Flotation of pyrrhotite and pyrite in saturated CaCO3 solution using a quaternary amine collector

In the production of ground calcium carbonate (GCC) for the paper industry, any colouring contaminants must be removed during processing (usually flotation) to achieve sufficient GCC brightness. Flotation of GCC feedstock is characterised by the presence of both particulate and dissolved calcite (CaCO₃), an alkaline pH and the use of amine collectors. This paper investigates the possibility of removing pyrrhotite and pyrite, under these conditions. Microflotation results show that the recoveries are highly dependent on pH. Pyrrhotite recovery decreases dramatically when going from pH 8 to pH 10. Pyrite display high recoveries at pH 8 and 9, and lower recoveries at higher pH. Recoveries were dependent on conditioning time. Pyrite floats considerably faster than pyrrhotite at all pH levels investigated, whereas pyrrhotite seems to be slow floating and could benefit from prolonged flotation times. Compared to the saturated CaCO₃ system, flotation recoveries decreased when using deionised water or CaCl₂ solution. The flotation results could not be fully explained by zeta potential. Bench scale flotation experiments on a sulphide bearing marble, confirmed the microflotation results at low collector concentrations (i.e. 200 g/t). By increasing the amine concentration, the flotation performance became independent of pH.     

The effect of heavy oxidation upon flotation and potential remedies for Merensky type sulfides

Surface oxidation of sulfide minerals, such as that found in the regions of a sulfide ore body near the water table, can have a significant impact upon flotation. This theme has been explored for Merensky ore type sulfides where an ore containing pyrrhotite, pentlandite and chalcopyrite was thermally oxidised and the role of potential remedies investigated. Back-scattered scanning electron microscope images are presented showing the oxidation layer which formed in the mineral surfaces. These oxidation layers were depleted in both sulfur and iron with incorporated oxygen. Flotation recovery rapidly decreased with increasing oxidation, particularly after 27 days and reached a plateau after 50 days. Up to 27 days, this effect could be partially overcome with higher collector additions. Oxidation had more impact upon the finer size fractions, particularly for pyrrhotite. For more heavily surface oxidised samples, ultrasonic treatment prior to collector conditioning was found to improve flotation recoveries. This treatment had the greatest effect upon chalcopyrite particles. Sulfidisation was successful in restoring the flotation recovery of the heavily oxidised sulfide minerals. Longer sulfidisation conditioning times were not conducive to good flotation recoveries of both oxidised pyrrhotite and pentlandite due to oxidation of the freshly formed sulfide surfaces. For maximum flotation recoveries of oxidised pyrrhotite, pentlandite and chalcopyrite, different sulfidisation conditions are indicated. It appears likely that in a mineral processing operation treating oxidised Merensky type ores, two stages of sulfidisation employing different conditions would be required.     

Separation of amine-insoluble species by flotation with nano and microbubbles

Amines (alkylamines–ether amines) are employed on a large scale to separate iron ores by reverse flotation of the gangue particles (mostly quartz and silicates). Quartz gangue particles coated with amine collector are dumped in tailings dams as concentrated pulps. Then, the fraction of the amines that detach from the surfaces and the portion that is soluble in water, contaminate surface and ground-water supplies. This work presents a novel flotation technique to remove decyl-trimethyl-ether-amine (collector employed in Brazilian iron mines) from water. This amine forms precipitates at pH > 10.5 which are removed by flotation with microbubbles (MBs: 30–100 μm) and nanobubbles (NBs: 150–800 nm). Bubbles were generated simultaneously by depressurization of air-saturated water (P_sat of 66.1 psi during 25 min) forced through a flow constrictor (needle valve). The flotation by these bubbles is known as DAF-dissolved air flotation, one of the most efficient separation technologies in water and wastewater treatment. Herein, best results (80% amine removal) were obtained only after selective separation of the MBs from the NBs exploring the fact that while the NBs remain dispersed in water, the MBs rise leaving the system. The MBs, because of their buoyancy, rise too rapidly and do not collide and adhere appropriately at the amine colloids/water interface, even causing some precipitates breakage. It was found that the “isolated” NBs attach onto the amine precipitates; aggregate (flocculate) them and entrain inside the flocs before rising by flotation. Because of the low residual amine concentration in water (6 mg L⁻¹), it is believed that this flotation technique have potential in this particular treatment of residual amine-bearing effluents.     

