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1.
Jan Drzymala Janusz Kapusniak Piotr Tomasik 《International Journal of Mineral Processing》2003,70(1-4):147-155
Potato starch and dextrins resulting from thermolysis of potato starch in the absence of reagents and presence of
-amino acids are promising depressants for separation of lead and copper minerals present in the Polish industrial copper concentrates. The polysaccharides were used for differential xanthate flotation of the final industrial concentrates produced by flotation with sulfhydryl collectors in the absence of depressants. The polysaccharides depressed galena and provided froth concentrate rich in chalcocite and other copper minerals as well as cell product containing lead minerals. The best results of separation were obtained in the presence of plain dextrin prepared by a thermal degradation of potato starch. The industrial concentrate containing 18.5% Cu and 5.5% Pb was divided into a froth product containing 38.1% Cu with 77% recovery of copper and a cell product assaying 7.3% Pb with 83% recovery of lead. It was accomplished using 2500 g/t of dextrin, 50g/t of potassium ethyl xanthate, and 50 g/t of frother (α-terpineol). The pH of flotation was 8.0–8.2. 相似文献
2.
Physical separation of bitumen from low-grade Utah tar sand deposits containing a relatively high viscosity bitumen phase (Sunnyside and Tar Sand Triangle deposits) has been accomplished by traditional size reduction and froth flotation techniques. At appropriate experimental conditions more than 90% of the bitumen can be recovered in a concentrate, containing more than 20 wt. % bitumen, which should be a suitable feed material for subsequent hot water or thermal processing. The efficiency of bitumen recovery depends on the extent of size reduction, as well as promoter and dispersant addition. Rejection of greater than 60% of the sand at ambient temperature and ease of water removal from the concentrate make such a process strategy both energy and cost effective. The energy required to achieve effective separation by the ambient temperature process is significantly less than the energy required for the recently developed hot water process which is being evaluated in a 100-tpd pilot plant this year.The flotation behaviour of the tar sand in this ambient temperature process has been correlated with contact angle measurements and the apparent point-of-zero-charge of the bitumen. The best flotation response at pH 7.8 to 9.0 occurs when the contact angle between the air bubble and bitumen surface is a maximum, corresponding to the apparent point-of-zero-charge of the bitumen as determined by titration. 相似文献
3.
《Chemie der Erde / Geochemistry》2021,81(4):125769
In this study, a new flotation approach, a low-alkaline and non-desliming process, was introduced for improving lead and zinc recoveries, lowering production cost and reducing environmental pollution. Lab-scale experiments results show that the new process contributed to the flotation of the complex mixed sulfide-oxide lead and zinc ore regarding two aspects: (1) High alkaline process (pH = 12±) was replaced by low alkaline process (pH = 9±) by using collector WS (a mixture of ethyl thiocarbamate, ammonium dibutyldithiophosphate and dithiophosphate-25) and combined depressant Na2S/ZnSO4/Na2SO3 for lead sulfide flotation; (2) Non-desliming process was successfully achieved by using collector MA (a mixture of ether amine, hydroxyethyl cellulose and polyacrylic acid) and combined depressant SHP/SS (sodium hexametaphosphate/sodium silicate) for zinc oxide flotation. And 43.37% Pb in the Pb concentrate was recovered, the corresponding Pb grade was 52.73%, total 84.42% Zn was recovered by the flotation of zinc sulfide minerals and zinc oxide minerals. Moreover, the two aspects of the new approach were systematically verified from lab-scale to industrial-scale application. The industrial-scale flotation tests show that Pb recovery in Pb concentrate increased by 1.86% compared with that of original system during industrial-scale tests period, and the Pb recovery increased by 4.09% compared with that of original system before industrial-scale tests period, while the Zn operating recovery in zinc oxide concentrate improved by 19.52%. Moreover, the total reagent cost of the whole new process significantly declined by 3.93 yuan per ton of ore. 相似文献
4.
