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铝土矿利用技术现状综述摘要：在我国当前工业化程度不断加深的情势下，铝土矿日益成为我国大宗紧缺矿产资源。本文主要简要的介绍当前铝土矿的利用技术的现状以及低品位的铝土矿的分选技术。关键词: 铝土矿 脱硫 脱硅 浮选Summarization bauxite utilization technologyAbstract: With degree of current industrialization deepening, bauxite increasingly become scarce among main mineral resources. This paper briefly describes the status of technology that people exploit and make use of bauxite mineral by some measures and separation technology for low-grade bauxite.Keywords: bauxite desulfurization desilication flotation1 引言铝是世界上被广泛利用的金属中仅次于铁的第二大金属，其主要由铝土矿提炼而成,具有多种优良性能，是国民经济发展的基础原料和战略金属。铝土矿广泛分布于世界各地区，主要分布于几内亚、澳大利亚、巴西、牙买加、中国等十多个国家。据有关数据表明,我国的铝土矿资源相对较丰富，虽然已探明储量达到了23 亿t1，但主要是产于古生界和下中生界古风化壳上的一水硬铝石2，品位、品级不高，电解铝能耗高，经济效益低，因此我国的铝土矿资源呈”富矿有限、贫矿比重较大”的特点。随着我国工业化、城镇化的快速发展，国内对铝资源的需求日益增加，且再生铝产业正处于起步阶段3，这样，铝资源开采和供给以及铝工业的发展就遭到了矿产资源条件的制约。此外，我国铝土矿勘查程度较低，勘查深度较浅，大型矿床少，基础储量少，后备资源严重不足，铝土矿成为了我国的大宗紧缺矿产。因此科学合理的开发和有效利用铝土矿“贫矿” 资源，对我国社会经济的可持续发展意义重大。下面浅谈几种铝土矿的分选技术。2. 低铝硅比铝土矿分选由于中国的铝土矿特点，以往提出的拜耳法制氧化铝是不适合的，因此，随着选矿-拜耳法技术思路的提出以及在中州铝业公司的成功应用，利用选矿方法脱硅的研究工作的不断取得拓展和深入，目前已在铝土矿的选择性碎解、一水硬铝石的选择性抑制、铝硅酸盐矿物的强化捕收、及矿泥的选择性分散4等多个方面取得了显著的成果。低铝硅比铝土矿的选矿关键是将一水硬铝石等主要成分有效富集, 使得含铝硅酸盐矿物和其它脉石有效脱除。2.1 采用粗细分选流程分选根据一水硬铝石的嵌布特性,曾克文5等人提出采用粗细分选流程分选的方案，即原矿在粗磨条件下分级, 粗粒部分快速浮选, 得粗粒铝土矿精矿, 粗粒浮选尾矿再磨后与分级的细粒部分合并进行浮选, 得细粒精矿, 两精矿合并为综合精矿。方案认为有部分一水硬铝石为富集合体, 其铝硅比已达精矿质量要求, 这部分可在粗磨条件下先快速浮出, 得到粗粒精矿, 其余部分可细磨使一水硬铝石尽量解离, 再浮选得到细粒精矿。经过试验，获得了良好的分选效果。2.2 利用微泡浮选柱分选浮选柱是细粒矿物浮选分离的有效设备。微泡浮选柱是专门为解决微细粒级矿物分离富集问题而设计的一种浮选柱，其微细物料与气泡的非惯性碰撞的矿化机理和小尺寸气泡比表面积大等特点有助于微细物料的分选。因此，一些研究工作者借助这种设备自身的特点对低品位铝土矿的分选进行了实验研究，通过对矿浆浓度、温度、药剂制度、设备工作参数以及浮选流程等诸多因素进行了考察，在铝硅比为5.31的情况下，利用一粗一精的流程取得精矿铝硅比提高一倍，回收率超过85%的良好指标，同时，利用微泡浮选柱还简化了铝土矿的分选流程6。2.3 旋流器反浮选联合分选随着铝土矿选矿技术的研究的深入，研究工作者们也将一些在选煤过程中用到的分选技术应用到铝土矿分选过程中。曹惠昌4等人利用旋流器可提供离心力场、有助于重选效果的特点，将其与反浮选联合起来对一水硬铝型中低品位铝土矿，以强化分选效果。当然，该过程中除矿浆性质及旋流器自身性质外，还特别考虑到抑制剂和捕收剂的影响。经过统的试验研究，结果表明这种联合流程分选可有效地改善分选效果。3 高硫铝土矿脱硫 我国铝土矿资源中，有超过1亿吨的含硫量大于0. 7%的高硫铝土矿，这部分高硫铝土矿很难得到利用，在铝土矿中约80%-90%的硫是以硫化铁的形式存在的，包括黄铁矿、胶黄矿、磁黄铁矿等。3.1 化学脱硫（1）石灰脱硫。即在浓度较低的条件下，向溶液中加入石灰，其中的硫酸根离子与石灰作用从而进入固相随赤泥排出，这种方法增加了A12O3的消耗量，赤泥排放量也增大，赤泥处理也遇到新的问题。（2）BaO脱硫。将BaO溶入铝酸钠溶液，产生的钡离子与铝酸钠溶液中的硫酸根离子反应，生成难容性BaSO4。这种方法脱硫率高、BaO利用率高、脱硫作用时间短、适应温度范围大，但成本比较高。传统的脱硫方法都是在氧化铝生产过程中的高温和铝酸钠溶液存在的条件下采取的化学方法脱硫, 其工艺流程长，过程复杂, 同时造成了氧化铝生产的热耗加大。随着人们对铝土矿开发利用的重视度的提高，为了解决化学脱硫带来的难题和提高经济效益，物理选矿方法成为了高硫铝土矿经济脱硫的重要途径之一。通过选矿将铝土矿中硫分降低，不仅为氧化铝生产提供低硫铝，还回收了硫精矿，实现了矿产资源的综合利用。3.2 浮选脱硫由于硫化矿的可浮性较好，黄铁矿容易用黄药等捕收剂浮选，而含铝矿物主要是以氧化物和氢氧化物形式存在的，矿物表面具有亲水性，不易被黄药捕收。因此，理论上用黄药可很容易实现黄铁矿和含铝矿物的分离。目前推荐的工业化生产的选矿脱硫流程有二产品全浮选和三产品全浮选。但浮选法需要加药剂，容易对环境造成污染，设备的成本也较高。3.3 细菌脱硫在选煤过程中会出现脱硫的问题，对于无机硫，采用普通的物理化学方法即可达到要求，但是，煤炭中所包含的有机硫的脱除是一个难题。随着研究的深入，生物脱硫已渐渐引起了人们的重视。近几年，很多研究学者都做了相关的研究试验，他们从高硫煤矿中分离的3 种氧化亚铁硫杆菌，用它们对重庆某高硫铝土矿石进行生物氧化浸出脱硫试验，在合适的工艺条件下，脱硫率均超过74%，其中SX1#菌的脱硫率达到83. 57%，并使矿石的硫含量由生物氧化浸出前的3. 83%降低到0. 69%，达到拜耳法生产氧化铝工艺对矿石硫含量的要求7。4. 铝土矿预脱硅目前铝土矿预脱硅的主要方法有化学方法、生物方法和物理方法。4.1化学脱硅在脱硅过程中, 含硅矿物在化学反应的作用下发生分解，报道的有焙烧（氢氧化钠溶出脱硅工艺）和氢氧化钠直接溶出（分选脱硅工艺）8-10。研究结果表明:铝土矿化学选矿的本质是在一定温度下矿石中含硅矿物(主要是高岭石)分解成SiO2和A12O3,然后用苛性钠溶液浸出,使矿物中的SiO2溶出而达到脱硅目的11。4.2 生物脱硅生物选矿脱硅利用微生物分解硅酸盐和铝硅酸盐矿物,可以将铝硅酸盐矿物分子分解成为A12O3和SiO2,并使SiO2转化为可溶物,而A12O3不溶，从而得以分离。生物脱硅是具有良好前景的铝土矿脱硅方法。该法可得到较高的工艺指标,并基本上消除对环境的污染。生物选矿脱硅法是用异养微生物来分解硅酸盐、铝硅酸盐矿物。4.3 物理脱硅物理选矿脱硅工艺的目的是将铝土矿中含硅矿物以天然矿物形式脱除,以降低铝土矿中SiO2的含量。根据分选目的的不同,物理选矿脱硅工艺又可细分为洗矿、筛分、浮选和选择性絮凝等，其中浮选法研究较多。按浮物不同,浮选法又分为正浮选和反浮选。正浮选脱硅（阴离子捕收剂浮选）即通过使用阴离子捕收剂捕收一水硬铝等有用矿物，以抑制剂抑制铝硅酸盐矿物。前苏联70年代的研究结果显示，以正浮选工艺进行三水铝石型铝土矿选矿脱硅，原矿物铝硅比3.06-4.2时，可获得铝硅比917的精矿，A12O3回收率40%-52%。1978年我国海南岛某三水铝石型铝土矿正浮选脱硅结果为：原矿铝硅比5.30时，精矿铝硅比8.32，A12O3回收率72.94%12。反浮选脱硅（阳离子捕收剂浮选脱硅）是通过抑制有用矿物，而以阳离子捕收剂浮选铝硅酸盐矿物。这种方法符合“浮少抑多”的原则，通过浮选捕收含量少的含硅脉石，节省了浮选药剂的用量，且过滤过程也很方便。正浮选过程中，当铝土矿磨矿达到一定细度要求时，铝硅酸盐矿物的粒度已经非常细了，这样不仅影响后续脱水作业，而且其容易吸附在一水硬铝表面，不利于有用矿物的浮选。而反浮选中，铝硅酸盐矿物易磨，一水硬铝可保持较粗的粒度，与正浮选相比，既降低了能耗，也有利于提高浮选效果。 5 结语总而言之，铝土矿利用技术的研究在理论和实践方面均已取得了一些成果，相比较之下，我国的铝土矿的利用技术与国外还存在一定的差距。作为一个铝生产和消费大国，没有先进的铝土矿选矿和生产技术是不能满足当前铝工业发展形式要求的。因此，对于铝土矿的物理化学选矿还有待优化和深入研究，而生物法选矿技术仍有很大的发展前景。参考文献1张凤林等. 高硫铝土矿脱硫研究现状与进展山西科技，2011，26(1):94-952 刘长龄论高岭石粘土和铝土矿研究的新进展沉积学报，2005，23( 3) : 467-4743张伦和我国再生铝产业现状及发展对策轻金属，2009，（6) : 3-64 曹惠昌等. 一水硬铝石型低品位铝土矿旋流 反浮选试验研究J5曾克文等。