某难选低品位铜矿的选矿工艺研究

在不同磨矿细度下,对云南某低品位铜矿原矿进行磨矿细度条件试验及流程对比试验,结果表明采用粗磨入选—粗精矿再磨流程,矿石入选细度70%-200目,可获得铜精矿铜品位22.00%、回收率83.72%的分选指标。     

提高大厂铟锌精矿产品质量的试验研究

研究的矿石属锡石多金属硫化矿类型,铁门锌矿与多种硫化矿紧密共生,浮选分离十分困难．试验采取在粗房条件下浮选硫化矿、硫化矿混合精矿再磨浮选铅锑、浮铅尾矿氧化浮锌的工艺,取得了满意的效果．     

Mineralogical characterization of a polymetallic sulfide ore to improve silver recovery

Based on the mineralogical characterization for the polymetallic sulfide ore, the way to improve silver recovery was studied. The results showed that silver was the most valuable metal whose grade was 448.82 g/t Ag, while 0.118% Cu, 1.65% Pb and 1.06% Zn may be comprehensively utilizated. The main silver-bearing minerals were argent and aregentite which accounted for 87.18% of total silver. Argentite and other metal minerals were distributed in the gangue minerals in complex forms. Argentite grains of 33.76% minus 50 μm indicated that a fine grinding scheme was necessary to enhance the degree of dissociation, and meanwhile selective grinding must be considered to prevent a complete grinding of coarse grains. The optimum regrinding fineness in the Cu flotation was determined as 73% minus 37 μm, while grains of 68.5% minus 74 μm in one-stage grinding remained unchanged as much as possible. Consequently, silver recovery increased to 2.68%, as well as the content of Pb simultaneously decreased from 7.26% to 2.68% in the Cu concentrate. From the lead pyrometallurgical point of view, recovering larger amounts of silver and lead at the expense of decreasing the grade of lead to a suitable level is not only economically viable for the plant, but also convenient for subsequent processing. Silver and lead recovery increased to 13.18% and 12.58%, respectively, while the Pb grade decreased from 53.1% to 46.12% for the Pb concentrate.     

Improvement of copper–molybdenum ore beneficiation using a combined flotation and biohydrometallurgy method

The authors present the results of analysis of material composition and experimental investigations of acid and biohydrometallurgical leaching of middlings on grain size, pH level, leaching process duration, temperature and slurry density. The rational parameters of flotation and acid-bacterial leaching of middlings providing an efficient release of valuable components from mineral complexes and recovery to flotation concentrate and leaching solution have been determined. A combined flowsheet and a beneficiation process for bulk flotation middlings of copper–molybdenum ore have been suggested, which include middlings grinding, sulfide minerals flotation, bacterial leaching of sulfide flotation tailings, liquid-phase extraction of dissolved copper and electrolysis of re-extraction eluates. The suggested combined method of cleaning of middlings of copper–molybdenum ores beneficiation provides the total copper recovery increase by 0.8% with a reduction of the cost price of saleable material by 0.5%.     

Removing iron by magnetic separation from a potash feldspar ore

A new permanent magnetic separator was introduced to treat the ores with the characteristics of weak magnetic iron minerals and in a fine size range. The new machine was applied to the iron removal from potash feldspar. The effects of the magnetic field intensity, pulp density and grinding fineness on the iron removal were investigated. The optimized operation parameters were achieved and listed as follows: the ?0.074 mm content is 85%, the pulp density is 45% and the magnetic field strength is 2T. A close test of middles regrinding was also carried out to improve concentrate yield. The data show that the grade of TFe(total iron) in potash feldspar product decreased from 1.31% to 0.21% and the concentrate yield reached 85.32%. All the results indicated that the traditonal high-intensity electromagnetic separators can be betterly substituted by the new permanent magnetic separator. This study may provide the theoretical evidence for iron removal from potash feldspar.     

