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浮选分级—抑制及再活化硫化矿混合精矿的分离浮选研究 总被引:1,自引:0,他引:1
对从苏州高岭土尾矿中用浮选法得到的硫化矿混合精矿进行了硫及铅锌混合精矿的分离浮选试验研究。在不磨矿的条件下,采用浮选分级-抑制及再活化浮选方法获得了铅、锌品位分别为19.95%、30.1%,回收率分别为82.00%、81.29%的铅锌混合精矿和硫品位和回收率分别为52.49%、75.5%的硫精矿。 相似文献
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某富银铅锌多金属矿, 银、铅、锌的品位分别为225 g/t、3.26%、1.14%,所含矿物以硫化矿为主,另含有少部分氧化矿。为更好的回收细粒嵌布的银矿石,本文通过选用BK809作为硫化银铅捕收剂、采用“硫化银铅浮选—锌硫混合浮选再分离—锌硫混浮尾矿再选氧化铅”工艺、并对硫化银铅精矿进行再磨处理,闭路试验获得了以下指标:铅总精矿中金品位3.56g/t、金回收率49.94%、银品位3777g/t、银回收率71.22%、铅品位55.57%、铅回收率71.73%;锌精矿中锌品位53.60%、锌回收率69.46%;硫精矿中硫品位40.90%、硫回收率45.79%,实现了矿石综合回收。 相似文献
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云南某铅锌矿选矿工艺试验研究 总被引:6,自引:0,他引:6
对云南某黄铁矿型含银铅锌多金属硫化矿选别的工艺流程及药剂条件进行了工艺试验研究。试验结果表明, 用优先浮选流程及所选药剂条件处理该试料可获得铅品位57.33%、铅回收率94.08%、银品位2 201.72 g/t、银回收率83.14%的铅精矿;锌品位48.28%、锌回收率88.38%的锌精矿和硫品位45.09%、硫回收率77.39%的硫精矿。 相似文献
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内蒙古某铜铅锌硫化矿石中铜、铅、锌含量分别为0.26%、0.72%、4.60%,硫、砷含量分别为13.14%、2.49%,属于高硫高砷难处理硫化矿石。为实现矿石中铜、铅、锌、硫的有效回收,避免传统高碱法带
来的一系列问题,开展了铜铅混浮、磁选脱硫、锌浮选条件试验研究。在此基础上,经“铜铅混浮(粗精矿再磨精选)—铜铅混合尾矿磁选脱硫—锌浮选”全流程闭路试验,最终可获得铜、铅、银品位分别为9.27%、
40.53%、4 397.76 g/t,铜、铅、银回收率分别为59.22%、88.93%、74.05%的铜铅混合精矿,及锌品位45.94%、锌回收率93.10%的锌精矿,选别指标良好,实现了铜、铅、锌及伴生银的有效回收,降低了精矿中有害
杂质砷的含量。 相似文献
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针对某低品位铜铅锌硫化矿,采用铜铅顺序优先浮选-锌硫混合浮选再分离工艺进行了浮选分离试验研究。选用高效选择性铜捕收剂BK916和铅捕收剂BK906进行了铜铅顺序优先浮选试验研究,并在锌硫分离试验研究中,利用环保型抑制剂BD和石灰的组合作用,有效抑制了锌硫混合精矿中的黄铁矿,获得了铜品位20.68%、铜回收率72.98%的铜精矿,铅品位61.38%、铅回收率73.57%的铅精矿,锌品位46.31%,锌回收率73.17%的锌精矿和硫品位48.54%的硫精矿。 相似文献
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对某黝铜矿型铜铅锌多金属矿进行了选矿试验研究。结合矿石性质及一系列探索试验研究结果,最终采用铜铅混浮-混浮精矿再磨-铜铅分离-混浮尾矿浮锌-锌尾矿浮硫的工艺回收该矿中的铜、铅、锌和硫,闭路试验获得了铜精矿铜品位18.25%、铜回收率73.88%,铅精矿铅品位59.91%、铅回收率82.06%,锌精矿锌品位50.15%、锌回收率91.82%,硫精矿硫品位49.96%、硫回收率74.14%。通过所确定的工艺流程与药剂制度对选矿工艺进行了改造,改造后铜精矿品位提高6.51个百分点,铜回收率提高8.68个百分点,铅、锌回收率分别提高6.59和2.36个百分点。 相似文献
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刘万峰 《有色金属(选矿部分)》2013,(3):14-17
通过对某银铅锌多金属矿进行系统的浮选试验研究,确定了采用"硫化银浮选—脱泥—氧化铅浮选—氧化锌浮选"的工艺流程进行有价金属矿物回收。闭路试验可以获得良好指标:银品位为2 300 g/t、回收率为40.87%的银精矿,铅品位为52.08%、回收率为64.45%的铅精矿,锌品位为32.00%、回收率为40.16%的锌精矿,银的总回收率为64.56%。浮选尾矿和矿泥中损失的银可以通过氰化浸出回收,银的浸出率为68.15%。 相似文献
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A complex process for the recovery of copper and zinc from mining and metallurgical wastes has been investigated and proposed. It includes sulfuric acid leaching of old pyrite flotation tailings to produce ferric containing leach solution; followed by ferric leaching of copper converter slag flotation tailings with the leach solution. A sample of old pyrite flotation tailings from the concentrator