Depressant Dosage Research Articles

Evaluation of disodium carboxymethyl trithiocarbonate as a depressant reagent for chalcocite: An electrochemical approach

The effect of the organic depressant disodium carboxymethyl trithiocarbonate (Orfom® D8) on the mineral chalcocite (Cu2S) was investigated with mixed potentials (Em), microflotation, UV–vis spectrophotometry, electrochemical impedance spectroscopy (EIS) and Fourier transform infrared spectroscopy (FTIR). The mixed potential measurements revealed a decrease in the potential as the depressant dosage was increased, which was attributed to adsorption of the depressant. Microflotation results showed that the mineral recovery percentage decreased from 56.37 % to 5.20 − 11.94 %, depending on the depressant concentration and solution pH. The fits of the experimental data indicated that the depressant was chemically adsorbed, and this was corroborated by EIS studies that indicated an electrochemical mechanism involving the depressant and the mineral. The proposed adsorption mechanism involved oxidation of the depressant for the formation of Cu(D8–)2 and CuD8– complexes on the chalcocite surface in addition to the formation of CuD8– complexes on the oxidized surface (Cu1.95S and Cu1.8S); thus, these complexes were responsible for depression of the mineral. AFM measurements revealed roughness changes associated with depressant adsorption and surface oxidation of the mineral. Finally, the FT-IR spectra showed that adsorption of the D8 depressant on chalcocite occurred through the trithiocarbonate group, which oriented the carboxylate group away from the mineral surface, thus conferring hydrophilic character. The observed behavior indicated that the Orfom® D8 depressant can be used as a chalcocite depressor in a selective flotation separation stage with an adsorption mechanism different from that proposed for chalcopyrite but with a similar depressor effect.

Flotation Enhancement of sedimentary phosphate ores by cornstarch as an environmental depressant: Modeling and analysis using full factorial design (FFD) and artificial neural network (ANN) approaches

The present work attempts to highlight different facets of corn starch as an environmentally friendly depressant using soybean oil as a collector. The important influence of parameters such as pH, collector and depressant dosage, and the role of the depressant are discussed. The grade (%P2O5) of the flotation products were analyzed by means of UV–visible spectroscopy, the recovery (%Re) and the efficiency (%E) of the flotation products are calculated based on the grade. The mechanism of action of starch depression was revealed through inductive coupled plasma (ICP), Fourier transform infrared (FT-IR) spectroscopy, and X-ray diffraction (XRD). In addition, full factorial design (FFD) and artificial neural network (ANN) were used to generate an evaluation approach for P2O5 content. Moreover, the results obtained confirm that starch has an influence on phosphate depression at more acidic and alkaline pH. Indeed, at pH ​= ​4, a P2O5 content of 28.29% was obtained with a recovery of 87.46% in the non-floating fraction. Similarly, at pH ​= ​12, a content of 27.60% P2O5 with a recovery of 92.10% was found at a CaO/P2O5 ratio equal to 1.6. These concentrates were obtained from a feed sample containing 22.09% P2O5 using 10.3 ​g/L of soybean oil and 15 ​g/L of corn starch. The results of the comparison between the ANN and FFD approaches show that the ANN model outperforms the FFD model in terms of performance, with a good and higher coefficient of determination (R2 ​= ​0.999).

Optimization of flotation conditions in the beneficiation of PGMs tailings

For several years, mining waste has shown a negative impact on both the environment and human health. The mining industry remains the backbone of the economic growth. Different technologies have been implemented to beneficiate and recover platinum group metals from tailings. The recycling of tailings has been a point of research interest due to their extensive applications. Flotation has been the primary process of upgrading and recovering PGMs. The focus of this study was to optimize flotation conditions in the beneficiation of PGMs for particular small-scale mine tailings. This was done to obtain the most favourable conditions for the small-scale mine tailings to improve operating conditions of specific particle sizes. PGMs tailings obtained from a small-scale mine were characterized using XRD, XRF, SEM/EDS, and ICP – OES to understand the properties of the tailings prior to mineral processing. Flotation batch tests were conducted. The results showed that the chosen particle size was 75 µm, and the favorable reagent dosages were 150 g/Mg and 100 g/Mg for SIBX (collector) and Starch(depressant), respectively. At favourable conditions, the recovery was 65.75% (Pt = 70.38%, Pd = 59.33%, Ru = 34.56%), and the grade was 31.46 g/Mg (Pt = 21.43 g/Mg, Pd = 9.62 g/Mg, Ru = 0.41 g/Mg). It can be concluded that all the flotation parameters are related; lower particle sizes yield high recoveries and better grades due to the exposure of the particle surface to the reagents responsible for the flotation of the PGMs. It was observed that the high collector dosages produce high recoveries with low grades of PGMs. High depressant dosages produce low recoveries with high grades of PGMs. The relationship between the collector and the depressant is of essential importance in the flotation process.