Effect of Leptospirillum ferrooxidans on the flotation kinetics of sulphide ores

The objective of this work was to determine the effect of Leptospirillum ferrooxidans on the floatability of chalcopyrite, sphalerite, and pyrrhotite by using xanthate as a collector. The tests were carried out in the absence and presence of bacteria in relation to the type of ore and contact time with bacteria. The results indicate that the chalcopyrite flotation rate significantly increased in the presence of L. ferrooxidans due to the formation of hydrophobic species. The bacteria function as a weak depressant for pyrrhotite after a conditioning time ?60 min. The behaviour of sphalerite remains without changes due to its low susceptibility to oxidation. It was concluded that L. ferrooxidans brings about superficial changes mainly due to the oxidation of minerals.     

Influence of bromine modification on collecting property of lauric acid

Bromine atom with strong electronegativity was introduced to α-carbon position of lauric acid (LA) by solvent-free method (Hell–Volhard–Zelinski reaction) at ambient pressure in laboratory, and the synthesized product α-Bromolauric acid (CH₃(CH₂)₉CHBrCOOH, α-BLA) was used as a new type collector for the flotation of quartz mineral. The flotation properties of pure quartz using α-BLA as a collector were investigated by single mineral flotation tests. The adsorption mechanism of α-BLA collector on quartz surface was established by zeta potential measurements, Fourier transform infrared (FT-IR) spectroscopy, and X-ray photoelectron spectroscopy (XPS), in conjunction with the results of quartz micro-flotation tests. Pure mineral flotation results showed that the collector α-BLA exhibited an excellent performance at alkaline conditions (pH ⩾ 11.50), activator CaCl₂ concentration 1.0 × 10⁻⁴ mol/L, and collector concentration 1.5 × 10⁻⁴ mol/L in a relatively lower temperature 15 °C, where about 99.8% of the quartz could be floated out. Compared with collector LA, the new synthesized collector α-BLA is more tolerant to lower pulp temperature and fluctuations of the reagents dosages. The study revealed that the α-BLA collector had adsorbed on the surface of pure quartz in the forms of chemical interaction, electrostatic adsorption and hydrogen bonding adsorption based on the results of zeta potential measurements, FT-IR spectra, and XPS.     

Flotation and adsorption of a new collector α-Bromodecanoic acid on quartz surface

A new type collector α-Bromodecanoic acid (CH₃(CH₂)₇CHBrCOOH, α-BDA) was synthesized by solvent-free method (Hell–Volhard–Zelinski reaction) in laboratory for the flotation of quartz mineral at a relatively low temperature of 16 °C. The adsorption mechanism of α-BDA collector on quartz mineral surface was established by zeta potential measurements, contact angle measurements, Fourier transform infrared (FT-IR) spectroscopy, and X-ray photoelectron spectroscopy (XPS), in conjunction with the results of quartz micro-flotation tests. The flotation results showed that the activator Ca(II) functioned as an indispensable and crucial factor on the recovery of pure quartz. When the quartz was floated with α-BDA alone at strong alkaline conditions (pH ⩾ 11.50), the recovery rate reached to only 11.9%. However, when using the activator Ca(II) at a concentration of 4 × 10⁻⁴ mol/L, the collector exhibited an excellent performance, where about 99.5% of the quartz had been floated at or above pH value 11.50 at 16 °C. The study revealed that the α-BDA collector had adsorbed on the surface of pure quartz in the forms of electrostatic interaction, hydrogen bonding and chemical adsorption based on the results of zeta potential measurements, contact angle measurements, FT-IR spectra, and XPS.     

Amine–oleate interactions in feldspar flotation

This study concerns the interaction between residual amine (tallow amine acetate) and sodium oleate for floating mica and metal-oxides respectively from a feldspar ore of Cine-Milas region of Turkey. Zeta-potential measurements show the co-adsorption of oleate onto feldspar surfaces only if amine is present. Zeta-potential measurements are in good compliance with laboratory scale flotation tests. Flotation tests show that the amine concentration ought to be kept lower than 1.47 × 10⁻⁵ mol/l (≈5 ppm) to prevent feldspar loss during metal-oxides flotation stage and hence dewatering/washing operation after mica flotation seems to be a crucial step. Feldspar recovery in the concentrate, being 86.67% without dewatering/washing increases to 94.58% after three stage dewatering/washing between mica and metal-oxides flotation stages. In order to get rid of the residual amine, as an alternative to dewatering, 300 g/t bentonite as the residual amine adsorber is used right after mica flotation and it helps to remove 97.7% of the residual amine in the cell. Bentonite addition provided almost the same feldspar recoveries compared with that of dewatering/washing.     

