Selective recovery of lead from zinc oxide dust with alkaline Na_2EDTA solution

The selective recovery of lead from the zinc oxide dust using an alkaline Na2EDTA solution was investigated. The effects of temperature, leaching time, Na2EDTA concentration and initial NaOH concentration on the leaching rates of lead and zinc were studied. The following optimized leaching conditions were obtained： liquid-to-solid ratio 5：1 mL/g, stirring speed 650 r/min, Na2EDTA concentration 0.12 mol/L, initial NaOH concentration 0.5 mol/L, leaching temperature 70 ℃, leaching time 120 min. Under the optimized conditions, the average leaching rates of lead, zinc, fluoride and chloride are 89.92%, 0.94%, 62.84% and 90.02%, respectively. The filtrate was used to electrowin lead powders. The average current efficiency of electrowinning is about 93% and lead content is higher than 98% under the conditions of temperature of 60 ℃, current density of 200 A/m2, H3PO4 concentration of 1.5 g/L, and lead ion concentration of above 5 g/L. The consumption of Na2EDTA and the direct current are about respectively 0.218 kg and 0.958 kW·h for per kilogram of lead powder.

Sulphidization flotation for recovery of lead and zinc from oxide-sulfide ores

A new flowsheet was developed to recover the valuable minerals from oxide or oxide-sulfide ores of lead and zinc. The flowsheet consisted of flotation of sulfide minerals, desliming and sulphidization-flotation of oxide minerals. The corresponding reagent system and techniques to the flowsheet were investigated. Batch and continuous tests show that the dosage of sodium sulfide, temperature, and collector type are main affecting factors on the recovery of smithsonite and cerussite. For the flotation of cerussite, there is an appropriate dosage of sodium sulfide at which the recovery reaches its maximum value. The required sodium sulfide for smithsonite flotation is higher than that for cerussite and the recovery of smithsonite flotation increases with the dosage of sodium sulfide at low level and becomes insensitive at high dosage. The appropriate temperature for smithsonite and cerussite flotation is found to be 2540℃. Amines are found to be the effective collectors for the flotation of smithsonite after sulphidization. Investigation also shows that desliming prior to sulphidization-flotation is essential to the effective recovery of smithsonite and cerussite, and the desliming process of two-stage hydrocyclon is well feasible and effective for the treatment of lead-zinc oxide ores. A further treatment on the cerussite flotation concentrate by shaking table is proposed to obtain higher lead grade.

Investigation on the deactivation cause of lead-zinc double oxide for the synthesis of diphenyl carbonate by transesterification

The deactivation cause of lead-zinc double oxide for synthesis of diphenyl carbonate （DPC） by transesterification of dimethyl carbonate （DMC） with phenol has been investigated. X-ray diffraction （XRD）, X-ray photoelectron spectroscopy （XPS）, infrared spectroscopy （IR）, thermogravimetry analysis （TG）, atomic absorption spectroscopy and elementary analysis are employed for the catalyst characterization. The results show that, the formation of Pb4O（OC6H5）6 through the reaction of phenol and lead species in the catalyst leads to the crystal phase change of active component and serious leaching of lead, which is the cause of the catalyst deactivation. In addition, the composition of the leached lead is ascertained to be a mixture of Pb4O（OC6H5）6 and PbO with the weight percentage of 62.7% and 37.3%, respectively.

Study on experiment and mechanism of thermal dissolved sulfuration of low grade lead-zinc oxide ore in lanpin

The thermal dissolved sulfuration technology is brought forward and performed based on the characteristic of low grade lead-zinc oxide ore in lanpin. Using sulfur as the sulphidizing agent in the experiment, the oxides in the sandstone and ignimbrite are changed into sulfides. The disproportionation reaction of sulfur in a solution is confirmed as 4S＋3H2O=2S^2-＋S2O3^2--＋6H^＋. The dynamics process is studied and the first-order reaction rate equation -1n（1-a）=ktt is obtained. The effects of the reactive products, stirring speed, dosage of sulfuration agent, value of pH and sulphidizing temperature on the sulfuration of oxide ore are investigated. The results indicate that the reactive apparent activation energy is 100.8 kJ/mol and the sulfuration ratio of lead-zinc oxide ore reaches 60% under the conditions of pH 5.9-7.5, the sulfuration temperature of 130 ℃, sulfuration time of 180 min and the stirring speed of 800 r/min.