A Novel Chalcopyrite Depressant for Selective Separation of Molybdenite from Cu-Mo Sulfide Ores and Its Interaction Mechanisms

In this study, GX2 was applied as a new high-efficiency chalcopyrite depressant to selectively separate molybdenite from Cu-Mo sulfide ores. The flotation performance and its interaction mechanisms with chalcopyrite and molybdenite were investigated using single-mineral and artificial-mixed-ore flotation, contact angle measurements, zeta-potential measurements, infrared spectroscopy, and X-ray optoelectronic spectrum analysis. The results indicated that molybdenite could be selectively separated from chalcopyrite under the optimal flotation conditions of pH 9.0, 80 mg/L GX2, 20 mg/L kerosene, 10 mg/L MIBC, and a flotation time of 3 min, while the molybdenite and chalcopyrite recoveries were around 90% and 5%, respectively. It was confirmed that GX2 could save ten times the depressant dosage compared to that of the Na2S baseline to achieve a similar separation efficiency. The contact angle test, zeta potential, infrared spectrum, and XPS results show that GX2 may be adsorbed on the surface of chalcopyrite via chemical adsorption, and the hydrophilic substances formed change its hydrophobicity, whereas the effect on the floatability of molybdenite is small, resulting in the excellent separation efficiency of chalcopyrite and molybdenite via flotation.

Flotation response of coal washery reject fines: Characteristics, process optimization, and oxidation

AbstractThe present paper outlines the characterization, electrokinetic behaviour, and flotation response of rejected coking coal fines with 32.5% ash generated in a coal washery in Eastern India. The response methodology and central composite rotatable design (RSM‐CCRD) were used for the process modelling and optimization of the flotation process using diesel, methyl isobutyl carbinol (MIBC), and sodium hexametaphosphate as a collector, frother, and depressant to maximize ash reduction, yield, and combustible recovery. At optimum condition, a 9.7% clean ash coal was achieved with a 63% yield at collector, frother, and depressant dosages of 0.78, 0.31, and 0.80 kg/ton, respectively. The model prediction and experimental data corroborated sufficiently. Subsequently, within 1 year, the fines oxidized and did not float with the collector. Fourier‐transform infrared spectroscopy (FTIR) confirmed surface oxidation on oxidized coal. The oxidized coal responded favourably to acid oil, a vegetable oil refinery waste. With acid oil as a collector, the oxidized coal can upgrade to 12% clean ash coal with a 60% yield; the combustible recovered is 80%. The work indicates successful upgradation of fresh and oxidized coal using the froth flotation process.

Development of a novel biocollector for flotation of low grade graphite ore

ABSTRACT Mineral processing plants operate in capacity of hundreds/thousands of tons per day. Accordingly, chemical reagents’ usage also increases proportionally. Stringent norms toward environment sustainability question the usage of chemical reagents, especially in large quantities and tailings disposal in open areas. Hence, bioreagents have gained great interest. Froth flotation is by far the most practiced processing route for fines beneficiation and low-grade ore upgradation especially for naturally hydrophobic minerals. Flotation being a physico-chemical separation technique, flotation reagents selection plays a pivotal role in the process performance. A novel environmental-friendly biocollector, an extract from the leaves of Vitex negundo, was used as flotation collector in the present investigation for beneficiating a low-grade graphite ore with 8.67% fixed carbon. A three-factor and three-level Box-Behnken Design (BBD) under Response Surface Methodology (RSM) was employed to study the effects of important process variables such as grinding time, depressant dosage, and collector dosage on the responses, namely, ash percent of final concentrate and its recovery. A final graphite concentrate with 4.24% ash and 14.42% yield was obtained using the developed biocollector by flotation of low-grade graphite ore with 89.47% ash content. The degree of significance of input variables was determined using ANOVA. Regression models for ash content, % of final concentrate, and its %recovery was obtained from BBD analysis. It showed that the grinding time has a significant influence on the process followed by depressant dosage on the grade of final concentrate and collector dosage on its recovery.