低铝硅比铝土矿选矿试验研究J有色金属(选矿部分)，2008,6:1-46欧乐明等. 利用微泡浮选柱分选中低品位铝土矿的试验研究J矿冶工程，2011,31（3）：40-437 周吉奎，李花霞. 高硫铝土矿中黄铁矿的细菌氧化试验研究J金属矿山，2011,（12）:67-708罗琳,何伯泉. 高硅铝土矿焙烧预脱硅研究现状评述 J . 金属矿山, 1999, (1) : 31 - 35. 9 罗琳. 一水硬铝石型铝土矿化学脱硅与综合利用研究D . 长沙:中南工业大学, 1999. 10 刘水红,方启学. 铝土矿选矿脱硅技术研究现状述评 J . 矿冶, 2004, (11) : 9 - 11.11 孙德四等.铝土矿预脱硅研究进展J.江西科学，2008,26（2）:256-26212 方启学等. 铝土矿选矿脱硅研究现状与展望J.矿产综合利用，2001（2）：25-31中国XX大学毕业设计（论文）任务书任务下达日期： 20* 年 2 月 20日设计（论文）日期：20*年 2 月20 日至20*年6 月10 日设计（论文）题目：大兴2.40Mt/a（矿区型）选煤厂初步设计及主厂房工艺布置设计（论文）专题题目：铝土矿利用技术现状综述设计（论文）主要内容和要求：（1）完成一座2.40Mt/a矿区型选煤厂初步设计及主厂房工艺布置，资料来自一矿（兴隆庄一矿）45%，二矿（兴隆庄二矿）55%；要求精煤灰分7.518.0%, 水分12.0%。具体工作：（a）对所给煤质资料进行审核、整理与分析；分组、分级判断与综合；可行方案设计与产品预测；方案比较与最佳流程确定；流程设计与计算；设备选型与计算。（b）进行工业广场总平面布置、车间布置以及图纸绘制。要求至少绘制7张图纸（包括车间工艺布置图4张、工艺流程图1张、设备流程图1张及工业广场总平面图1张）。（c） 详细编制设计说明书及投资经济概算说明书。（2）撰写30005000字的专题论文1篇。（3）翻译专业外语文章1篇（3000字以上）。院长（系主任）签字： 指导教师签字：1 英文翻译原文Optimization of a Multi Gravity Separator to produce clean coal from Turkish lignite fine coal tailingsSelcuk Ozgen, Ozkan Malkoc, Ceyda Dogancik, Eyup Sabah, Filiz Oruc SapciA B S T R A C TIn this study, the beneficiation of two lignite tailings by Multi Gravity Separator (MGS) was investigated. The tailings samples from the Tunbilek/Ktahya and Soma/Manisa regions have ash contents of 66.21% and 52.65%, respectively. Significant operational parameters of MGS such as solid ratio, drum speed, tilt angle, shaking amplitude, wash water rate, and feed rate were varied. Empirical equations for recovery and ash content were derived by a least squares method using Minitab 15. The equations, which are second- order response functions, were expressed as functions of the six operating parameters of MGS. The results showed that it is possible to produce a coal concentrate containing 22.83% ash with a recovery of 49.32% from Tunbilek coal tailings, and a coal concentrate containing 22.89% ash with a recovery of 60.01% from Soma coal tailings.1. INTRODUCTIONCoal washing plants in Turkey use gravity separation to beneficiate coal particles larger than 500 lm, and discharge the fine fraction as tailings. In most coal washeries, around 20% of totalrun-of-mine coal is found to be less than 500 lm. The increased use of highly mechanized mining methodologies to enhance the productivity is the major cause for the generation of large uantitiesof coal fines 1. This situation not only causes economical loss, but coal areas also encounter serious environmental issues. In the past, beneficiation of fine coal tailings used the most commercially viable concentration processes such as jigs, densemedium cyclones, spirals, haking tables, and flotation, either individually or in combination 24. However, the processing of finecoal fraction is relatively difficult due to high processing cost, low process recovery, and high moisture content of the product. Recently, a new gravity based processor, Multi Gravity eparator(MGS), have appeared on the market with an operating principle which seems to be very promising for processing of fine particles 5. A detailed description of the MGS is given elsewhere 6,7. MGS may be visualized as a cylindrical version of a conventional shaking table 1,810. Briefly, the principal of MGS concentrating particles is based on the combined effects of centrifugal acceleration and forces acting on a conventional table. This device was developed for selective separation of fine and ultra-fine particles mostly based on the differences in their densities. The use of centrifugal forces in a MGS enhances the relative settling velocity differential between particles different in size and density 1,11. In addition, shearing force created by shaking motion of the drum enhances the particle separation process 12. The early applications of the unit for concentrating heavy minerals like tin, tungsten, tantalum, chromites, and celestite have been reported elsewhere 5,9,1220. Recently, most of the studies have focused on its use in processing of fine coal 1,7,10,2126. In this