云南高泥尾矿铅锌分离实验研究

针对云南高泥尾矿铅锌嵌布粒度细、泥化严重、性质相对复杂的技术难点,进行了铅锌分离实验研究,采用泥砂分级别浮选-重选联合工艺流程,有效分离了铅锌,实现了该复杂尾矿资源的综合回收利用。结果表明,该矿中铅含量为4.29wt%,锌含量为4.99wt%,铅主要以白铅矿和铅铁矾的形式存在,铅和铁相互交代形成不同的包裹形式,分离难度极大;锌主要以氧化锌的形式存在,氧化程度较深,锌氧化率达99%,且主要为难选的异极矿。最终通过闭路选矿流程,获得铅品位33.87%,回收率62.53%的铅精矿;铅精矿中银品位142.50 g/t,银回收率30.92%;获得锌品位15.21%、回收率47.82%的锌精矿。     

综合回收某低品位磷铁矿石中磷和铁的试验研究

辽宁某低品位磷铁矿石中可回收元素为铁和磷,铁品位8.69%、磷品位2.79%,铁和磷主要以磁铁矿和磷灰石形式存在。试验采用磁浮联合流程,磁选为一粗一精、浮选为一粗一扫三精正浮选流程,在磨矿细度为-0.074mm占70%,磁场强度12kA/m,碳酸钠为pH值调整剂,调pH至7.5,水玻璃用量1500g/t,捕收剂改性氧化石蜡皂用量1000g/t的条件下,获得铁精矿铁品位为66.97%、铁回收率为67.28%,磷精矿磷品位为35.97%、磷回收率为79.03%的选矿试验指标。     

磁选技术在车河选矿厂的应用研究

针对车河选矿厂原矿中铁丝、铁块等杂物多、枱浮摇床及+74摇床精矿锡品位较低且铁含量高、细粒级锡石全浮选流程回收率低且药剂消耗量大、锌硫分离尾矿未得到综合利用的问题,开展了原矿除铁试验研究、枱浮摇床给矿及精矿磁选试验研究、+74摇床精矿磁选试验研究、30米脱硫浮选给矿及一次浮锡精矿磁选试验研究、锌硫分离尾矿磁选试验研究等一系列研究,并在生产中得到应用,最终磁出原矿中铁质杂物50 kg/班,减少了流程堵塞及设备磨损;枱浮摇床精矿铁含量由17.1%降至13.2%;+74摇床精矿锡品位由42.5%提高至47.1%、铁含量由23.9%降至11.6%;细粒级锡石回收率提高1.01个三分点;锌硫分离尾矿得到综合回收利用。本次研究效果显著。     

从高岭土尾矿中浮选铅锌

根据硫化矿浮选理论,研究了高岭土尾矿中铅锌硫矿物分离的药剂制度和工艺流程.结果表明,方铅矿和闪锌矿在磨矿至粒度74mm颗粒占70%时单体基本解离,在捕收剂乙硫氮30g/t、组合抑锌剂ZnSO_4 1000g/t及Na_2SO_3 300g/t的药剂制度下铅锌矿物可有效分离;在捕收剂异丁黄药50g/t、锌活化剂CuSO_4 300g/t的药剂制度下锌硫矿物可有效分离.通过磨矿-优先浮铅-锌硫矿物分离的闭路流程,得到铅品位64.55%、回收率84.47%的铅精矿,锌品位45.04%、回收率91.94%的锌精矿.     

响应曲面法优化某氧化铜矿硫化浮选

对云南某地难选氧化铜矿进行了SEM,XRD表征和铜物相分析,确立了硫化浮选的选矿工艺,进行了单因素实验,并通过响应曲面法优化浮选条件.结果表明,响应曲面法优化精矿铜品位和回收率模型p值均小于0.05.磨矿细度和磷酸乙二胺用量对铜精矿回收率有显著影响,磷酸乙二胺用量对铜精矿品位影响显著.响应曲面法优化的最佳浮选条件为磨矿细度小于0.074mm颗粒占86.07%、硫化钠用量2012.75 g/t、磷酸乙二胺用量132.19g/t,该条件下得到回收率79.007%、品位22.156%的铜精矿,浮选实验结果与响应曲面法优化结果基本一致.根据优化的浮选条件进行浮选闭路流程实验,所得铜精矿含铜21.93%,回收率为76.23%.     