The Use of Acid Mine Drainage (AMD) in the Flotation of a Platinum-Group-Minerals-Bearing Merensky Ore

Water scarcity is compelling mining houses to not only recycle process water but to also identify alternative sources of make-up water in concentrators. South Africa has significant volumes of acid mine drainage (AMD) generated from vast mining operations. This study investigated the viability of using AMD as a replacement for potable water in the flotation of a platinum-group-minerals (PGM)-bearing Merensky ore. Rougher and cleaner flotation testwork was conducted at laboratory scale to compare the performances of potable water (baseline water), AMD treated with Ca(OH)2, and AMD treated with the Veolia process. Water analysis showed that the three water types differed in pH, water hardness, conductivity, and total dissolved solids. The results showed the AMD treated with Ca(OH)2 was detrimental to PGM recovery compared to potable water at depressant dosages of 50 g/t. Specifically, AMD treated with Ca(OH)2 achieved a PGM rougher recovery of 67.8%, while potable water achieved a PGM rougher recovery of 88.4%. Depressant dosage optimisation and treatment of the AMD using the Veolia process were investigated as potential strategies to mitigate the detrimental effects of the AMD treated with Ca(OH)2 on the flotation performance of a Merensky ore. The AMD treated with the Veolia process achieved a PGM rougher recovery of 70.8%. Thus, treatment of the AMD was beneficial, though the PGM and base metal sulphides (BMS) recoveries were still lower than those achieved in potable water. Reducing the depressant dosage to 25 g/t in AMD treated with Ca(OH)2 resulted in the highest PGM, Cu, and Ni rougher recoveries of 91%, 60.2%, and 58%, respectively. The AMD treated with Ca(OH)2 at lower depressant dosage outperformed the potable water in terms of PGM and BMS recoveries and concentrate grades, indicating that AMD has the potential to replace potable water as make-up water in Merensky ore processing plants. The results showed that depressant optimisation is important to achieve superior metallurgical results when using AMD treated with Ca(OH)2. The use of AMD in Merensky ore processing plants not only conserves freshwater in minerals processing plants but also reduces high volumes of contaminated effluents.

Impact of flotation operational parameters on the optimization of fine and coarse Itabirite iron ore beneficiation

This paper investigates the impact of operational parameters on the reverse flotation of Timbopeba itabirite iron ore in a pneumatic (MBE Pneuflot) flotation cell. To this end, the effects of physical and chemical parameters (i.e., collector and depressant dosage, superficial gas velocity (Jg), solids concentration, and feed rate) on the reverse flotation of iron ore are investigated in detail. Results show that increasing collector dosage increases quartz recovery, particularly for the coarser quartz where recovery increases by nearly 30%. Increasing depressant dosage results in a decrease in both quartz recovery and hematite loss, for the coarser hematite particles. In addition, increasing the feed rate does not affect quartz recovery or hematite loss, suggesting that the pneumatic flotation cell is not constrained by residence time. Furthermore, increasing the superficial gas velocity results in an increases in quartz recovery to an optimum at a Jg of around 0.6 cm/s after which it decreases significantly, particularly for the coarser particles where recovery decreases by 50%. Decreasing the solids concentration does not affect quartz recovery but results in a significant reduction in hematite loss. Here, fine and coarse hematite loss decrease from around 25% to 12% and 10% to 2%, respectively. Solids concentration and Jg are operating parameters that have the largest effect on the flotation performance of the pneumatic flotation cell. Overall, the obtained results suggest that this flotation cell should be operated at relatively low superficial gas velocities and solids concentrations. However, these findings are for a pilot scale and would need to be examined at larger scale (e.g., an industrial scale) where different operating conditions may be optimal.

