Use of centrifugal-gravity concentration for rejection of talc and recovery improvement in base-metal flotation

The possibility of using a centrifugal-gravity concentrator to reject Mg-bearing minerals and minimize metal losses in the flotation of base metals was evaluated. Sample characterization, batch scoping tests, pilot-scale tests, and regrind-flotation tests were conducted on a Ni flotation tailings stream. Batch tests revealed that the Mg grade decreased dramatically in the concentrate products. Pilot-scale testing of a continuous centrifugal concentrator（Knelson CVD6） on the flotation tailings revealed that a concentrate with a low mass yield, low Mg content, and high Ni upgrade ratio could be achieved. Under optimum conditions, a concentrate at 6.7% mass yield was obtained with 0.85% Ni grade at 12.9% Ni recovery and with a low Mg distribution（1.7%）. Size partition curves demonstrated that the CVD also operated as a size classifier, enhancing the rejection of talc fines. Overall, the CVD was capable of rejecting Mg-bearing minerals. Moreover, an opportunity exists for the novel use of centrifugal-gravity concentration for scavenging flotation tailings and/or after comminution to minimize amount of Mg-bearing minerals reporting to flotation.

Effect of Flotation Reagents on the Cake Moisture of Copper Concentrate

The effect of reagents used in separating chalcopyrite from pyrite on the cake moisture of the copper concentrate at Daye Iron Mine Mineral Processing Plant was investigated. The results showed that the dosage of lime used for depressing pyrite was the main factor that increased the filter cake moisture of copper concentrate. With increasing the dosage of lime, the cake moisture of copper concentrate increased sharply. The cause was concluded to be the addition of lime to the pulp, which resulted in the formation of floc and a high pH value. The collector Z-200#, used for collecting chalcopyrite, had, as well, an adverse effect on the cake moisture of copper concentrate, but its effect was inferior in respect to that of lime. The cake moisture of copper concentrate can be decreased by changing the method with which lime is added and the pH value of pulp is regulated. The experiment results showed that the sulfuric acid was the best regulator. When the clarified liquor of lime was used as a depressant and the pH value of the pulp was regulated to 6.5€*7.0 by adding sulfuric acid, the cake moisture of copper concentrate was reduced from 15.49% to 13.13%. The examination of chalcopyrite surface by using ESCA (Electron Spectroscopy for Chemical Analysis) showed that calcium sulfate and iron hydroxide had formed on the surface of chalcopyrite when lime was added to the pulp. The formation of calcium sulfate and iron hydroxide on its surface increased the hydrophilicity of chalcopyrite so that its cake moisture increased. The addition of sulfuric acid to the pulp not only removed the calcium sulfate, but also reduced the concentration of iron hydroxide on the surface of chalcopyrite so that the cake moisture of copper concentrate was decreased.

Flotation-Calcination-Magnetic Separation Hybrid Process for Concentration of Rare Earth Minerals Contained in a Carbonatite Ore

A hybrid process consisting of flotation and magnetic separation has been developed to concentrate multi-phase rare earth minerals associated with a carbonatite ore that contains a significant amount of niobium. The deposit is known to contain at least 15 different rare earth minerals identified as silicocarbonatite, magnesiocarbonatite, ferrocarbonatites, calciocarbonatite, REE/Nb ferrocarbonatite, phosphates and niobates. Although no collector exists to float all the different rare earth minerals, the hydroxamic acid-based collectors have shown adequate efficiency in floating most of these minerals. 92% recovery of total rare earth oxide (TREO) and niobium in 45% mass was possible at d80 of <65 microns grind size. It was also possible to reduce the mass pull to 28%, but TREO and Nb’s recovery dropped to 85%. Calcination of the concentrate followed by quenching and fine grinding to <25 μm allowed upgrading the flotation concentrate by magnetic separation. It was demonstrated that at least 87% TREO and 85% Nb could be recovered in 16% of the feed mass. The paper discusses the overall concept of the flowsheet and the experimental strategies that led to this process.

