Rheological investigations of sulphide mineral slurries

The flotation separation of ultrafine (−5 μm) sulphide mineral particles is inherently problematic, since even strongly hydrophobic fines have relatively low flotation rate constants. Selective aggregation of fines provides a route for enhanced flotation selectivity, but requires detailed knowledge and ultimately control of the electrostatic and hydrophobic inter-particle forces acting in concentrated sulphide mineral slurries. Rheological investigations offer insight into these inter-particle interactions. Rheological studies of ultrafine sphalerite particle slurries at pulp densities akin to those experienced in flotation practice were undertaken; the influence of pH, electrolyte concentration, activating copper (II) ions and ethyl xanthate ions on the level of pulp aggregation, as defined by the yield values and viscosities, are reported.The rheological behaviour of oxidised, and effectively hydrophillic, sphalerite particles (in the pH range 4 to 10) are controlled by electrostatic repulsive forces and yield values scale inversely with electrophoretic mobilities. Copper (II) activation, ethyl xanthate treatment and addition of slime particles dramatically affect pulp rheology and, depending on the particular concentrations and conditions employed, cause either aggregation or dispersion of the pulp. We show that an understanding of the theological behaviour of sulphide mineral pulps enhance our knowledge of particle-particle (and bubble-particle interactions) and may enable fine particle flotation performance to be optimised.     

The dependency of the critical contact angle for flotation on particle size - Modelling the limits of fine particle flotation

The flotation behaviour of fine particles is studied in this work. Fine methylated quartz particles within the size range from 0.2 to 50 μm, and with varying contact angles, were floated in a mechanical flotation cell. Results indicate that particles of a given size need to possess a minimum critical contact angle, which increases in value as particle size decreases, for flotation to be initiated. As a consequence, a non-floating component exists within a given size fraction. This is interpreted as a fraction consisting of particles below the critical contact angle for flotation for that size. The critical contact angle for flotation is explained in terms of the existence of an energy barrier for bubble-particle attachment. The flotation results are interpreted by means of [Scheludko et al., 1976] and [Drelich and Miller, 1992] models for the floatability of fine particles. The experimental data compared very well with calculations using the Drelich and Miller equation, allowing extension to the prediction of the critical contact angle for flotation down to particle sizes well below the previous limits investigated, bridging the gap existing in the literature.     

粗颗粒矿物在矿浆相中的悬浮特征

浮选过程中颗粒的悬浮是气泡与颗粒碰撞和黏附的基础条件。粗颗粒矿物由于尺寸较大,在矿浆相中的悬浮特征有别于常规粒级和细粒级矿物。为了分析粗颗粒在浮选机矿浆相中的悬浮影响因素,以层析度函数为分析手段,选取金矿、钛铁矿、黄铜矿三种浮选系统中的四种规格浮选机中的粗颗粒为研究对象,研究了三种矿物在单槽浮选机内、沿作业分布、随浮选机规格变化等情况下的悬浮特征。结果表明:在单体浮选机竖直截面内,粗颗粒矿物在叶轮上方悬浮效果比其它粒级差,随着高度的增加重量层析度迅速减小;沿作业分布的三台同规格浮选机内的粗颗粒矿物悬浮特征相近;而浮选机的规格大小对粗颗粒的悬浮特征影响不敏感,给矿粒级分布和矿物的比重对粗颗粒的悬浮特征影响强烈。在同一浮选机槽体,固相颗粒重量分布层析度与金属量分布层析度两者变化并不一致,粗颗粒分选环境尤其明显,这为粗颗粒浮选机的设计提供了新的思路。     

Effect of impeller rotational speed on the size dependent flotation rate of galena in full scale plant cells

The first three rougher cells in the lead circuit of the Elura concentrator (formerly Pasminco Australia Limited) were selected as the plant cells for investigation. Metallurgical surveys were performed and various hydrodynamic measurements taken, allowing the galena flotation rate constant and the bubble surface area flux (S_b) in these cells to be calculated over a wide range of gas flow rates, and at two impeller rotational speeds. It was determined that altering the impeller rotational speed did not significantly change the rate constant dependency on S_b when flotation was considered on an unsized basis.The analysis was further extended to examine the same cells parameters on a size-by-size basis. The results obtained have been used to identify differences in the flotation behaviour of the various particle size fractions, independently of surface hydrophobicity. It is shown that the physical conditions for effective flotation of fine (<9 μm) and coarse (>53 μm) particle size fractions differ substantially, suggesting that a specific hydrodynamic environment will favour a high flotation rate for fine galena, which may be detrimental to the recovery of coarse galena, and vice versa. These observations are in accord with metallurgical practice that suggest that it is difficult to improve fine particle flotation without also compromising coarse particle stability efficiency simply by modifying the cell hydrodynamics alone. A fundamental flotation model was applied to quantify differences in the flotation rate of the various particle size fractions with impeller rotational speed.     

