Investigating fine galena recovery problems in the lead circuit of Mount Isa Mines Lead/Zinc Concentrator part 1: Grinding media effects

A plant survey was carried out on the lead secondary rougher and scavenger banks of the Lead/Zinc Concentrator of Mount Isa Mines Limited. Sizing analysis of the survey samples demonstrated that a major limiting factor to overall lead recovery in this section of the plant was the diminished recovery of the fine galena in the minus 5 microns particle size fraction. Batch flotation experiments were carried out on a plant sample of lead secondary rougher feed and a sample of rod mill feed ore. Mineral recovery-size data for these tests showed similar fine galena flotation behaviour to that observed in the plant. Increased collector addition did not improve either the maximum recovery or the flotation rate constant of the fine galena but did reduce the selectivity of galena against sphalerite. Changing of the grinding media used for the ore sample from a high carbon steel to a high chromium alloy steel resulted in a significant increase in the maximum recovery and flotation rate constant of the fine galena. EDTA (ethylene diaminetetraacetic) extractable iron measured for the high carbon steel media were similar in magnitude to those measured within the plant and were higher than those measured for the high chromium alloy steel media. The increased surface concentration of hydrophilic layers of oxidised iron species on the fine galena was a likely reason for their diminished flotation behaviour both in the laboratory and in the plant.     

Effect of grinding media and NaHS on copper recovery at Northparkes Mines

The flotation behaviour of copper minerals in the Northparkes Mines copper-gold ore is dependent on grinding media (and hence grinding Eh) and aeration during conditioning. Laboratory grinding and flotation tests were conducted using mild steel and stainless steel rods with and without the addition of sodium hydrosulphide (NaHS) in the conditioning stage. In the absence of NaHS, grinding using stainless steel media increased the kinetics of subsequent copper mineral flotation compared to the use of mild steel media. Flotation kinetics after grinding using mild steel media were increased by aeration prior to reagent addition. The addition of NaHS after grinding using mild steel rods increased the froth stability by increasing froth mineralisation, thus increasing the flotation kinetics of the copper minerals. When stainless steel grinding media was used, the copper flotation recovery was very high (>90 % in the rougher stage) without the addition of NaHS, so its use was not warranted. The flotation of the bornite was not reduced by oxidation, in fact copper mineral recovery was the highest under the most oxidising grinding conditions.     

Flotation scale up: use of separability curves

Separability curves (mineral recovery versus yield) have been used to characterize the copper flotation process both at batch laboratory scale and industrial plant scale (rougher banks). Then, to approach the scale-up problem the rougher bank operation and the batch were compared using the corresponding separability curves. Comparison was made at the maximum separation efficiency point in both operations. Thus, a time factor was established for optimal technical separation. The time factor can then be used for kinetic scale-up models, together with the ratio between minerals recovery in both operating scales. Experience from several tests recorded over a period of 10 months in an industrial concentrator showed a good consistency for scaling-up the rougher flotation recovery from batch tests within a 1% absolute error range.The effect of particle size and air flowrate in laboratory batch tests was evaluated in the space of separability curves, regarding their effect on recovery at the optimum separability point. Also, the effect of pulp level and particle size on the bank flotation kinetics was evaluated in an industrial flotation circuit. Thus, estimation of recovery changes due to variations in mineral characteristics and operating conditions was explored.     

国外某大型铜矿选矿厂流程考查与分析

对国外某大型铜矿选矿厂铜钼混合浮选生产流程进行了考查。结果表明, 选矿厂矿石处理量略低于设计产能, 铜精矿品位41.63%、回收率88.73%, 但浮选给矿粒度组成不合理, 难选粒级+212 μm粒级和-10 μm粒级含量较高; 磨矿分级优化空间较大, 通过优化入选粒度组成可提高金属回收率。     

某低品位金铜矿石浮选柱应用试验研究

某低品位金铜矿石含铜0.46%、金0.18 g/t,矿石中铜矿物主要以蓝辉铜矿、辉铜矿、铜蓝、硫砷铜矿等次生铜矿物存在,其可浮性好但容易过磨,造成浮选时细粒级损失较高,试验采用浮选柱+浮选机联合选别与单独采用浮选机相比,其它指标相当的情况下,铜精矿品位提高9.6%,硫精矿回收率提高9.23%,试验表明浮选柱对提高精矿品质、简化流程和强化细粒级回收方面具有较为明显地优势。     