滇东某多金属氧化铅锌矿高效回收选矿工艺

对滇东某多金属氧化铅锌矿采用先铅后锌的工艺浮选,浮选流程均为两粗一精一扫.结果表明,矿石中有价元素为铅、锌、银,铅主要赋存于白铅矿和铅矾中,锌主要赋存于菱锌矿和异极矿中,银以伴生形式存在,目的矿物嵌布粒度较细.浮选所得铅精矿铅品位为61.45%,铅回收率为86.41%,银品位为451.58 g/t,银回收率为66.73%,含锌3.68%;锌精矿锌品位为42.32%,锌回收率为90.63%,含铅1.39%.两性捕收剂R_(144)对锌的捕收能力和选择性比十二胺、十八胺和二者混合胺更好.     

菱铁矿磁化焙烧与磁选分离制备铁精矿

对国内某地菱铁矿进行无还原剂全粒级磁化焙烧-磁选研究. 结果表明,菱铁矿在无还原剂条件下于800℃焙烧15 min,所得焙烧矿在磨矿粒度小于0.074 mm占90%、磁场强度0.10 T条件下磁选,得到铁品位63.15%、铁回收率92.52%的铁精矿. 磁选精矿中的锰、镁和部分钙与铁元素以类质同像共存.     

云南某硅酸盐型铁矿选矿试验研究

随着富矿开采日渐枯竭,贫矿开发的重要性越发显著.本文以云南具有代表性的某硅酸盐型低品位铁矿为研究对象,对其进行了详细的工艺矿物学研究,并根据原矿性质特点进行选矿试验研究,分别进行了弱磁选以及弱磁尾矿的强磁选、摇床重选及还原焙烧—弱磁选等试验,并最终确定了弱磁选提磁铁矿和还原焙烧—弱磁选提赤铁矿、褐铁矿的选矿流程.通过两步磁选,获得的铁精矿平均铁品位为60.07％,铁回收率为74.35％.     

硫酸钠作用下褐铁型红土矿选择性还原富集镍铁

以Na2SO4为添加剂,采用煤基还原-磁选工艺,对含Ni 1.17%和Fe 35.71%的褐铁型红土矿选择性还原富集镍铁的行为进行了研究.结果表明,原矿中27.40%的镍以氧化物形式存在,69.35%的镍以硅酸盐形式存在;在还原温度1100℃、还原时间60 min、硫酸钠用量20%(?)、无烟煤用量8%(?)的最佳还原条件下,在200 m T的磁场强度下湿式磁选,可富集得到镍品位13.26%、镍回收率88.4%、铁品位67.34%、铁回收率14.71%的镍铁精矿,镍富集比达11.33.红土镍矿还原过程中加入Na2SO4破坏了原矿中硅酸盐的结构,释放出其中的镍与铁,进一步生成Fe S,抑制铁深度还原并降低了体系熔点,有利于金属离子传质,促进镍铁颗粒长大.     

辽宁某含金铁矿石综合回收的选矿试验研究

辽宁省本溪市某铁矿在生产过程中发现含有金,原矿含金品位为1.47g/t,含铁品位为18.82%。通过浮选回收金+磁选回收铁的联合工艺流程,获得了比较理想的选矿工艺指标。试验矿石在磨矿细度为-0.075mm占65%的条件下,采用硫酸铜作为金载体矿物的活化剂,丁基黄药和丁铵黑药作为捕收剂,采用一次粗选三次精选二次扫选的浮选工艺流程,试验取得的工艺指标为,金精矿含金品位为50.85g/t,金回收率为75.49%。浮选尾矿进行湿式弱磁场回收磁铁矿,粗精矿再磨至细度为-0.075mm 97%再选得铁精矿,试验取得的工艺指标为,铁精矿含铁品位为65.52%,铁回收率为29.42%。     

云南某富银锌精矿锌银浮选分离实验研究

云南某富银锌精矿中银主要以类质同象形式存在于白铅矿中,本工作以该矿样为研究对象,根据其性质,采用抑铅浮锌的工艺流程进行浮选,考察了磨矿细度、抑制剂、活化剂及捕收剂等因素对浮选分离指标的影响。结果表明,–19+10 μm粒级中银含量最高。在磨矿细度–74 μm占90%,硅酸钠用量为2000 g/t,硫酸铜用量为200 g/t,丁基黄药用量为300 g/t,松醇油用量为30 g/t的条件下,1次粗选、1次精选、1次扫选,中矿顺序返回的全流程闭路实验,可获得含锌61.08%,回收率95.89%的锌精矿和含银1548.32 g/t,回收率为71.17%的银精矿,实现了锌银的浮选分离及伴生银的高效富集。     

硫酸渣磁重选联合工艺回收铁精矿研究

研究了从硫酸渣中回收铁精矿的工艺流程.硫酸渣分选最佳工艺流程为:预先分级、磨矿后在120kA/m条件下磁选,磁选尾矿用螺旋溜槽重选,混合精矿的品位61.32%,回收率83.28%,产率72.86%.硫酸渣不经磨矿直接磁选得不到高品位精矿,全部磨矿后分选,精矿品位略有提高,但回收率下降较多.     