The Processing Mineralogy for Lead and Zinc Oxide Ore in Sichuan

Using variety of modern testing methods, the processing mineralogical characteristics for a lead and zinc oxide ore in Sichuan were studied systematically. The chemical analysis result shows that the lead and zinc oxide content exceeding the minimum industrial grade and iron ore, total iron content reaches a minimum industrial grade and associated with gold and silver; The mineralogical analysis result shows that lead and zinc mineral composition and configuration are very complexity. The zinc minerals and zinciferous minerals are sphalerite, hemimorphite, Smithsonite, Hydrozincite, zinc chlorite, limonite, zinc dolomite and zincocalcite; lead minerals and plumbiferous minerals are mainly galena, cerussite, anglesite, limonite and Coronadite; The minerals disseminated grain size are very fine and mineral dissemination characteristics are very complex ; expected theoretical recoveries for lead and zinc were 72% and 67% respectively. The results of this study provide basic data and theoretical basis for ore dressing.

New Constraints on the Environmental Impact of Iron Oxides and Lead-Zinc Mines in Red Sea Hills,Egypt

Iron oxides and lead-zinc deposits in host rocks located in ten drainage basins in the coastal plain of the Red Sea,Egypt,have been subjected to important mining activities.The drainage basins were analyzed to estimate the transportations of these minerals. Fourteen soil samples and fifteen plant samples were collected from sites located in the basin and also in neighboring area and chemically analyzed.In

Extracting indium and preparing ferric oxide for soft magnetic ferrite materials from zinc calcine reduction lixivium

A hydrometallurgical process for indium extraction and ferric oxide powder preparation for soft magnetic ferrite material was developed. Using reduction lixivium from high-acid reductive leaching of zinc oxide calcine as raw solution, copper and indium were firstly recovered by iron powder cementation and neutralization. The recovery ratios of Cu and In are 99% and 95%, respectively. Some harmful impurities that have negative influences on magnetic properties of soft magnetic ferrite material are deeply removed with sulfidization purification and neutral flocculation method. Under the optimum conditions, the content of impurities like Cu, Pb, As, Al in pure Zn-Fe sulfate solution are less than 0.004 g/L, but those of Cd, Si, Ca and Mg are relatively high. Finally, thermal precipitation of iron is carried out at 210 ℃ for 1.5 h. The precipitation ratio of Fe is 93.33%. Compared with the quality standard of ferric oxide for soft magnetic ferrite materials, the contents of Al and Mg in obtained ferric oxide powder meet the requirement of YHT1 level of ferric oxide, and those of Si, Ca meet the requirement of YHT3 level of ferric oxide. XRD and SEM characterizations confirm that the obtained sample is well-dispersed spindle spherule with regular a-Fe2O3 crystal structure. The length-to-diameter ratio ofa-Fe2O3 powder is （3-4）：1 with an average particle size of 0.5 μm.

Metallogeny of the Baiyangping Lead-Zinc Polymetallic Ore Concentration Area, Northern Lanping Basin of Yunnan Province, China