加压溶气气浮法浮选微细颗粒的理论分析

分析了加压溶气浮选设备选别微细颗粒的理论依据,并提出设计中应注意的问题.从理论上证明了溶气释放析出将会优先选择在疏水矿粒表面上发生,解决了传统浮选设备进行细粒浮选时碰撞几率过小的问题.     

An evaluation of the role of particle size in the flotation of coal using different cell technologies

A large variety of new flotation cells has been introduced in the last few years, probably as a result of the successful introduction of column flotation in the minerals processing industry. In common with the column cell, a number of these new cells employ an essentially quiescent separation zone. However, a number of novel cell designs have been introduced that use agitation mechanisms similar to those employed on conventional flotation cells. The aim of this investigation was to evaluate flotation behaviour in both an agitated and non-agitated environment, particularly with respect to particle size. A hybrid ‘agitated’ column cell was designed for the investigation, and the operation of this unit was compared to that of a column cell, and to a batch flotation cell, on a laboratory scale. The testwork was conducted on coal fines, as problems with the flotation of coarse coal particles in a column cell had previously been identified. It was demonstrated that the addition of an agitated stage to a column cell can significantly improve the coarse particle recovery in comparison to the conventional column cell, while maintaining good selectivity in the fine sizes.     

炼焦中煤的尺度效应及对浮选的影响

为了考察煤泥粒度变细对浮选效果的影响,从不同细度颗粒的几何特征、润湿热及浮选行为3方面探讨了炼焦中煤的尺度效应。使用BET比表面测定仪、微量热仪分别研究了炼焦中煤粒度减小对其比表面积、总孔容和平均孔径及润湿热的影响,用Young-Laplace方程分析了基于毛细管作用捕收剂吸附的变化,通过半经验公式计算了颗粒粒径及润湿性对上浮概率的影响,并进行了不同细度超低灰精煤的浮选试验。研究表明,随着炼焦中煤粒度的变细,其比表面积和总孔容呈数倍增大,Ⅱ类孔增多,平均孔径增加;基于毛细管作用的捕收剂吸附增强,对去离子水的润湿热值急剧降低;存在最佳粒度范围使颗粒的上浮概率最大,中等紊流强度的浮选环境中,控制疏水颗粒粒径在-200+45 μm或亲水性颗粒粒径在-10 μm或+45 μm均可强化浮选回收。相同浮选药剂条件下,煤泥粒度越细,煤的回收率越难以保障,并且水回收率呈增加趋势,浮选效率降低。     

Application of flotometry for characterizing flotation in the presence of particles aggregation

Different size fractions of coarse magnetite were floated in water in a monobubble Hallimond tube with an increasing dosage of sodium oleate (Na01), diethyl dixanthate ((EtX)₂), and hexadecane. All maximum recovery (R_m) vs collector concentration curves plotted for various size fractions were similar in shape having zero recovery at a low and high collector concentration with a maximum recovery in between at a characteristic collector dosage c_m, which provides maximum hydrophobicity of the system. Properties of the studied flotation system and literature data indicated that the cessation of magnetite flotation with a high dosage of Na01 is due to hydrophilic layers of oleate ions sorption, while in the other two systems it is due to oil agglomeration of the solids. The values of the maximum particle size which can still be floated (a^m₅₀), taken from the maximum recovery vs particle mean size at the concentration equal to c_m and R_m=50%, seems to be the parameter which well characterizes any solid-collector-water flotation system because this value derives mainly from the hydrophobic properties of the flotation system and particle density in water, ^′. The a^m₅₀ value and the so-called “flotometry equation” (for our Hallimond tube in the form L_m = a^m₅₀^′) were used for calculation of the parameter L_m which in turn was used for a comparison of the floatability of magnetite with different collectors and subsequently with other flotation systems. The values of parameter L_m indicated that magnetite floatability increases in the order: collectorless (mechanical carryover), flotation with (EtX)₂, flotation in the presence of hexadecane, and oleate flotation. It was established that the floatability of magnetite with Na01 is equal to the floatability of galena with (EtX)₂ and very similar to the collectorless floatability of Teflon. Floatability of the above three systems is most likely the maximum possible floatability of a solid in an aqueous environment.     