study, the applicability of MGS to recover two different fine coal tailings has been investigated, and the results from thisstudy have been used to create an empirical equation of recovery and ash content for each coal tailings. Furthermore, it aims optimization of six operational variables for a MGS (pulp solid ratio, drum speed, tilt angle, shaking amplitude, wash water rate, feed rate), which are expected to be important in coal upgrading. Therefore the use of regression analysis with a mathematical software package, already successfully applied in beneficiation test and it is well suited to the main and interaction effects of the variables on cleancoal using a MGS. 2. MATERIALS AND METHODS2.1. MaterialsTwo lignite coal tailings obtained from Tunbilek Coal Preparation Plant of G.L.I of Turkish Coal Enterprises (Ktahya-Turkey) and Soma Dereky Coal Preparation Plant of Aegean Lignite Enterprises (Manisa-Turkey) were used in this study. The samples were taken from slurry waste with standard of TS ISO 5667-10 27.2.2. Methods2.2.1. Characterization testsA number of qualitative and quantitative analysis techniques were used to characterize the coal tailings. The chemical composition of the tailings was defined by X-ray fluorescence (XRF). The mineral composition of the tailings was determined by X-ray diffraction (XRD) method using a Rigaku-Giger Flex analyzer. The particle size distribution of the tailings was obtained using a Retsch AS200 Sieve Shaker and Fritsch-Analysette 22 Particle Size Analyzer. The specific gravity of the tailings was determined by Quantachrome Ultrapycnometer 1000. The ash and sulfur contents of the tailings were determined according to ISO 1171 and ISO 351, respectively. In addition, the calorific value was determined based on ISO 1928. Following the haracterization tests, the beneficiation studies were immediately initiated before any physical and chemical decomposition of tailings. 2.2.2. Multi Gravity Separator (MGS) experiments First, a series of classification tests were performed using a hydrocyclone to separate the clay and/or carbonate minerals from the coal before the MGS studies. In these tests, a small diameterhydrocyclone (44 mm) was selected for the experiments due to large amounts of fine and ultra fine particles in the tailings, which cause a low cut point. Many classification tests were carried out to identify the effect of pulp ratio, feed pressure, apex diameter, and hydrocyclone diameter. The underflow product from the hydrocyclone classification tests was concentrated using a MGS (Fig. 1). The effects of operating parameters such as drum speed, tilt angle, shaking amplitude, wash water rate, pulp feed rate, and pulp solid ratio on the separation efficiency was investigated in detailed. 2000 g of the dry coal tailings was used for each test. When changing a parameter in each test, the other parameters were kept constant, and the optimization results were used for other tests. The shaking frequency of the MGS was fixed at 4.9 cps for all experiments. The parameters used in these tests are presented in Table 1. To obtain a required solids ratio in the feed, measured quantities of solids and water were mixed in a slurry tank. The MGS variableswere adjusted at necessary levels. The feed slurry was pumped into the MGS drum at the required flow rate using a peristaltic pump while the MGS was in operation. The products fromthe clean coal and tailing streams were collected at a steady-state condition. After the separation, the products were filtered, dried, and analyzed for ash content, and to calculate the combustiblerecovery which, was calculated from:Combustible recovery（%） = Mc(1-Ac)/Mf(1-Af) 100 （1）where Ac is ash content of clean coal, Af is ash content of feed, Mc is mass of clean coal, and Mf is mass of feed.3. RESULTS AND DISCUSSION3.1. Characterization of coal tailingsThe results of the chemical analysis for the coal tailings used in this study are presented in Table 2. As seen