The limitations of pulp chemistry measurements in the flotation of a platinum group mineral (PGM) bearing Merensky ore were demonstrated in Part 1 of this article. In this paper the importance of the contribution of the froth structure due to changing froth stability is analysed using the batch flotation data. The effects of mild steel (MS) and stainless steel (SS) milling media and the addition of copper sulphate on the flotation performance of the sulphide minerals in Merensky ore have been evaluated in relation to the changes in stability of the froth phase. The effects of pulp chemistry and froth stability on the flotation of sulphide minerals were distinguished by using two different rate constants (kt and kw). The rate constant (kw) calculated as a function of cumulative water recovery was used to describe characteristics of froth phase and kt was calculated as a function of flotation time. The results revealed that the type of grinding media and copper sulphate addition had an interactive effect on the froth stability. While mild steel (MS) milling increased the froth stability due to the presence of hydrophilic iron hydroxides and colloidal metallic iron, the addition of copper sulphate reduced the stability, especially with stainless steel (SS) milling. Copper sulphate addition had a dual role in the flotation of Merensky ore in that it caused destabilisation of the froth zone as well as activation of selected sulphide minerals. The dominant effect was found to depend on the type of milling media and floatability of the mineral in question and this work has demonstrated the importance of using a combination of measurements to evaluate flotation performance holistically. 相似文献
5.
A laboratory flotation column using Venturi aerators and a vacuum system to remove froth was used to investigate the contribution of gas flow, pulp flow, cell volume and froth retention time on the ink removal efficiency and on cellulose fibres and mineral fillers loss. The increase in the gas flow from 4 to 8 L/min gave a general rise of particle transport from the pulp slurry to the froth with an ensuing strong increase in ink removal, from 75% to 85%, and water and total loss, from 10% to 40% and 15% to 30%, respectively. Whereas, the increase of the cell volume from 14 to 24 L improved ink removal from 72% to 80% without considerably affecting flotation loss. The rise of the froth retention time in the flotation cell from 5 to 20 s before removal gave a general decrease in the flotation loss from 20% to 11% without a corresponding decrease in ink removal. This trend was interpreted as reflecting poor ink drainage through the froth. The increase of both pulp and froth retention time in the flotation cell appeared as the most favourable way to improve ink flotation selectivity. A mathematical model, describing particle removal during flotation in terms of true flotation, entrainment and drainage, was proposed and used to fit experimental data. 相似文献
6.
《International Journal of Mineral Processing》2005,77(2):65-79
A model is developed by taking into account the simultaneous mechanisms of true flotation and entrainment in a conventional flotation cell. The total volume of the cell is divided into three compartments: pulp collection zone, pulp quiescent zone and froth region, with the mechanisms being modeled as occurring at the same time but originating at different places: true flotation from the collection zone and entrainment from the quiescent one. A particle is referred to as suspended in water or attached to an air bubble, depending upon its original state before crossing the pulp–froth interface (whether or not it remains in that state all the way to the concentrate launder). The model is obtained by solving a set of equations describing the mass conservation of solids and water between adjacent compartments. The principal mass transfer factors are identified as: the flotation rate constant, the mean residence time in the collection zone, the froth recovery of attached particles, the degree of entrainment through the froth and the water recovery from the feed to the concentrate. The development presented here allows the intricate nature of the mass transfer in a flotation cell to be reduced to one single equation, overcoming the need of numerical methods for simulation purposes. Moreover, it is shown that reliable prediction of grade and recovery can be obtained without detailed information on the pulp hydrodynamics or on any froth sub-process either than drainage, bubble bursting and bubble coalescence. 相似文献
7.
8.
Leonard J. Warren 《International Journal of Mineral Processing》1985,14(1):33-44
Laboratory batch flotation tests were carried out on a deslimed (+6 μm) sulfiderich cassiterite ore, an ultrafine fraction (?6 μm) of a cassiterite ore and a bituminous coal. Chemical conditions were kept constant but the water recovery was varied by changing the height of the froth column and the rate and depth of froth removed. The recovery of the floatable mineral in each system was then found to be linearly related to the weight of water recovered. The intercept of the regression line on the mineral recovery axis, where the water recovery is zero, was interpreted as the recovery due to true flotation. The entrainment contribution was proportional to the slope of the line. In this way the contributions of entrainment and true flotation to overall recovery were separated. 相似文献
9.