CFD model simulation of bubble surface area flux in flotation column reactor in presence of minerals

Bubble surface area flux(S_b) is one of the main design parameter in flotation column that typically employed to describe the gas dispersion properties, and it has a strong correlation with the flotation rate constant. There is a limited information available in the literature regarding the effect of particle type,density, wettability and concentration on Sb. In this paper, computational fluid dynamics(CFD) simulations are performed to study the gas–liquid–solid three-phase flow dynamics in flotation column by employing the Eulerian–Eulerian formulation with k-e turbulence model. The model is developed by writing Fortran subroutine and incorporating then into the commercial CFD code AVL FIRE, v.2014.This paper studies the effects of superficial gas velocities and particle type, density, wettability and concentration on Sband bubble concentration in the flotation column. The model has been validated against published experimental data. It was found that the CFD model was able to predict, where the response variable as indicated by R-Square value of 0.98. These results suggest that the developed CFD model is reasonable to describe the flotation column reactor. From the CFD results, it is also found that Sb decreased with increasing solid concentration and hydrophobicity, but increased with increasing superficial gas velocity. For example, approximately 28% reduction in the surface area flux is observed when coal concentration is increased from 0 to 10%, by volume. While for the same solid concentration and gas flow rate, the bubble surface area flux is approximately increased by 7% in the presences of sphalerite.A possible explanation for this might be that increasing solid concentration and hydrophobicity promotes the bubble coalescence rate leading to the increase in bubble size. Also, it was found that the bubble concentration would decrease with addition of hydrophobic particle(i.e., coal). For instance, under the same operating conditions, approximately 23% reduction in the bubble concentration is predicted when the system was working with hydrophobic particles. The results presented are useful for understanding flow dynamics of three-phase system and provide a basis for further development of CFD model for flotation column.

Surface nanobubbles on the hydrophobic surface and their implication to flotation

Nanobubbles play a potential role in the application of the flotation of fine particles.In this work,the identification of nanoentities was performed with a contact mode atomic force microscope(AFM).Moreover,the influences of setpoint ratio and amplitude of the cantilever and the responses of the formed surface nanobubbles to the fluctuation of pH,salt concentration,and surfactant concentration in the slurry were respectively studied.Nanobubbles were reported on the highly oriented pyrolytic graphite(HOPG)surface as the HOPG was immersed in de-ionized water under ambient temperature.The coalescence of nanobubbles occurred under contact mode,which provides strong evidence of the gaseous nature of these nanostructures on HOPG.The measuring height of the surface nanobubbles decreased with the setpoint ratio.The changes in the pH and concentration of methyl isobutyl carbinol(MIBC)show a negligible influence on the lateral size and height of the preex-isting surface nanobubbles.The addition of LiCl results in a negligible change of the lateral size;however,an obvious change is noticed in the height of surface nanobubbles.The results are expected to provide a valuable reference in understanding the properties of surface nanobubbles and in the design of nanobubble-assisted flotation processes.

Correspondence of bubble size and frother partitioning in flotation

The size of bubbles created in the flotation process is of great importance to the efficiency of the mineral separation achieved.Meanwhile,it is believed that frother transport between phases is perhaps the most important reason for the interactive nature of the phenomena occurring in the bulk and froth phases in flotation,as frother adsorbed in the surface of rising bubbles is removed from the bulk phase and then released into the froth as a fraction of the bubbles burst.This causes the increased concentration in the froth compared to the bulk concentration,named as frother partitioning.Partitioning reflects the adsorption of frother on bubbles and how to influence bubble size is not known.There currently exists no such a topic aiming to link these two key parameters.To fill this vacancy,the correspondence between bubble size and frother partitioning was examined.Bubble size was measured by sampling-for-imaging(SFI)technique.Using total organic carbon(TOC)analysis to measure the frother partitioning between froth and bulk phases was determined.Measurements have shown,with no exceptions including four different frothers,higher frother concentration is in the bulk than in the froth.The results also show strong partitioning giving an increase in bubble size which implies there is a compelling relationship between these two,represented by CFroth/CBulk and D32.The CFroth/CBulkand D32 curves show similar exponential decay relationships as a function of added frother in the system,strongly suggesting that the frother concentration gradient between the bulk solution and the bubble interface is the driving force contributing to bubble size reduction.

Effect of some operational variables on bubble size in a pilot-scale mechanical flotation machine

This work aims to provide a relationship of how the key operational variables of frother type and impeller speed affect the size of bubble(D32). The study was performed using pilot-scale equipment(0.8 m3) that is up to two orders of magnitude larger than equipment used for studies performed to date by others, and incorporated the key process variables of frother type and impeller speed. The results show that each frother family exhibits a unique CCC95-HLB relationship dependent on n(number of C-atoms in alkyl group) and m(number of propylene oxide group). Empirical models were developed to predict CCC95 from HLB associated with other two parameters α and β. The impeller speed-bubble size tests show that D32 is unaffected by increased impeller tip speed across the range of 4.6 to 9.2 m/s(representing the industrial operating range), although D32 starts to increase below 4.6 m/s. The finding is valid for both coalescing and non-coalescing conditions. The results suggest that the bubble size and bubble size distribution(BSD) being created do not change with increasing impeller speed in the quiescent zone of the flotation.

