Surface chemical mechanisms of inadvertent recovery of chromite in UG2 ore flotation: Residual layer identification using statistical ToF-SIMS analysis

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.     

Study of celestite flotation efficiency using sodium dodecyl sulfonate collector: factorial experiment and statistical analysis of data

Flotation tests of synthetic mixtures of celestite (SrSO₄) and calcite (CaCO₃) minerals using sodium dodecyl sulfonate as collector were carried out using a factorial experimental design 2³. The independent experimental variables included celestite grade in the feed (50% and 90% SrSO₄), conditioning pH (6.5 and 9) and sodium silicate depressant concentration (10⁻⁵ and 10⁻³ M). All experiments were performed at a constant collector concentration (10⁻⁴ M) and conditioning time of 15 min. The response variables were recovery and grade of SrSO₄ in the concentrate.The factors that had the greatest effect on the grade and recovery of celestite were the celestite grade in the feed (L) and the depressant concentration (C); also the combination of these factors significantly affected the response variables. The highest celestite yield (96% recovery, 98% grade) was obtained when a 90% of celestite grade in the feed was used.     

新型捕收剂RA-92在低品位碳酸锰矿选矿中的应用

微细粒低品位锰矿由于颗粒间的非选择性聚集、浮选药剂用量大、浮选效率低等技术难题而致使其利用困难,造成大量浪费。在品位低于13%的锰矿浮选技术研究中,捕收剂最受关注,前人已研究了多种类型的捕收剂,所得精矿品位在16.9%~18.3%之间,回收率为56%~97%,回收率比较理想,但精矿品位总体不高。本文将新型捕收剂RA-92应用于湖南凤凰-花垣地区低品位碳酸锰矿(锰品位为10.7%)的选矿工艺中,实验研究了磨矿细度、pH值、抑制剂和捕收剂用量对浮选效果的影响,在最佳工艺条件下,精矿品位由原矿的10.7%提升至17.4%,回收率达到80.2%。研究表明RA-92对碳酸锰矿具有良好的捕收性能,浮选工艺相对简单且捕收剂用量少,浮选成本较低,可为此种捕获剂在微细粒低品位碳酸盐锰矿选矿中的应用得到推广。     

An experimental study on the partition of zinc and lead between the silicate and sulfide liquids

On the basis of experimental works in the FeS-FeO-SiO₂(-Fe₃O₄ or -Na₂O) system with synthetic ZnS or PbS, the partition of zinc and lead between silicate and sulfide liquids is shown to be affected by the oxygen content of the sulfide liquids. The partition coefficients K, metal wt. % in sulfide liquid / metal wt. % in silicate liquid, for zinc and lead go through a minimum at relatively low oxygen contents of the sulfide liquids. K_Zn and K_Pb for natural sulfide liquids in equilibrium with basic magmas near the earth's surface are estimated at 0.1–0.5 and greater than 10, respectively. Although K_Zn and K_Pb change appreciably with oxygen content of the sulfide liquids, they never become sufficiently high to concentrate zinc and lead in economic amounts in magmatic sulfide ores.     

Selective flotation of scheelite from calcium minerals with sodium oleate as a collector and phosphates as modifiers. I. Selective flotation of scheelite

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.     

Removal of lead minerals from copper industrial flotation concentrates by xanthate flotation in the presence of dextrin

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.     

ToF-SIMS-derived hydrophobicity in DTP flotation of chalcopyrite: Contact angle distributions in flotation streams

Particle hydrophobicity has been derived from Time of Flight Secondary Ion Mass Spectrometry (ToF-SIMS) measurements and its impact on the flotation behaviour of chalcopyrite investigated. Batch flotation tests were performed using a dithiophosphate-type of collector in different concentrations. Three flotation regimes were studied using particle size ranges of 20–38 μm, 75–105 μm and 150–210 μm. The individual particle contact angle, and hence, the distribution of contact angles of chalcopyrite within feed, concentrate and tail flotation samples has been determined using ToF-SIMS secondary ions. The effects of particle size and hydrophobicity on the flotation behaviour have been investigated using this new approach. The hydrodynamic effects of the particle size were highlighted by the different distributions of contact angles obtained for each concentrate size fraction, with fine and coarse sizes requiring higher average contact angles to float. This effect was overtaken by hydrophobicity when a high collector concentration was used. The broad distribution of contact angles observed in all samples, i.e. heterogeneity in hydrophobicity, has significant consequences for interpreting flotation behaviour. The methodology of analysis conducted in this study was applied to real ore and can be used as a quantitative, diagnostic tool for examining surface chemical factors affecting hydrophobicity. This new technique has promise and may advance the understanding of mechanisms, which may lead to better control strategies for improving flotation performance. Furthermore, any mineral–collector system can be targeted, provided appropriate calibration is performed.     

