黄铁矿烧渣氯化法回收有价金属

黄铁矿制酸烧渣含金、银、铜、铅和锌等有价金属,通过焙烧深度脱硫,焙砂细磨,氯化浸出,回收其中的金、银、铜、铅和锌.确定最佳焙烧时间和温度,浸出时间、温度、液固比、氯离子浓度、磨矿粒度等,并进行浸出渣再浸,以及浸出液多方案回收金银等的研究.结果表明,氯化浸出金属回收率高,废水经处理后能达标排放.     

某伴生金银铅锌矿的工艺矿物学与浮选工艺

某铅锌矿含锌2.96%、铅0.85%、铜0.13%、金1.01 g/t、银33.70 g/t,选厂仅将其当作铅锌矿分选,铜金银等伴生金属未获得综合回收。工艺矿物学研究表明,闪锌矿是锌的主要矿物,其嵌布粒度粗、嵌布关系简单,理论回收率可达97.84%;方铅矿是主要的铅矿物,其次是铅矾和白铅矿,方铅矿与闪锌矿嵌布关系密切,部分因氧化蚀变被铅矾包裹,影响铅的浮选,铅的理论回收率仅为88.32%;银主要赋存在游离银矿物和方铅矿中,理论回收率为72.81%;金主要以裸露金和硫化矿形式存在,可随银铅回收;铜主要赋存在黄铜矿中,黄铜矿与硫化矿嵌布关系密切,粒度细、品位低,可随方铅矿回收。基于矿石特性,制定“粗磨铅优先-锌硫混选-锌硫分离”工艺,以GW-221为铅铜金银捕收剂,强化金银回收。新工艺获得铅精矿含铅24.73%、铜3.71%,含银558.64 g/t、金13.52 g/t,回收率分别为77.69%、76.24%、44.26%和35.74%;锌精矿含锌48.99%,锌回收率为89.56%,铅锌铜金银均获得了综合回收;后续需强化游离银矿物和金的回收,以提高金银回收率。     

四川某硫化铅锌矿选矿工艺试验研究

针对四川某多金属硫化铅锌矿中方铅矿、闪锌矿嵌布粒度较细,选矿现场铅、锌分离效率低的问题,研究采用"优先选铅-铅粗精矿再磨精选-铅尾选锌"的工艺流程对其展开选矿试验。结果表明:在磨矿细度为-0.074 mm占75%、硫酸锌作抑制剂、25#黑药作捕收剂的条件下,经1粗2扫3精可获得铅品位为45.58%,铅回收率为84.11%,锌品位为5.43%,锌回收率为6.00%,银品位为861.72g/t,银回收率为81.84%的铅精矿;选铅尾矿采用石灰进行调浆,硫酸铜作活化剂,丁基黄药作捕收剂,经1粗2扫3精可获得锌品位为54.10%、锌回收率为87.14%的锌精矿。试验指标良好,该工艺既解决了铅锌分选效率低的问题,又为其他类似复杂多金属矿物的综合回收提供了一定的借鉴意义。     

某复杂难选铜铅锌多金属硫化矿选矿试验

辽宁葫芦岛地区某金、银品位较高的铜铅锌多金属硫化矿石结构构造复杂,铜、铅、锌分离难度较大。为高效开发利用该矿石,按优先混浮铜铅-混浮精矿铜铅分离-混浮尾矿抑硫浮锌的原则流程对该矿石进行了系统的选矿试验。结果表明,采用2粗1扫2精铜铅混浮、1粗2扫3精铜铅分离、1粗2扫2精选锌、中矿顺序返回的闭路流程处理该矿石,最终获得了铜、金、银品位分别为20.88%、2.37 g/t、1 808 g/t,铜、金、银回收率分别为85.72%、46.27%、22.46%的铜精矿,铅、金、银品位分别为63.13%、0.99 g/t、5 973 g/t,铅、金、银回收率分别为80.00%、19.57%、75.16%的铅精矿,锌、金、银品位分别为55.96%、0.35 g/t、37.80 g/t,锌、金、银回收率分别为84.21%、10.47%、0.72%的锌精矿,较好地实现了铜、铅、锌的分离回收。     