in Table 2, the lost of ignition (LOI) of Tuncbilek and Soma tailings are about 37% and 49%, respectively. This result indicates that the amount of organic material (coal) in the tailings is significant in comparison to the amount of inorganic impurities in the tailings. Meanwhile, the higher LOI for Soma tailings compared to Tuncbilek tailings can be attributed to organic matter (coal) and CaCO3. These result also showed that the Al2O3 content of Tuncbilek tailings is higher than that of Soma tailings, which suggests the clay content of the Tuncbilek tailings is higher. Fig. 2 shows the particle size distribution of the tailings with a P80 of _28 lm and 45 lm for Tuncbilek and Soma tailings, respectively. The maximum particle size the bothtailings are under 500 lm. However, according to the particle size distribution of the Tuncbilek and Soma tailings have slime size (20 lm) material with a percentage of 76 and 60, respectively. Table 3 shows the ash content of the tailings with respect to particle size. As seen from Table 3, the ash content of the tailings increases with an increase in the size.The XRD analysis results for the tailings indicated that the main minerals for Tuncbilek tailings are kaolinite, illite, mica, smectite, quartz, feldspar, siderite, dolomite, and pyrite (Fig. 3a). The dominant contaminating minerals of Soma are calcite and quartz (Fig. 3b). Besides of these minerals, there is kaolinite, amorphous silica, mica, vermiculite, cristobalite, sepiolite, smectite, zeolite, dolomite, and pyrite. The mineralogical analysis for the tailings presented shows that the main minerals in the tailings are clay, silt, and sand at different percentages. The total sulfur content and calorific value of the Tuncbilek and Soma coal tailings were determined as 1.38%-1835 kcal/kg and 1.03%-2258 kcal/kg, respectively. Figure 1. The concentration schematic of fine coal tailings slurries.3.2. Studies on MGS with coal tailingsA 44 mm diameter hydrocyclone was selected for separating the ultra-fine particles in the tailings. The original tailings containing 66.21% and 52.65% ash was pre-concentrated with the hydrocyclone for Tuncbilek and Soma coal tailings, respectively. The optimum result was achieved at 5% and 10% solids, 2.2 and 3.2 mm of apex diameter for Tuncbilek and Soma coal tailings, respectively. And 14.3 mm of vortex diameter, 1 bar of feed pressure in both. The partition curves for these conditions are presented in Fig. 4. Once the optimum hydrocyclone conditions were identified and applied, the concentrate was further processed in the MGS. Classification experiments with the hydrocyclone obtained a product with an ash value of 45.9% and 38.35% for Tuncbilek and Soma coal tailings, respectively. The variables and results obtained from the MGS beneficiation tests for Tuncbilek and Soma coal tailings are given in Tables 4 and 5, respectively. In these tests, a regression analysis was performed to determine the relationship between six variables and two response functions. The response functions representing the ash content and the recovery of clean coal can be expressed as functions of the pulp solid ratio (s), drum speed (r), tilt angle (), shaking amplitude (sh), wash water rate (w), feed rate (f). The relationship between response (ash content and recovery of clean coal) and variables obtained with Minitab 15 for coded unit are asfollowing: Ash content of the clean coal on MGS for Tunbilek:（Ash）1 = 65.9 - 0.529x1 0.233x2 + 2.27x3 + 0.824x4+ 1.81x5 - 1.53x6 （2）Recovery of the clean coal on MGS for Tunbilek:（Recovery）1= 181- 2.89x1 0.633x2 + 2.30x3 + 1.91x4+ 8.04x5 3.79x6 （3）Ash content of the clean coal on MGS for Soma:（Ash）2= 20.4 - 0.165x1 0.0077x2 + 0.215x3 + 0.0875x4+ 0.592x5 0.287x6 （4）Recovery of the clean coal on MGS for Soma:（Recovery）2= 93.3- 0.53x1 0.576x2 + 0.13x3 + 2.87x4+ 8.48x5 0.75x6 （5）The results obtained from the experiments were summarized below.3.2.1. Effect of feed solid concentrationIt