Efficient separation of zinc from zinc containing copper sulfide concentrate by chemical leaching

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.     

Zinc extraction from Iranian low-grade complex zinc–lead ore by two native microorganisms: Acidithiobacillus ferrooxidans and Sulfobacillus

A mesophilic iron oxidizing bacterium, Acidithiobacillus ferrooxidans, has been isolated (33 °C) from a typical, chalcopyrite concentrate of the Sarcheshmeh copper mine in the region of Kerman located in the south of Iran. In addition, a thermophilic iron oxidizing bacterium, Sulfobacillus, has been isolated (60 °C) from the sphalerite concentrate of Kooshk lead and zinc mine near the city of Yazd in the center point of Iran. Variation of pH, ferrous and ferric concentration on time and effects of some factors such as temperature, cell growth, initial ferrous concentration and pH on bioleaching of low-grade complex zinc–lead ore were investigated. The results obtained from bioleaching experiments indicate that the efficiency of zinc extraction is dependent on all of the mentioned variables; especially the temperature and initial Fe(II) concentration have more effect than other factors for these microorganisms. In addition, results show that the maximum zinc recovery was achieved using a thermophilic culture. Zinc dissolution reached 58% with Sulfobacillus while it was 51% with A. ferrooxidans at pH = 1.5, initial Fe(II) concentration = 7 and 9 g/L for A. ferrooxidans and Sulfobacillus, respectively, after 30 days.     

Exploiting the dual functions of polymer depressants in fine particle flotation

Fine hydrophilic particles are known to be entrained with water in flotation of many ores. Flocculation of the hydrophilic particles by polymer depressants could potentially reduce the mechanical entrainment of these particles. This paper reports testwork completed on fine particles of several solids, iron oxide, hydroxyapatite and sphalerite, as well as on a relatively coarse quartz sample (− 75 + 38 μm). Dodecylamine was used as a collector for quartz, and several dispersants and polymer depressants, including sodium silicate, sodium metaphosphate, zinc sulfate, cornstarch, corn dextrin and carboxymethyl celluloses (with molecular weights of both 700,000 and 80,000) were used as flotation modifiers. The major part of the testwork involved flotation tests in a 200 mL flotation column. It was observed that flocculation of the fine hydrophilic particles significantly reduced their mechanical entrainment, while dispersion severely aggravated it. Thus, in the flotation separation of synthetic mixtures of the − 75 + 38 μm quartz and fine (reagent grade) iron oxide or hydroxyapatite, polymer depressants that caused flocculation performed better than those that did not cause flocculation.     

The flotation of gersdorffite in sulphide nickel systems — A single mineral study

The flotation response of gersdorffite has been determined in a series of batch flotation tests in which a sulphide sample containing about 76% gersdorffite was floated from mixtures with quartz.Using xanthate as collector, the arsenide floated strongly at pH 9, irrespective of whether lime, sodium hydroxide or sodium carbonate was used as pH modifier. When the pH was raised further to 10 or more using lime, the flotability decreased sharply.Changing the pulp potential at pH 9 also strongly affected the flotation response. Above about − 230 mV SHE, the mineral floated strongly, but below this potential it was essentially non-flotable. As far as it was tested (up to + 400 mV SHE), no upper limiting potential was found.Adding cyanide to the grinding mill brought about only a weak depressant effect on subsequent gersdorffite flotation at both pH 9 and 10. However, adding the same amount of cyanide to the flotation cell produced a strong depressant effect at the same pH values. At both pH 9 and pH 10, a threshold cyanide addition existed below which gersdorffite floated strongly, and above which it did not. For the conditions used in this study, this addition equates to about 100 g/t at pH 9 and about 80 g/t at pH 10. A hypothesis for the depression effect is presented which is consistent with both the available flotation data and a series of diagnostic analyses of grinding and flotation pulps completed during the test work.     