不同磨矿介质对金川铜镍矿浮选的影响研究

以金川硫化铜镍矿石为研究对象,通过浮选试验研究了不锈钢球、锆球、铸铁球三种介质磨矿时硫化铜镍矿石的浮选行为。结果表明,锆球磨矿可获得最高的回收率,铸铁球磨矿时脉石上浮量大。使用AFM、ICP等研究了磨矿产品表面形貌、电化学性质等的差异,结果表明,锆球磨矿时颗粒表面较为平整,矿浆电位较高,铸铁球磨矿时,颗粒表面较为粗糙,矿浆中检出较多铁离子,矿浆电位较低,这些差别导致了矿石浮选行为的差异。     

云南某铅锌多金属矿选矿试验研究

在对云南某铅锌多金属矿进行简单工艺矿物学研究的基础上,按拟定的铜铅混浮-铜铅分离-锌硫混浮-锌硫分离原则流程进行了磨矿细度、药剂种类及用量条件试验,采用1粗1扫2精混浮铜铅、1粗1扫2精铜铅分离、1粗1扫2精混浮锌硫、1粗1扫2精锌硫分离、中矿顺序返回的闭路试验流程处理该矿样,获得了铅品位45.26%、回收率81.33%的铅精矿,锌品位45.97%、回收率88.29%的锌精矿,分选指标理想,但综合回收产品铜精矿和硫精矿的指标有待提高。     

Integrated simulation of grinding and flotation application to a lead-silver ore

The paper outlines a procedure to simulate a grinding -flotation plant. Population balance models are used to describe the grinding, classification and -flotation processes. In the grinding and classification models particles are characterized by their size, while in the flotation models particles are characterized by their size and mineral composition. The link between the grinding and flotation circuits is made by an empirical model that generates the mineral-size particle population from the size distribution of the grinding circuit product. The simulator is calibrated using plant data collected from the Peñoles-Fresnillo PbAg concentrator in Mexico.     

新疆某铜镍矿石工艺矿物学研究

为给新疆某大型低品位强氧化铜镍硫化矿石的开发利用提供技术依据,进行了工艺矿物学和混合浮选研究。结果表明：①矿石铜品位0075%、镍品位057%,铜、镍均主要以硫化矿的形式存在,其中硅酸镍难以回收;②矿石中的主要目的矿物为黄铜矿和镍黄铁矿,均可通过浮选回收,脉石以橄榄石为主;③镍黄铁矿在镜下呈自形、半自形粒状均质体,其中呈不规则颗粒状、与磁黄铁矿或黄铜矿以多种不同形态嵌连紧密的镍黄铁矿能较好地通过浮选回收,呈微细粒分布、形状不一和呈不规则粒状或蠕虫状及浸染状的镍黄铁矿因嵌布粒度微细而难以实现单体解离,从而不易通过浮选回收;黄铜矿则常呈不规则粒状、浸染状零星嵌布在脉石中;④磨矿（-0.074 mm 80%）-1粗1精2扫、中矿顺序返回闭路浮选流程可获得镍品位为9.17%、铜品位为1.57%,镍回收率68.01%、铜回收率87.37%的混合精矿,铜、镍富集效果较好。     

Flotation of copper sulphides assisted by high intensity conditioning (HIC) and concentrate recirculation

This paper describes the effect of the partial concentrate (rougher floated product) recirculation to rougher flotation feed, here named concentrate recirculation flotation – CRF, at laboratory scale. The main parameters used to evaluate this alternative approach were flotation rate and recovery of fine (“F” 40–13 μm) and ultrafine (“UF” <13 μm) copper sulphide particles. Also, the comparative effect of high intensity conditioning (HIC), as a pre-flotation stage for the rougher flotation, was studied alone or combined with CRF. Results were evaluated through separation parameters, grade-recovery and flotation rates, especially in the fine and ultrafine fractions, a very old problem of processing by flotation. Results showed that the floated concentrate recirculation enhanced the metallurgical recovery, grade and rate flotation of copper sulphides. The best results were obtained with concentrate recirculation flotation combined with high intensity conditioning (CRF–HIC). The kinetics rate values doubled, the Cu recovery increased 17%, the Cu grade increased 3.6% and the flotation rates were 2.4 times faster. These were accompanied by improving 32% the “true” flotation values equivalent to 2.4 times lower the amount of entrained copper particles. These results were explained and proved to proceed by particle aggregation (among others) occurring after HIC, assisted by the recycled floatable particles. This “artificial” increase in valuable mineral grade (by the CR) resulted in higher collision probability between hydrophobic particles acting as “seeds” or “carrier”.     