Surface characterisation,collector adsorption and flotation response of enargite in a redox potential controlled environment

We previously investigated oxidation of the surface of natural enargite (Cu₃AsS₄) under potentiostatic control and the formation of oxidation species at the mineral surface at selected applied potentials in the oxidative range. Here we further extended the research by incorporating flotation collectors into the system. Electrochemical techniques, X-ray photoelectron spectroscopy (XPS) and microflotation in a redox potential controlled environment were applied to examine surface properties, collector adsorption and flotation response of enargite in pH 10 solutions of sodium ethyl xanthate (SEX) and sodium dialkyl dithiophosphinate (3418A). The spectral details of XPS analysis of electrochemically treated enargite surfaces show significant adsorption of SEX and 3418A collectors onto enargite at an applied voltage of +516 mV, but no adsorption of both collectors at −400 mV. The results of XPS analysis agree with the floatability of enargite determined by microflotation, showing that the flotation recovery was highest at high oxidative potential (+516 mV), then decreased at low oxidative potential (+100 mV) and was very poor at −400 mV. These results confirm that enargite floatability can be efficiently controlled electrochemically.     

Flotation behaviors of ilmenite,titanaugite, and forsterite using sodium oleate as the collector

The flotation behaviors of ilmenite, titanaugite, and forsterite using sodium oleate as the collector were investigated using microflotation experiments, zeta-potential measurements, Fourier transform infrared (FT-IR) analyses, X-ray photoelectron spectroscopy (XPS) analyses and the artificially mixed minerals flotation experiments. The results of the microflotation experiments indicate that ilmenite exhibits good floatability when pH > 4.0. Titanaugite possesses a certain floatability at pH 4.0–6.0 and pH > 10.0, and forsterite possesses certain floatability at pH 5.0–7.0 and pH > 9.0. The results of FT-IR and XPS analyses indicate that sodium oleate mainly interacts with Fe, resulting in ilmenite flotation; that the Ca and Mg on the titanaugite surface chemically reacted with sodium oleate, and that the Mg on the forsterite surface chemically reacted with sodium oleate under acidic condition. However, sodium oleate mainly reacted with the Ca and Mg on the titanaugite surface, whereas sodium oleate mainly reacted with the Mg on the ilmenite and forsterite surfaces under alkaline conditions. The results of the artificially mixed minerals flotation experiment demonstrate that the concentrate of TiO₂ grade increases from 16.92% to 30.19% at pH 5.4, which represents the appropriate conditions for the flotation separation of ilmenite from titanaugite and forsterite under weak acidic conditions.     

A mixed collector system for phosphate flotation

Flotation has been used in industry for more than a half century as the primary technique for upgrading phosphate. While the flotation of phosphate was inefficient when oleic acid was used alone as a collector, therefore a mixed collector of oleic acid (HOl), linoleic acid (LA) and linolenic acid (LNA) was employed to improve the recovery of phosphate flotation. The batch flotation results showed that the optimal composition of the mixed collector was 54 wt.% HOl, 36 wt.% LA and 10 wt.% LNA. Additionally, the effect of pH on the mixed collector application was studied while considering the surface tension, contact angle and micro-flotation. The results showed that the mixed collector should be used at a pH of 9.5. Above a pH of 9.5, the adsorption of fatty acids dimers on the apatite surface hindered phosphate flotation. The influence of the mixed collector assembly on apatite flotation was also investigated. It was demonstrated that due to its low critical micelle concentration, a sufficiently hydrophobic apatite surface could be generated at a collector concentration of 60 mg/L. In addition, zeta potential experiments suggested that collector adsorption was governed by chemisorption. FTIR and XPS spectra studies further indicated that the chemical reaction involved the carboxyl groups of fatty acids and Ca species at the apatite surface for each fatty acid in the mixed collector.     