Mineral processing and extraction of rare earth elements from the Wadi Khamal Nelsonite Ore, Northwestern Saudi Arabia

A technological sample (50 kg) from Wadi Khamal Nelsonite ore was subjected to magnetic and flotation concentration techniques. Excellent recovery percentages of 72.95% and 71.22% were achieved by the dry/wet magnetic and flotation concentration techniques, respectively. The weight of the apatite concentrate reached a reasonable percentage of approximately 23.5% with an overall 40.23% P₂O₅ total content. Analytical data of the apatite concentrate after digestion in concentrated sulfuric acid revealed that the total content of the rare earth elements (REE) constitutes about 0.2% of the total apatite content. The REE content (0.2%) was partitioned between phosphoric acid liquor (65%) and gypsum precipitate (36%). The extraction of the REEs from the phosphoric acid liquor using oxalic acid and sodium carbonate–bicarbonate mixture (1:10?w/w) yielded the RE oxide cake which constitute about 1.2% (w/w). The produced rare earth oxide cake contains traces of various metal oxides, e.g., SrO, Na₂O, etc. in addition to rare earth oxides. Attempts to determine quantitatively the constituents of the cake will be considered in future work.     

Effect of collector, pH and ionic strength on the cationic flotation of kaolinite

The effect of amine collector type, pH, and ionic strength on the flotation behaviour of kaolinite was investigated in a series of laboratory batch flotation tests. In distilled water, ether diamine, a strong collector for silica, does not induce any flotation or only very weak flotation of kaolinite over a wide pH range from pH 3 to pH 10.5. Ether monoamine causes strong flotation of kaolinite in distilled water, especially in acidic solutions, but high dosages of the collector are required. Such observations are in contrast to the flotation behaviour of oxide minerals such as silica for which ether diamine shows stronger collecting power than ether monoamine. The pH dependence of kaolinite flotation is also opposite to that of oxides, with lower flotation recovery obtained at higher pH. In contrast to oxides, the flotation recovery of kaolinite increases with ionic strength. It was demonstrated that the enhanced flotation of kaolinite in NaCl solutions cannot be attributed to the frothing ability of NaCl or the intercalation of kaolinite by alkylamines. It is proposed that the screened zeta potential of kaolinite particles in a high ionic strength environment causes random aggregation of kaolinite particles exposing hydrophobic (001) silica plane in the presence of ether amines.     

Flotation of galena. Influence of grain size,of pulp deoxygenation,and of cleaning with ammonium acetate

The flotation of < 10, 10–20, and 20–40 μm galena fractions was studied. For uncleaned galena a given collector coverage produced better floatability with increasing grain size. Nitrogen had a detrimental effect only for the < 10 μm fraction, producing at a given collector coverage a recovery smaller than that obtained with air.Galena cleaned with 400 g/l ammonium acetate had very poor floatability, although xanthate abstraction was fairly high; this confirms that strong xanthate adsorption is necessary for flotation. Formation of monothiocarbonate was small in all cases, which points to a very minor influence, if any, of this compound in the flotation process.In blank flotation tests, or for very low residual xanthate concentrations, a peak at 208 nm and a shoulder at 255 nm were observed. The former was assigned to the uncomplexed Pb²⁺ ion, and the latter was tentatively attributed to the PbOH⁺ ion.Lead in solution results from dissolution of the oxidation products of galena, as galena itself has an exceedingly low solubility. The curve for total lead in solution vs. initial xanthate concentration, had a minimum for an initial xanthate concentration of 10^?5M, the further increase in dissolved lead is attributed to formation of complexes such as PbX⁺ (X = xanthate). Dissolved lead concentrations were nearly as high for cleaned as for uncleaned galena, which indicates a high oxidation rate of the mineral.     

Proposed upgrading techniques for East Sibaiya phosphorites,Aswan area,southern Egypt

The Late Cretaceous (Campanian-Maastrichian), low-grade phosphorite sequence of East Sibaiya, Aswan area, which is known as the Duwi (phosphate) or Sibaiya Formation, is usually intercalated with marl, oyster limestone, and chert beds. These strata, which crop out in a generally east–west trending belt spanning the middle latitudes of Egypt, are phosphorite-rich sediments and of great economic importance. Representative samples were collected from the investigated area at several localities (e.g., Umm Tundubah-2 and Wadi El-Batur) and 200 kg of phosphate rocks was used for upgrading processes of low-grade phosphorites of the East Sibaiya area. The upgrading processes included two techniques: the first technique (gravity separation) comprises crushing, sieving, gravity separation by a shaking table, and magnetic separation. This technique raised the P₂O₅ in the head sample from 24.73 to 31.91% as a phosphate concentrate. The second technique (flotation technique) depends on certain flotation parameters such as pH, grain size, and phosphate collector dose (i.e., oleic acid). The flotation technique increased the P₂O₅% from 24.73 to 31.16% as a final product of the phosphate concentrate. These data were confirmed by X-ray fluorescence analyses of major elements.     

