聚苯乙烯载体浮选微细粒白钨矿研究

微细粒百钨矿质量小,表面能大,在常规浮选过程中流失严重。文章研究了基于聚苯乙烯为载体粒子,对-19μm的白钨矿载体浮选行为及作用机理。结果表明:聚苯乙烯可显著提高微细粒白钨矿的浮选富集效果。在碱性溶液环境下,使用油酸钠作为捕收剂,粒径为-100μm+50μm的聚苯乙烯粒子在强搅拌条件下可实现对难回收的-19μm的白钨矿最大限度地回收;聚苯乙烯载体浮选工艺能从白钨矿-石英人工混合矿中有效分离白钨和石英;通过扫描电子显微镜观察,可发现微细粒白钨矿在聚苯乙烯表面发生单层及多层疏水性黏附,并通过白钨矿与载体的界面相互作用势能EDLVO理论计算证明了该白钨矿矿粒与载体之间存在巨大的疏水力作用。     

临汾盆地褐土剖面磁化率与粒度垂直变化特征及相关性研究

[目的]研究临汾盆地褐土剖面磁化率与粒度垂直变化特征及相关性.[方法]在临汾市尧都区农田取120 cm深的土壤剖面,每5 cm连续采样,对所取的24个土样进行磁化率和粒度的分析.[结果]磁化率随土壤深度增加而递减,>50 μm粒级含量随着土壤深度的增加而递减,10～50、1～10、<1 μm粒级含量随着土壤深度的增加而递增.地表以下0～30 cm为砂土,30～120 cm为粉土.磁化率与1～10、<1 μm粒级含量均呈极显著负相关,与>50 μm粒级含量呈极显著正相关,与10～50 μm粒级含量不相关.频率依赖系数XFD1.0与>50 μm粒级含量呈负相关,与10～50、1～10、<1 μm的粒级含量呈正相关;XFD0.1与10～50 μm粒级含量呈极显著负相关,与>50 μm粒级含量呈显著负相关.[结论]磁性颗粒主要赋存于粗颗粒物质中;磁化率及粒度对沉积环境和古气候的变化有一定的响应.     

用纯碱焙烧—浸出法从白钨矿精矿中提取钨

研究了用纯碱焙烧-浸出法处理白钨矿精矿。把碳酸钠掺入浸出剂,从而成功地消除了浸出时白钨矿从钨酸钠溶液中返沉。把浸出残渣经简单粒度分离,使粗粒级(+37μm)返回使用,可进一步提高钨的回收率。     

浮选柱法从浮选尾矿中回收微细粒级白钨矿的研究

为提高钨资源利用效率,针对湖南安化湘安钨业公司白钨浮选尾矿中微细粒级在浮选机中未能有效分选的特点,利用微泡技术开发了CMPT微泡浮选柱,利用专家系统控制浮选柱关键参数,确保浮选柱处于较好的工作状态.半工业试验和工业试验表明,微泡浮选柱能从浮选尾矿中回收微细粒级白钨.通过半工业试验获得了浮选柱的较优的工作参数,工业试验表明其平均精矿质量和回收率分别为:24.52%和43.41%,富集比35.03.水析试验表明5～10 μm,lO～19μm,19～38 μm3个粒级的回收率均达到65%.     

白钨粗选工艺流程优化研究及应用

某钼钨矿白钨浮选粗选工艺原采用一次粗选和一次扫选回收白钨矿,回收率较低。实验室筛析试验表明,在尾矿回收率较高的粒级中存在较多能被有效回收的白钨矿。为进一步提高回收率,通过对原有流程进行改造,采用一次粗选、两次扫选以及分段添加浮选剂的方式,延长浮选时间的同时强化浮选剂的作用效果,可以有效提高回收率。改造完成后,白钨矿浮选时间增加15.57 min,回收率提高1.06%,通过加温精选后,每年最终可增加白钨金属量83.22 t,增加净利润800万元左右,当年即可收回改造投资成本,经济效益显著。     

润磨活化机理初探

通过SEM、X衍射分析、激光粒度分析以及比表面积和润湿热的测定 ,对润磨预处理铁精矿的活化机理进行了研究。研究表明 ,润磨后精矿平均粒径明显减小 ,细粒含量增加 ,- 4 5 μm粒级增加近 11% ,而 1～ 10 μm粒级由 13%提高到 2 0 %。粒径的对数值接近正态分布。精矿的润湿热由 1.4 5× 10 - 5J cm2 提高到 4 .77× 10 - 5J cm2 。衍射峰趋于变宽 ,衍射强度趋于减弱 ,晶格变形增加     

某低品位难选白钨矿常温浮选试验研究

江西某白钨矿石含钨0.23%,原矿钨品位低、嵌布粒度较细、萤石含量高,是典型的白钨-方解石-萤石（或重晶石）型难选白钨矿.针对该难选白钨矿中有价金属钨无法有效回收的现状,对其进行了试验研究,结果表明,采用组合药剂（GYW＋731）做捕收剂,新型药剂WH做活化剂,实现了该低品位难选白钨矿的常温浮选,闭路试验获得了品位为35.11%,回收率为72.20%的钨精矿.     