Flotation of rare earth minerals from silicate–hematite ore using tall oil fatty acid collector

The flotation of rare earth (RE) minerals (i.e. xenotime, monazite-(Nd), RE carbonate mineral) from an ore consisting mainly of silicate minerals (i.e. primary silicate minerals and nontronite clay) and hematite was investigated using tall oil fatty acids (Aero 704, Sylfat FA2) as collector. The RE minerals are enriched with Fe. The effects of tall oil fatty acid dosage, pH, temperature, and conventional depressants (sodium lignin sulfonate, sodium metasilicate, sodium fluoride, sodium metasilicate and sodium fluoride, and soluble starch) were determined at grinding size of P80 = 63 μm. At this grinding size, the grain size of the RE minerals ranges from 2 to 40 μm, percentage liberation is 9–22%, and percentage association with nontronite and quartz is 30–35%. Results indicated that Sylfat FA2 at 22450 g/t concentration was the more efficient tall oil fatty acid collector at natural pH (pH 7) to basic pH (pH 10.0–11.5). Flotation at the room temperature (25 °C) gave higher selectivity than 40 °C temperature flotation. The results on the effect of depressants showed similar selectivity curves against the gangues SiO₂, Al₂O₃, and Fe₂O₃ suggesting that the chemical selectivity of the depressants has been limited by the incomplete liberation of the RE minerals in the feed sample. High recoveries at 76–84% (Y + Nd + Ce)₂O₃ but still low (Y + Nd + Ce)₂O₃ grade at 2.1% in the froth were obtained at flotation conditions of 63 μm, 25 °C, pH 10.5, 1,875 g/ton sodium metasilicate and 525 g/ton sodium fluoride or 250 g/ton soluble starch as depressant for the silicates and hematite, and 22,450 g/t Sylfat FA2 as collector for the RE minerals (initial (Y + Nd + Ce)₂O₃ feed grade = 0.77%). The recoveries of gangue SiO₂, Al₂O₃, and Fe₂O₃ in the froth were low at 25–30%, 30–37%, and 30–36%, respectively. The mineralogical analysis of a high grade froth and its corresponding tailing product showed that the RE minerals have been concentrated in the froth while the primary silicate minerals and hematite have been relatively concentrated in the tailing. However, the clay minerals, primary silicate minerals, and hematite still occupy the bulk content of the froth. This suggests that incomplete liberation of the RE minerals led to the poor grade result, supporting likewise the selectivity curve results by the different depressants. This study showed that liberation is important in achieving selective separation.     

Selective separation of fine albite from feldspathic slime containing colored minerals (Fe-Min) by batch scale dissolved air flotation (DAF)

In this study, the separation of feldspar minerals (albite) from slimes containing feldspar and iron containing minerals (Fe-Min) was studied using dissolved air flotation (DAF) technique whereby bubbles less than 100 μm in size are produced. Before the flotation experiments with slimes, single flotation experiments with albite and Fe-Min were carried out using DAF in order to obtain optimum flotation conditions for the selective separation of feldspar from the slimes. Flotation experiments were performed with anionic collectors; BD-15 (commercial collector) and Na-oleat. The two methods of reagent conditioning were tested on the flotation performance; traditional conditioning and charged bubble technique. In addition, the effect of pH, flotation time, rising time, and drainage time which influence the selective separation in the DAF system were studied in detail. Overall, the flotation results indicated that the separation of albite from Fe-Min can be achieved with DAF at 5 min of rising time and 5 min of drainage time. Interestingly, these results also showed that the conditioning of the particles with the charged bubbles increased the flotation recovery of Fe-Min compared to the traditional conditioning. Furthermore, the flotation tests with the feldspathic slime sample were carried out under the optimum conditions obtained from the systematic studies using the single minerals. The charged bubble technique produced an albite concentrate assaying 0.33% Fe₂O₃ + TiO₂ and 11.07% Na₂O + K₂O from a slime feed consisting of 1.06% Fe₂O₃ + TiO₂ and 10.36% Na₂O + K₂O.     

