新疆某硫化铅矿石选矿试验研究

新疆某硫化铅矿石铅品位为2.63%,锌品位为0.37%。铅主要赋存在方铅矿中,部分呈白铅矿、铅矾、铅铁矾形式存在。为给该矿石的开发利用提供依据,对其进行了选矿试验。结果表明:采用硫化钠+碳酸钠+硫酸锌为抑制剂(给入球磨机中),在磨矿细度为-0.074 mm占74%条件下,以25#黑药+乙黄药为捕收剂,经1粗2精3扫闭路浮选,得到的铅精矿铅品位为63.66%、回收率为92.71%,锌品位为1.02%。试验结果可以为该矿石选矿工艺流程的确定提供依据。

硫化矿氧化率测定问题的研究Ⅱ.硫化铅矿中氧化铅的分离测定

本文研究了有机酸在化学物相分析中的应用问题。提出用抗坏血酸－乙酸铵溶液作为分离测定硫化铅矿中氧化铅的选择性溶剂，对抑制方铅矿的部分溶解而产生的干扰，具有良好效果。

微波辐照下硫化铅矿常压溶解动力学

本文研究了微波辐照下FeCl_3溶液常压溶解硫化铅矿动力学。结果表明,微波辐照加热下铅溶解速率较传统加热下铅溶解速率快。考查了微波辐照下温度、FeCl_3浓度和粒度对铅溶解率的影响。根据非恒温动力学方程,求得反应活化能。表观总速率方程为:

新疆某伴生银硫化铅矿石选矿试验研究

新疆某硫化铅矿石中铅品位为13.35%、伴生银品位473g/t,铅主要赋存在方铅矿中,部分呈白铅矿、铅矾、磷氯铅矿形式存在。针对该矿石特点采用一次粗选、两次扫选、两次精选的闭路浮选试验,得到的铅精矿品位为铅66.12%、银1913g/t;铅精矿中铅、银回收率分别为:铅93.47%、银76.63%。试验结果可以为合理开发该矿石资源提供依据。

硫化铜矿与硫化铅矿闪速熔炼的差异分析

对硫化铜矿与硫化铅矿闪速熔炼过程的差异进行了,包括化合物物理化学性质、原料性质、冶炼原理与工艺、炉体结构方面的差异。研究表明,铜、铅精矿成分的差异决定了"两步炼铜"和"一步炼铅"的工艺差异,炉体结构也有所不同,但是二者采用闪速熔炼工艺均能取得较好的经济效益和环保效益。

硫化铅矿悬浮电解研究

叙述了四种硫化铅矿石悬浮电解直接产铅的试验研究结果,探讨了各种参数对过程和指标的影响,考察了、金、银及硫等的走向和回收方案。按10～25升电解槽综合试验指标,对于含铅30～60%的硫化铅矿,铅的溶出率在96%以上,所产阴极铅粉熔铸所得粗铅的品位>99%;硫转变为元素硫,全部入渣,转化率为90%。

某硫化铜铅矿选矿试验研究

针对某硫化铜铅矿石进行了选矿试验研究,试验采用混合浮选工艺流程,即将铜、铅一起浮出,再进行铜铅精矿的分离。通过一系列调整剂、捕收剂种类及用量条件试验,确定了较佳浮选流程。闭路试验最终获得的铜精矿和铅精矿指标如下:铜精矿品位24.59%,铜回收率86.11%;铅精矿品位52.81%,铅精矿回收率73.11%。

云南澜沧某铅矿选矿工艺试验研究

以云南澜沧某铅硫化矿为研究对象,在工艺矿物学研究基础上采用优先浮选流程抑锌浮铅。当原矿含铅2.46%、锌0.45%时,采用一次粗选、一次扫选、两次精选的简单流程,获得了铅品位为54.2%,回收率为83.45%的铅精矿。试验结果表明铅优先浮选流程适合该矿石的选别。

云南石屏某氧化铅矿石选矿试验

云南石屏某氧化铅矿石铅品位为3.14%,Ag含量为56.24 g/t,矿石中的主要铅矿物为方铅矿、白铅矿,铅矾等少量,银呈类质同象赋存于白铅矿、方铅矿中,主要有害元素为碳和砷。矿石在磨矿细度为-0.074 mm占91.1%的情况下,以石灰为矿浆p H调整剂、水玻璃为矿泥分散剂、丁铵黑药+乙硫氮为捕收剂,1粗2精流程浮选硫化铅矿物,以乙二胺磷酸盐+硫化钠为组合活化剂,丁基黄药为捕收剂,1粗2精1扫流程浮选氧化铅矿物,最终获得铅品位为53.21%、含银1 003.54 g/t、铅回收率为33.38%、银回收率为35.15%的硫化铅精矿,以及铅品位为30.88%、含银561.24 g/t、铅回收率为56.74%、银回收率为57.58%的氧化铅精矿,铅总回收率为90.12%、银总回收率为92.73%,较好地实现了铅、银矿物的回收。