The effect of the design of a secondary grinding circuit on platinum flotation from a UG-2 ore

Platinum concentrator plants experience significant losses in their overall Platinum Group Elements (PGE) recoveries due to the inefficiencies of their secondary grinding circuits. This study involves an investigation of selective grinding of the platinum-bearing silicate particles present in UG-2 platinum ores found in the Bushveld Igneous Complex (BIC).Batch-scale laboratory test work was done to investigate the effect of a secondary milling circuit configuration, using a hydrocyclone underflow sample from a UG-2 concentrator plant as feed material. The envisaged secondary milling circuit consists of a conventional hydrocyclone to de-slime the feed followed by density separation with a spiral concentrator to separate the ore into lights (silicates-rich) and heavies (chromite-rich) fractions, followed by separate milling of the two fractions in parallel ball mills, and combined rougher flotation. A full-scale spiral was run in batch mode, followed by separate milling of samples in a 200 mm diameter mill and combined flotation in a 4.2 l cell. The milling energy inputs were re-distributed between the lights and heavies mills to determine the effect on the platinum mineral rougher flotation recovery and the Cr entrainment.The most promising results were found with 88% of the energy input to the lights mill and 12% to the heavies mill. The results indicated that under batch conditions, the secondary rougher flotation recovery (69% 4E) was similar to the conventional mill-float circuit (70%) however the Cr entrainment was significantly reduced by approximately 40% (2.3–1.4% Cr).This test work has confirmed the benefit of separate milling in the secondary milling circuit for a UG-2 ore. Spiral concentrators have shown potential as an effective density separating device to produce a silicate-rich and chromite-rich fraction for milling; further test work will be conducted to confirm its viability on an industrial scale.     

优化冬瓜山铜矿磨矿产品粒度组成提高浮选指标研究

针对冬瓜山入浮颗粒粒度较粗且粒度组成分布不合理问题,基于磨矿产品的粒度分布及矿石力学性质对磨矿介质配比进行调整以优化入浮颗粒的粒度组成,结果表明:冬瓜山一段磨矿介质尺寸方案为m（φ60）:m（φ40）:m（φ30）:m（φ25）=40:10:30:20,采用推荐方案可提高磨矿产品中-0.1+0.01 mm颗粒产率2.28%。推荐方案与现场方案磨矿产品经一粗两精两扫的浮选闭路对比试验,推荐方案铜精矿回收率90.11%,较现场方案提高1.34%,精矿品位提高了0.94%。对浮选尾矿筛分并检测分析可知推荐方案磨矿产品在-0.1+0.01 mm颗粒中铜的回收效果优于现场方案,利用推荐的介质配比方案优化磨矿产品粒度组成,有效提高了冬瓜山选铜浮选指标。      

The effect of particle breakage mechanisms during regrinding on the subsequent cleaner flotation

Stirred mills have been widely used for regrinding, and are acknowledged to be more energy efficient than tumbling mills. These two types of mills present different particle breakage mechanisms during grinding. In this study, the effect of regrinding by both mills on surface properties and subsequent mineral flotation was studied, using chalcocite as the mineral example. A rod mill and a stirred mill with the same stainless steel media were used to regrind rougher flotation concentrates. Different chalcocite flotation recovery was achieved in the cleaner stage after regrinding in tumbling and stirred mills. The factors contributing to the different recovery included particle size, the amount of created fresh surfaces, surface oxidation and the redistribution of collector carried from rougher flotation. All the factors were examined. It was determined that the predominating factor was the different distribution of collector resulting from different particle breakage mechanisms in the stirred and tumbling mills, in line with ToF-SIMS analysis. In the tumbling mill, the impact particle breakage mechanism predominates, causing the collector to remain on the surface of newly produced particles. In the stirred mill, the attrition breakage removes collector from the surface, and decreases particle floatability. Furthermore, the type of grinding media in the stirred mill also influences the subsequent flotation, again due to the change of particle breakage mechanisms. The results of this study demonstrate that the selection of regrinding mills and grinding media should not only depend on the required energy efficiency, but also on the properties of the surfaces produced for subsequent flotation.     

微细粒金矿石选矿试验研究

主要对微细粒金矿石进行了选矿试验研究。通过粗选磨矿细度试验和药剂条件的正交试验，确定了该矿粗选的最佳操作条件。在此基础上进行一粗二精三扫的实验室闭路试验．可获得金品位48．04g／t，回收率85．63％的金精矿，这一结果表明，可通过浮选硫化矿进行富集金。     

某天青石矿浮选工艺研究

论述了某天青石矿浮选试验研究结果。研究表明,采用SHS为调整剂、以BK-12S为捕收剂、以PG为起泡剂,经两次粗选、三次精选、一次扫选流程,可以取得较好的浮选分离指标。天青石性脆易过磨,部分天青石矿物嵌布粒度较细,包裹于方解石矿物中,细磨难以单体解离。     