某难选高硫含铜白钨矿选矿试验研究

某难选高硫含铜白钨矿中钨主要以白钨矿的形式存在,硫化铁主要以磁黄铁矿的形式存在。为给该矿石的开发利用提供技术支持,采用磁选-铜硫混合浮选-白钨浮选原则流程进行条件试验。结果表明,原矿磨矿至-74μm占65%时进行磁选,可获得品位为38.33%、回收率为51.14%的硫精矿,而磁选尾矿经铜硫混合-铜硫分离浮选,可分别获得品位为20.06%、回收率为73.12%的铜精矿和品位为35.20%、回收率为42.11%的硫精矿;其中铜硫混合浮选尾矿以碳酸钠为调整剂、水玻璃为抑制剂、731氧化石蜡皂为捕收剂,进行一粗一扫三精白钨常温浮选,可得到WO_3品位为63.93%、回收率为89.60%的白钨精矿,有效地实现了铜硫的分离和白钨矿的回收。     

钨矿表面特性与可浮性关系研究现状

钨矿选别过程中,黑、白钨矿分别主要采用重选法和浮选法回收.但因为重选难以高效回收微细粒黑钨矿以及浮选过程中白钨矿难以与含钙矿物有效分离等原因,导致钨矿回收率低或品位不理想.因此,有必要对黑、白钨矿的表面特性进行深入研究.介绍了黑、白钨矿表面特性的研究现状,主要评述了黑、白钨矿表面键的断裂、表面化学组成、表面异向性、表面电性以及表面溶解组分对钨矿可浮性的影响.对未来钨矿表面特性的研究方向和研究方法进行了展望.     

广西某白钨矿选矿试验研究

对广西某白钨矿进行了选矿试验研究。试验采用浮选脱硫-脱硫尾矿浮选选钨工艺。试验表明:采用碳酸钠做pH调整剂,硫酸铜做活化剂,丁黄药与丁铵黑药做捕收剂,脱硫效果好,硫脱除率90%以上;钨的浮选采用水玻璃、碳酸钠做抑制剂和G-O1做捕收剂效果佳。闭路试验采用一粗二扫二精浮选脱硫-脱硫尾矿一粗一扫五精选钨的浮选流程,可获得含硫51.22%的硫精矿,含WO371.27%、回收率为84.55%的钨精矿。     

赤铁矿的自载体作用及对浮选的影响

通过单矿物浮选试验、光学显微镜分析、E-DLVO理论计算、团聚动力学分析等研究了油酸钠浮选体系下赤铁矿浮选过程中的自载体作用。单矿物浮选试验表明，粗粒赤铁矿（?106 + 45 μm）的可浮性较好，当油酸钠用量超过15 mg·L?1时，回收率可达到90%以上，而细粒赤铁矿（?18 μm）的浮选回收率、浮选速率则较低；当粗?细赤铁矿中粗粒和细粒的质量近似相等时，粗粒的“自载体”效果最强，浮选回收率增加的也最明显，但粗粒过量则会导致粗粒对细粒赤铁矿浮选的强化作用减弱。光学显微镜分析和E-DLVO理论计算表明，粗?细赤铁矿颗粒间的相互作用能高于细粒赤铁矿间的相互作用能，与细粒赤铁矿相比，粗?细赤铁矿间更容易发生团聚，这也是粗粒能够强化细粒赤铁矿浮选（自载体作用）的主要原因。但过量的粗粒赤铁矿会增强其浮选过程中的“磨削、剪切”作用，导致粗粒的“自载体”效果减弱，浮选回收率降低。      