Beneficiation of low-grade limestone by flotation

Low-grade limestone samples from Northern Telangana, India were put through to beneficiation by flotation to upgrade the CaO value by minimizing the unwanted (gangue) minerals for utilizing in the various industries. Analyzation of the sample unfolds the minerals such as 50.85% CaO, 4.86% SiO2, 0.986% Al2O3, 91.31% TC, and 39.40% LOI. XRD characterization reveals that it predominantly contains calcite, quartz minerals, and minor traces of alumina mineral. Wet Sieve analysis of the sample shows 80% of the feed material was 53% microns or finer and mineral sample was found to be limestone having density of 2697.16 kg/m3 from gas pycnometer analysis. Direct flotation was espoused to improve calcium oxide grade by minimizing the gangue minerals (alumina and silica) in laboratory experiments by variations of grinding time, depressant dosage, collector dosage, and pH value. This study comes out with optimum results at pH 6, 12 min grinding time, 0.67 kg/ton of sodium oleate (collector), 1.25 kg/ton of sodium silicate (depressant) and concentrate assayed 52.03% CaO, 3.89% SiO2, and 0.77% Al2O3.

Selective separation of chalcopyrite and talc using pullulan as a new depressant

As a gangue mineral, talc is usually associated with chalcopyrite. Therefore, the depression of talc is required in the field of sulphide mineral flotation. Herein, the depression mechanism of pullulan in the flotation separation of chalcopyrite from talc was investigated via flotation tests and characterisation techniques, such as Fourier transform infrared analysis (FTIR), X-ray photoelectron spectroscopy (XPS) and atomic force microscopy (AFM). The flotation test results showed that chalcopyrite can be selectively separated from talc with 20 mg/L pullulan as a depressant, and that the separation effect remains unaffected when the depressant dosage is increased to 80 mg/L. Additionally, the zeta potential, FTIR and adsorption measurements indicated that pullulan is selectively adsorbed on talc surfaces, and that talc surfaces adsorb more depressants than chalcopyrite. The XPS results indicated that the pullulan adsorption on talc surfaces is physical. The AFM results indicated that pullulan adsorption on talc surfaces forms polymer layers. Therefore, pullulan is considered a potential depressant of talc in the flotation industry.

Development of impurities removal process for low-grade Sanje iron ore using mineral processing technologies

In this study, the impurities removal process for low-grade Sanje iron ore was developed using Wet High-intensity magnetic separation (WHIMS) and Reverse flotation (RF). Sanje Iron Ore is the low-grade hematite ores found in Nampundwe area of Zambia from which Iron is to be extracted and used as the feed in the steelmaking process. The ore contains 34.18 mass per cent Iron grade, 31.10 mass% of Silica (SiO2) and 7.65 mass per cent Alumina (Al2O3). Magnetic Separation experiments were done using Series L Model 4 laboratory magnetic Separator (L-4 Machine) as the first stage impurity removal process and the effect of various magnetic separation parameters such as magnetic flux density, particle size density and pulp density of the feed were studied. The results showed that 10 T was optimal magnetic flux density which enhanced the recovery of 93 per cent of iron with 53.22 mass per cent grade. The iron concentrate produced from magnetic separation contained 12.04 massper cent Silica and 3.94 massper cent Alumna and therefore, it was further treated using Reverse flotation. In reverse flotation, various parameters such as pH, collector dosage, Iron depressant dosage and quartz activator dosage were investigated. The results showed that 81.94 per cent was recovered at the concentrate’s pH of 6.8 using 200 g/T of 0.1 per cent calcium oxide (CaO) as silica activator and one kg/T of 0.1 per cent causticised starch as Iron depressant. Sodium Oleate (NaOL) and Dodceylamine Acetate (DAA) each with discrete dosage, were used as Anionic and Cationic collectors respectively. Alumina was consequently reduced to 1.04 mass per cent and Silica to 2.04 mass per cent at optimum respective collector’s dosage of 0.250 kg/T using 0.02 kg/T of Methyl Isobutyl Carbinol (MIBC) frother. Additionally, phosphorous was also observed to be reduced from 0.05 mass per cent to 0.01 mass per cent. The designed multi-stage process involving feeding the concentrate from WHIMS into RF process therefore produced a high-grade iron concentrate iron with 67.27 mass per cent grade, 2.04 mass per cent silica, 1.04 mass per cent alumina.