The Lanping Basin in the Nujiang-Lancangjiang-Jinshajiang （the Sanjiang） area of northeastern margin of the Tibetan Plateau is an important part of eastern Tethyan metallogenic domain. This basin hosts a number of large unique sediment-hosted Pb-Zn polymetallic deposits or ore districts, such as the Baiyangping ore concentration area which is one of the representative ore district. The Baiyangping ore concentration area can be divided into the east and west ore belts, which were formed in a folded tectogene of the India-Asia continental coUisional setting and was controlled by a large reverse fault. Field observations reveal that the Mesozoic and Cenozoic sedimentary strata were outcropped in the mining area, and that the orebodies are obviously controlled by faults and hosted in sandstone and carbonate rocks. However, the oreforming elements in the east ore belt are mainly Pb-Zn -Sr-Ag, while Pb-Zn-Ag-Cu-Co elements are dominant in the west ore belt. Comparative analysis of the C-O-Sr-S-Pb isotopic compositions suggest that both ore belts had a homogeneous carbon source, and the carbon in hydrothermal calcite is derived from the dissolution of carbonate rock strata; the ore- forming fluids were originated from formation water and precipitate water, which belonged to basin brine fluid system; sulfur was from organic thermal chemical sulfate reduction and biological sulfate reduction; the metal mineralization material was from sedimentary strata and basement, but the difference of the material source of the basement and the strata and the superimposed mineralization of the west ore belt resulted in the difference of metallogenic elements between the eastern and western metallogenic belts. The Pb-Zn mineralization age of both ore belts was contemporary and formed in the same metaliogenetic event. Both thrust formed at the same time and occurred at the Early Oligocene, which is consistent with the age constrained by field geological relationship.

甘肃某细粒铅锌矿工艺优化试验研究

甘肃某铅锌矿中铅品位为0.74%,锌品位为3.72%,为了实现该矿产资源的高效回收利用,针对铅锌矿物嵌布关系紧密,粒度较细,造成铅锌精矿互含量较高,影响浮选指标的问题,在原矿性质研究的基础上,进行了4种不同方案的工艺流程对比试验。试验结果表明:采用铅锌优先浮选—铅、锌粗精矿均再磨的工艺流程,最终获得的铅精矿铅品位为58.42%,铅回收率为79.95%,锌精矿锌品位为56.27%,锌回收率为93.05%。相较于方案2,锌精矿锌品位提高了3.03个百分点,锌回收率提高了2.75个百分点;相较于方案1现场生产工艺,铅、锌主品位提高,铅、锌互含量显著降低,锌回收率提高了4.15个百分点。

铜熔炼渣的火法回收工艺与铅锌锑富集规律

铜精矿火法冶金后产生的熔炼渣经浮选回收铜后,尾渣含Pb、Zn和Sb,仍具有一定回收经济价值。浮选法对铜熔炼渣中Pb、Zn和Sb的回收率低,而火法熔炼通过高温和碳热还原可以有效分离这些元素。对直接还原熔炼、造锍还原熔炼和氧化-还原熔炼3种典型的火法回收工艺进行了对比中试试验。结果表明:火法回收工艺可达到理想的铜回收率,但不同工艺对Pb、Zn和Sb的回收效果不同;直接还原熔炼提高了Pb的回收率;造锍还原熔炼提高了Zn的回收率;熔炼法需消耗大量还原剂和能源,经济性需慎重考虑。火法熔炼适合作为铜回收的备用工艺进行研究。

铅锌氧化物还原锌及动力学研究

本文以铅锌烧结矿为原料,研究了还原温度、还原时间、配炭量、石灰添加量、料球直径等条件对氧化锌还原过程的影响。结果表明:锌还原率随着还原温度的升高、还原时间的延长、配炭量的提高、料球直径的减小而增加,石灰添加量对锌还原率影响较小。在还原温度为1423 K、还原时间为20 min、3.0倍理论配炭量、料球直径为10 mm的较优条件下,还原渣锌含量为1.98%,锌还原率为96.22%。在1273~1423 K的温度范围内,氧化锌还原过程受化学反应控制,还原反应的表观活化能为264.551 kJ/mol。