Beneficiation studies on recovery and in-depth characterization of ilmenite from red sediments of badlands topography of Ganjam District,Odisha, India

This paper deals with the recovery of ilmenite mineral from red sediments of badlands topography and suggested flowsheet with material balance. The results of these investigations reveal that the red sediment samples contain 33.2% total heavy mineral, in which ilmenite mineral concentrate is 28.71% (by weight). The ilmenite concentrate recovered from red sediment sample by physical beneficiation process, which included scrubbing, desliming, gravity concentration, magnetic and electrostatic separation, contains 99.41% grade with 97.3% recovery. The ilmenite mineral concentrate recovered from red sediments is also suitable for industrial applications. The characterization studies on ilmenite reveal that the TiO₂ percentage is marginally increasing from 46.69% to 47.86% with increasing magnetic intensity from 0.46 to 1.55 T.     

Separation of hematite from banded hematite jasper(BHJ) by magnetic coating

The separation of iron oxide from banded hematite jasper(BHJ) assaying 47.8% Fe, 25.6% Si O2 and 2.30%Al2O3 using selective magnetic coating was studied. Characterization studies of the low grade ore indicate that besides hematite and goethite,jasper, a microcrystalline form of quartzite, is the major impurity associated with this ore. Beneficiation by conventional magnetic separation technique could yield a magnetic concentrate containing 60.8% Fe with 51% Fe recovery. In order to enhance the recovery of the iron oxide minerals, fine magnetite, colloidal magnetite and oleate colloidal magnetite were used as the coating material. When subjected to magnetic separation, the coated ore produces an iron concentrate containing 60.2% Fe with an enhanced recovery of56%. The AFM studies indicate that the coagulation of hematite particles with the oleate colloidal magnetite facilitates the higher recovery of iron particles from the low grade BHJ iron ore under appropriate conditions.     

Selective separation of silica from a siliceous-calcareous phosphate rock

Selective separation of silica from a siliceous-calcareous phosphate ore that had been sieved into different size fractions is investigated by a combination of chemical analysis, zeta potential measurement and FTIR and XPS techniques. Scrubbing is a better choice than flotation for removing silica from the coarse fractions. The P2O5 grade of the coarse fractions is increased to about 30% by scrubbing and the product yields are higher than those obtained by flotation. The silica in the fine fraction is separated by reverse flotation. An alkyl amine salt (DAH) is an effective collector and the P2Os grade of the fine fraction can be increased by 7% to beyond 30% under acidic conditions. The higher zeta potential obtained using DAH suggests that it is more strongly absorbed onto the ore panicles than the other cationic collectors.FTIR and XPS results confirm physical absorption of the cationic collector onto the ore surface. They also indicate that calcite is dissolved at low pH values, which increases the Si concentration on the ore surface.     

磁化矿石颗粒模型及磁选过程分析

基于磁选过程中颗粒尺寸、磁场强度和磁选精矿品位三者之间的关系,建立磁化矿石颗粒模型,对其进行理论分析与计算,确定最佳磁场强度,并进行磁化矿石的磁选研究。结果表明：在配煤量4%（质量分数）,焙烧温度850℃,焙烧时间60 min,磨矿细度-0.074 mm占60%（质量分数）,磁场强度为40 mT的条件下,得到铁品位57.7%（质量分数）,铁回收率90.3%（质量分数）的铁精矿,较好地实现了铁精矿的富集和回收。     

哈海岗铜多金属矿矿石组分及工艺性质

哈海岗铜多金属矿床Ⅱ号矿体的有用组分为铁、铜、铅、锌、钨,Ⅳ号和Ⅲ号矿体有用组分为铁、钨、钼。该矿床的主要金属矿物为磁黄铁矿、闪锌矿、方铅矿、磁铁矿、黄铜矿、黄铁矿、白钨矿、辉钼矿。矿石破碎后,〉0.25 mm粒级的有用元素产率最大（〉50%）,其次是〈0.075 mm粒级（〉19%）,其他粒级的有用元素产率较少,特别是0.096～0.075 mm粒级,有用元素产率仅为1%。在各个粒级中铁、铜、铅、锌、钼等元素的品位变化不大,仅钨在Ⅲ矿体中细粒级品位高于粗粒级品位近2倍。     