Minimalization of Ash from Iranian Gilsonite by Froth Flotation

Gilsonite is a natural, brittle, and glisten tar hydrocarbon, which widely uses in the chemical, paint, and oil industry, besides asphalt production and thermal insulation. Major gangues of Gilsonite are gypsum, pyrite, silica, dolomite, calcite, and shale which are known as ash. Due to the fine liberation degree of gangue minerals from Gilsonite, most of the physical mineral processing methods are not applicable. Gilsonite has partially a behavior like coal in flotation due to its high carbon content and the similar composition of ash. As a result, flotation is one of the best methods to decrease the ash content of Gilsonite. In this study, four factors i.e. the dosage of collector, frother, and depressant, and solid to liquid ratio in three levels were examined in flotation to reduce the ash content and increase the recovery of Gilsonite. These tests were designed with the Taguchi method by Design-Expert software. Our results showed the lowest ash content of 5.2% was obtained in the condition which 200 g/t Gasoil as collector, 100 g/t MIBC as a frother, 300 g/t sodium silicate as a depressant, and pulp density was 5% solid to liquid weight ratio.

Studi Konsentrasi Pada Bauksit Asal Tayan Dengan Menggunakan Metode Flotasi Kebalikan

This paper is one of a study that implements reverse flotation methods as an alternative beneficiation on bauxite processing. This study is aimed to increase alumina content in bauxite so its qualified as feed of Bayer Process (Al2O3> 51%, SiO2 <7%, Al2O3 / SiO2 > 10 ratio). The experiments was focused to increase alumina content from bauxite were carried out in a 1375 mL of flotation cell tank. Dodecylamine (DDA) was used as the collector, starch as the depressant, and Dowfroth 1012 as frother. The overall experiments were conducted on solid content to solution 25% w/w and speed rotation of impeller were 900 rpm. The characterization results showed that the feed ore had Al2O3 and SiO2 content, respectively 61,24% and 27,9%. Experiments were carried out in several variations to determine what variables were most affected in the study. There were 4 variations used in this study, size fraction of particle, collector dosage, pH of slurry and with the addition of / without the addition of depressants. Alumina content as results of flotation was analyzed by XRF. The results of XRF of flotation products show that increasing of alumina content. The optimum pH range of bauxite flotation using DDA as collectors can be achieved at pH 4 – 10. From the results, The increasing of collector dosage at pH 10 gives a good result on Al 2 O 3 content is 62% with 50,44% recovery of Al 2 O 3 obtained at dosage of collector at 500 g/ton. The Al 2 O 3 content which higher in conditions without the addition of depressant is 66,8%. While for the size fraction variation, it shows that separation better in the finer particle which gives an increasing of alumina and decreasing of silica as a result. The best condition in this study were achieved by 400 g/ton of DDA dosage, 400 g/ton of depressant dosage and size fraction of particle in -325 mesh at pH 10 with increasing alumina to 79.1%.

Utilization of polyepoxysuccinic acid as the green selective depressant for the clean flotation of phosphate ores

In flotation, the application of some flotation reagents may cause serious water pollution and damage the ecological environment and human health. Large consumption of flotation reagents also restricts the sustainable production of mines. Therefore, the development and application of environmentally friendly and economical flotation reagents are essential for the sustainable production of phosphate ores and phosphate fertilizer. In this work, a non-nitrogen and phosphorus-free green, economical and environmentally friendly reagent, polyepoxysuccinic acid was used to evaluate its selective inhibition on calcite during apatite flotation. Micro-flotation experiments were used to investigate the floatability of the two minerals. Bench-scale flotation experiments were used to verify the efficiency of industrial application of polyepoxysuccinic acid in the flotation phosphate ores. Depression mechanism of polyepoxysuccinic acid was revealed through zeta potential measurements, Fourier transform infrared spectroscopy, X-ray photoelectron spectroscopy, and atomic force microscope. The flotation separation of apatite and calcite can be achieved with polyepoxysuccinic acid using sodium oleate as collector. Bench-scale flotation experiments show that a good flotation index with a P2O5 grade of 32.74% and a recovery of 84.82% can be obtained at a depressant dosage of 80 g/t. Mechanism experiment proved polyepoxysuccinic acid had a stronger adsorption on calcite than apatite. The adsorption of polyepoxysuccinic acid on calcite was mainly chemical adsorption through the chelation of its carboxyl group and calcium atoms. Therefore, this work provides a clean and economic reagent for the sustainable production of phosphate ores and phosphate fertilizer, which is of great significance for the sustainable development of clean mines.

