Effects of frother type and froth height on the flotation behaviour of chromite in UG2 ore

About 70% of the UG2 reef consists of the gangue mineral chromite (FeO · Cr₂O₃). In the processing of UG2 ore by flotation for the recovery of platinum group elements (PGEs) the presence of chromite in the concentrates can cause serious downstream processing problems and a grade of less than 3% Cr₂O₃ is sought. This constrains operating procedures and compromises optimum recovery of the PGEs.In this study, the influence of the froth phase on the recovery of chromite was investigated by changing both frother type and dosage and froth height in batch scale flotation tests. The results obtained showed that it was possible to obtain concentrates with less than 3% Cr₂O₃ content by increasing the froth height, allowing for better drainage of both entrained gangue particles and coarse particles with low hydrophobicity. At a 3 cm froth height, very low water and mass recovery were obtained and thus low entrainment. Nevertheless a small amount of chromite particles coarser than 45 μm was persistently recovered which may be attributed to the true flotation of these particles.The mechanism of chromite recovery was discussed on the basis of the difference in the appearance of the froth structure and water recovery.     

Bank profiling and separation efficiency

Various authors have discussed methods of optimising a bank of flotation cells. In this paper, JKSimFloat is used to investigate the effect of recovery profiling and mass pull profiling (i.e., mass distribution to cells in a bank) on the separation efficiency between floatable minerals and against entrained gangue.In the case of two floatable minerals, a balanced recovery profile was found to be optimal: supporting and extending previous analysis. In the case of separation of a floatable mineral from entrained gangue, the entrainment model that links water overflow rate to solids overflow rate was employed. When the value of b in the entrainment model is greater than one, a balanced mass pull profile was found to be optimum. The evidence for b > 1 is briefly reviewed; no example has been found where b < 1. Most of the profiles were controlled in the software by altering the bubble surface area flux distribution; a sensitivity analysis was performed using other variables.Recovery profiling was tested as part of a bank optimisation campaign at a talc operation in Timmins, Canada. Using air and frother as manipulated variables, it was found that as the rougher bank was moved toward a balanced profile the final plant product showed improvement in grade and yield.     

The effect of the design of a secondary grinding circuit on platinum flotation from a UG-2 ore

Platinum concentrator plants experience significant losses in their overall Platinum Group Elements (PGE) recoveries due to the inefficiencies of their secondary grinding circuits. This study involves an investigation of selective grinding of the platinum-bearing silicate particles present in UG-2 platinum ores found in the Bushveld Igneous Complex (BIC).Batch-scale laboratory test work was done to investigate the effect of a secondary milling circuit configuration, using a hydrocyclone underflow sample from a UG-2 concentrator plant as feed material. The envisaged secondary milling circuit consists of a conventional hydrocyclone to de-slime the feed followed by density separation with a spiral concentrator to separate the ore into lights (silicates-rich) and heavies (chromite-rich) fractions, followed by separate milling of the two fractions in parallel ball mills, and combined rougher flotation. A full-scale spiral was run in batch mode, followed by separate milling of samples in a 200 mm diameter mill and combined flotation in a 4.2 l cell. The milling energy inputs were re-distributed between the lights and heavies mills to determine the effect on the platinum mineral rougher flotation recovery and the Cr entrainment.The most promising results were found with 88% of the energy input to the lights mill and 12% to the heavies mill. The results indicated that under batch conditions, the secondary rougher flotation recovery (69% 4E) was similar to the conventional mill-float circuit (70%) however the Cr entrainment was significantly reduced by approximately 40% (2.3–1.4% Cr).This test work has confirmed the benefit of separate milling in the secondary milling circuit for a UG-2 ore. Spiral concentrators have shown potential as an effective density separating device to produce a silicate-rich and chromite-rich fraction for milling; further test work will be conducted to confirm its viability on an industrial scale.     

Post-regrind selective depression of pyrite in pyritic copper–gold flotation using aeration and diethylenetriamine