A fully automated semi-commercial flotation column incorporating state of the art instruments was designed to study the amenability of flotation column for the beneficiation of different minerals. In the present study, beneficiation of sillimanite was investigated by installing the flotation column in the flotation circuit of Orissa Sands Complex, Indian Rare Earths Limited, Chatrapur, Orissa. The effect of process parameters on sillimanite grade and recovery was investigated. At optimum conditions, the flotation column was operated continuously with a feed rate of one ton per hour and demonstrated the efficiency of the technology for the beneficiation of sillimanite. The results show that a concentrate assaying 96% sillimanite at 90% recovery can be obtained in a single column flotation stage. 相似文献
10.
In the UG2 ore (Bushveld Complex, South Africa) flotation, normally more than 3% of the gangue minerals, principally chromite with talc and pyroxene, report to the concentrate diluting the PGM recovery and contributing to subsequent processing costs. Previous studies have identified residual talc-like layers on orthopyroxene surfaces in Merensky ore flotation contributing to inadvertent flotation of relatively large particles (20–150 µm) of this mineral. Chromite (75–150 µm) from flotation of UG2 ore has been similarly examined. Statistical comparison of ToF-SIMS analysis of particles from concentrate and tails reveals no significant difference in Cu, Pb, Ni and collector (IBX and DTP) signals between these streams but surface exposure of Mg and Si is favoured in the concentrate. The flotation rate of coarse chromite correlates with the exposures of magnesium and silicon in patches on the chromite surface; higher exposures give earlier flotation. Conversely, there is a negative correlation with signals corresponding to the chromite surface, i.e. Cr, Fe, Al. Flotation of chromite without collector has confirmed this statistical discrimination. Hydrophobic talc-like residual layers, similar to those found on orthopyroxene surfaces, probably from partial alteration, explain this flotation mechanism. 相似文献
11.
It is well known that the chemical environment determines the success of the flotation process, however its characterisation and control is difficult to achieve. This paper, as two parts, Part I and Part II, evaluates the use of various measurements and their interpretation to gain an understanding of the influence of varying parameters such as the type of milling media and copper sulphate addition on the flotation performance of sulphide minerals from a platinum group mineral (PGM) bearing Merensky ore. It shows the complexity of interpretation and the importance of analysing flotation performance holistically. Part I focuses on the pulp chemistry and mineral potential measurements have been used to show the differences in the response of the various mineral electrodes to different conditions. The final flotation recoveries of the sulphide minerals in the ore followed the same trend as the decrease in mineral potential due to collector addition viz. chalcopyrite > pentlandite > pyrrhotite. Type of milling media and copper sulphate addition slightly affected the mineral electrode potential and flotation recovery of chalcopyrite. Addition of copper sulphate increased the recovery of pentlandite and particularly pyrrhotite due to activation by copper (II) ions. The copper activation mechanism was likely to be in the form of initial adsorption of copper hydroxide followed by reduction to Cu+ at the surface. However, the changes in flotation performance of the different minerals in the ore could not be completely described by the electrochemical changes, demonstrating the limitations of these measurements. Part II addresses the effect of froth stability as demonstrated by the variations in the mass and water recovery data resulting from the different milling conditions and addition of copper sulphate which emphasised the importance of considering the froth phase in the evaluation of flotation data. 相似文献
12.
Jan Drzymala Andrzej Luszczkiewicz Piotr Simiczyjew 《International Journal of Mineral Processing》1983,10(4):289-296
Ilmenite from Polish magnetite-ilmenite ores has been floated with tall oil. Chemical and mineralogical analysis of the flotation products established that the ores containing hercynite are difficult to upgrade. For ore A (18.7% hercynite and 32.8% ilmenite) and ore B (9% hercynite and 41.3% ilmenite) poor concentrates containing much less than the required 45% TiO2 have been obtained. It has been found that the higher concentration of hercynite in the feed, the lower the grade of concentrates in respect to TiO2 is obtained. Ilmenite ore containing a minor amount of hercynite (ore C, 0.15% hercynite and 12.3 ilmenite) gives a good concentrate. By means of microscopic observation it was established that hercynite floats together with ilmenite in the rougher flotation but is partially depressed in the cleaning and scavenging flotation. 相似文献
13.
