湖南某高钙型钨萤石矿综合回收新工艺研究

为了综合回收湖南郴州某高钙矿石资源中的钨和萤石,针对矿石中钨、萤石、碳酸钙三者都是含钙矿物可浮性相似的问题,进行了原矿化学分析和工艺矿物学研究。根据化学分析及工艺矿物学研究结果且进行了选矿工艺试验研究,采用钨、萤石等可浮—钨、萤石加温分离—萤石浮选的工艺流程,获得的钨粗精矿WO_3含量为2.11%,CaF_2含量为78.45%,钨回收率为78.45%,萤石回收率为38.74%;萤石精矿中CaF_2含量为90.21%,萤石合计总回收率为84.52%的试验结果,该工艺较好地实现了钨和萤石的综合回收。     

福建某低品位共伴生白钨、萤石矿选矿试验研究

对福建某WO3品位0.10%、CaF2品位25.45%的低品位共伴生白钨、萤石矿,以矿冶科技集团有限公司自主研发的高效选矿药剂BK418作为白钨捕收剂,BK410作为萤石捕收剂,采用“白钨常温浮选-常温浮选钨精矿加温精选-白钨常温浮选尾矿浮选萤石”的工艺流程处理该矿石,获得钨精矿中WO3品位为60.48%,WO3回收率为65.29%;萤石精矿中CaF2品位为90.64%,CaF2回收率为55.34%的良好指标。为开发利用该类型低品位共伴生白钨、萤石矿提供了技术依据。     

从湖南某钨多金属矿尾矿中回收伴生萤石试验研究

对湖南某钨多金属尾矿进行了综合回收萤石的浮选试验研究.实验室试验采用一粗一扫四精浮选流程,结合萤石高效浮选药剂,获得了萤石品位96.39％、总回收率62.65％的萤石精矿;新工艺工业试验获得萤石平均品位87.62％、平均回收率59.64％的萤石精矿指标,浮钨尾矿中伴生的萤石资源得到了高效综合回收.     

湖南某钨尾萤石脱药浮选新工艺研究与应用

湖南某钨选矿厂采用选钨新工艺,使得下游萤石选场采用原浮选工艺不能得到合格的萤石精矿产品。因此在选萤石矿之前加入新的脱药工艺,脱出钨尾残留的选钨药剂,消除药剂影响,让萤石浮选更容易。加入脱药新工艺后,在实验室、半工业实验和工业试验均达到理想指标。最终工业试验萤石精矿的CaF2品位稳定在93%以上,回收率稳定在55%左右。另外,脱药新工艺的加入使得萤石精矿杂质更低,有效硫含量更低,降低了萤石精矿深加工的成本。     

悬振锥面选矿机对柿竹园萤石精矿再选钨试验研究

本次试验研究对象为柿竹园萤石精矿。原矿含CaF295.5%,Sn0.059%,WO_30.13%。为了回收萤石精矿中的钨,对原矿进行粒度分析后,采用新型选矿设备—悬振锥面选矿机进行试验研究。通过试验研究,确定了设备的最佳进行参数,经过一次粗选,可获得品位10.56%,回收率32.23%的钨精矿,钨品位富集了81倍,而萤石精矿选钨后,萤石品味提高了2.8个百分点。试验结果证明,悬振锥面选矿机对柿竹园萤石精矿再选钨有良好的回收效果。     

湖南某低品位萤石矿浮选工艺研究

湖南某萤石矿含CaF2为10.65%,属于低贫萤石矿,试验采用预先脱硫—浮选萤石工艺,在弱酸性介质精选的工艺制度下,通过一次粗选八次精选二次扫选,获得了CaF2品位为97.80%、回收率为49.25%的萤石精矿产品,萤石浮选尾矿采用重选综合回收白钨矿,重选白钨精矿含WO367.97%,回收率53.42%。     

湖南某低品位白钨矿加温浮选改常温浮选试验研究

针对湖南某低品位白钨浮选生产中加温浮选能耗大、成本高、加温精选作业回收率不理想等问题，开展了白钨加温浮选工艺改常温浮选工艺试验研究。试验结果表明：采用CK-5为白钨矿捕收剂，碳酸钠为pH调整剂，水玻璃为硅酸盐脉石抑制剂，CF为萤石、方解石以及细泥等脉石特效抑制剂，通过实验室闭路试验取得了白钨常温浮选精矿WO₃品位50.20%、回收率为72.54%的浮选指标，成功实现了白钨加温浮选工艺改为常温浮选工艺，同时白钨回收率较矿山2021年现场生产回收率提高了7.86个百分点，研究结果有助于国内外选厂白钨加温浮选工艺改常温浮选，对加温浮选矿山节能降耗有重要意义。     