粒度大小对赤铁矿和石英浮选分离的影响

通过浮选试验、DLVO理论计算、聚焦光束反射测量（FBRM）等研究了油酸钠浮选体系下粒度大小对赤铁矿和石英浮选分离的影响。人工混合矿浮选试验表明,窄粒级粗粒或中等粒级的赤铁矿?石英混合矿（CH&CQ和MH&CQ）的浮选效果较好,其中CH&CQ和MH&CQ的分选效率分别为85.49%和84.26%,明显高于全粒级混合矿（RH&RQ）的分选效率74.94%;但窄粒级的细粒赤铁矿?石英混合矿（FH&FQ）的浮选效果则较差,其分选效率只有54.98%。浮选动力学试验表明,赤铁矿的浮选速率和回收率不仅与赤铁矿的粒度有关,还受石英粒度的影响,细粒脉石矿物石英会降低赤铁矿的浮选速率和回收率。DLVO理论计算表明,当矿浆pH值为9.0时,石英与赤铁矿颗粒间的相互作用力为斥力,此时细粒石英很难“罩盖”在赤铁矿表面并通过这种“直接作用”的方式抑制赤铁矿浮选,这也与聚焦光束反射测量（FBRM）的测定结果基本一致;颗粒?气泡碰撞分析表明,在浮选过程中细粒石英可能通过“边界层效应”的方式跟随气泡升浮（夹带作用）,影响赤铁矿颗粒与气泡间的碰撞及黏附,从而降低了赤铁矿的浮选速率和回收率。      

Flotation Kinetics

Abstract

The physical variables that influence the rate of flotation are examined. The probabilistic model of flotation is used to establish the effect of the particle si2e and density, bubble size and agitation on the rate of flotation

In quiescent flotation, it appears that the flotation rate is limited by the particle-bubble collision and subsequent attachment of the particle to the bubble. For fine (<20 μm) or low density particles the remedy for low recovery rates would be to either use small bubbles of the order of 100 μm, or to use moderate to high agitation with larger bubbles

In the usual turbulent conditions, the limit is set by the destruction of the bubble-particle aggregates. Broadly speaking, the same parameters favour both attachment and detachment so that the ultimate flotation rate is a compromise between these two competing mechanisms

The bounds which define the best agitation level and bubble size to use are strong functions of the particle size and density. This results in conflicting requirements for the optimum flotation of the fine and the coarse particles. Best conditions for the flotation of each are indicated.     

Recovery Enhancement of Ultrafine Wolframite through Hydrophobic Flocs Magnetic Separation

The recovery of ultrafine wolframite (<10 μm) by using flocs magnetic separation was investigated. Magnetic-separation results showed that recovery was closely correlated with the particle size of wolframite, where ultrafine particles were difficult to capture. Hydrophobic particles induced by octyl hydroxamic acid (OHA) could generate flocs, which enlarged the apparent size of particles. Furthermore, the recovery of ultrafine wolframite by flocs magnetic separation was higher than that by conventional magnetic separation. These findings indicated that the recovery was related to the increase in the magnetic force due to the particles’ size induced by hydrophobic flocculation.     

水玻璃在白钨浮选中的适用环境研究及机理分析

水玻璃是一种在浮选中被广泛应用的调整剂,研究其在白钨浮选中的适应条件和机理,对生产实践具有理论指导意义。通过单矿物浮选试验研究表明,水玻璃在白钨与方解石、萤石分离时的最适宜pH值均在9¹0;碳酸钠与水玻璃的配合使用,最有利于白钨的选择性分离浮选。溶液化学分析表明,对方解石产生抑制作用的主要组分是Si(OH)4,而对萤石、白钨起抑制作用的主要组分是SiO(OH)3-,金属离子的添加可以改变水溶液中有效组分的含量,从而起到增强水玻璃选择性抑制的能力。     

Physical Methods and Flotation Practice in the Beneficiation of a Low Grade Tungsten-Bearing Scheelite Ore

In this paper, beneficiation studies were carried out on a low-grade tungsten-bearing scheelite from Nezam Abad ore with total WO₃ grade of 0.11%. Mineralogical studies showed that scheelite is mainly distributed in the ore and gangue minerals include Quartz and Tourmaline. Liberation degree (d₈₀) of tungsten- bearing scheelite is achieved around particles size 150 μm. Gravity concentration, magnetic and flotation methods were conducted by using experimental designs including fractional factorial and response surface methodology. Gravity concentration results indicated that jig separator could not be able to improve tungsten grade in size fraction +600–1750 μm; however, shaking table increased feed grade up to 27.05% with total recovery more than 50% by using four stages concentration in the size range of +125–600 μm. Multi Gravity Separator (MGS) applied on the intermediate products, improved efficiently the total tungsten recovery of the circuit. The results of flotation practice on the pre-concentrated product demonstrated that WO₃ grade could be increased up to 9.2% with total recovery of 27.04% by using one stage rougher and four stages of cleaning. Different methods including MGS, wet and dry magnetic separation were considered for upgrading fines from grinding stages; however, only MGS result was satisfactory. The MGS produced a product with WO₃ grade 0.64% and total recovery 93%.     