G.D. Senior L.K. Smith W.J. Bruckard 《International Journal of Mineral Processing》2010,94(1-2):92-96
High arsenic levels in nickel sulphide concentrates can often present technical and environmental problems at the smelter. In some ores a flotation separation between the nickel sulphides and the arsenide minerals is required in order to meet smelter specifications for arsenic. There is very little information in the literature about such separations and the flotability of nickel arsenides in general. Data are presented here from a single mineral flotation investigation into the flotability of niccolite (NiAs), one of the key nickel arsenide minerals present in nickel sulphide ores in Australia.Niccolite is only weakly flotable with xanthate collector at pH 9 and this behavior is independent of the grinding environment. This suggests that high arsenic levels in nickel concentrates are not due to niccolite flotation, but are more likely to be due to flotation of more floatable arsenic-bearing minerals such as gersdorffite. Cyanide diminishes the poor recovery of niccolite further. 相似文献
14.
Phosphate rock contains various gangue minerals including silicates and carbonates which need to be reduced in content in order to meet the requirements of the phosphate industry. Froth flotation has become an integral part of phosphate concentration process. In this study, double reverse flotation was applied to recover apatite from phosphate ore. H3PO4 and CaO were used as phosphate depressants, in acidic and alkaline conditions. Fatty acids and amines were added as carbonate and silicate collectors respectively. An experimental protocol devised to optimize the grade and recovery of phosphate using anionic–cationic method was found effective. Consequently, a required high quality of phosphate concentrate containing 30.1% P2O5 was obtained, with a recovery of 94%. X-ray diffraction and optical microscopy studies were performed to define the main minerals. 相似文献
15.
以云南大屯选矿厂锡粗精矿为研究对象,采用化学分析、X射线衍射分析及光学显微镜分析等手段对该粗精矿的化学组成、矿物组成、矿物嵌布粒度特征等进行了详细的研究。结果表明,锡粗精矿中有价元素锡的品位为13.80%,锡矿物主要以锡石形式产出。锡粗精矿中TFe含量为30.78%,主要以褐铁矿、磁黄铁矿的形式存在,磁黄铁矿是导致粗精矿含硫高的主要原因。锡粗精矿中主要的脉石矿物有白云石、透闪石、电气石、石英、白云母、萤石等,且脉石矿物与锡石均有不同程度的毗邻连生、包裹共生关系。本次工艺矿物学研究认为,大屯选矿厂锡粗精矿宜采用浮选预先脱硫,除去其中的硫化物,再对浮选尾矿采用重选工艺提高锡品位和回收率。该研究结果可以为大屯选矿厂工艺流程改造和合理开发利用锡资源提供科学依据。 相似文献
16.
针对福建某低品位钼矿矿石性质,确定了粗磨粗选、粗精矿再磨精选的浮选方案。采用该浮选方案及合理的选别条件,对含钼为0.08%的钼原矿选别,获得钼精矿品位52.45%、钼精矿产率为0.14%、钼回收率90.19%的良好指标。 相似文献
17.
The effects of Na2SiO3, Na3PO4, Na4P2O7, (NaPO3)6, quebracho, tannic acid and S 808 (sulphonated product of rough phenantrene) on the floatability of the following five pure minerals: scheelite, calcite, fluorite, garnet and quartz, with sodium oleate as collector were investigated in detail as well as the role of pH on these effects. The results obtained indicate that Na4P2O7 and (NaPO3)6 were effective modifiers for the selective flotation of scheelite. The results of the batch flotation tests on mixtures of these minerals showed that the recovery of scheelite from scheelite-silicate mixtures (31% WO3) with (NaPO3)6 or Na4P2O7 increased by 20% as compared with sodium silicate and the WO3 grade of the concentrate by 5%. At room temperature, the scheelite-calcium mineral mixtures could not be separated with sodium silicate. In the separation of these mixtures with the phosphate modifiers, a concentrate grade of 47–60% WO3 was obtained at 70–90% recovery. This showed that the flowsheet of the selective flotation of scheelite with phosphate modifiers may replace the conventional Petrov's process. 相似文献
18.