The effect of grinding on mill performance at Division Salvador,Codelco-Chile

This paper presents a sensitivity analysis of the impact the ore grinding level (% +212 microns) has upon the rougher flotation performance, where the main copper losses are related to fine particles (less than 12 microns) with high content of soluble copper and coarse particles (larger than 212 microns) that are less liberated.Firstly, a one year grinding data set, considering average daily shifts, was analysed and it was shown that the classical Bond correlation properly describes the average trend of the grinding circuit operation in terms of ore tonnage, operational work index, product particle size d₈₀ and power availability. On the other hand, the rougher flotation kinetic was characterized from plant testing. Thus, cropper and moly recoveries were correlated by Klimpel's model to describe the rougher flotation performance, at different particle size classes, in terms of design and operating variables. The copper recovery was found to be critically dependent on the ore grinding level, % +212 microns, and the soluble copper content.Using the grinding and flotation correlations, a plant simulator, that integrates the grinding and flotation capacities, was built. The simulator was validated with plant data for a range of ore tonnage, for different grinding levels and at two power levels. The simulator is useful in selecting the grinding reference, to identify critical requirements of instrumentation, bottlenecks limiting plant capacity, and to complement the supervisory control strategy. Thus, a powerful tool to estimate the best compromise between ore tonnage and grinding level, in order to maximize the values recovery, was established.     

Flocculation and flotation response of Rhodococcus erythropolis to pure minerals in hematite ores

Rhodococcus erythropolis as a flotation collector for hematite was evaluated in the study. The surface morphology and cell wall composition of R. erythropolis was analyzed. Zeta potentials for four pure minerals from hematite ores were measured before and after adsorption by R. erythropolis. Pure mineral flotation tests and mixed mineral separation tests were performed to describe adsorption characteristics and mechanisms. A rod-shaped bacterium was detected with CH₂, CH₃ and COO groups on its cell wall that imparted hydrophobicity and negative charges. The ability of R. erythropolis to collect hematite was stronger than its ability to collect quartz, kaolinite and apatite. For a pulp pH of 6 and a cell concentration of 75 mg/L, recovery rate of hematite was 89.67%. The recovery differences between hematite and quartz, kaolinite and apatite were 66.43%, 60.36% and 54.30%, respectively. These data indicated that electrical properties of hematite surface were more suitable for adsorption of R. erythropolis than other three minerals. The adsorbed hematite particles appeared as large agglomerates after interaction with R. erythropolis. The quartz, kaolinite and apatite particles, however, were in the form of dispersed particles or small agglomerates. The chemical adsorption of hematite on bacterial cell wall resulted in agglomeration. The effects of flocculation and flotation of R. erythropolis on micro-fine hematite particles were characterized for the first time. The results showed that R. erythropolis can act as a flotation collector for hematite from hematite ores.     

Factors affecting the floto-elutriation process efficiency of a copper sulfide mineral

Coarse mineral particles exhibit poor conventional flotation efficiency because of many factors, including the low carrying capacity of bubbles, bubble/particle adhesion problems due to cell turbulence, and low degrees of liberation (low hydrophobicity). Many attempts to improve the recovery of coarse fractions have been explored, such as floto-elutriation operating at a high solid content while dispersed in a fluidized (or expanded) bed formed with a continuous injection of compressed air and an uprising water flow. This work analyzed the comparative performances of floto-elutriation (FE) and conventional flotation (CF) on a classified copper sulfide mineral feed as an example of a difficult-to-liberate low-grade ore. Contrary to expectations, CF and FE (Hydrofloat) displayed similar particle recovery rates with feed size distributions for P80s of 130, 240 and 280 μm. However, metallurgical recoveries from classified fractions of −297+210 μm were 25% higher in FE than in CF and as expected, coarse (+297 μm) particles were not recovered in the CF, but in the FE. The recovery of fine fractions in the FE process was due to high hydraulic entrainment and surprisingly the recovery of intermediate and liberated fractions (+74−149 μm) was very low, due to its low air hold-up. However, the enhancement of the holdup in FE increased the recovery of these mid-sized fractions. Because of the hydraulic carryover caused by the bubbles and water elutriation, the metallurgical grades obtained in all cases were very low compared to conventional bench flotation. It is believed that this FE equipment works better with coarse, narrowly classified particles and high-grade feeds and that performance decreases for low-grade ores requiring high liberation. Certain features of these findings are visualized.     