提高某铅矿金铅回收率的选矿试验研究

针对某铅矿矿石特点,试验通过提高矿石磨矿细度,使部分细粒金矿物有效单体解离,加强对黄铁矿的活化及对矿泥的分散,采用碳酸钠替代石灰,丁基黄药与sk9011合理配比使用,延长粗选浮选时间,控制氧化铅矿物浮选时硫化钠的用量等措施,使金铅回收率有大幅度提高。     

Effect of particle size distribution on recovery of coarse chalcopyrite and galena in Denver flotation cell

Abstract

The flotation recovery by particle size of single mineral chalcopyrite and galena was studied in a Denver flotation cell, using sodium dicresylthiophosphate (DTP) and sodium isopropyl xanthate (SIPX) as collectors and polypropylene glycol (PPG) as a frother. The study was extended to very coarse particle size (up to 1·6 mm). Froth stability was also measured in parallel to the batch flotation tests, in a specifically designed froth stability column, following the Bikerman approach. It is shown that particles up to 850 μm can be floated successfully, provided they are liberated and hydrophobic. However, the recovery of both chalcopyrite and galena was strongly influenced by the overall particle size distribution, decreasing sharply as the fraction of fines (?106 μm) in the feed also decreased. Rheology measurements showed negligible differences in pulp viscosity, and therefore in the collection zone hydrodynamics, between the different conditions tested. Froth stability, on the contrary, decreased as the feed particle size distribution became coarser. Correlation was found between the amount of fines in the pulp, froth stability and flotation recovery. The recovery of mineral particles is critically dependent on froth stability, which in turn is highly influenced by the overall particle size distribution of the feed material. For these reasons, the study also suggests that it is not possible in batch flotation to determine the rate and recovery of the coarse particle size fractions floating them independently from the fine size fractions.

Dans une cellule de flottation de Denver, on a étudié la récupération par flottation en fonction de la taille de particule d’un minéral unique de chalcopyrite ou de galène, en utilisant du dicrésyle thiophosphate de sodium (DTP) et de l’isopropyle xanthate de sodium (SIPX) comme agents collecteurs et du polypropylène glycol (PPG) comme agent moussant. On a étendu l’étude à la taille de particule très grossière (jusqu’à 1·6 mm). On a également mesuré la stabilité de la mousse en parallèle aux essais de flottation discontinue, dans une colonne de stabilité de la mousse spécialement conçue, d’après l’approche de Bikerman. On montre que l’on peut faire flotter avec succès des particules ayant jusqu’à 850 μm, à la condition qu’elles soient libres et hydrophobes. Cependant, la récupération, tant de la chalcopyrite que de la galène, était fortement influencée par la distribution globale de la taille de particule, diminuant sévèrement à mesure que la fraction de particules fines (?106 μm) dans l’alimentation diminuait. Les mesures de rhéologie montraient des différences négligeables dans la viscosité de la pulpe et ainsi dans l’hydrodynamique de la zone de collection, parmi les différentes conditions évaluées. Au contraire, la stabilité de la mousse diminuait à mesure que la distribution de la taille de particule de l’alimentation devenait plus grossière. On a trouvé une corrélation entre la quantité de particules fines dans la pulpe, la stabilité de la mousse et la récupération par flottation. La récupération des particules minérales dépend, de façon critique, de la stabilité de la mousse qui, à son tour est hautement influencée par la distribution globale de la taille de particule du matériel d’alimentation. Pour ces raisons, l’étude suggère également qu’il n’est pas possible, en flottation discontinue, de déterminer la vitesse et la récupération des fractions de taille de particules grossières, en les faisant flotter indépendamment des fractions de taille fine.     

RECOVERY OF CLEAN COAL FINES THROUGH A COMBINATION OF GRAVITY CONCENTRATOR AND FLOTATION PROCESSES

Flotation feed is a mixture of coarse and ultra-fine fractions. During conditioning of the flotation feed with collector and frother, the finer fraction consumes more reagents as compared to coarser particles. This is mainly due to more specific surface area of the ultra fine than the coarse fraction. This favors the adsorption of reagents toward ultra-finer fractions leads to less complete surface coverage of coarse particles and more entrainment of finer gangue particles. This results in the lower yield of coarse fractions from the flotation circuit and loss in selectivity. Hence, the major challenge is to improve the recovery of the coarser fraction and selectivity of ultra-fine fractions by improving flotation kinetics of all size fractions. This article deals with an approach to overcome the improper reagent adsorption by fine and coarse coal fractions in the flotation circuit through an innovative washing circuit containing gravity operation and flotation processes. Flotation performance between a new washing circuit having stub cyclone and flotation and normal single-stage reagent addition flotation process is compared in terms of selectivity, separation efficiency, rate constant, and size-wise recovery. The washing circuit having stub cyclone and flotation processes improves the fine clean coal yield by 10% and reduces the consumption of reagent compared to the normal single-stage reagent addition flotation process.