贵州织金含稀土磷矿石的浮选富磷降镁实验

针对贵州织金含稀土磷矿石进行了选矿富磷降镁实验,原矿P₂O₅品位21.80%,MgO质量分数8.31%,采用一段反浮选流程,可获得精矿P₂O₅品位33.35%,MgO质量分数为1.36%,P₂O₅回收率86.53%;考察了选矿产品中稀土元素总量（∑REE）和重金属元素Pb、Cd的质量分数及走向,并分析精矿中∑REE、重金属元素Pb与P₂O₅品位之间的相关性。结果表明:重金属元素Pb和稀土元素在精矿中富集,而重金属元素Cd在尾矿中富集,为后续含稀土磷矿石的分离利用提供了依据。     

贫磁铁矿反浮选降硅试验

湖北省某地具有较为丰富铁矿资源,矿石中铁含量较低,原矿中全铁(TFe)含量约15%,属于贫磁铁矿,铁矿物的嵌布粒度较细,通过单一弱磁选很难得到全铁品位超过60%的铁精矿,针对该矿弱磁选精矿进行反浮选提铁脱硅研究,一粗一精开路反浮选流程精矿品位可达60%以上,铁回收率60%,产率50%左右.通过小型闭路试验,反浮选最终获得较好指标:精矿产率为68.57%,品位为58.62%,回收率为82.83%.     

贫磁铁矿反浮选降硅试验

湖北省某地具有较为丰富铁矿资源,矿石中铁含量较低,原矿中全铁（TFe）含量约15%,属于贫磁铁矿,铁矿物的嵌布粒度较细,通过单一弱磁选很难得到全铁品位超过60%的铁精矿,针对该矿弱磁选精矿进行反浮选提铁脱硅研究,一粗一精开路反浮选流程精矿品位可达60%以上,铁回收率60%,产率50%左右.通过小型闭路试验,反浮选最终获得较好指标：精矿产率为68.57%,品位为58.62%,回收率为82.83%.     

细粒难选锐钛矿浮选试验研究

对TiO2品位为10.97%的某细粒复杂难选锐钛矿进行浮选试验研究.研究表明,试样磨矿细度为-0.043 mm占85.96%时,以碳酸钠将pH调至8.5,粗选采用羧甲基纤维素和氟硅酸钠为组合抑制剂,精选采用硫酸铝和淀粉为组合抑制剂,用乙酸铅作活化剂,用苄基胂酸和羟肟酸钠作组合捕收剂,经两次粗选、三次精选的浮选工艺,可得到产率为28.51%、TiO2品位为30.91%、回收率为80.32%的锐钛矿精矿,初步实现了细粒难选锐钛矿的选矿富集,为该矿的进一步深入研究及开发奠定了一定的基础,同时为伴生的V、Cr、Co等稀土元素的综合回收利用提供了一定的参考.研究结果表明：①锐钛矿的可浮性与金红石相似,Pb2＋离子同样是锐钛矿浮选的有效活化剂;②对嵌布粒度细、矿物组成复杂的难选矿,药剂的组合使用是使目的矿物以连生体的形式浮选富集的有效手段.     

难处理高结合率氧化铜矿选冶联合工艺研究

采用选冶联合工艺对含铜1.53％、氧化率47.06％、结合率21.57％的高结合率氧化铜矿进行回收.原矿的砂光片分析结果表明,矿石中大部分铜矿物嵌布粒度极细,多呈星点状和不均匀浸染状分布,与硅、钙、镁、铝等脉石共生严重,导致浮游性较差.针对该矿石的特点,研究了工艺参数及流程结构对指标的影响,确定了“三次粗选—粗精矿再磨—三次精选”的硫化浮选工艺流程,获得了含铜品位为23.43％、回收率为53.72％的铜精矿.对尾矿的形貌及矿物组成表征发现：铜矿物呈细粒浸染状或被硅酸盐矿物包裹,导致这部分铜损失在尾矿中.在最佳的酸浸工艺条件下,对浮选尾矿进行酸浸试验,获得了相对原矿的浸出率为33.21％的试验指标;铜综合回收率为86.93％.     