Grinding operation optimization of the CODELCO-Andina concentrator plant

This paper presents results following the application of a sub-optimal control scheme, both through simulation and in situ, from the operation of Section C of the CODELCO-Andina copper concentrator plant. The algorithm permits the determination of the necessary control action at each instant of time in order to maximize a defined plant performance index. The main objective of the algorithm is to maximize the mineral tonnage processed by the section, subject to it not exceeding a predetermined value establishes! for the operation conditions of the mills, while at the same time maintaining constant the mass fraction over 65 mesh (212 [microns]) in the overflow of the hydrocyclones, at a value within the operational requirements of the flotation stage. The performance index is defined in terms of; the percentage of pulp solids fed to the hydrocyclones of each line of ball mills, penalty functions to prevent electric power to the ball mills ,falling below the lower limit (so as not to enter the overload region), the tonnage processed in the section, and, since water is a scarce resource, a term considering the water added to each sump is also included. The scheme is first studied and adjusted in a simulator of a concentrator plant similar to that used in the industrial application. For plant implementation, a PC software program, denoted CONMOL, is developed in TurboPascal for Windows. This software allows plant applications to be carried out through a communications interface. Finally, the results of two tests performed on Section C of the CODELCO-Andina copper concentrator plant are shown where the control is applied over 3 and 5.5 hours respectively.     

澳斯麦特炉渣选铜工艺研究与生产实践

某公司澳斯麦特炉渣中的铜主要为硫化铜,其次为少量的金属铜,还有微量的氧化铜、易溶铜盐和其它铜。铜矿物嵌布粒度细且不均匀,呈粒状、浸染状、星点状分布。通过缓冷工艺、磨矿和浮选药剂等的试验研究,确定了两段磨矿分级后进行铜浮选的原则流程,并在原诺兰达炉渣磨浮生产工艺基础上进行技术改造。澳斯麦特炉渣选铜多年生产实践的结果表明,当原渣品位1.152%时,获得的精矿品位19.31%,尾矿品位0.243%,选铜回收率79.91%,生产实践取得成功。     

贵州某铅细粒嵌布的高硫铅锌矿浮选工艺研究

针对贵州某铅和硫嵌布粒度细、硫含量较高的铅锌矿开展浮选工艺研究。结果表明,磨矿细度-0.074mm占60%,采用优先浮选流程,铅浮选流程为"一粗三精三扫"、锌浮选流程为"一粗三精三扫"、硫浮选流程为"一粗一精二扫",能获得合格精矿,铅精矿中铅品位43.29%、回收率78.33%,锌精矿中锌含量为44.90%、回收率91.21%,硫精矿硫含量为45.85%、回收率为58.99%。     

The separation of chalcopyrite and chalcocite from pyrite in cleaner flotation after regrinding

In this study, the differences between the separation of chalcopyrite and chalcocite from pyrite in cleaner flotation after regrinding were investigated. In the rougher flotation prior to regrinding, high chalcopyrite and chalcocite recovery were obtained in conjunction with high pyrite flotation recovery due to the activation of pyrite by copper ions during primary grinding. The rougher flotation concentrate was reground in a rod mill before cleaner flotation. It was found that chalcopyrite and chalcocite exhibited different flotation behavior and also affected pyrite flotation differently in cleaner flotation. The mechanism underpinning these phenomena was investigated by a range of techniques including the polarization of mineral electrodes, X-ray photoelectron spectroscopy (XPS) analyses and ethylene diamine tetraacetic acid (EDTA) extraction. It was found that the flotation behavior of both copper minerals and their effect on pyrite flotation after regrinding were governed by their electrochemical activities and galvanic coupling with pyrite.     

悬振选矿机在弓长岭选矿厂铁尾矿再选中的应用

弓长岭选矿厂铁浮选尾矿,品位高,粒度细,-0.074 mm含量约65%,铁矿物在细粒级-0.019 mm富集明显。根据弓长岭选矿厂铁浮选尾矿的矿石性质,利用微细粒级重选设备-悬振选矿机对该尾矿进行再选试验研究,通过分级分选,细粒级部分一次悬振选别可获得品位64.35%,回收率30.93%的铁精矿,粗粒级通过磨矿后（磨矿细度-0.074 mm 85%）再悬振分选,获得的精矿铁品位为59.93%,回收率9.80%,综合铁精矿品位63.22%,回收率40.73%,综合尾矿铁品位降至12.58%,有效的回收了该尾矿中的铁,为弓长岭选矿厂的铁浮选尾矿回收与再利用提供可选方案,其社会及经济效益显著。