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”.     

Critical contact angle for coarse sphalerite flotation in a fluidised-bed separator vs. a mechanically agitated cell

This work investigates the critical contact angle for the flotation of coarse (850–1180 μm, 425–850 μm and 250–425 μm) sphalerite particles in an aerated fluidised-bed separator (HydroFloat) in comparison to a mechanically agitated flotation cell (Denver flotation cell). In this study, the surface chemistry (contact angles) of the sphalerite particles was controlled by varying collector (sodium isopropyl xanthate) addition rate and/or purging the slurry with either nitrogen (N₂) or oxygen (O₂) before flotation. The flotation performance varied in response to the change in contact angle in both the aerated fluidised-bed separator and the mechanically agitated cell. A critical contact angle threshold, below which flotation was not possible, was determined for each particle size fraction and flotation machine. The results indicate that the critical contact angle required to float coarse sphalerite particles in a mechanically agitated cell was higher than that in the fluidised-bed separator, and increased as the particle size increased. At the same particle size and similar contact angles, the recoveries obtained by the aerated fluidised-bed separator in most cases were significantly higher than those obtained with the mechanically agitated flotation cell.     

Effect of mechanical and chemical clay removals by hydrocyclone and dispersants on coal flotation

Fine minerals, mostly clays, are known to have a detrimental effect on coal flotation. This paper focuses on the effect of mechanical and chemical removals of fine minerals by hydrocyclone and dispersants on coal flotation. The experimental results showed that the flotation recovery slightly increased from medium acidic to medium alkaline ranges. The flotation experiments carried out with dispersants at different dosages showed that the dispersants did not enhance the flotation recovery significantly. However, the removal of the fine fraction from the feed using a hydrocyclone significantly increased the flotation recovery. The bubble–particle attachment tests also indicated that the attachment time between an air bubble and the coal particles increased in the presence of clay particles. These attachment time results clearly showed that the clay particles adversely affected the flotation of coal particles by covering the coal surfaces which reduced the efficiency of bubble–coal attachment. An analysis based on the colloid stability theory showed that the clay coating was governed by the van der Waals attraction and that the double-layer interaction played a secondary role. It was also concluded that the best way to increase the flotation recovery in the presence of clays was to remove these fine minerals by mechanical means such as hydrocylones.     

提高微细粒嵌布金矿石浮选回收率的试验研究

针对某金矿载金矿物分散,金嵌布粒度微细,采用常规浮选方法金回收率不高的问题,采用细磨、添加分散剂、加入Mac-12新型捕收剂等措施对该金矿石进行强化浮选试验。试验研究结果表明,该工艺对此金矿石具有很好的适应性,在pH值为9左右的弱碱性条件下,可提高金回收率8.27个百分点,提高金品位21.28 g/t。     

Flotation in a novel oscillatory baffled column

This paper presents an evaluation of an oscillatory baffled column (OBC) as a novel flotation device. The cell is based on a standard column design but employs a novel agitation mechanism where a series of baffle plates are oscillated sinusoidally through the fluid. This type of agitation has been shown to produce a more evenly distributed shear rate in the cell and allows the effect of agitation on particle–bubble contacting to be decoupled from gas dispersion effects. The column was first characterised in terms of mixing and gas dispersion, before being flotation tested using a quartz-amine system. Results indicated that the OBC was able to improve the flotation rate constant by up to 60% for fine particles (<30 μm) and by between 30% and 40% for coarser particles, relative to a standard flotation column. Interestingly, optimal flotation performance was obtained at power intensities orders of magnitude lower than those found in similar studies in stirred systems. This is believed to arise from the even distribution of shear in the OBC together with the oscillatory motion of the fluid in the cell. This oscillatory motion does not contribute to the average power intensity in the fluid and therefore results in more fluid motion per unit energy than would be obtained in a conventional stirred system. The OBC was therefore able to significantly improve flotation rates at power intensities orders of magnitude lower than those found in conventional cells.     