The effect of reagents on selective flotation of smithsonite–calcite–quartz

In this paper the effects of sodium sulphide, sodium hexa methaphosphate (SH), sodium fluoric, starch and sodium silicate adsorption on smithsonite, quartz and calcite surfaces at various pH values, and using Armac C and oleic acid as collectors were investigated through microflotation. Also, the effects of various primary amine collectors (Armac C, Armac T, Flotigam SA, Flotigam TA and Armeen TD) were investigated for smithsonite flotation. The flotation tests were performed using purified samples from Angooran mine by the microflotation technique. The cationic flotation results showed that the maximum recovery of smithsonite could be improved to 92% using 400 g/t Armac C and 500 g/t sodium sulphide at pH 11. Also, the quartz and calcite recoveries reached 98% and 89%, respectively, at the above mentioned conditions. Moreover, using 1250 g/t SH and 1500 g/t sodium silicate as a depressant, the quartz and calcite recoveries decreased to 70% and 20%, respectively, and also the smithsonite recovery was reduced to 82%. Furthermore, the experiments showed that the behavior of sodium fluoric as a quartz depressant is similar to that of sodium silicate. Flotation results using oleic acid revealed that the maximum recovery of 90% occurs at pH 9 and 500 g/t oleic acid. Also, the quartz and calcite recoveries reached 26% and 87%, respectively, in the anionic flotation conditions. Increasing amount of sodium silicate to 2000 g/t caused a decrease in the smithsonite recovery to 87% and also decreased the calcite and quartz recoveries by 10% and 15%, respectively.     

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 amine and starch dosages on the reverse cationic flotation of an iron ore

In iron ore concentration, reverse cationic flotation of quartz has been successfully employed for particles below 150 μm previously deslimed. Amine and starch are used, respectively, as quartz collector and iron oxides depressant. Understanding the mechanisms of reagents interaction is relevant to improve the separation selectivity, especially for high amine dosages. The term clathrate was used to explain this interaction, meaning a molecular compound in which molecules of one species occupy the empty spaces in the lattice of the other species, resulting in the depression of hydrophobic minerals. Laboratory scale experiments were carried out with itabirite iron ore in three different size ranges. The clathrate formation between molecules of amine and starch may explain the increase of SiO₂ content in the concentrates of the coarse size range (−150 + 45 μm) due to an increase in amine dosage.     

Adsorption of copper sulphate on PGM-bearing ores and its influence on froth stability and flotation kinetics

Copper sulphate is used as an activator in the flotation of base metal sulphides as it promotes the interaction of collector molecules with mineral surfaces. It has been used as an activator in certain platinum group mineral (PGM) flotation operations in South Africa although the mechanisms by which improvements in flotation performance are achieved are not well understood. Some investigations have suggested these changes in flotation performance are due to changes in the froth phase rather than activation of minerals by true flotation in the pulp zone. In the present study, the effect of copper sulphate on froth stability was investigated on two PGM containing ores, namely Merensky and UG2 (Upper Group 2) ores from the Bushveld Complex of South Africa. Froth stability tests were conducted using a non-overflowing froth stability column. Zeta potential tests and ethylenediaminetetraacetic acid (EDTA) tests were used to confirm the adsorption of reagents onto pure minerals commonly found in the two ores. The results of full-scale UG2 concentrator on/off copper sulphate tests are also presented. The UG2 ore showed a substantial decrease in froth stability in the order of reagent addition: no reagents > copper > xanthate > copper + xanthate, while Merensky ore showed a slight decrease. It was shown through zeta potential measurements that copper species were to be found on plagioclase, chromite, talc and pyrrhotite surfaces and through EDTA extraction that this copper was in the form of almost equal amounts of Cu(OH)₂ and chemically reacted copper ions on the Merensky and UG2 ore surfaces. In certain cases, the presence of copper sulphate and xanthate substantially increased the recovery, and therefore the implied hydrophobicity, of pure minerals in a frothless microflotation device. It was, therefore, proposed that increases in hydrophobicity beyond an optimum contact angle for froth stability, were the cause of instabilities in the froth phase and these were found to impact grade and recovery in a full-scale concentrator. Differences in the extent of froth phase effects between the different ores can be attributed to differences in mineralogy.     

Bulk flotation of auriferous pyrite and arsenopyrite by using tertiary dodecyl mercaptan as collector in weak alkaline pulp

Laboratory and industrial scale experiments were conducted to investigate the effect of tertiary dodecyl mercaptan (TDM) as a collector on the flotation of auriferous pyrite and arsenopyrite. The optimum recovery of gold associated with auriferous sulphides was obtained by adding a mixture of TDM and sodium butyl xanthate, together with only a little CuSO₄ as an activator in a weak alkaline pulp adjusted by NaOH. A two-month industrial trial at the Liumei plant in Guangxi, China showed that an average gold recovery of 90.8% into a concentrate assaying 81.1 g/t Au from a feed assaying 2.9 g/t Au could be achieved at pH 8–8.5 using TDM as a collector.     