Improving recovery of iron using column flotation of iron ore slimes

In general, the industrial flowsheets for the beneficiation of itabirite iron ores from the Quadrilátero Ferrífero (Minas Gerais state) in Brazil involves grinding, desliming, magnetic separation and cationic reverse flotation using both mechanical and columns flotation cells. At Brazilian plants, it is estimated that Fe losses generated at the slimes stage range between 8% reaching values higher than 20%. Slimes generated at hydrocyclone containing particle sizes of approximately 99% < 45 μm, 25% > 10 μm, 45% < 5 μm and 15% < 1 μm, 3% of solids by weight and between 35% and 45% of Fe grade. This work aims at evaluating the technical feasibility of concentration of two slimes samples from two different plants in operation at the Quadrilátero Ferrífero using the column flotation technology without prior desliming. Despite the similarity in terms of mineralogy, grade and particle size distribution, the flotation results indicate a very distinct behavior considering similar process conditions, including variables such as collector type and dosage, depressant dosage, pH and circuit configuration. A final concentrate containing an Fe grade higher than 60%, 4% SiO2 and 3.1% Al2O3 with an Fe recovery up to 75% in rougher column flotation can be obtained from Sample 1 with 45.2% Fe, 28.6% SiO2 and 3.1% Al2O3, when a collector type amide-amine (Flotinor 5530) was used in substitution of a conventional amine collector. On the other hand, it was not possible to obtain a final concentrate containing an Fe grade higher than 55% from Sample 2 with 39.6% Fe, 33.4% SiO2 and 5.1% Al2O3, using both types of collector (amide-amine and amine) even when applying different types of column flotation circuits (rougher or rougher/cleaner). Based on literature review and the results obtained in this study, it can be observed the non-existence of a standard solution for an efficient slimes processing in the iron ore industry. Considering flotation as the most suitable technology for the treatment of such type of material, future research should focus on a deep fundamental study of the collector-depressant mechanism interaction with the iron-bearing and gangue minerals, especially in those ores with a high amount of alumina-containing minerals. Considerations about the effect of bubble generator device to flotation columns and its impact on the metallurgical performance are also discuss regarding to application at industrial scale to slimes concentration.

Beneficiation of a Low-grade Phosphate Ore Using a Reverse Flotation Technique

ABSTRACT This work aimed to upgrade a low-grade sedimentary phosphate ore through attrition scrubbing/desliming and an anionic/cationic reverse flotation technique. The mineralogical and chemical examination showed that the sample contained low-grade apatite minerals (P2O5 ~ 22.5%), and the main associated gangue minerals were gypsum and quartz. The effects of both collector and depressant dosage were investigated and optimized in each individual flotation process. The results of attrition scrubbing accompanied by desliming showed that a high content of P2O5 (27.97%) with a recovery of 72.1% was obtained in the coarse size fraction (+53 µm), and the majority of quartz and gypsum separated in the fine size fraction (−53 µm). The effect of using a new tall oil collector mixture was investigated during the anionic reverse flotation. Sodium tripolyphosphate (STPP) was used as a phosphate depressant during anionic and cationic reverse flotation. The flotation results of the deslimed sample demonstrated that cationic reverse flotation was much more desirable in terms of achieving a high grade and recovery of P2O5 compared to anionic reverse flotation. Considering the statistical design, a maximum grade of 31.23% P2O5 with a recovery of 95.22% was obtained in the non-floated fraction using 950.88 g/t of the collector (amine) and 500 g/t of the STPP.

Parametric Optimization in Rougher Flotation Performance of a Sulfidized Mixed Copper Ore

The dominant challenge of current copper beneficiation plants is the low recoverability of oxide copper-bearing minerals associated with sulfide type ones. Furthermore, applying commonly used conventional methodologies does not allow the interactional effects of critical parameters in the flotation processes to be investigated, which is mostly overlooked in the literature. To tackle this issue, the present paper aimed at characterizing the behavior of five key effective factors and their interactions in a sulfidized copper ore. In this context, dosage of collector (sodium di-ethydithiophosphate, 60–100 g/t), depressant (sodium silicate, 80–120 g/t) and frother (methyl isobutyl carbinol (MIBC), 6–10 g/t), pulp pH (7–11) and agitation rate (900–1300 rpm) were examined and statistically analyzed using response surface methodology. Flotation experiments were conducted in a Denver type agitated flotation cell at the rougher stage. The experimental results showed that increasing the pH (from 8 to 10) at low agitation rate (1000 rpm) enhanced the recovery from 80.36% to 85.22%, while at high agitation rate (1200 rpm), a slight declination occurred in the recovery. Meanwhile, increasing the collector dosage at a lower frother value (7 g/t), caused a reduction of about 4.44% in copper recovery owing to the interactions between factors, whereas at a higher frother level (9 g/t), the recovery was almost unchanged. The optimization process was also performed using the goal function approach, and maximum copper recovery of 92.75% was obtained using ~70 g/t collector, 110 g/t depressant, 7 g/t frother, pulp pH of 10 and 1000 rpm agitation rate.