can be observed from Table 4 that the change in the percentage solids in the feed from 10% to 20% by weight (Exp. 7, 8, and 9) increased the ash content from 19.72% to 32.48%, and increased the recovery from 41.34% to 70.28% for Tuncbilek tailings. It can be observed from Table 5 that the change in the percentage solids in the feed from 10% to 20% by weight (Exp. 6, 12, and 13) increased the ash content from 22.89% to 24.65%, and changed the recovery from49.88% to 60.01% for Soma tailings. From these experiments, 15% optimum solid concentration was obtained for both coal tailings. At higher solid concentrations, the separation was not successful due to inadequate MGS surface, and hence schist mixed to the concentration. Therefore, the higher ash content was obtained at 20% solid concentration.3.2.2. Effect of drum speedThe effect of drum speed is shown in Table 4 for Tuncbilek tailings (Exp. 2, 6, and 7) and in Table 5 for Soma tailings (Exp. 1, 2, and 3). As seen from Tables 4 and 5, the ash content and the clean coal recovery of Tuncbilek coal tailings increased with increasing the drum speed. The reason for this was an increase is an increasing centrifugal force. Therefore, the schist minerals such as kaoline, illite, mica, which are heavier than coal, easily moved into the upper stream easily. On the other hand, at high speeds, coal leakages increased in the upper stream, which resulted in a decrease in the recovery of clean coal. In the case of Soma coal tailings, the drum speed showed no significant effect on the ash content, however, the recovery of clean coal decreased due to increasing centrifugal forces at high speeds.3.2.3. Effect of tilt angleThe effect of the tilt angle on the separation was studied at three levels, i.e., 0_, 2_, and 4_ to the horizontal. If a rising column of water is flowing at a velocity between the settling velocities of the two minerals then the heavier and bigger minerals will be able to sink and the lighter and smaller minerals will be lifted by the water and hence a separation of the two minerals can be made. It should be noted from Tables 4 and 5 that the ash content of the concentrate increased with an increase in the tilt angle for Tuncbilek and Soma tailings, respectively. This effect may be explained by the fact that the downward flow velocity of the wash water and particles increased with an increase in the tilt angle. As a result, the residence time of the particles inside the drum decreased, which decreased the separation time between the heavy and light particles in the drum. Both smaller and heavier particles will go to concentrate with coal particles. This will increased ash content of clean coal. In this case, both the coal tailings have been observed in the same way. For coal fines, the heavier particles, i.e., the ash-forming particles had more chance to report into the lighter fraction, which ultimately increased the ash content of the clean coal. Similarly, due to lower residence time, the lighter particles, i.e., the combustible particles will have a higher probability of being carried away with the excess water flowing over the particlebed to the lighter fraction, which ultimately increases the recovery of the clean coal 1.Figure 3. XRD diagram of Tunbilek (a) and Soma tailings (b).3.2.4. Effect of shaking amplitudeThe effect of the shaking amplitude on the separation was also studied at three levels, i.e., 10, 15, and 20 mm. It can be seen from Tables 4 and 5 that an increase in the shaking amplitude increased the ash content and recovery of clean coal (Exp. 7, 10, and 11 in Table 4; Exp. 6, 10, and 11 in Table 5). This observation may be explained by the fact that enhanced shearing force created by the strong shake disrupts the layer of heavy particles stratified onto the drum surface. Therefore, only the coarse and the heavy particles are expected to be dragged by the scrappers. Although this is a desired phenomenon for separation of mineral particles, this phenomenon