某铅锌浮选厂颗粒物浓度分布及其影响研究

铅锌浮选厂在生产过程中产生大量颗粒物,污染工作环境。以某铅锌浮选厂为研究对象,对厂房内、外颗粒物浓度分布情况进行测试分析,并对其污染状态及健康影响进行深入研究。结果表明:粒径小于10μm颗粒物其厂房内数量浓度高于厂房外,且数量浓度随粒径变小而增加,厂房内、外颗粒物的数量浓度比平均值大多大于1。PM_(10)、PM_(2.5)和PM1的厂房内质量浓度显著高于厂房外,厂房内PM_(10)、PM_(2.5)分别超过一级标准5.333倍和6.313倍,超过二级标准2.614倍和2.134倍。厂房内对厂房外颗粒物浓度有较大影响,颗粒物的分布和浓度也有所不同,环境因素对粉尘污染有较大影响。研究结果可为矿山环境颗粒物的有效控制提供数据参考价值。

“生化+高级氧化”组合工艺处理铅锌选矿废水技术研究

铅锌矿多采用高碱快速浮选工艺进行选矿,其产生的选矿废水水质特征复杂,具有COD、重金属、固体悬浮物,以及盐分浓度高,pH值波动大的特点,废水处理难度大,对环境和人体健康具有巨大的威胁。常规选矿废水处理方法难以达到最佳的废水处理效果,效果不稳定,无法保证选矿废水达标排放。为实现铅锌选矿废水处理稳定达标排放,本文首先对某铅锌矿选矿废水水质特征开展调研,针对废水特性,采用“生化+高级氧化”组合工艺,对生化反应时间、不同高级氧化方法、药剂投加量等各工艺条件参数进行优化并验证,核算废水处理运行成本。试验结果表明:“生化+臭氧催化氧化”组合工艺处理铅锌选矿废水可实现出水稳定达到《铅、锌工业污染物排放标准》(GB 25466—2010)中水污染物特别排放限值要求。最佳工艺参数为:SBR工艺缺氧好氧反应时间为2 h∶10 h,臭氧催化剂体积为反应器体积的30%,臭氧投加量为90 mg/L,直接运行吨水成本为3元。该工艺可有效解决铅锌矿选矿废水达标处理难题,为同类型铅锌矿山选矿废水处理提供借鉴。

高浓度快速浮选技术在铅锌硫化矿浮选中的研究与应用

采取高浓度快速浮选技术,解决了在不改变选矿工艺的情况下原矿处理量增大对生产指标的影响。技术改造后,铅、锌金属回收率分别提高了1.75%和1.61%,伴生银回收率提高了8.60%,药剂成本下降了1.36元/t,降低了选矿生产用水量,年创造经济效益约511.80万元,保护了生态环境。

广西某硫化铅锌矿浮选分离试验研究

以广西某多金属硫化铅锌矿为研究对象,针对其“锌高铅低”、铅锌分离困难的问题,采用新型高效抑制剂JFR-1,经一粗两精两扫铅锌混合浮选和一粗两精两扫铅锌分离,最终获得了Pb品位21.00%、回收率72.82%的铅精矿和Zn品位6.96%、回收率94.21%的锌硫化矿,尾矿中Pb、Zn品位及对应的回收率分别为0.04%、0.13%及10.85%、4.79%,实现了铅锌的富集和分离。

铅锌选矿废水处理工艺研究进展

有色金属选矿废水中残留的化学药剂、固体悬浮物、重金属离子及COD往往都严重超标,直接排放不仅对矿区周边环境造成严重污染,还会导致水资源浪费,降低矿企经济效益。铅锌矿选矿厂在我国大量存在,每年产生大量的铅锌选矿废水,只有得到妥善处理,才能保障矿企可持续发展。本文以铅锌选矿废水处理工艺为对象,详细综述了当前沉降法、化学沉淀法、吸附法、氧化法、生物法、人工湿地法等不同方法处理铅锌选矿废水的现状,对比剖析了不同方法存在优劣势,为后续更经济、高效处理复杂铅锌选矿废水提供借鉴。

