,具有易于浸出且锌回收率高的特点,但随着氟、氯等元素增多,对后期的锌浸出及电积工艺控制造成很大的影响。高效去除次氧化锌烟灰中的氟氯成为锌冶炼行业研究的重点内容。通过单因素试验,以水和碳酸钠溶液作为浸出剂,分别考察液固比、浸出温度、浸出时间、pH等因素对铅冶炼生产过程产生的次氧化锌烟灰中氯离子脱除率及锌离子损失率的影响,并确定最佳工艺参数。结果表明,采用碳酸钠碱性浸出,液固比8∶1,浸出温度45℃,浸出时间100 min,pH=11.0,在此优化参数下,氯离子脱除率可达到87.43%,锌损失率为0.5%。

微细浸染型金矿石二氧化氯预氧化-浸出动力学

针对微细浸染型金矿石的二氧化氯预氧化-浸出过程动力学进行了研究,分别考察了磨矿细度、搅拌速度、浸出温度、亚氯酸钠浓度等因素在硫酸体系下对该矿石中金预氧化-浸出速率的影响。结果表明,适当地增加磨矿细度、提高浸出温度、增加亚氯酸钠浓度有利于提高浸出速率。浸出过程的动力学计算表明:采用二氧化氯预氧化-浸出微细浸染型金矿石的过程符合收缩核模型,浸金反应速率受混合过程控制,计算得到表观活化能为12.74 kJ/mol,并最终建立了该浸出过程的动力学方程式。

改性电炉钛渣酸解法制备沸腾氯化炉料新工艺研究

以改性电炉钛渣为原料,采用硫酸浸出-氟化氢铵浸出-盐酸浸出三段联合新工艺制备满足沸腾氯化要求的炉料,从热力学和浸出行为两方面对沸腾氯化炉料新工艺进行研究。研究结果表明:硫酸浸出可分解黑钛石相,并生成Ti OSO_(4),经水洗后TiOSO_(4)进入到滤液中。在酸矿比为1.7,熟化温度为250℃、熟化时间为2 h的条件下,Al、Ca、Mg、Si和Fe的浸出率分别为84.90%、26.90%、82.70%、12.03%和89.13%,Ti O_(2)品位和回收率为72.58%和42.41%。钙镁硅酸盐可以被氟化氢铵分解并生成CaMg_(2)Al_(2)F_(12)沉淀。在NH_(4)HF_(2)质量分数为15%、液固比为10:1、浸出时间为2 h、浸出温度为25℃的条件下,Al、Ca、Mg、 Si和Fe的浸出率分别为29.21%、 2.25%、 2.32%、 98.01%和26.74%, Ti O_(2)品位和回收率分别为82.31%和89.60%。氟化氢铵浸出渣中的钙镁氟化盐可以被盐酸浸出分解。在盐酸质量分数为15%、液固比为10:1、酸浸温度为25℃、酸浸时间为4 h的条件下,Al、Ca、Mg、Si和Fe的浸出率分别为91.24%、93.15%、92.89%、1.29%和70.12%,获得最终产品二氧化钛品位超过94.36%,杂质氧化钙和氧化镁质量分数分别为0.32%和0.14%的沸腾氯化炉料。

黏土型锂矿氯化焙烧—酸浸提锂工艺试验研究

全球“碳中和”背景下,清洁能源的需求高涨,锂已成为战略资源。黏土型锂矿储量大且未大规模开发利用,未来将成为锂产品供应的重要来源。针对氧化锂含量为0.53%的黏土型锂矿,进行了氯化焙烧—酸浸试验研究。结果表明,在氯化钙用量18%、焦炭用量5%、焙烧温度800℃、焙烧60 min、10%硫酸浓度作浸出剂、液固比3:1、浸出时间120 min的条件下,获得了锂浸出率92.16%的技术指标。通过氯化焙烧,黏土型锂矿中的铝硅酸盐矿物转化为氯硅铝钙石,有利于锂的浸出;焙烧温度高于800℃时,氯硅铝钙石逐渐分解,硅灰石与硅酸铝钙生成,阻碍了锂的浸出。

铀水冶废渣氯化焙烧强化预处理提铀工艺研究

针对某铀水冶企业生产过程中产生的废渣(铀品位12.000%)浸出效果不佳、铀回收率低、至今尚未回收利用的现状,开展了低温氯化焙烧强化预处理-硝酸浸出试验研究,考察了焙烧温度、氯化钠添加量、焙烧时间、样品粒度等因素对铀浸出行为的影响。结果表明:在氯化钠添加量为样品量的20%、焙烧温度200℃、焙烧时间2 h、硝酸浓度7.4 mol/L、浸出温度80℃、浸出时间2 h的条件下,铀浸出率达到95.97%,与传统酸浸工艺相比,提高了19.52%。物相分析结果表明:经过氯化钠焙烧预处理后的废渣比表面积和孔隙率明显增加,促进了铀的解离,加大了铀与酸的接触面积,从而使得铀浸出率大幅提升。该工艺为铀水冶工艺废渣的回收利用提供了技术支撑。