含杂质方铅矿的电子结构和光学性质的第一性原理计算

采用基于密度泛函理论(DFT)的计算方法,研究银、铟和铊杂质对方铅矿晶体电子能带结构和光学性质的影响。结果表明,银和铊杂质使方铅矿的带隙变窄,而铟杂质相反。当方铅矿晶体中的铅原子被银和铊取代时,费米能级向低能方向移动;且在价带中出现了铟和铊的杂质能级。银和铊杂质没有改变方铅矿的半导体类型,铟杂质的掺入使方铅矿由直接带隙p型半导体转变为间接带隙n型半导体。这有利于电子的转移和提高方铅矿的电化学反应活性。光学性质的计算结果表明,银、铟和铊杂质使方铅矿的吸收带红移。特别是银杂质的存在使方铅矿的吸收系数增加了三个数量级。

硫砷铜矿和砷黝铜矿与不含砷的硫化铜矿物选择性氧化-溶解分离法

研究了用矿物选择性氧化法浮选分离硫化铜矿物 (辉铜矿、铜蓝和黄铜矿 )和含砷的硫化铜矿物 (硫砷铜矿和砷黝铜矿 )的可行性。试验表明 ,在弱酸性 pH介质中选择性氧化这些矿物表面 ,或碱性介质中先氧化再用络合剂选择性溶解表面氧化产物后 。

Electrochemical behavior of galena and jamesonite flotation in high alkaline pulp

In order to effectively separate galena and jamesonite and improve the recovery during the mixing flotation, the interaction mechanisms between the minerals and the collector of diethyl dithiocarbamate （DDTC） were investigated. Single mineral flotation test was organized to research the effect of pulp pH value on the flotation behavior of galena and jamesonite. Electrochemistry property of the interaction of these two minerals with DDTC was investigated by cyclic voltammetry and Tafel tests. Flotation test shows that the recovery of jamesonite in high alkaline pulp is strongly depressed by lime （Ca（OH）2）. The cyclic voltammetry and Tafel tests results show that the interaction between galena and DDTC is an electrochemical process. High pH value has little influence on the interaction between galena and DDTC, while it has great effect on jamesonite due to self-oxidation and specific adsorption of OH^- and CaOH^＋ on jamesonite surface. Non-electroactive hydroxyl compound and low-electroconductive calcium compounds cover the surface of jamesonite, which impedes electron transfer and DDTC adsorption, thus leads to very low floatability of jamesonite.

四川冕宁稀土矿床硫同位素地球化学

分析了四川冕宁大型稀土矿床硫同位素组成。结果显示成矿期脉石矿物重晶石与矿化期后硫化物(黄铁矿和方铅矿)的硫同位素组成明显不同,前者富集34S,其3δ4S为+1.8‰^+6.7‰,后者富集32S,其3δ4S在-10.9‰^-2.1‰之间,表明本区成矿流体中的硫和矿化期后富含硫化物流体中的硫具有不同的来源。矿区各种类型矿石中的重晶石均普遍遭受过风化作用,且δ34S随风化作用的增强而增加,暗示物理分馏效应是本区重晶石硫同位素组成差异的主要原因,重晶石风化过程实际上是一个贫32S、富34S过程,未风化重晶石的3δ4S与典型幔源硫的δ34S(0‰左右)相近,成矿流体中硫主要来源于地幔。矿石中重晶石风化作用过程中被淋滤的32S可能是矿化期后富含硫化物流体中的硫的重要来源,但有待深入研究。

有色金属火法冶炼用耐火材料及其发展(续)

（六）铅、锌火法洽炼用耐火材料 我国铅、锌矿储量丰富。炼铅、炼锌用的矿物有：方铅矿（PbS）、铅锌矿与闪锌矿（ZnS）。若为铅锌共生矿，则需采用能同时生产出粗铅与粗锌的工艺。若为单一的硫化铅矿则采用直接炼铅法。

Separation of sulfide lead-zinc-silver ore under low alkalinity condition

A complex lead-zinc-silver sulfide ore containing 2.98% Pb, 6.49% Zn and 116.32×10^-4 % Ag （mass fraction） from Yunnan Province, China, was subjected to this work. Research on mineral processing was conducted according to the properties of the lead-zinc-silver ore. Under low alkalinity condition, the lead minerals are successfully separated from the zinc minerals with new reagent YZN as zinc depressant, new reagent BPB as lead collector, CuSO4 as zinc activator and ethyl xanthate as zinc collector. The associated silver is mostly concentrated to the lead concentrate. With the process utilized in this work, a lead concentrate of 51.90% Pb with a recovery of 82.34% and a zinc concentrate of 56.96% Zn with a recovery of 81.98% are produced. The silver recovery in the lead concentrate is 80.61%. Interactions of flotation reagents with minerals were investigated, of which the results indicate that the presence of proper amount of Na2S can precipitate Pb^2＋ and has a sulfidation on oxidized lead minerals. The results also show that NazCO3 and YZN used together as combined depressants for sphalerite can signally improve the depressing effect of new reagent YZN on sphalerite.