江西香炉山钨尾矿中萤石综合回收技术的研究

江西香炉山钨尾矿含氟化钙7.10%、碳酸钙8.00%,属萤石含量低而方解石含量高的尾矿,尾矿中萤石的经济价值不高,当前售价约在2 000元/t以内,但为了提高矿产资源综合回收利用率和尽量延长尾矿库使用寿命,开展了钨尾萤石的综合回收技术研究。采用预处理—萤石浮选—酸浸工艺,获得产率为3.88%,Ca F291.88%,回收率为50.26%的萤石精矿,说明香炉山钨尾中的萤石可得到有效回收。本文对萤石综合回收工艺及条件试验进行了论述,并根据尾矿的矿物学特征,对影响萤石回收指标的因素进行了分析,对三氧化钨在萤石分离工艺中的走向进行了查定。     

柿竹园白钨浮选尾矿综合回收萤石试验研究

柿竹园白钨浮选采用733-烧碱法,矿浆pH值在12以上,不利于后续的萤石综合回收。本研究针对柿竹园白钨浮选尾矿,采用硫酸为活化剂、水玻璃为抑制剂、733为捕收剂,进行了综合回收萤石的试验研究,最后采用一次粗选、两次扫选和五次精选工艺流程,可获得萤石精矿品位94.31%、回收率70.06%的试验指标。     

某钨、钼、铋、萤石复杂多金属矿分选尾矿中钨的选矿试验

通过采用弱磁选-黑白钨混合浮选-黑白钨分离浮选-白钨精选-黑钨摇床选别-黑钨细泥浮选的工艺流程回收某钨、钼、铋、萤石复杂多金属矿经等可浮硫化矿浮选尾矿中钨,可得到白钨精矿WO₃品位68.79%,回收率53.27%,黑钨精矿WO₃品位52.49%,回收率17.57%,钨总的回收率70.84%的选矿技术指标。同时指出白钨精矿酸浸可以除掉磷,溶去方解石等杂质,白钨精矿品位提高了2.46个百分点。     

山西某金红石矿选矿试验研究

山西某金红石矿采用重选主干流程进行选别,精矿产品TiO2品位为90%左右,但金红石(TiO2)的回收率不足50%。为提高金红石的选矿回收率,开展了以浮选为主干流程的选矿工艺研究。确定的选矿方案为两次浮选抛尾─金红石浮选(一次粗选、两次精选)─浮选精矿除杂(弱磁选—强磁选—重选)。全流程试验结果表明:采用浮选主干流程大大提高了精矿TiO2的回收率,总精矿TiO2回收率为69.25%,金红石矿物的回收率达到86.42%,其中精矿1含TiO289.58%、TiO2回收率46.84%;精矿2含TiO280.53%、TiO2回收率22.41%。同时综合回收了磁铁矿和钛铁矿。     

Short-cut method for flotation rates modelling of industrial flotation banks

The knowledge of the distributed performance of a flotation bank, consisting of a number of cells in series, is a key factor for different purposes such as process design, scale-up, diagnosis, operation, control and optimization. A common practice in plant operation is to develop mass balances around the whole flotation bank in order to characterize the overall recovery, typically in rougher flotation. However, testing to fit flotation rate models are seldom developed on industrial flotation banks because they are high consumers of human labor during sampling, mineral samples preparation and chemical analysis development. In this paper a short-cut method is proposed which allows obtaining the relevant information for flotation rate modeling in a flotation bank with minimum effort and cost, and within a reasonable accuracy (less than 1–2% error in estimating cell recovery along the bank). The procedure considers two mass balances, one around the first cell of the bank and the second is the overall mass balance around the whole flotation bank, with a total of only 5 sampling streams. Examples developed in four rougher flotation banks located in three industrial concentrators illustrate the merit of this procedure.     

贵州水银洞低品位卡林型金矿矿石选矿试验

对贵州水银洞低品位卡林型金矿进行选矿试验研究,在对氰化法和浮选法进行比较的基础上,采用浮选方法,取得了满意的试验效果:金的回收率为91.64%,浮选金精矿品位为42.6g/t,然后对金精矿进行预氧化-氰化试验,金的氰化浸出回收率提高到88.76%,金矿石选冶总回收率达到了81.34%。     

Selective detachment process in column flotation froth

The selectivity in flotation columns involving the separation of particles of varying degrees of floatability is based on differential flotation rates in the collection zone, reflux action between the froth and collection zones, and differential detachment rates in the froth zone. Using well-known theoretical models describing the separation process and experimental data, froth zone and overall flotation recovery values were quantified for particles in an anthracite coal that have a wide range of floatability potential. For highly floatable particles, froth recovery had a very minimal impact on overall recovery while the recovery of weakly floatable material was decreased substantially by reductions in froth recovery values. In addition, under carrying-capacity limiting conditions, selectivity was enhanced by the preferential detachment of the weakly floatable material. Based on this concept, highly floatable material was added directly into the froth zone when treating the anthracite coal. The enriched froth phase reduced the product ash content of the anthracite product by five absolute percentage points while maintaining a constant recovery value.     