Estimation of Copper and Molybdenum Grades and Recoveries in the Industrial Flotation Plant Using the Artificial Neural Network

In this paper, prediction of copper and molybdenum grades and their recoveries of an industrial flotation plant are investigated using the Artificial Neural Networks (ANN) model. Process modeling has done based on 92 datasets collected at different operational conditions and feed characteristics. The prominent parameters investigated in this network were pH, collector, frother and F-Oil concentration, size percentage of feed passing 75 microns, moisture content in feed, solid percentage, and grade of copper, molybdenum, and iron in feed. A multilayer perceptron neural network, with 10:10:10:4 structure (two hidden layers), was used to estimate metallurgical performance. To obtain the optimal hidden layers and nodes in a layer, a trial and error procedure was done. In training and testing phases, it achieved quite correlations of 0.98 and 0.93 for Copper grade, of 0.99 and 0.92 for Copper recovery, of 0.99 and 0.92 for Molybdenum grade and of 0.99 and 0.94 for Molybdenum recovery prediction, respectively. The proposed neural network model can be applied to determine the most beneficial operational conditions for the expected Copper and Molybdenum grades and their recovery in final concentration of the industrial copper flotation process.

Investigation on the Flotation Mechanism of Tantalite-Niobite Minerals with Styrenephosphonic Acid Collector

The flotability and mechanism of tantalite-niobite minerals which have different contents of Ta, Nb, Fe Mn using styrenephosphonic acid (SPA) were investigated. The optimum pH value in pulp is less than 5, flotability varies with the different contents of mineral, and flotability of minerals with higher Nb2O5 content is better than that with higher Ta2O5 content. In wide range of pH in solution, the Fe, Mn on the surface are the main active species for flotation. Research using Auger electron spectroscopy, infrared spectroscopy and ultraviolet spectroscopy for determining surface conditions has been done. SPA interacts with Fe, Mn ions on the surface by a chemical bond force. On the outer layer of the collector, which is a non-uniform multilayer, adsorption of SPA on surface is both chemico-adsorption and physico-adsorption, but chemico-adsorption is dominant.

2^kFactorial Experiments on Factors that Influence the Recovery of Gold during the Upgrade of Ilesha-Itagunmodi Gold Ore through Froth Flotation

A low-grade gold ore from Ilesha-Itagunmodi, south western Nigeria was panned, dried and ball-milled before sieving into fractions. The morphological, mineralogical and chemical composition was studied by optical micro-scopy (Reflected and Transmission), X-ray diffraction (XRD) and Energy Dispersive X-ray fluorescence (XRF), respectively. The sieved fractions were subjected to chemical analysis (AAS). The +106 μm sieve had the highest concentration of the mineral and was then selected for the upgrade through froth flotation using standard reagents. In this paper a report on a 2k factorial experiment that provides an understanding of the impact of operational variables on the quantity of gold mineral obtained from the ore during froth flotation is presented. Analysis results showed that Ilesha-Itagunmodi gold ore is non-refractory with fine grain particles, amenable to froth flotation and contained about 20 other associated minerals, gold had a concentration of about 0.0024%. A com-bination of P-Xanthate and amine glycol collectors at a pH of 9.2 only produced a considerable increase in gold yield. This translated to about 87.13% increase in recovery of gold from the ore. Analysis of variance (ANOVA) was carried out and the model equation obtained was subsequently optimised to obtain a model equation that could be used in predicting the recoverable quantity of gold, indicating that F11,1-values for Collector concentration, Frother concentration, pH and Conditioning time were 156.86, 6.96, 43.81, and 56.77 respectively. A model with an F value of 88.41 was obtained which indicated that the model was significant. The model equation obtained was subsequently optimised to be able to predict the recoverable quantity of gold. A “Pred R-Squared” value of 0.9365 (93.65%) was also obtained and is in reasonable agreement with the “Adj R-Squared” value of 0.9534 (95.34%). It was established that Ilesha placer gold ore is amenable to froth flotation using standard reagents.