will result in increased recovery and ash content of the clean coal 1.3.2.5. Effect of wash water rateThe effect of the wash water rate on the separation was studied at three different flow rate, i.e., 1, 3, and 5 lpm. As shown in Tables 4 and 5 that while all other variables were kept constant, an increase in the wash water rate increased the ash content and recovery of the clean coal concentrate (Exp. 1, 2, and 3 in Table 4; Exp. 1, 4, and 5 in Table 5). The increase in the wash water rate increased the wash water flow velocity. Consequently, more feed material was transported towards the overflow, which ultimately increases the recovery of the clean coal. On the other hand, the ash forming particles also move towards the overflow end along with light non-ash forming particles, which, increases the ash content of clean coal.3.2.6. Effect of feed rateThe effect of the feed rate was studied at three different flow rates, i.e., 1, 2, and 3 lpm. Tables 4 and 5 show that an increase in the feed rate increased the ash content and recovery of the clean coal concentrate (Exp. 7, 12 and 13 in Table 4; Exp. 6, 8, and 9 in Table 5). This effect is the same as the wash water rate. While the feed flow velocity increases, more feed material was transported towards the overflow. Therefore, when the feed rate was increased, the ash content and recovery of clean coal increased. Eqs. (2)(5) were derived from Tables 4 and 5 for the response factors. The response factors at any regime in the interval of our experiments can be calculated from these equations. The observed (experimental) results and the predicted values btained from these model equations are presented in Figs. 5 and 6. The predicted values and the observed data points indicate a very good fit (R2 value of 0.807 and 0.843 for ash content of the clean coal, R2 value of 0.944 and 0.831 for recovery of the clean coal on MGS for Tuncbilekand Soma, respectively). In the final MGS beneficiation tests of Tuncbilek coal tailings, the outcome of test number 7 is chosen optimum in the basis of ash content and the recovery ratio of the clean coal and the results are 19.72% and 44.38%, respectively. However, the clean coal which has the lowest ash content of 16.75% is produced by test number 4; and the highest clean coal recovery ratio of 78.57% is produced by test number 6. On the other hand final MGS beneficiation tests with Soma coal tailings showed that the optimum result according to the ash content and recovery rate of clean coal is test number 6 and the results are 22.89% and 60.01%, respectively. With Soma coal tailings the test number 4 resulted lowest ash content of clean coal with 22.89% and at test number 6 the highest recovery with 60.01% were obtained.4. CONCLUSIONSIn this study, the recovery of fine coal from the lignite tailings (Soma and Tuncbilek) was studied with enhanced gravity methods after classification with a hydrocyclone. The utilization tests have been supported by a mathematical model. The following conclusions may be drawn from the results: The characterization studies for the coal tailings showed that the LOI of Soma tailings is higher than Tuncbilek. This was attributed to the amount of organic material (coal) and CaCO3 in Soma tailings. The amount of the clay in Tuncbilek tailings was more than that of Soma tailings. The particle size distribution of Tuncbilek was finer than Soma tailings. While the tilt angle was determined as a major variable on the reduction of ash content of Tuncbilek coal tailings, the dominant parameter for Soma tailings was found to be the feedsolids. The mathematical model of Minitab 15 has been found more effective for Tuncbilek tailings than Soma tailings. The following equations were supposed to utilize clean coal from tailings for Tuncbilek (Eqs. (6) and (7) and Soma (Eqs. (8) and (9) coalplants:Ash of Tunbilek = 65.9 0.529s - 0.233r + 2.27a+ 0.824sh + 1.81w - 1.53f （6）Recovery = 181 - 2.89s - 0.633r + 2.30a + 1.91sh+ 