会泽铅锌硫化矿异步浮选新技术研究

通过单因素和正交浮选试验研究了会泽方铅矿、黄铁矿与闪锌矿之间的浮选分离。根据浮选动力学基本原理,对方铅矿和黄铁矿的浮选动力学特性进行了分析。结果表明,基于总体平衡理论的分速浮选模型可以较好地模拟方铅矿和黄铁矿的浮选过程,浮选回收率模型拟合值与试验值之间的相关系数R2均达到0.999。研究认为,异步浮选新技术充分利用不同矿物及同种矿物可浮性和浮游速度特性,实现矿物的个性化、差异性浮选。进一步探讨了异步浮选新技术的理论背景,对选矿人员完善已有工艺及开发新技术具有一定的参考意义。     

The role of a flash flotation circuit in an industrial refractory gold concentrator

In order to determine the contribution of the flash flotation circuit to the overall plant performance of the Kanowna Belle concentrator, two survey campaigns both with and without the flash circuit in operation have been conducted on two distinctly different ore types: a very high grade ore, and a very low grade ore of higher hardness. Using two different ores with the same target valuable mineral species (gold and pyrite) through the same treatment route allows any trends in performance to be more easily identified. As both survey campaigns involved running the plant with and without the flash flotation circuit in operation, the significant contribution of the flash flotation cell to overall plant recovery and final concentrate grade is highlighted. The flash circuit on this plant may be considered as the primary rougher, contributing in excess of 42% of the valuable material that is recovered to the final concentrate stream, at a grade of approximately 35% sulphur; and in-so-doing reducing the overall plant footprint that would otherwise be required to achieve the same recoveries at the target concentrate grade.Mineralogical analysis of survey samples shows that the feed to the flash flotation cell (cyclone underflow) is of a much higher grade and contains a higher proportion of well liberated valuable material as compared to the conventional flotation circuit feed (cyclone overflow). Maximising the recovery of this material before it re-enters the milling circuit should be of paramount importance to optimising overall plant performance.When the flash flotation circuit is taken off-line the recovery of sulphur (and hence pyrite) is observed to decrease dramatically, and whilst the recovery of gold also decreases, it is to a much lesser extent. The difference in the recoveries of gold and pyrite that is observed without the flash flotation circuit in operation is most likely attributable to a change in the way the gold is being liberated as a function of the change in grinding circuit operation that is required when the flash circuit is taken off-line. The distribution of valuable material in the cyclone overflow stream (conventional flotation feed) undergoes a step change when the flash circuit is taken off-line with an increase in the amount of valuable fines being generated, which is further reflected in the flotation tails with a higher proportion of both pyrite and gold being present in the intermediate and fine size classes. This increase in the amount of pyrite fines in particular may have contributed to the loss in recovery that was observed when the flash flotation circuit was taken off-line.Pulp chemistry data from various points around the flotation circuit highlight the different processing conditions in the flash cell, compared to the conventional circuit, which will impact on the type of minerals able to be recovered by flotation, as well as reagent selection for this type of processing application.     

某低品位难浸金矿石选矿试验研究

在实验室条件下对某低品位难浸金矿石进行了选矿试验研究。结果表明,采用一粗三精三扫的浮选闭路流程,可获得品位118.5g/t、回收率80.34%的浮选金精矿。浮选金精矿在600℃下焙烧1h后再氰化浸出,可获得金浸出率90.94%、综合回收率73.06%的分选指标。     

管段高紊流强化煤泥柱浮选的应用及作用机理

为了强化煤泥的柱浮选与细粒级回收,调整了旋流-静态微泡浮选床的管流段长度,分析了不同管流段长度下的煤泥浮选效果、产品特性和承载能力变化规律,并结合流体动力学和浮选动力学理论探讨了其强化作用机理。结果表明:管流段长度由1.5 m延长至3.0 m后,在精煤灰分相当的情况下,精煤可燃体回收率提高6.46%;细粒级0.074~0.045 mm和0.045 mm回收率分别提高7.96%和8.41%,设备承载能力提高0.09 t/(h·m2);同时各自增幅都随入料干煤泥量的增加逐渐变大。管流段延长可以增大管流段的紊流动能值,提高颗粒-气泡的碰撞速率,促进整体浮选指标的提升。     

Marmatite depression in galena flotation

Laboratory flotation tests were conducted to investigate the role of pH, zinc sulphate and potassium metabisulphite in the flotation of galena and marmatite in the Broken Hill orebody. Marmatite was depressed under alkaline conditions alone, however the addition of the two depressants, either individually or combined, resulted in the greatest selectivity. The use of zinc sulphate under alkaline conditions resulted in good selectivity whilst maintaining a high lead recovery. The addition of potassium metabisulphite under alkaline conditions reduced zinc recovery, but high additions reduced lead recovery. The addition of both depressants at alkaline pH produced the best result, with a lead concentrate of 72.6% at 98% recovery with less than 4% zinc recovery. The overall aim of the study was to minimise the flotation of marmatite into the galena rougher concentrate, which was achieved under the conditions given above.     

缅甸某难选锌氧化矿浮选工艺探索

针对缅甸某高硅、高钙、高泥化率难选氧化锌矿进行了浮选探索。在条件试验的基础上采用分步试验工艺,在磨矿细度为-200目占85%的条件下,经过两段分步浮选处理,获得一段精矿品位32.15%,二段精矿品位27.07%,综合回收率达84.77%的结果。