云南大屯锡粗精矿工艺矿物学研究

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。     

河北某斑岩型钼矿提高选矿指标试验研究

河北省某钼矿为单一斑岩型钼矿,主要金属矿物为辉钼矿。为了进一步提高钼精矿的品位和回收率,试验采用混合捕收剂(煤油∶2号油=2∶1)和新型捕收剂PE-100相结合的方法,粗选时可使粗精矿的回收率提高2个百分点,品位也略有提高。为节约生产成本,试验采用阶段磨矿阶段选别的选矿工艺,即原矿磨矿(-0.074mm占60%)后,经一次粗选,一次扫选,粗精矿再磨(-0.038mm占85%)后再进行5次精选,最终获得钼精矿品位w(Mo)=50.007%,回收率为89.90%的较好指标。     

Dissolved air flotation (DAF) of fine quartz particles using an amine as collector

Experimental studies concerning the dissolved air flotation (DAF) of fine (d_p < 100 μm) quartz particles, using two different flotation cells (setups), are presented. Pure and well characterised quartz samples were treated with a commercial amine as collector prior to flotation and bubbles were characterised by the LTM-BSizer technique. Bubble size distribution showed 71% (by volume) and 94% (by number) of the bubbles having sizes (d_b) lower than 100 μm (i.e. microbubbles). The Sauter and arithmetic mean diameters were 79 μm and 56 μm, respectively, for the bubbles generated at 300 kPa (gauge) saturation pressure (after 30 minute saturation time). Quartz particle size distribution (obtained by laser diffraction) showed a volume-moment diameter of 13 μm. The Rosin–Rammler–Bennett, Gates–Gaudin–Schumann and log-normal distribution functions were well fitted (R² > 0.96) to the bubble size distribution and quartz particle size distribution data. Values of total quartz recovery ranging from 6% to 53% (by mass) were obtained for the DAF experiments under different collector concentrations (up to 2 mg g^− 1), with an optimal collector concentration found at 1 mg g^− 1. These results are significant considering that 27% (by volume) of the quartz particles are ultrafine (d_p < 5 μm), demonstrating the widely-known efficiency of DAF to remove small particles when applied in the field of water and wastewater treatment. The true flotation behaviour, as a function of particle diameter (d_p), exhibits a local minimum when particles are approximately 3–5 μm in size. The results contribute to the discussion in the literature about the existence of such a minimum, which is generally interpreted as a change in the mechanism of particle collection from convection (collision) to diffusion at lower particle sizes.     

Flotation of calcareous Mussorie phosphate ore

Separation of Mussorie rock phosphate (P₂O₅ = 20%) from Uttar Pradesh, India, containing pyrite, calcite and other carbonaceous impurities by flotation has been successfully attempted to upgrade the phosphate values. Based on Hallimond cell flotation results of single and synthetic mineral mixtures of calcite and apatite using oleic acid and potassium phosphate, conditions were obtained for the separation of calcite from apatite which is considered to be the most difficult step in the beneficiation of calcareous phosphates. Further studies using 250 g of the mineral (?60 +150 and ?150 mesh fractions, deslimed) in laboratory size Fagergren subaeration machine employed a stagewise flotation viz. carbonaceous materials using terpineol, pyrite using potassium-ethyl xanthate and calcite using oleic acid respectively. Separation was, however, found to be unsatisfactory in the absence of a depressant.Among starch, hydrofluosilicic acid and dipotassium hydrogen phosphate, which were tried as depressants for apatite in the final flotation stage, dipotassium hydrogen phosphate proved to be superior to others. However, the tests with the above fractions did not yield the required grade. This was possibly due to insufficient liberation of the phosphate mineral from the ore body and different experimental conditions due to scale up operations. Experiments conducted using ?200 mesh deslimed fractions has yielded an acceptable grade of 27.6% P₂O₅ with a recovery of about 60%. The results have been explained in terms of the specific adsorption characteristics of phosphate ions on apatite and the liberation size of the mineral.     

Two-stage reverse flotation process for removal of carbonates and silicates from phosphate ore using anionic and cationic collectors

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.