难浮煤泥浮选速率试验研究

针对主导粒级为细粒级的难浮煤泥,进行了窄粒度级和全粒度级的浮选速度试验。结果表明：浮选速度由快到慢的粒级依次为0.125～0.074、0.250～0.125、〈0.074、0.500～0.250、〉0.500 mm。粗粒级煤粒因其质量大、脱附概率高、部分连生体表面疏水性较差,需在高能量输入条件下,通过较长时间的矿化作用...     

Interaction mechanism of miscible DDA–Kerosene and fine quartz and its effect on the reverse flotation of magnetic separation concentrate

In this study, a modification of oil assisted flotation processes of quartz particles has been proposed, which is based on introduction of miscible Dodecylamine (DDA)-Kerosene as collector with DDA cationic surfactant coated on kerosene to the hydrophilic quartz particles in the pulp. The property of miscible DDA–Kerosene emulsion was investigated. Due to the adsorption of DDA at kerosene/water interface, a smaller and uniform kerosene emulsion formed. Addition of cationic surfactant to the kerosene emulsion changed the zeta potential value from negative to positive, which resulted in enhancing the adhesion of the oil droplets to negatively charged quartz. The results showed that agglomeration and flotation process can be realized simultaneously with DDA–Kerosene. The agglomeration of fine quartz minerals in the presence of miscible DDA–Kerosene led to the formation of very large compact agglomerates resulting in increasing hydrophobicity of the particles and inducing a higher probability of collision and adhesion to air bubble. Experimental data indicated that miscible DDA–Kerosene had better selectivity and stronger collectability to quartz than DDA–HCl, which can be used as an efficient collector in the reverse flotation of magnetic separation concentrate of TISCO. At the same DDA dosage (60 g/t), separation efficiency got to 18.53% when using DDA–HCl as collector; while a better result was obtained with DDA–Kerosene, the efficiency of separation reached 59.07% which was identical with 120 g/t DDA–HCl.     

Effect of entrainment in bubble load measurement on froth recovery estimation at industrial scale

Froth recovery was calculated in a 130 m³ mechanical cell of a rougher flotation circuit. This was done by bubble load determinations along with mass balance surveys. Valuable grade in the bubble load decreased in the −38 μm due to fine particles entrained to the chamber of the device. The effect of fine particle entrainment on froth recovery was evaluated. A comparison between results from the raw bubble load data (assuming all particles were transported by true flotation) with those from corrected bubble load information (subtracting fine particle entrainment) was carried out. Entrainment occurred due to hydraulic transport in the bubble rear, which corresponds to the worst case scenario for froth recovery estimation. Results showed that the relative error was less than 0.3%, which allowed validation of the bubble load measurement as an effective methodology for froth recovery estimation at industrial scale.     

Characterisation of coarse composite sphalerite particles with respect to flotation

The flotation response of a typical zinc-lead (Zn/Pb) ore, with respect to coarse composite (sulphide/non-sulphide) particles is reported. The flotation tests were carried out on a selected feed particle size range (−600 + 75 μm, at P₈₀ of 390 μm) and the recovery of Zn composite particles analysed on a size by size basis. The best results were achieved with the use of 75 g/t sodium isopropyl xanthate (SIPX), obtaining a Zn recovery of 77%, with a significant improvement at the coarse end of the particle size distribution. Computerised scanning electron microscope (QEMSCAN) was used to characterise value mineral grain size and degree of liberation, as well as gangue and sphalerite association in particles reporting to both concentrate and tailings. A new characterisation function (Locking ratio, LR) was developed based on the data from the automated mineralogical analysis to characterise particles into two-phase composites with different degree of locking texture (simple and complex). The function, which is based on the mode of occurrence of sphalerite, grain size, proportion and composition of the constituent minerals in each particle, was used to study the flotation response of the particles with different degrees of locking. The results highlight the difference in recoverability of the sphalerite bearing particles with different degrees of locking, with simple locking texture giving higher recovery than complex locking texture, for the same overall liberation.