Molybdenum removal from copper ore concentrate by sodium hypochlorite leaching

The removal of molybdenum from a copper ore concentrate by sodium hypochlorite leaching was investigated. The results show that leaching time, liquid to solid ratio, leaching ternperature, agitation speed, and sodium hypochlorite and sodium hydroxide concentrations all have a significant effect on the removal of molybdenum. The optimum process operating parameters were found to be: time, 4 h: sodium hydroxide concentration, 10%; sodium hypochlorite concentration, 8%; liquid to solid ratio, 10:1; temperature, 50℃; and,agitation speed, 500 r/min. Under these conditions the extraction of molybdenum is greater than 99.9% and the extraction of copper is less than 0.01%. A shrinking particle model could be used to describe the leaching process. The apparent activation energy of the dissolution reaction was found to be approximately 8.8 kJ/mol.     

某铜铁矿选矿工艺研究

主要针对某铜铁矿矿石性质,研究其选矿工艺流程,最终确定选铜回路采用浮选工艺流程,浮选药剂为石灰和丁基黄药;选铁回路采用磁选工艺流程方案。最终铜精矿品位为20.53%、回收率94.50%,铁精矿品位58.54%、回收率72.30%,获得了较好的试验指标。     

硫酸渣磁化焙烧—磁选提铁降硫

硫酸渣铁品位为55.08％,其中有害元素硫的含量为1.3％.为高效利用硫酸渣,必须提高铁含量、降低硫磷等有害元素.硫酸渣试样直接进行弱磁选,得到铁精矿品位60.54％,精矿回收率仅为54.46％,采用磁化焙烧-弱磁选的方法来进行选铁试验,通过对磁化焙烧时间、磁化焙烧温度、还原剂的质量配比等条件试验,确定了在焙烧时间40 min,焙烧温度750℃,还原剂10％的最佳焙烧条件.焙烧矿磨矿至-0.074 mm 97.02％,用弱磁选管进行磁选的最佳试验条件,在此焙烧条件下,进行一粗一精的磁选,获得了铁品位64.57％,精矿回收率86.99％,硫含量降低到0.13％.     

钙硅质胶磷矿浮选工艺研究

针对湖北某中低品位钙硅质胶磷矿,分别进行单一反浮选、双反浮选和常温正反浮选工艺研究及药剂费用对比,结果表明：采用单一反浮选,可得到磷精矿P2O530.37%,MgO 0.36%,回收率88.38%的较好指标,且最为经济,每吨原矿药剂费用仅为15.28元;若希望精矿质量更好些,回收率更高些,建议采用双反浮选,可得到磷精矿P2O531.60%,MgO 0.43%,回收率90.63%的好指标,每吨原矿药剂费用为23.40元;正反浮选相对来说精矿质量一般、回收率较低、药剂费用较高、磨矿细度较细,指标分别为磷精矿P2O530.54%,MgO 0.70%,回收率84.62%,每吨原矿药剂费用为26.98元.     

云南罗平史家寨硫铁矿的浮选研究

云南省罗平县史家寨硫铁矿矿石是一种细粒嵌布的低品位矿石.原矿含硫仅14.30%.为制定该矿石的最佳选矿工艺流程,以生产出品位高于35%的硫精矿,进行了一系列研究.研究结果表明,为产出合格精矿,保证85%-200目以上的磨矿细度是关键措施.除此之外,用乙基黄药而不是丁基黄药作为扑收剂、使用适量的硫酸铜作为活化剂以及添加足够的起泡剂2号油,也是获得良好选矿指标的重要条件.按所拟定的流程和药剂制度,产出了品位为35.20%的硫精矿,回收率高达84.43%.