某铜铅锌矿工艺矿物学及选矿试验研究

某铜铅锌矿具有矿石嵌布关系复杂、嵌布粒度不均匀的特点,属于难选的复杂多金属硫化矿。该矿石中主要的回收对象为黄铜矿、方铅矿和闪锌矿,其铜、铅、锌的品位分别为0.20%、0.78%和1.64%。通过系统的工艺矿物学研究,全面地了解了该铜铅锌矿的矿石性质。最终确定采用"铜铅部分混合浮选—选铜铅尾矿活化选锌"的原则工艺流程。获得了含铜6.01%,回收率为77.54%,含铅21.26%,回收率达到88.85%的铜铅精矿;及含锌44.27%,回收率达到74.75%的锌精矿。金、银大部分富集在铜铅精矿中。含金、银分别为37.27 g/t、1539.50 g/t,较好地实现了铜、铅、锌、金、银有价元素的综合回收。     

某含银复杂铜铅锌多金属硫化矿浮选试验

本文介绍了某铜铅锌多金属硫化矿的浮选试验。矿石中有用矿物为闪锌矿、方铅矿、黄铜矿等,其他矿物有黄铁矿、磁黄铁矿、白铁矿以及少量菱锌矿、铜蓝、铅钒等矿物,并伴生大量的银。本试验采用了铜铅混合优先浮选-混合精矿再磨-铜铅分离-铜铅尾矿再选锌的优先浮选流程,并且综合回收了银。试验中探索了药剂的组合使用,在保证选矿指标的前提下,又节省了成本。。原矿经过一粗-两次混合精选-铜铅分离流程得到铜精矿品位25.65%,铜回收率73.25%,银回收率2.47%;铅精矿品位46.59%,铅回收率87.78%,银回收率82.23%。铜铅尾矿经过一粗二精一扫的流程,得到了锌精矿品位38.19%,锌回收率86.64%,银回收率7.44%,银的综合回收率达到92.14%。     

安徽某低品位铜锌多金属矿工艺矿物学研究

安徽某低品位铜锌多金属矿石中锌、铜、铁和硫的品位分别为1.93%、0.35%、7.85%、4.03%;金品位为1.01 g/t,银品位为13.05 g/t。为合理开发利用该资源,对该矿石开展了系统的工艺矿物学研究,查明了各有价金属元素的赋存状态、各目的矿物的嵌布特征和嵌布粒度,以及影响它们回收的最主要因素,为确定合理的工艺流程、实现其综合回收提供了理论依据。采用优先选铜、铜中矿再磨精选,锌硫混选再分选的原则流程,并将金、银富集于铜精矿中,实现了矿石中锌、铜、硫以及金、银的综合回收。      

两段焙烧工艺有望突破黄金冶炼渣再利用瓶颈问题

正日前,中国地质调查局郑州矿产综合利用研究所黄金冶炼渣中有价金属提取技术取得新进展,有望破解长期以来制约黄金冶炼渣再利用的瓶颈问题。黄金冶炼渣是采用焙烧预处理一氰化提金工艺提金后得到的尾渣。我国金矿资源富矿少、贫矿多、共伴生矿多的突出特点使得黄金冶炼渣中常含有金、银、铜、铅、锌、铁等金属元素,具有较高的综合回收价值。郑州综合所科研人员在充分总结前人工作经验的基础上进行了大量试验探索,最终提出采用两段焙烧工艺综合回收黄金冶炼渣中金、银、铜、铅、锌、铁等金属。     

西藏某高硫铜铅锌硫化矿浮选试验研究

西藏某铜铅锌硫多金属硫化矿,矿物嵌布粒度细、共生关系复杂,且含硫量较高。采用铜铅混合浮选→铜铅分离→尾矿抑硫浮锌浮选工艺流程。最终获得铜精矿铜品位28.22%、回收率85.29%,铅精矿铅品位57.49%、回收率85.61%,锌精矿锌品位44.17%、回收率62.96%,银在铜、铅精矿中的总回收率达到89.7%,实现了矿物的综合回收。     

日本大江锰矿选矿厂工艺流程考察与分析

一、概况大江锰矿选矿厂位于北海道积丹半岛中部,属余市郡,已有90年的历史。生产能力为月产一万吨锰精矿。除锰外,尚综合收回了金、银、铜、铅、锌及硫化铁等六种金属。全厂职工25名。该矿属浅热液裂隙充填矿床,矿物组成如表1所示。金、银赋存于硫化物中,各种硫化物的粒度在50微米以上。菱锰矿嵌布粒度较粗,但     