The present study investigates the effect of aeration and diethylenetriamine (DETA) on the selective depression of pyrite in a porphyry copper–gold ore, after regrinding (at grind sizes, d₈₀ = 38 and 8 μm) with respect to Au recovery and grade using oxygen demand tests, flotation, QEMSCAN, X-ray spectroscopy (XPS) and EDTA extraction analysis. It was found that pyrite depression increases after aeration and with decreasing grind size. This was observed to be due to the markedly higher oxygen consumption rate of pyrite at the 8 μm (kla = 0.10 min⁻¹) than at the 38 μm grind size (kla = 0.02 min⁻¹). The addition of DETA improved pyrite depression (9% with aeration only versus 39% with aeration + DETA) at the 38 μm grind size. Gold and copper flotation recovery followed pyrite recovery for the two grind sizes using XD5002 in the presence of air and DETA.The surface analysis (XPS and EDTA extraction) revealed that the significant pyrite depression at the 8 μm grind size was due to increased amount of surface iron oxides, oxy-hydroxides (FeO/OH), sulphate species and increased liberation of mineral phases (QEMSCAN analysis), whilst the poorer pyrite depression at the 38 μm grind size was due to insufficient liberation of mineral phases and the persistence of activating Cu on the pyrite surface. The addition of DETA increased pyrite depression at the coarser grind size due to a significant reduction in Cu(I)S and increased Cu(II)O species, correlating with the flotation results of pyrite under this test condition. Two-stage copper and pyrite flotation, followed by Au cleaning after regrinding to 38 μm grind size, under high pH or aerated condition is proposed as the recommended route to optimise Au flotation.     

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

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

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.     

The role of citric acid in the flotation separation of rare earth from the silicates

The Nechalacho project is the most advanced large heavy rare earth elements (HREE) project outside of China. Open circuit and locked cycle flotation tests along with pilot plant testing of rare earth elements (REE) concentration from the host rocks are accomplished with collectors of alkyl phosphates and the modifier of citric acid. In this study, the function of citric acid in the separation of rare metals against silicates is investigated by a combination of micro-flotation tests and time of flight secondary ion mass spectrometry (ToF-SIMS) surface chemical analysis. It was observed that there was little effect of citric acid on the REE recovery in the micro-flotation tests conditioned with de-ionized water (DIW). To evaluate the flotation response with excess secondary ions in the pulp, micro-flotation tests were performed to look at changes in recovery as a result of adding Al ions and the subsequent presence of citric acid. The results from three micro-flotation tests (DIW, DIW with the addition of 100 mg/L Al and DIW + 100 mg/L Al and 500 g/t citric acid) revealed that the addition of Al ions led to a decrease of REE grade, a lower REE minerals recovery and/or an unexpected promotion of silicates to the concentrate. Citric acid reduced the negative effect generated by the Al ions in the flotation, which was shown by an improvement in REE grade. ToF-SIMS surface analysis of undifferentiated grains from the tests with and without citric acid revealed that grains reporting to the concentrate are doing so in response to collector attachment in combination with having more secondary Al on their surface. Citric acid may partially form aqueous soluble metal–ligand complexes resulting in less Al ions on the grains surface, which were rejected to the tailings. Citric acid also may form chelation competing for adsorption on gangue minerals, resulting in a diminished effectiveness of the activation site.     

Characterisation of coarse composite sphalerite particles with respect to flotation

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

Flocculation/flotation of hematite fines with anionic temperature-responsive polymer acting as a selective flocculant and collector

A novel approach to the recovery of valuable fines is the use of temperature-responsive polymers such as poly (N-isopropyl acrylamide) (PNIPAM). These polymers act as dual-function flocculants and collectors to form hydrophobic aggregates improving particle–bubble collision and attachment. The aim of this study is to investigate the flotation performance of anionic PNIPAM for an iron ore sample containing fines compared to sodium oleate, an industrial collector for hematite. PNIPAM conditioned at room temperature (25 °C, below the lower critical solution temperature (LCST)) and floated at 50 °C (above the LCST) was found to provide improved hematite grade and recoveries for particles above 20 μm, compared to sodium oleate. This was attributed to the increased selectivity and hydrophobicity of PNIPAM. Turbidity testing confirmed the flocculation of fines with PNIPAM, which deslimes the surface of the coarser particles. Below 20 μm, the hematite fines were almost completely recovered with PNIPAM. However, this recovery was not selective, attributed to the entrapment of gangue in the hydrophobic aggregates. Furthermore, conditioning of the polymer above the LCST resulted in significant losses in grade and selectivity, as the polymer self-aggregates hydrophobically and precipitates unselectively onto the closest surface.     

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.     