提高硫酸尾渣磁选铁精粉品位的方法

介绍硫酸尾渣磁选制取铁精粉的不同生产工艺;瑞和化肥有限公司13万t/a硫铁矿制酸装置,利用其尾渣的矿尘生产铁精粉的工艺流程及提高铁精粉品位及产率的措施,使铁精粉产量从200t/d提高到300t/d,w(Fe)从53%提高到58%,w(S)从0.9%降到0.3%,获得较好的经济效益.     

Iron recovery from zinc mine tailings by magnetic separation followed by carbothermal reduction of self‐reducing briquettes

The volume of tailings produced by the extractive industry has been increasing due to the processing of the low‐grade ore. This issue can cause environmental accidents and require significant investment to control the disposal of tailings. Therefore, this study aims to recover iron from zinc mine tailings by wet magnetic separation followed by the carbothermal reduction of self‐reducing briquettes. Two magnetic separation routes were investigated to concentrate iron. Zinc mine tailings were processed by route I, in a rougher stage followed by a scavenger stage; and route II, in a rougher stage followed by a cleaner stage. The carbothermal reductions were performed using self‐reducing briquettes composed of Fe concentrate from the route with high Fe content and charcoal. The products were analyzed by scanning electron microscopy with energy dispersive spectroscopy (SEM‐EDS), x‐ray diffraction (XRD), inductively coupled plasma optical emission spectrometry (ICP‐OES), and volumetric chemical analysis. Magnetic separation route II provided the highest‐grade Fe concentrate, 52% Fe, while route I provided 33% Fe. In the carbothermal reductions, a metallization degree of 98% in the Fe concentrate briquette, 97% in the briquette with a 10% replacement of its raw material by Fe concentrate, and 99% in the hematite briquette was reached. The replacement of raw material by Fe concentrate showed no significant change in Fe recovery. Considering the whole process, magnetic separation and carbothermal reduction, the recovery of Fe from the zinc mine tailings was 67%. Therefore, the process route suggested in this study will not only reduce tailings disposal and consequently the risk of environmental accidents, but it will also provide profitable raw material for the steel industry.     

山东某铅锌银多金属矿选矿试验研究

针对山东某铅锌银多金属矿嵌布关系复杂多变的特点,采用银铅锌混合浮选后铅锌分离和铅优先浮选两种技术路线进行试验.采用铅优先浮选技术路线使铅锌分离,得到了较好的选别指标,铅精矿中铅品位为70.04％,铅回收率97.75％,银品位1541 g/t,银回收率79.56％;锌精矿中锌品位43.73％,锌回收率82.55％.     

硫铁矿烧渣提纯制铁精矿的试验研究

在研究硫铁矿烧渣性质的基础上，进行了烧渣提纯的磁选、重选和工艺流程的条件试验，原料预处理后磁选可制得TFe＞60％、回收率70％以上的铁精矿，可用作炼铁原料。     

Beneficiation of micro-fine magnetic minerals from reductive iron ore with ultrafine grinding-magnetic flocculation separation

Beneficiation of micro-fine magnetic minerals from reductive iron ore was investigated. Sample characteristics and main force analysis of magnetic floc were conducted. The results indicated that the iron phase in reductive iron ores was predominantly metallic iron (below 20 μm). By applying ultrafine grinding-magnetic flocculation separation (MFS) to the raw ore (29.85% Fe), a concentrate assaying 74.12% Fe with 81.45% iron recovery was obtained. The iron recovery increased by 6.68% compared with the conventional magnetic separation (CMS). The high efficiency in beneficiation may be attributed to an increase in magnetic force on the micro-fine iron minerals in the form of flocs.     

Separation of Phosphorus and Magnetic Mineral Fines from Siderite Reductive Ore by Applying Magnetic Flocculation

The characteristics of siderite reductive ore and the ultrafine grinding-magnetic flocculation separation (MFS) of this ore were investigated in the present work. The results indicated that the iron phase in the raw ore was predominantly metallic iron with an iron particle size below 30 μm, and the phosphorus compound was apatite. By applying MFS to siderite reductive ore containing 37.14% Fe and 0.52 P, a concentrate assaying 66.37% Fe and 0.19 P with 74.32% recovery was produced. The iron recovery increased by 5.77% compared with the results of the conventional magnetic separation. The high efficiency in phosphorus removal and iron recovery achieved by the MFS process may be attributed to the adequate liberation of iron particles and the increase in magnetic force on the iron mineral fines in the form of flocs in a magnetic field.