基于浮选泡沫图像预测精矿品位的研究进展

随着人工智能技术在矿业生产的广泛应用,利用计算机视觉技术提高精矿品位预测的准确性和效率已成为必然趋势。在综述了传统图像处理算法和深度学习算法在精矿品位预测中的应用与发展历程基础上,并探讨了未来的发展趋势和挑战。传统图像处理技术通过提取泡沫图像的尺寸、颜色、纹理和流速等特征,结合分水岭分割、颜色矩、灰度共生矩阵和局部点特征匹配等算法进行特征提取。这些特征在计算资源有限的场景中具有一定的应用价值,但在应对精矿品位预测任务时精度较低。深度学习技术通过构建合适的模型架构并利用大量数据进行训练,能够提取高层语义特征,具有较高的预测精度,与图形处理单元(GPU)等高效运算设备配合使用,可实现高性能和高效率的统一。介绍了支持向量机(SVM)、极限学习机(ELM)等机器学习算法以及多层感知器(MLP)、全连接层和多尺度特征融合等深度学习算法在特征映射和品位预测中的应用,以及深度学习模型的发展历程。最后综述了工业界视觉检测系统的应用现状,并从数据驱动模型、多模态数据融合、算法实时性和数据集规模等方面分析了该领域所面临的挑战和未来发展趋势。

硫精矿中有价金属铜金银的综合回收利用实验

这是一篇矿物加工工程领域的论文。对某含铜金银硫精矿进行了选矿综合回收实验研究,实验结果表明,采用异丁基黄药+乙硫氮组合捕收剂,氧化钙做抑制剂,2^(#)油做起泡剂,经过一次粗选-三次精选-二次扫选的浮选工艺流程,最终可得到铜品位为16.66%、回收率为86.96%,金品位为278.95 g/t,回收率为75.56%;银品位为1848.74 g/t,回收率为78.55%的精矿产品,有效地实现了硫精矿中铜的综合回收,同时使金、银很好地富集在铜精矿中,便于回收利用,选矿指标理想。

酒钢悬浮焙烧磁铁矿选矿工艺优化研究

对酒钢悬浮焙烧磁铁矿进行了磨矿-选矿工艺优化研究。采用110 mT低场强磁选机预提精、中矿再磨后采用磁选富集抛尾、富集精矿经阳离子反浮选,最终获得TFe品位60.06%、SiO_(2)含量5.17%、金属回收率84.27%的综合精矿;与原工艺指标相比,精矿TFe品位提高了1.96个百分点,金属回收率提高了1.38个百分点,SiO_(2)含量降低了1.06个百分点。

甘肃某细粒铅锌矿工艺优化试验研究

甘肃某铅锌矿中铅品位为0.74%,锌品位为3.72%,为了实现该矿产资源的高效回收利用,针对铅锌矿物嵌布关系紧密,粒度较细,造成铅锌精矿互含量较高,影响浮选指标的问题,在原矿性质研究的基础上,进行了4种不同方案的工艺流程对比试验。试验结果表明:采用铅锌优先浮选—铅、锌粗精矿均再磨的工艺流程,最终获得的铅精矿铅品位为58.42%,铅回收率为79.95%,锌精矿锌品位为56.27%,锌回收率为93.05%。相较于方案2,锌精矿锌品位提高了3.03个百分点,锌回收率提高了2.75个百分点;相较于方案1现场生产工艺,铅、锌主品位提高,铅、锌互含量显著降低,锌回收率提高了4.15个百分点。

基于ISSA-HKLSSVM的浮选精矿品位预测方法

针对浮选过程变量滞后、耦合特征及建模样本数量少所导致精矿品位难以准确预测的问题,提出了一种基于改进麻雀搜索算法(Improved Sparrow Search Algorithm,ISSA)优化混核最小二乘支持向量机(Hybrid Kernel Least Squares Support Vector Machine,HKLSSVM)的浮选过程精矿品位预测方法.首先采集浮选现场载流X荧光品位分析仪数据作为建模变量并进行预处理,建立基于最小二乘支持向量机(Least Squares Support Vector Machine,LSSVM)的预测模型,以此构建新型混合核函数,将输入空间映射至高维特征空间,再引入改进麻雀搜索算法对模型参数进行优化,提出基于ISSA-HKLSSVM方法实现精矿品位预测,最后开发基于LabVIEW的浮选精矿品位预测系统对本文提出方法实际验证.实验结果表明,本文提出方法对于浮选过程小样本建模具有良好拟合能力,相比现有方法提高了预测准确率,可实现精矿品位的准确在线预测,为浮选过程的智能调控提供实时可靠的精矿品位反馈信息.