An experimental study on the effects of some process parameters on lignite flotation

ABSTRACT Flotation, as an efficient method for separating fine coal particles (−500 µm), is making slow progress for use with lignite, which basically remains at the laboratory stage. Lignite flotation is more difficult than flotation of other coals, and flotation efficiency here is very low. The difficulty in floating lignite primarily includes two aspects, abundant oxygen containing groups on the coal surface and porous surface of the coal particles. Therefore, it is of paramount importance to clean lignite by flotation. In this article, the effects of some process parameters for lignite flotation were studied. Laboratory experiments were carried out, and the effects of depressant (sodium silicate) dosage (0–1000 g/t), collector (kerosene) dosage (50–3000 g/t), frother (methyl isobutyl carbinol) dosage (15–1000 g/t), pulp pH (2–12), solid ratio (10%–20%), flotation time (1–7 minutes) and particle size (500–63 µm) on flotation performance were investigated. A concentrate with 37.18% yield and 40.11% combustible recovery was obtained at 250 g/t sodium silicate (Na2SiO3), 250 g/t kerosene and 100 g/t methyl isobutyl carbinol (MIBC). Within the studied range, the optimum pulp pH was 2, the solid ratio was 10%, the flotation time was 5 minutes, and the particle size was −500 µm. Under these optimum conditions, the ash content of lignite coal was reduced from 19.08% to 12.70%.

Influence of operating parameters on the flotation of the Khibiny Apatite-Nepheline Deposits

Phosphate as fertilizer is so essential for agriculture plants growth. Most of the phosphate resources cannot be directly treated due to the low grade of P2O5 and high content of impurities. Therefore, upgrading of this type of ore is so essential to achieving a grade of concentrate suitable for the production of fertilizer and phosphoric acid. Flotation is one of the most efficient techniques applied to phosphate upgrading. This paper aimed to investigate the effect of operating parameters, including collector dosage, depressant dosage, particle size, and pulp pH, on the rougher stage flotation of apatite ore. The results indicated that at the optimum conditions of flotation, a concentrate of 30% P2O5 with a recovery of a proximately 75% was obtained from a feed of particle size less than 250 microns and 10.8% P2O5.

Use of citric acid and Fe(III) mixture as depressant in calcite flotation

Calcite, as a gangue mineral, needs to be removed from the target minerals in the flotation. Single depressant citric acid (CA) or Fe(III) species has a limited depressive effect on calcite in the sodium oleate (NaOL) flotation system. In this paper, the floatability of calcite with single depressant (CA or Fe(III)) and mixed depressant (CA /Fe(III)) was studied by microflotation experiments. Solution chemical calculation and X-ray photoelectron spectroscopy (XPS) analysis were used to explain the relevant depressive mechanism. Microflotation experiments indicate that the addition of Fe(III) greatly improves the depressive effect of the depressant CA on calcite. When mixed depressant Fe(III)/CA is used and the depressant dosages of Fe(III) and CA are all 1 × 10−4 mol/L, the recovery of calcite is only 3.9% at pH 8.5. The mixed depressant Fe (III)/CA can be co-adsorbed on the calcite surface. The Fe(III) species in the slurry react with CA species first and then co-adsorbed on the calcium atomic sites of the calcite. Moreover, the Fe(III) species is first adsorbed on the oxygen atom sites of the calcite, and then the Fe(III) species provides the adsorption sites for further adsorption of citric acid. Therefore, the mixed depressant Fe(III)/CA can be adsorbed at both the calcium atomic sites and the oxygen atomic sites of the calcite. The mixed depressant greatly improves the depressive effect on calcite, which is beneficial to the flotation separation of calcite and related useful minerals.