A modification in the flotation process of a calcareous–siliceous phosphorite that might improve the process economics

The amenability of a low-grade Egyptian phosphorite to flotation for separation of both calcareous and siliceous gangue minerals by just pH control was investigated. The ore, assaying 19.39% P₂O₅, 16.1% L.O.I. and 12.41% A.I. is mainly composed of francolite and hydroxy apatite minerals consolidated into three different phosphatic varieties according to texture and origin, i.e. coarse phospho-chem, sharp-edged phospho-clast and fine cementing phospho-mud. This was endorsed by microscopic investigation of thin sections. X-ray diffraction analysis of the ore sample showed that the main gangue minerals are calcite and quartz with minor dolomite and some gypsum.Anionic flotation of calcite, under pH4.5, was successfully conducted on the −0.25 + 0.074 mm phospho-chem fraction without any use of phosphate depressants. This was followed by direct flotation of phosphate after raising the pH to 9. Mechanical cleaning of the phospho-concentrate was carried out, without any addition of the collector to get rid of the entrained silica. About 3 kg/t of oleic acid was required for the whole process which was added step-wise 0.5 kg/t each except for the first step which was 1.0 kg/t to activate the flotation pulp. Phospho-concentrate assaying 30.54% P₂O₅, 8.7% L.O.I. and 5.76% A.I. with a P₂O₅ recovery of 64.34% was finally obtained without the use of expensive depressants, e.g. phosphoric acid or sodium silicate.A trial to explain the results in view of others’ findings and in terms of the ore mineralogical characteristics was shown.     

Froth flotation of sphalerite: Collector concentration,gas dispersion and particle size effects

This experimental work on sphalerite flotation investigated the effect on flotation performance of three particle size fractions, namely, coarse (d₈₀ = 100 μm), medium (d₈₀ = 39 μm) and fine (d₈₀ = 15 μm), bubble size distribution, superficial air velocity, and collector dosage. Bubble size distributions were characterized with the image analysis technique. The two-phase (liquid–gas) centrifugal pump and frother addition (MIBC, 5–30 ppm) allowed generating bubble diameters between 150 and 1050 μm, and air holdup ranging from 0.2% and 1.3%. Main results showed that each particle-size distribution required an optimal bubble-size profile, and that sphalerite recovery proceeded from mechanisms involving true flotation (when J_g = 0.04 cm/s and 1.9 × 10⁻⁴ M SIPX). However, cluster-flotation occurs at high collector dosage (when J_g = 0.04 cm/s and d₃₂ between 285 and 1030 μm), and requiring further investigation.     

Selective flotation of cassiterite with benzohydroxamic acid

The flotation of cassiterite mineral from gangue with a collector benzohydroxamic acid (BHA), and the interactions between the BHA and cassiterite have been investigated. It is shown through microflotation that the BHA is able to flot cassiterite very well, calcite quite limitedly, and quartz not at all, so the selective separation of cassiterite–quartz mixture was readily achieved; while for the efficient separation of cassiterite–calcite mixture containing 48.94% SnO₂, sodium hexametaphosphate (SHMP) was needed as a depressant for the gangue, and under the condition of the BHA 100 mg L⁻¹, SHMP 3.5 mg L⁻¹, a cassiterite concentrate with the grade of 85.50% SnO₂ was obtained with the recovery of SnO₂ 95.5%. Batch flotation further demonstrated that for an industrial tin slime, which contained 0.42% Sn, 13.65% SiO₂, 24.14% CaO, 16.60% MgO, 4.50% Al₂O₃ and 6.58% Fe, the tin recovery of 84.5% after one separation was reached with the concentrate grade of 1.84% Sn under the condition of the BHA 178 mg L⁻¹, SHMP 27 mg L⁻¹. In terms of zeta potential and infrared spectra studies the main interactions between the collector BHA and the mineral cassiterite in a flotation system are chemisorption with the formation of Sn–BHA compounds rather than electrostatic attractions between them.