含砷银铅锌矿工艺矿物学研究

以呼伦贝尔某含砷银铅锌矿石为研究对象,利用化学多元素分析、物相分析、光学显微镜、MLA等多种分析手段对矿石进行工艺矿物学研究。结果表明:矿石中有用元素为铅和锌,质量分数分别为1.63%和1.72%,伴生贵金属银含量约46.31 g/t,有害元素As质量分数为0.34%;矿石中金属矿物主要为黄铁矿、闪锌矿、毒砂、磁黄铁矿、方铅矿,脉石矿物主要为石英、云母、长石、绿泥石、方解石等;矿石矿物粒径中等,+0.105 mm的方铅矿、闪锌矿和毒砂的占比分别为73.56%、79.2%、83.43%,属中粒粒级嵌布范畴,方铅矿和闪锌矿与其他矿物整体以毗邻镶嵌为主,少量方铅矿和闪锌矿沿黄铁矿、磁黄铁矿、毒砂、黄铜矿等矿物的裂隙或粒间充填交代,构成较为复杂的镶嵌关系,可能造成方铅矿-毒砂连生体和闪锌矿-毒砂连生体进入到铅精矿或锌精矿,导致铅精矿、锌精矿中含砷偏高。

某铅锌冶炼厂废水COD氧化降解实验研究

随着铅锌冶炼技术持续进步,其产生的废水成分愈加复杂,尤其国家近年来加强对COD(化学需氧量)排放总量的管控,企业外排废水COD的提标深度处理急需高效、绿色处理技术。针对当前COD处理需求,本文采用新型催化高级氧化技术(KYCAOP)处理铅锌冶炼废水,重点对比了其与常用氧化去除技术(如芬顿氧化法和次氯酸钠氧化法)对废水中COD的处理效果,并探索了KYCAOP的最佳使用条件。结果表明,对同一批次冶炼废水(原水COD 151 mg/L),芬顿氧化法和次氯酸钠氧化法的COD去除率仅有47.2%和22.1%,而KYCAOP的去除率高达75%以上。此外,KYCAOP的反应速度快,反应20 min后出水COD浓度可稳定在40 mg/L以下,证明此技术具有深度处理铅锌冶炼废水中COD的应用潜力。

某复杂铅锌矿锌精矿降碳试验研究

针对某复杂铅锌矿选矿所得锌精矿含碳量高的问题,进行了锌精矿降碳试验。采用可溶性淀粉为碳抑制剂,闭路试验获得了锌品位48.71%、含碳0.86%、锌回收率90.28%的锌精矿;与原工艺指标相比,在提高锌精矿品位及回收率的同时,碳含量降至1%以下,降碳效果显著。

高锌含量铅锌氧化渣直接还原试验研究

铅锌氧化渣熔融还原是得到金属铅和锌的重要步骤,但氧化渣中锌含量升高和铅含量降低会导致熔化温度增加,使得还原工艺条件发生较大变化。本文对高锌含量铅锌渣(Zn≈25%)的还原工艺条件和渣型制度进行了研究,主要考察了还原煤量、还原温度、还原时间、Fe/SiO_(2)、CaO/SiO_(2)对铅锌氧化渣中Pb、Zn、Cu回收率的影响。试验结果表明,铅锌渣Pb、Zn、Cu金属回收率随着还原煤量、还原温度、还原时间的增加而增加,但过高的煤比和还原温度不利于Pb、Cu回收率的提高;铅锌渣Pb、Zn、Cu金属回收率随着Fe/SiO_(2)和CaO/SiO_(2)的增加而增加,但Fe/SiO_(2)增加至0.78,CaO/SiO_(2)增加至0.8时,继续增加Fe/SiO_(2)和CaO/SiO_(2)将降低Pb、Zn、Cu金属回收率。试验用铅锌渣的适宜还原条件为煤比1.2~1.4、还原温度1623~1673 K、还原时间90~120 min、Fe/SiO_(2)=0.78~1.17,CaO/SiO_(2)=0.5~0.8。在上述工艺条件下,铅锌氧化渣中Pb、Zn、Cu回收率分别可达84.01%、94.51%、85.8%以上。