8.04w - 3.79f （7）Ash of Soma = 20.4 - 0.165s + 0.0077r + 0.215a+ 0.0875sh + 0.592w + 0.287f （8）Recovery = 93.3 - 0.53s - 0.576r + 0.13a + 2.87sh+ 8.48w - 0.75f （9）Consequently, a coal product containing 22.83% ash was obtained with a recovery of 49.32% for Tuncbilek coal tailing in pulp solid ratio of 15%, drum speed of 201 rpm, tilt angle of 2 , shaking amplitude of 10 mm, wash water rate of 3 lpm, and feed rate of 2 lpm. In the case of Soma coal tailings, a coal product containing 22.89% ash was obtained with a recovery of 60.01% in pulp solid ratio of 15%, drum speed of 201 rpm, tilt angle of 0_, shaking amplitude of 20 mm, wash water rate of 3 lpm, and feed rate of 1 lpm. The results from this study clearly indicated that it is possible to recover the coal from the tailings using a MGS.REFERENCES1 Majumder AK, Bhoi KS, Barnwal JP. Multi-gravity separator: an alternate gravity concentrator to process coal fines. Minerals & Metallurgical Processing 2007;24:1338.2 Rao LS, Bandopadhyay P. Application of a mozley mineral separator for treatment of coal washery rejects. International Journal of Mineral Processing 1992;36:13750.3 Luttrell GH, Venkatraman P, Yoon RH. Development of a combined flotation/gravity separation circuit for the fine coal cleaning. In: Proceedings of the 12th International Coal Preparation Congress. Krakov, Poland; 1994.4 Luttrell GH, Honaker RQ, Phillips DI. 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In: TS ISO 5667-10. 2002.2 英语翻译译文从土耳其褐煤细粒尾煤中回收洁净煤的多重力分选机的优化Seluk zgen, zkan Malko, Ceyda Dogancik, Eyp Sabah , Filiz Oru Sapi摘要:本次研究中，主要考察了多重力分选机(MGS)对取自Tunbilek/Ktahya和Soma/Manisa地区的两种褐煤尾矿的分选回收。这两种尾矿样灰分分别为66.21%和52.65%。MGS的一些重要的操作参数如矿浆固体含量，转鼓转速，倾斜角度，振幅，给料速度和洗水流量等都是可变的。计算回收率和灰分的经验公式通过Minitab 15软件由最小二乘法得出。这个第二阶指标函数的公式用来描述上述MGS的操作参数的变化规律。结果表明，从Tunbilek和Soma的褐煤尾矿中分选出回收率为49.32%,灰分为22.83%和回收率为60.01%,灰分为22.89%的精煤。关键词：细粒煤尾矿 废物处理 多重力分选机（MGS） 建模1.引言土耳其选煤厂采用重选的方法分离出粒度大于500m的煤并排出细碎粒级作为尾矿。大多数洗煤厂中，毛煤都要产生约20%的500m以下的细粒原煤。为提高生产率而越来越多地使用高度机械化的开采方法是产生大量煤粉的主要原因1。这种情况不仅造成经济损失，还使矿区也遭遇严重的环境问题。在过去，细粒尾煤的分选主要采用最经济可行的洗选方法，如跳汰选，重介质旋流器分选，螺旋分选，摇床选和浮选等，独立或联合处理均可2-4。但是，由于处理成本高，回收率低以及产品水分高，细粒级煤的洗选相对较难。最近，市面上出现了一种新型的基于重力场的分选设备，多重力分选机(MGS)，其工作原理对细粒煤处理似乎很有效果，有关它的一些详细介绍可见文献6,7。MGS可视为常规摇床的圆柱体形式1,810。简单地说，MGS分选矿粒的原理是基于离心加速度和传统摇床床面的作用力的交互作用。这种设备主要是根据矿物密度差异逐渐被用于细粒和超细粒矿物的选择性分离。MGS的离心力的采用强化了不同粒度和密度矿粒的相对沉降速度的差异1,11。此外，由滚筒摇动产生的剪切力也加强了矿粒的分离 12。重矿物如锡，钨，钽，铬，天青石的分选设备的早期应用已有报道5,9,12-20。目前，大多数研究都集中于其在细粒煤加工方面的应用 1,7,10,21-26。本研究已考察了MGS对两种细粒尾煤的回收的适用性,考察结果用以建立一个计算得出各尾煤灰分和回收率的经验公式，此外，该公式是要优化MGS的几个主要操作变量以使得细粒煤得到更好地富集和分离。因此，选矿试验中已经通过一种数学软件应用了回归分析方法，这种方法很适用于考察采用MGS选煤过程中变量的主效应和交互效应。2.实验矿样与研究方法2.1.实验矿样研究实验分别采用Tuncbilek的土耳其煤炭企业Kutahya-Turkey公司G.L.I和Soma Derekoy的Aegean褐煤企业Manisa-Turkey公司旗下选煤厂的两种褐煤尾矿。这两种矿样都是根据TS ISO 5667-10标准27从尾矿浆采得。2.2 研究方法2.2.1 煤质特性试验在矿样煤质特性检验过程中采用了许多定量和定性分析技术。如利用X-射线荧光（XRF）确定尾矿样的化学组成，使用Rigaku-Giger Flex分析仪器过X射线衍射（XRD）测定尾矿样的煤岩组成,利用Retsch AS200振动筛和Fritsch- Analysette 22粒度分析仪测得矿样的粒度分布，由Quantachrome Ultrapycnometer 1000仪器测定褐煤尾矿样的比重，尾矿样的灰分、硫分及发热量分别根据ISO1171，ISO351和ISO 1928确定。随着煤质特性测试实验获得尾矿样物理化学组分资料后，褐煤尾矿回收试验马上随之展开。2.2.2. 多重力分选机（MGS）实验在研究MGS之前,首先进行了一系列的分级试验,即利用水力旋流器将粘土或碳酸盐矿物从煤中分离出来。这些试验中，由于褐煤尾矿中存在大量的细粒和超细粒矿物，导致粒群分选密度较低，因此试验采用小直径水力旋流器（44mm）。许多分级试验验证了矿浆浓度，给料压力，底流口直径以及旋流器直径等因素的影响效果。水力旋流器分级试验的底流产品经MGS精选（图1）。同时对转速，倾斜角，振幅，洗水流速，矿浆进料速率和矿浆的固体含量等操作参数对分离效率的影响进行了详细的研究，每次试验用2000g干矿样，在其他操作参数不变，每次改变其中一个参数时，所得的优化结果作为其他试验的实验条件。所有试验中MGS的振动频率均为4.9r/s，试验的操作参数值见表1.为了获得合适的入料矿浆固体含量，待测的干物料和水应在搅拌桶中混合调浆，同时MGS的工作参数也调到适当的状态。当MGS运转时，入料矿浆经蠕动泵以适宜的流速泵入MGS转鼓中，在相对稳定状态条件下收集所得的精矿和尾矿矿浆。两产品分离之后需过滤，干燥和测定灰分，同时计算其可燃体回收率，计算公式：可燃体回收率,(%)= Mc(1-Ac)/Mf(1-Af) 100 （1）式中：Ac精煤灰分Af入料灰分Mc精煤质量Mf入料质量3结果与讨论3.1.尾煤特性本次研究采用的尾煤样的化学分析结果见表2，表中显示Tunbilek和Soma的尾矿样的（LOI）值分别约为37%和49%，这一结果表明尾矿样中有机质（煤）含量与无机矿物相比还是相当可观的，同时可知Soma尾矿比Tuncbilek尾矿具有较高的LOI值是归因于有机物质（煤）和CaCO3含量的差异。这些分析结果还表明Tuncbilek尾矿样的Al2O3含量比Soma尾矿样高，这说明Tuncbilek尾矿样的粘土含量较高。图2表示尾矿样的粒度分布，当粒级含量为80%时，Tuncbilek尾矿样和Soma尾矿样粒级范围分别为-28m和-45m，两种矿样的最大粒度均为500m，由粒度分布曲线可知，Tuncbilek和Soma尾矿样粒度小于20m的煤泥含量分别为76%和60%。表3表示物料灰分与粒度之间的关系，由表可知，物料灰分随粒度的增大而增大。