Sodium sulfide leaching of low-grade jamesonite concentrate in production of sodium pyroantimoniate

Sodium sulfide leaching of a low-grade jamesonite concentrate in the production of sodium pyroantimoniate through the air oxidation process and the influencing factors on the leaching rate of antimony were investigated. In order to decrease the consumption of sodium sulfide and increase the concentration of antimony in the leaching solution, two-stage leaching of jamesonite concentrate and combination leaching of high-grade stibnite concentrate and jamesonite concentrate were used. The experimental results show that the consumptions of sodium sulfide for the two-stage leaching process and the combination leaching process are decreased by 20% and 60% compared to those of one-stage leaching process respectively. The final concentrations of antimony in the leaching solutions of both processes are above 100g/L.

Sulfidation mechanism of cerussite in the presence of sulphur at high temperatures

In this paper,sulfidation mechanism of cerussite in the presence of sulphur at high temperatures was investigated based on micro-flotation,X-ray powder diffractometry(XRD),electron probe microanalysis(EPMA)and X-ray photoelectron spectroscopy(XPS).The micro-flotation test results showed that flotation recovery of the treated cerussite increased to above 80%under a suitable flotation condition.It was found that the S/PbCO3 mole ratio and pH obviously affected flotation recovery.XRD analysis results confirmed that the cerussite was decomposed into massicot and then was transformed into mainly PbS and PbO·PbSO4 after sulfidation roasting.EPMA analysis results demonstrated that surface of the obtained massicot was smooth,but surface of the artificial galena was rough and even porous.Content of oxygen decreased,whereas content of sulphur increased with an increase in the S/PbCO3 mole ratio.XPS analysis results revealed that various lead-bearing species,including mainly PbS,PbSO4 and PbO·PbSO4,were generated at the surface.Formation of PbS was advantageous to flotation of the treated cerussite.Based on these results,a reaction model of the cerussite sulfurized with sulphur was proposed.

Electrochemical mechanism and flotation of chalcopyrite and galena in the presence of sodium silicate and sodium sulfite

The electrochemical mechanism involved in the selective separation of chalcopyrite from galena was investigated by flotation and electrochemical methods in the presence of sodium sulfite and sodium silicate,respectively,as a single depressant and their mixture as a combined depressant.Flotation tests revealed that the floatability of chalcopyrite was unaffected by depressants and its recovery remained constant(>80%)within the studied dosage range.Galena flotation was severely depressed with descending depressing order as follows:combined depressant﹥sodium silicate﹥sodium sulfite.Electrochemical analysis confirmed the high affinity of depressants on the galena surface,resulting in the formation of hydrophilic species,such as lead sulfite,lead sulfate,and lead orthosilicate.The oxidation of chalcopyrite surface and depressants did not exhibit any signals;conversely,the self-oxidation of chalcopyrite was depressed.The results of cyclic voltammograms well agreed with flotation results,demonstrating that chalcopyrite primarily reacted with the collector O-isopropyl-N-ethyl thionocarbamate and that galena mostly reacted with depressants.

Heavy Metals in a Sulfldic Minespoil: Fractions and Column Leaching

Fractions of various heavy metals in a sulfidic minespoil were investigated. Column leaching experimentwas also conducted to simulate 'acid mine drainage' (AMD) from the minespoil. The results show thatleaching of heavy metals from the minespoil was extremely significant during the initial water flushing.The amounts of heavy metals leached out dramatically reduced after leaching twice. It is worthwhile tonote that in this study, Zn, Mn, Fe, As and Ni in the first leachate exceeded the total amount of eachcorresponding water-extractable (1:5, soil:water) metal contained in the minespoil sample. This appears tosuggest that 1:5 water extraction did not allow accurate estimation of water-leachable concentrations of theabove heavy metals. This work has implications for the management of sulfidic minespoils. Acid drainageof great environmental concerns is likely to occur only during heavy rainfall events after substantial solubleand readily exchangeable acid and metals are accumulated in the minespoils. The slow-reacting fractionsother than water-soluble and readily exchangeable fractions may pose little environmental hazards. This isparticularly true for Pb, As and Ni.

Studies on Environmentally Friendly Leaching Processes in China

The newly developed green leaching processes for chromium, lead and gold extraction from ores or concentrates are described. The chromium is extracted from the iron chromite ore with fused sodium hydroxide at 500-550℃ as sodium chromate. The galena in lead sulfide concentrate is converted into lead carbonate in ammonium or sodium carbonate solution at 50-80℃ followed by the separation of lead carbonate formed from the unconverted sulfide ores by flotation. Gold associated with sulfide ore (such as pyrite and chalcopyrite) can be extracted into sodium thiosulfate solution without any pretreatment such as roasting, high pressure aqueous oxidation or bacteria pre-leaching.