《矿物学报》2013,(Z1)
Effects of particle size and chain length on flotation of quaternary ammonium salts (QAS) onto kaolinite have been investigated by mico-flotation tests. The two kinds of quaternary ammonium salts [RN(CH3)3] with different chain lengths, dodecyltrimethylammonium chloride (DTAC) and cetyltrimethylammonium chloride (CTAC) were used as collectors for kaolinite in different particle size fractions (0.075–0.01 mm, 0.045–0.075 mm, 0–0.045 mm). The anomalous flotation behavior of kaolinite have been further explained based on crystal structure considerations by adsorption tests and molecular dynamics (MD) simulation. The results show that the flotation recovery of kaolinite in all different particle size fractions decreases with an increase in pH when DTAC and CTAC are used as collectors. As the concentration of collectors increases, the flotation recovery increases. The longer the carbon chain of QAS is, the higher the recoveries of coarse kaolinite (0.075–0.01 mm and 0.045–0.075 mm) are. But the flotation recovery of the finest kaolinite (0–0.045 mm) decreases with chain lengths of QAS collectors increasing, which is consistent with the flotation results of unsifted kaolinite (0–0.075 mm). It is explained by the froth stability related to the residual concentration of QAS collector. In lower residual concentration, the froth stability becomes worse. Within the range of flotation collector concentration, it's easy of CTAC to be completely adsorbed by kaolinite in the particle size fraction (0–0.045 mm), which led to lower flotation recovery. Moreover, it is interesting that the particle size of kaolinite is coarser, the flotation recovery is higher. The anomalous flotation behavior of kaolinite is rationalized based on crystal structure considerations. The results of MD simulations show that the (001) kaolinite surface has the strongest interaction with DTAC, compared with the (001), (010) and (110) surfaces. On the other hand, when particle size of kaolinite is altered, the number of basal planes and edge planes is changed. It is observed that the finer kaolinite particles size become, the greater relative surface area of edges is, and the more the number of edges is. It means that fine kaolinite particles have more edges to adsorb less cationic colletors than that of coarse kaolinite particles, which is responsible for the poorer floatability of fine kaolinite. 相似文献
19.
鄂西高磷鲕状赤铁矿原矿全铁品位47.56%,含P 0.93%,主要脉石矿物为绿泥石、磷灰石、石英、方解石、铁白云石,属难选铁矿石。通过磁化焙烧-磨矿-磁选优化工艺,最佳磁化焙烧条件为:焙烧温度800℃、焙烧时间90min、还原剂用量12%,焙烧矿磨矿细度-0.074mm占85.15%,经弱磁选可得到全铁品位为58.13%、磷含量0.70%,铁回收率为90.41%的粗精矿。对磁化焙烧-磁选过程的各产物组成分析表明,焙烧矿和粗精矿中主要矿物为磁铁矿,占比分别为65%和85%;主要脉石矿物为绿泥石、磷灰石、石英、铁白云石等。粗精矿矿物的嵌布粒度较细,-0.074mm粒级占85.15%,但部分矿物仍以相互浸染、包裹、鲕状碎屑、连晶等形式存在,矿物仍未完全单体解离,从而导致粗精矿中杂质磷、铝等含量较高。粗精矿细磨后粒度-0.022mm含量为80%时,磁铁矿的解离度为84.63%,可实现磁铁矿充分单体解离,经过深选可提高铁精矿质量。 相似文献
20.
《Chemie der Erde / Geochemistry》2021,81(4):125773
As the most abundant copper containing resource and zinc containing resource, chalcopyrite and sphalerite/marmatite commonly coexist as Cu-Zn mixed ores in deposits. However, it is difficult to completely separate sphalerite and chalcopyrite by flotation, thus resulting in the existence of zinc impurity in copper concentrate. Sphalerite/marmatite existed in copper sulfide concentrate as impurity may lead to severe damage of the smelting equipment, and cause the waste of copper and Zn resources, it will also decrease of the sale price of copper concentrates. Therefore, the deep separation of zinc from zinc bearing copper sulfide concentrate is of great significance. In this work, selective chemical leaching was developed to efficiently remove zinc from zinc containing copper sulfide concentrate. Under the optimal condition (i.e., sulfuric acid concentration exceed 100 g/L, temperature of 80 °C, pulp density of 10%, leaching time of 48 h), over 85% Zn was extracted into the leaching solution together with only about 10% Cu and Fe, according to the leaching experiment. Leaching slurry had good solid-liquid separation characteristics, and zinc can be further effectively recovered from the leaching solution. According to X-ray diffraction (XRD) and scanning electron microscope/energy dispersive spectrometer (SEM/EDS) analysis, chalcopyrite was the main mineralogical phase in the residues, which can be regarded as high quality copper concentrate for metallurgy. Accordingly, a new process for deep and efficient separation of Cu-Zn mixed ores has been proposed. 相似文献