containing 0.36% of copper and 0.23% of zinc was leached with 10% sulfuric acid in the column. Recovery of copper and zinc reached 47.1% and 47.2%, respectively. The pregnant leach solutions contained 15.9 g/L of ferric iron. The subsequent ferric leaching of copper converter slag flotation tailings containing 0.53% copper and 2.77% zinc with the pregnant leach solution was conducted. The effects of various process parameters on the leaching dynamics of metals under batch conditions were investigated. Under the best conditions (temperature 70 °C, pulp density 30%, ferric iron concentration 15.9 g/L, initial pH of the pulp 0) the recovery of copper and zinc reached 79.6% and 43.7%, respectively. It was concluded that acid leaching of base metals from old pyrite flotation tailings with pregnant leach solution for the ferric leaching of copper converter slag flotation tailings is a prospective and promising technique for the complex treatment of mining and metallurgical wastes. 相似文献
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内蒙古某锌冶炼厂利用锌电解工业回水,浮选分离硫化锌氧压浸出渣中的硫磺和含银矿物。由于工业回水具有高酸、高杂质离子浓度的特点,生产指标较实验室小型浮选试验(自来水调浆)差得多。本文基于对工业回水性质的测定,深入研究了高浓度的硫酸、锌离子和铁离子对浸出渣中硫磺和含银矿物分选回收的影响。根据研究结果,针对性地提出了浮选工艺优化思路,即在工业回水调浆条件下,对矿浆流体性质进行调控,提高矿浆分散性,并采用适应性较强的丁铵黑药和煤油组合捕收剂代替原有黄药体系。在优化的条件下,获得了硫精矿硫品位79.52%,S回收率91.06%;尾矿银品位446.4 g/t,Ag回收率82.63%。相比生产指标,硫精矿品位和回收率分别提升了9.40%和22.10%,银的回收率提升了35.88%。研究结果为生产指标的提升提供了理论和技术支撑,同时未改变利用工业回水的现状。这对于提高企业经济和环境效益具有实际意义。 相似文献
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Heavy-metal-containing neutralization sludge (NS) and sulfur-containing zinc leaching residue (ZLR), both of which are difficult to dispose of, are two of the main solid wastes produced in the Pb/Zn smelting process. This study focused on the application of ZLR as a sulfur source to sulfidize NS, which could then be separated by flotation for metal recovery. A mixture of NS, ZLR and additional sulfur was first mixed by ball milling, and then successively treated by hydrothermal sulfidation and flotation. Based on the chemical properties of NS and ZLR, the effects of the mass ratio of NS-to-ZLR, the amount of sulfur and ball milling time on sulfidation and the floatability and stability of the sulfidation product were investigated. The sulfidation percentages of Zn and Pb were as high as 82.6% and 95.6%, respectively. Flotation tests revealed that Zn and Pb can be enriched with a concentrate grade of 21.3% Zn and 3.4% Pb. Toxicity characteristic leaching procedure (TCLP) results indicated that stabilization of NS and ZLR occurred after sulfidation. 相似文献
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