从某金精矿中回收金银铜铅锌的试验研究

山西某复杂多金属硫化矿石采用混合浮选获得的金精矿含Au34.22g/t、Ag904.4g/t、Pb8.78%、Cu1.32%、Zn3.35%,混合精矿直接外销,但其铜、铅、锌基本不予计价,造成了有价金属的流失。采用浮选精矿氰化浸金—氰化渣铅、铜、锌依次优先浮选流程,获得金总回收率96.60%、银95.51%、铅85.39%、铜72.37%、锌83.51%,实现了高效综合回收该矿石中的有价元素,经济效益和社会效益显著。     

Thiosulphate leaching of gold in the presence of ethylenediaminetetraacetic acid (EDTA)

The addition of low levels of ethylenediaminetetraacetic acid (EDTA) in the ammoniacal thiosulphate gold leach system lowered the catalytic cupric/cuprous redox equilibrium potential, hence the mixed solution potential and reduced the consumption of thiosulphate. In the leaching of pure gold, gold dissolution was enhanced in the presence of EDTA at a relatively low concentration, but excessive EDTA decreased gold dissolution. Raman analysis of the leached gold foil indicated that the stabilisation of thiosulphate by EDTA decreased the formation of the passivation layers of elemental sulphur and copper sulphide at the gold surface. In the leaching of a sulphide ore, the leaching kinetics and overall extractions of gold and silver were enhanced substantially, while the consumption of ammonium thiosulphate was reduced from 9.63 kg/t to 3.85 kg/t in the presence of 2.0 mM EDTA after 24 h leaching. This beneficial effect became more pronounced at a higher EDTA concentration. The enhanced gold and silver extractions by EDTA were attributed to the increase in the dissolution of gold and silver bearing sulphides, the stabilisation of copper and thiosulphate in leach solutions, the prevention of leaching passivation and the decrease in the interference of foreign heavy metal ions. The use of EDTA in the ammoniacal thiosulphate leaching system makes it practical to achieve satisfactory gold extraction over extended periods of leaching under low reagent concentrations, where the consumption of thiosulphate is low.     

Interactions between sulphides and manganese dioxide in thiosulphate leaching of gold ores

The interactions between manganese dioxide and pyrite or chalcopyrite were investigated in both the ammoniacal and ammoniacal thiosulphate leaching systems. Pyrite and chalcopyrite dissolved at enhanced rates in the presence of manganese dioxide in both ammoniacal and ammoniacal thiosulphate solutions. The interactions between manganese dioxide and sulphides were also applied in the thiosulphate leaching of gold from a pyrite concentrate and a sulphide ore. The addition of a small amount of manganese dioxide in the thiosulphate leaching of the sulphide bearing gold ores improved both the kinetics and the overall gold extractions without much impact on thiosulphate consumption.A pre-treatment process with copper ammoniacal solutions enhanced gold extractions from the sulphide bearing ores and reduced thiosulphate consumption to a large extent. The sulphides partially degraded in the pre-treatment process likely exposing gold to leach solutions. SEM images of the surface corrosion of pyrite and chalcopyrite in contact with manganese dioxide showed that the pre-treatment with manganese dioxide enhanced the degradation of the sulphide matrices, thus achieved a better gold leaching from the sulphide bearing ores.     

某低品位铅锌硫化矿浮选试验研究

某硫化铅锌矿含铅锌原矿品位低、嵌布粒度细、伴生关系复杂。通过多种方案的比较,采用优先浮选抑锌浮铅的选别流程,试验采用乙硫氮作为优先选铅的捕收剂,石灰作为调整剂以及黄铁矿的抑制剂,硫酸锌和亚硫酸钠作为闪锌矿的抑制剂,之后利用硫酸铜作为闪锌矿的活化剂,用丁基黄药作为捕收剂来实现铅与锌的有效分离。试验获得铅精矿含铅51.00%、铅回收率86.63%、含银518 g/t、银回收率47.41%,锌精矿含锌51.20%、锌回收率85.27%、含银234 g/t、银回收率38.38%。     

硫化铜铅锌矿石伴生金银的综合回收

硫化铜铅锌矿石伴生金银的综合回收,在我国有重要经济意义,针对不同矿石的特性,采用相应的选别流程和技术措施,均能取得较好的综合回收技术指标和经济效益.     