尾矿样的XRD分析结果表明，Tunbilek尾矿主要矿物是高岭石，伊利石，云母，蒙脱石，石英，长石，菱铁矿，白云石，黄铁矿（图3a）；Soma尾矿中主要的杂质为方解石和石英（图3b），除了这些矿物，还有高岭石，无定形二氧化硅，云母，蛭石，方英石，海泡石，蒙脱石，沸石，白云石，和黄铁矿。先前的尾矿矿物学分析表明，尾矿中主要矿物是不同比例的粘土，淤泥和沙子。Tunbilek和Soma尾煤的总硫含量和发热量分别为1.38、-1835大卡/千克和1.03、-2258大卡/公斤。Figure 3. XRD diagram of Tunbilek (a) and Soma tailings (b).3.2. MGS的尾煤处理研究尾煤中超细粒颗粒的分选采用的是直径为44mm的水力旋流器， Tunbilek和Soma原尾煤的灰分分别为66.21和52.65，两种尾煤均经小直径水力旋流器预先浓缩处理。研究的优化结果是Tunbilek和Soma尾煤入料固体含量分别为5%和10%,两煤样试验水力旋流器底流直径分别为2.2mm，3.2mm，同时，两者均具有14.3mm的旋流直径和1个工程大气压的入料压力，这些条件下的分配曲线如图4。水力旋流器最佳的条件被确定和应用后，旋流器浓缩底流就去MGS进一步精选。Tunbilek和Soma尾煤的分级试验经水力旋流器得到的产品分别有45.9%和38.35%的灰分。从Tunbilek和Soma尾煤的MGS分选试验中获得的参数和优化结果分别在表4 和表5中给出，分选试验过程采用回归分析来确定MGS的6个操作参数之间的关系及两个响应函数，这些表征精矿灰分和回收率的指标函数看作是矿浆固体含量（s），转速（r），倾斜角（），振幅（sh），细水流量（w）和进料速率（f）的函数。指标（洗选精煤的灰分和回收率）和由Minitab15编码单元软件获得的参数之间的关系如下：Tunbilek 尾煤的MGS分选精矿灰分（Ash）1 = 65.9 - 0.529x1 0.233x2 + 2.27x3 + 0.824x4+ 1.81x5 - 1.53x6 （2）Tunbilek 尾煤的MGS分选精矿回收率（Recovery）1= 181- 2.89x1 0.633x2 + 2.30x3 + 1.91x4+ 8.04x5 3.79x6 （3）Soma尾煤的MGS分选精矿灰分（Ash）2= 20.4 - 0.165x1 0.0077x2 + 0.215x3 + 0.0875x4+ 0.592x5 0.287x6 （4）Soma尾煤的MGS分选精矿回收率（Recovery）2= 93.3- 0.53x1 0.576x2 + 0.13x3 + 2.87x4+ 8.48x5 0.75x6 （5）从研究实验中得到的结果总结如下。3.2.1入料固体浓度的影响从表4中可以发现，Tunbilek 尾煤试验过程中入料中固体百分比从10至20（实验7，8和9）的变化使得精矿灰分由19.72%增加至32.48%，回收率由41.34%增大至70.28%。表5表明Soma尾煤试验过程中入料中固体百分比从10至20（实验6，12和13）的变化使得精矿灰分由22.89%增加到24.65%，回收率由49.88%增大至60.01%。经过试验,两种尾煤分选的最佳固体浓度均为15%，当固体浓度大于15%时，由于MGS分选面积不足和片岩易混入精矿中，则分离效果有所下降，因此，当入料固体浓度为20%时，灰分更高。3.2.2转鼓转速的影响转速对Tunbilek 和Soma尾煤分选的影响可见于表4（实验2，6和7）和表5（实验1，2和3）。从表4和表5可知，随转鼓转速的增大，Tunbilek 尾煤的精矿灰分和回收率不断增大，其原因在于转鼓转速增大导致离心力的增大，因此，比煤重片岩矿物如高岭土，伊利石，云母很容易混入到溢流（精矿）中，另一方面，高速运转状态下溢流中混入底流的煤量也增加，引起回收率降低。至于Soma尾煤，精矿灰分不受转鼓转速影响，但回收率随转速加大而降低。3.2.3倾斜角的影响试验针对倾斜角对分选的影响对其采用3种水平，即与水平面夹角0，2和4，如果一股上升水流以介于两种矿物沉降末速的流速流动时，则较大或较重的矿物将能下沉，而较小或较轻的矿物将随水流上升，这样就实现了两种矿物的分离。应当指出表4和表5中Tunbilek和Soma尾矿精矿的灰分均随倾斜角增大而增加，这是由于矿粒和洗水的下降流速随倾斜角的增大而增大，导致矿粒在分选室中的停留时间缩短，从而也缩短了重矿粒与轻矿粒的分离时间。较细且较重的矿粒会进入回收精煤中，从而引起精矿灰分偏高，这种情况在Tunbilek和Soma尾矿分选过程中均存在。对煤粉来说，重矿粒即带灰矿粒更容易混入到轻产物中，最终使得分选精矿的灰分增大，同样，由于矿粒停留时间短，可燃体含量高的轻矿粒很可能被料层表面过量的水流带到轻产物溢流中，最终使分选精矿的回收率增大1。3.2.4.振幅的影响本研究针对振动幅度对分选效果的影响对其采用3种水平，10mm,15mm,20mm。 从表4和表5中可以看出，随振动幅度的增加，分选精矿的灰分与回收率均是增加的（表4中实验 7，10，和11; 表5中实验6，10，和11）。导致这一观察结果的原因可能是由于强烈震动而产生强大的剪切力使得在转鼓表面分层的重颗粒物料层分离。因此，只有粒度大的重粒子才能被刮刀刮走。虽然这对矿粒分选来说是理想的现象，但这种现象会导致分选精矿的灰分和回收率的增加1。3.2.5.洗水流量的影响本研究对分选的洗水流量的1，3，5L/min三个水平进行了试验。如表4和表5所示，当其他所有参数保持恒定时，洗水流量的增加将引起精矿灰分和回收率的增加（表4中实验1，2，和3；表5中实验1， 4，和5），细水流量增大使水流速度增大，因此，更多的进料被水流带向溢出，最终使精矿回收率增大。另一方面，带灰颗粒也随带煤颗粒朝着溢流端运动，从而增加分选精矿的灰分。3.2.6.入料速率的影响试验取1，2，和3lpm三个流速水平研究了入料速率对分选的影响。表4和表5中，当入料速率增大时，富集精矿的回收率和灰分均增加（表4中实验7，12和13; 表5中实验6，8和9）。同洗水流量的影响相同当入料矿浆流速增大时，很多物料会被运输到溢流中，因此，精矿回收率和灰分量增加。为计算试验指标，式（2）-（5）是由表4和表5推导得来。实验过程中任一阶段的试验指标都可由这些方程式计算得来，所观察到的实验结果和从这些模型方程得到的预测值的关系如图5和图6。预测值与试验数据点表明一条很好的关系曲线（R2值0.807和0. 843分别为Tuncbilek和Soma尾煤的MGS分选精矿灰分修正系数，R 2值0.944和0.831分别为Tuncbilek和Soma尾煤的MGS分选精矿回收率修正系数）。在Tunbilek尾煤最后的MGS分选试验中，试验7的结果被作为最佳优化结果，精矿灰分和回收率分别为19.72和44.38，但是，试验4得到16.75的最低灰分;试验6得到了78.57的最高回收率。另一方面，Soma尾煤最后的MGS分选试验中，试验6的精矿灰分和回收率是最佳的优化结果，精矿灰分和回收率分别为22.89和60.01，试验4得到22.89的最低灰分;试验6得到了60.01的最高回收率。4.结语本研究利用强化重力方法对利用水力旋流器分级后的（Soma和Tunbilek）褐煤细粒尾煤的回收进行研究，且依据试验结果建立了数学模型。由试验结果可得以下结论：（1） 尾煤特性研究表明，Soma尾煤的LOI高于Tunbilek尾煤。这是由于Soma尾煤含有大量的有机物质（煤）和CaCO3，而Tunbilek尾煤中含粘土的量较Soma尾煤多。至于粒度分布，Tunbilek尾煤比Soma尾煤更细。（2） 降低Tunbilek尾煤分选精矿灰分的主要因素是倾斜角，而就Soma尾煤而言，是入料固体含量。（3） 对于nitab 15的数学模型Tunbilek尾煤比Soma尾煤更适合。下面的方程式中，分别用于Tunbilek选煤厂（式（6）和（7）和Soma选煤厂（式（8）和（9）的尾煤回收精矿的指标计算：Ash of Tuncbilek= 65.9 0.529s - 0.233r + 2.27a+ 0.824sh + 1.81w - 1.53f （6）Recovery = 181 - 2.89s - 0.633r + 2.30a + 1.91sh+ 8.04w - 3.79f （7）Ash of Soma = 20.4 - 0.165s + 0.0077r + 0.215a+ 0.0875sh + 0.592w + 0.287f （8）Recovery = 93.3 - 0.53s - 0.576r + 0.13a + 2.87sh+ 8.48w - 0.75f （9）因此，以Tunbilek尾煤作为原料煤，在入料矿浆固体含量15，转鼓转速201r/m，倾斜角2 ，振幅10mm，洗水流量3lpm和入料速率2lpm的情况下，可得回收率为49.32、灰分为22.83%的精矿产品，而对于Soma尾煤来说，在入料矿浆固体含量15，鼓转速201r/m，倾斜角0 ，振幅20mm，洗水流量3lpm和入料速率1lpm的情况下，可得回收率为60.01、灰分为22.89%的精矿产品。研究结果清楚地说明了利用MGS从尾煤中回收精煤是可行的。参考文献：1 Majumder AK, Bhoi KS, Barnwal JP. 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