Flotation of copper sulphides assisted by high intensity conditioning (HIC) and concentrate recirculation

This paper describes the effect of the partial concentrate (rougher floated product) recirculation to rougher flotation feed, here named concentrate recirculation flotation – CRF, at laboratory scale. The main parameters used to evaluate this alternative approach were flotation rate and recovery of fine (“F” 40–13 μm) and ultrafine (“UF” <13 μm) copper sulphide particles. Also, the comparative effect of high intensity conditioning (HIC), as a pre-flotation stage for the rougher flotation, was studied alone or combined with CRF. Results were evaluated through separation parameters, grade-recovery and flotation rates, especially in the fine and ultrafine fractions, a very old problem of processing by flotation. Results showed that the floated concentrate recirculation enhanced the metallurgical recovery, grade and rate flotation of copper sulphides. The best results were obtained with concentrate recirculation flotation combined with high intensity conditioning (CRF–HIC). The kinetics rate values doubled, the Cu recovery increased 17%, the Cu grade increased 3.6% and the flotation rates were 2.4 times faster. These were accompanied by improving 32% the “true” flotation values equivalent to 2.4 times lower the amount of entrained copper particles. These results were explained and proved to proceed by particle aggregation (among others) occurring after HIC, assisted by the recycled floatable particles. This “artificial” increase in valuable mineral grade (by the CR) resulted in higher collision probability between hydrophobic particles acting as “seeds” or “carrier”.     

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.     

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.     

Evaluation of a MicrocelTM sparger in the Red Dog column flotation cells

Column flotation cells have been installed in numerous base metal operations around the world. The majority of these cells utilize conventional air-only spargers to introduce air into the bottom of the column. The recent development of instruments to measure the bubble characteristics in these columns has provided a renewed understanding of column behaviour. These new tools provided insight into why the columns at Red Dog Mine had never performed up to expectations. After efforts to optimise the spargers failed to substantially change the bubble size and air efficiency, alternate sparger systems were investigated. The Metso Minerals CISA Microcel sparger system appeared promising and was selected for a full-scale plant trial. The Microcel sparger system was originally developed for the coal industry at the Virginia Centre for Coal and Minerals Processing.In October 2003, a Microcel was retrofitted into one of two 3.66 m diameter flotation columns in the zinc retreat circuit at Red Dog Mine. The operation of the two different sparger systems in parallel allowed a detailed comparison.This paper discusses the performance of the Microcel based on the data collected during several detailed surveys. Bubble size measurements carried out in the pulp zone using the McGill University bubble viewer showed a significant difference in bubble size. The mean Sauter diameter of the bubbles decreased from 3.4 mm for the jetting-type sparger to 1.9 mm for the Microcel sparger. The overall recovery and the recovery by size fraction for both valuable and gangue minerals were compared. Paired t-tests demonstrated that the Microcel column produced a higher concentrate grade (0.6% zinc absolute) and a higher unit recovery (2.8% zinc absolute) than the existing Canadian Process Technologies Inc (CPT) SlamJet column. These improvements provided a payback period of 1.5 months for the $109,000 investment.     

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.     

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.     

Prediction of the metallurgical performances of a batch flotation system by image analysis and neural networks

It is now generally accepted that froth appearance is a good indicative of the flotation performance. In this paper, the relationship between the process conditions and the froth features as well as the process performance in the batch flotation of a copper sulfide ore is discussed and modeled. Flotation experiments were conducted at a wide range of operating conditions (i.e. gas flow rate, slurry solids%, frother/collector dosage and pH) and the froth features (i.e. bubble size, froth velocity, froth color and froth stability) along with the metallurgical performances (i.e. copper/mass/water recoveries and concentrate grade) were determined for each run. The relationships between the froth characteristics and performance parameters were successfully modeled using the neural networks. The performance of the developed models was evaluated by the correlation coefficient (R) and the root mean square error (RMSE). The results indicated that the copper recovery (RMSE = 2.9; R = 0.9), concentrate grade (RMSE = 1.07; R = 0.92), mass recovery (RMSE = 1.94; R = 0.94) and water recovery (RMSE = 3.07; R = 0.95) can be accurately predicted from the extracted surface froth features, which is of central importance for control purposes.     

Effect of gas rate and impeller speed on bubble size in frother-electrolyte solutions

In a flotation cell, bubble size is a function of both coalescence and breakup phenomena. Two phase tests, conducted in a conventional 5.5 L Denver mechanical flotation cell, studied the effect of impeller speed, gas flow rate and frother concentration on bubble size in various electrolyte-frother solutions. The addition of frother to a synthetic sea salt did reduce the measured bubble size (at certain mechanical conditions); whereas the effect of frother addition to NaCl was too small (when compared to measurement errors) to make significant conclusions. This led to more detailed CCC curves (0–50 ppm MIBC) for NaCl, NaCl + MgCl₂, NaCl + CaSO₄, and NaCl + KCl solutions, at constant electrolyte concentrations, to be conducted. They showed an increase in bubble size with the addition of MIBC. This was attributed to the saturation of frother at the air-water interface, reducing local surface tension gradients that help produce smaller bubbles. This occurrence is typically masked in traditional CCC curves due to the dominance of coalescence effects at low frother concentrations.     

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.