某银精矿氰化尾渣中铅和锌的浮选回收试验

某浮选银精矿经常温常压碱式氧化预处理-氰化浸出金、银后的氰化尾渣中,含有铅、锌、金、银等有价元素,金属矿物主要为黄铁矿、方铅矿、闪锌矿和毒砂,并含有少量含银矿物。该尾渣粒度很细,含泥量大,铅、锌矿物被氧化,使铅、锌的选别回收受到影响。对该尾渣进行铅、锌的浮选试验,结果表明,铅矿物不能得到有效富集而形成铅精矿,但可以获得锌品位为55.62％,锌回收率为66.15％的合格锌精矿,锌精矿中金、银品位为66.94 g/t和538.9 g/t,金、银回收率为47.96%和25.67%。     

云南某含金银低品位硫化铅锌矿浮选试验

云南某含金银硫化铅锌矿石铅品位为0.77%，锌品位为2.13%，并且伴生大量金、银等贵金属，金、银的嵌布粒度微细。为给该矿石开发利用提供依据，采用优先浮选硫化铅，选铅尾矿再选锌的优先浮选流程进行试验。结果表明：在磨矿细度为-0.074 mm占81.33%，以碳酸钠为pH调整剂，以硫酸锌+焦亚硫酸钠为抑制剂，以乙硫氮+3418A为捕收剂，经过2粗3精1扫选铅，选铅尾矿以硫酸铜+氯化铵为活化剂，以丁基黄药为捕收剂，经1粗2精1扫流程选锌，获得了铅精矿铅品位50.36%、金品位28.79 g/t、银品位965.47 g/t、铅回收率82.41%、金回收率77.18%、银回收率78.69%，锌精矿锌品位41.21%、锌回收率87.45%的指标，实现了矿石中有用金属的高效回收。     

某复杂铜铅锌多金属硫化矿选矿试验研究

根据某铜铅锌矿矿石中铜、铅、锌等硫化矿物嵌布关系复杂、嵌布粒度极不均匀的特点,采用"铜铅混合浮选—混合精矿再磨—铜铅分离—混合浮选尾矿选锌"的工艺流程及合理的药剂制度,闭路试验获得良好的铜、铅、锌选矿技术指标,同时,矿石中的伴生银也得到了较好回收,铜、铅、锌及银的回收率分别达到65.98%、88.83%、85.31%、84.98%。     

Zinc and silver recoveries from zinc–lead–iron complex sulphides by pressure oxidation

Acid pressure oxidation followed by cyanide leaching of the residue is a promising process for the treatment of complex sulphides and the recovery of precious metals along with the base metals will improve the economy of the process. However, silver is incorporated into the jarosite specie during the pressure oxidation and cyanide leaching of the residue yields very low silver extraction.In this work, iodide was added to the pressure oxidation of zinc–lead–iron complex sulphides to prevent the deportment of silver ions into the jarosite phase. At low temperature range (110–130 °C), the silver ions were completely sequenced into the silver iodide phase because of the fast precipitation kinetics of silver iodide and its stability at low temperatures. The leaching of the residue in cyanide solution yielded high silver extraction (above 90%).Silver extraction from the residue decreased when the pressure oxidation was conducted at high temperatures (140–150 °C). At this temperature range, the enhanced stability and the precipitation kinetics of the jarosite specie posed a challenge by competing (with iodide) for silver ions. This competition was minimised by using moderately high initial acid for the pressure oxidation.High zinc extraction was achieved during the pressure oxidation. Also, there were appreciable iron precipitation and acid neutralisation of the slurry. The resulting pregnant solution is suitable for zinc recovery by electrowinning and the residue can be leached for silver and gold extraction.     

Limits of the CIL circuit in copper–gold plants

This study examined the performance of the CIL (Carbon-in-Leach) circuit at Telfer, a copper–gold plant treating porphyry copper deposits containing gold associated with both copper and iron sulphides, with an objective to identify factors normally limiting the gold recovery in the CIL circuit in the presence of a small amount of copper after copper flotation, and then propose a means to improve it. Diagnostic leaching assessment and mineralogical analysis by MLA revealed that the occlusion of gold by other minerals and the fine grain size of gold associated with them may be the contributing factors to the low gold recovery in the CIL circuit. Fine grinding of the CIL feed increased gold recovery significantly from the leaching process. However, it is interesting to find that fine grinding increased the amount of released copper ions which complex with cyanide resulting in significantly higher cyanide consumption. It is therefore proposed that regrinding of the CIL feed followed by copper flotation is an appropriate pre-treatment method for the CIL circuit.