酒钢肃南宏兴矿业铜尾矿中有价元素回收试验研究

酒钢肃南宏兴矿业铜矿石中伴生有价元素Fe、Ba,Fe主要以菱铁矿的形式赋存、Ba以重晶石的形式赋存,浮选选铜后有价元素Fe、Ba富集于尾矿中。试验确定铜尾矿有价元素回收工艺为强磁选-悬浮焙烧-磨矿-弱磁选回收铁精矿,强磁选尾矿浮选回收重晶石。采用此工艺,在铜浮选尾矿含Fe 10.8%、BaO 4.7%的情况下,试验回收铁精矿含Fe 56.4%,回收重晶石含BaSO 488.79%。由于Fe回收只需新建强磁选和脱水工序,强磁选回收铁矿粉进酒钢选矿厂现有悬浮磁化焙烧选别系统即可,投资小、易实施。

熟化银精矿氧压强化酸浸行为研究

考虑到浮选银精矿极具回收价值及现有处理工艺存在不足,本文提出了浓硫酸熟化-氧压酸浸的工艺路线。首先,以湿法炼锌过程产出的浮选银精矿为对象,利用浓硫酸的强氧化性破坏浮选残留有机药剂矿物包裹体;其次,利用加压湿法冶金独特的技术优势,氧化溶解低价金属硫化物;最后,重点研究了工艺过程原理及氧压酸浸过程中Zn、Cu、In的浸出行为及浸出渣中物相演变规律。研究结果表明:银精矿颗粒表面残留浮选药剂的P—O、—OH极性基团被浓硫酸氧化破坏,避免产生水化膜及矿物包裹体,促进有价元素的高效浸出;在高温氧化性条件和Fe^(3+)/Fe^(2+)变价离子对的助浸作用下,银精矿物相组成由闪锌矿(Zn S)、锌铁尖晶石(Zn Fe_(2)O_(4))、雌黄铁矿(FeS)、氯化亚铜(Cu Cl)为主,转变为以单质硫(S8)、铁矾(MeFe_(3)(SO_(4))_(2)(OH)_(6),Me:1/2Pb、Na、K)为主,并残留少量的闪锌矿(Zn S);适当提高反应温度和初始质量浓度可促进Zn、Cu、In的浸出,但温度过高会加剧铟铁共沉淀问题,降低锌、铟的浸出率,而质量浓度过高会降低氧气对Zn S及Fe^(2+)的氧化能力,抑制银精矿的溶解;延长反应时间、提高液固比均有利于银精矿的解离浸出;增大氧分压可以加强Fe^(3+)的氧化能力,促进对Zn S及ZnxFe(1-x)S的氧化作用,有利于银精矿的溶解。在反应温度为155℃、初始质量浓度为160 g/L、反应时间为180 min、氧分压为0.8 MPa、液固比为7 m L/g的氧压酸浸条件下,银精矿中Zn、Cu、In的浸出率分别为94.84%、98.36%和91.13%,浸出渣中铅、银的富集率分别达2.88%和7754 g/t。

云南某贫细硫化铜矿石选矿工艺优化研究

云南某低品位细粒嵌布的铜矿石铜、硫品位分别为0.161%和9.704%,有价铜矿物嵌布粒度细,铜硫分离富集难度较大。为实现该硫化铜矿石资源的高效开发利用,在选厂原选矿流程及原矿工艺矿物学研究基础上,进行了系统的磨浮试验。结果表明:矿石采用磨矿—1粗2扫铜硫混合浮选—混浮粗精矿再磨—1粗3精2扫铜硫分离浮选闭路流程处理,可获得铜品位14.69%、铜回收率63.15%的铜精矿,硫品位29.16%、硫回收率75.10%的硫精矿,试验铜精矿品位较现场提高近3个百分点,铜回收率提高15个百分点以上,显著优于现场生产指标,达到了理想的综合回收铜硫效果。研究结果可作为工艺流程优化的依据。

HYIC捕收剂强化鞍千混合磁选精矿反浮选效果试验

鞍千选矿厂为了优化选别工艺中的反浮选效果,取浮选入选的混合磁选精矿进行浮选试验,进行了抑制剂、活化剂、捕收剂用量及常规捕收剂TD-Ⅱ与新型捕收剂HYIC浮选效果对比试验。试验结果表明:固定矿浆pH值为11.5、抑制剂淀粉用量1200 g/t、活化剂CaO用量500 g/t、粗选捕收剂用量800 g/t、精选捕收剂用量300 g/t,采用常规捕收剂TD-Ⅱ时,闭路试验获得了精矿铁品位68.20%、铁回收率88.73%的选别指标;采用新型捕收剂HYIC时,闭路试验获得了精矿铁品位68.25%、铁回收率90.04%的选别指标;使用捕收剂HYIC比采用捕收剂TD-Ⅱ所获得的浮选精矿产率、铁品位、铁回收率分别高0.91,0.05,1.31个百分点,捕收剂HYIC可以强化反浮选效果,提高选厂的经济效益。