Process mineralogy as a predictive tool for flowsheet design to advance the Kamoa project

The Kamoa resource, located in the Democratic Republic of the Congo, contains an array of copper sulphide minerals which are present as small grains, averaging 10–27 μm. An initial flowsheet was developed in 2011/12 for the prefeasibility study that was robust enough to handle flotation of all the copper sulphide minerals. Copper recoveries of the flowsheet were 85.4% for the hypogene ore and 83.4% for the supergene ore. Further work on the flowsheet required reduction of the SiO₂ grade of the concentrate, which at 19.1% negatively affected the downstream smelter processing, and also required improvement to copper grades and recoveries given the high grade of the ore. When new sample material became available as part of the Phase 6 drilling program, a fundamental reassessment of the ore and its flotation behaviour was conducted. Although mineralogical characterisation of the ore and liberation of the sulphides was quantified in previous phases of work, there was little understanding of the kinetics of each of the copper sulphide minerals and how they performed in the flowsheet. Comprehensive flotation kinetic tests at various primary grind sizes were performed. The corresponding timed concentrates of the three best performing grinds were characterised by QEMSCAN on a size-by-size basis to fully understand the flotation kinetics and liberation characteristics of the various copper sulphides. A simple and practical recovery model using minerals, particle size and liberation and association was developed from these data, and various flowsheet configurations were simulated. These simulations led to some robust process implications completely rearranging the flowsheet from the previous iteration into a more simple and economic configuration with better performance. The modelled data was confirmed with practically achieved data, extending the use of process mineralogy as a valid, predictive tool in process design. Additionally, the simulations using mineralogical, reduced empirical flotation testing needed to develop the new flowsheet.     

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”.     

Selective sulphidation and flotation of nickel from a nickeliferous laterite ore

The sulphidation of a nickeliferous lateritic ore was studied at temperatures between 450 and 1100 °C and for sulphur additions of 25–1000 kg of sulphur per tonne of ore. The experiments demonstrated that the nickel could be selectively sulphidized to form a nickel–iron sulphide. It was found that both the grade and the sulphidation degree largely depended upon the temperature and the sulphur additions, with temperatures above 550 °C exhibiting the highest nickel sulphidation degrees and grades. A DTA/TGA with mass spectrometer was used to further elucidate the nature of the phase transformations that occurred upon heating of the ore in the presence of sulphur.It was found that at low temperatures, the Fe–Ni–S phase was submicron in nature and heating to temperatures between 1050 and 1100 °C allowed for the growth of the particles, due to the increased sulphide mobility associated with the formation of a liquid sulphide matte phase, containing dissolved oxygen. Flotation studies conducted on 60 g samples showed that the sulphides responded to flotation with maximum grades of up to 6–7 wt.% nickel being achieved. Recoveries were approximately 50% on a sulphide basis and it was determined that the low nickel grades were due to the entrainment of magnetite fines.     

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.     

Complex treatment of mining and metallurgical wastes for recovery of base metals

A complex process for the recovery of copper and zinc from mining and metallurgical wastes has been investigated and proposed. It includes sulfuric acid leaching of old pyrite flotation tailings to produce ferric containing leach solution; followed by ferric leaching of copper converter slag flotation tailings with the leach solution. A sample of old pyrite flotation tailings from the concentrator containing 0.36% of copper and 0.23% of zinc was leached with 10% sulfuric acid in the column. Recovery of copper and zinc reached 47.1% and 47.2%, respectively. The pregnant leach solutions contained 15.9 g/L of ferric iron. The subsequent ferric leaching of copper converter slag flotation tailings containing 0.53% copper and 2.77% zinc with the pregnant leach solution was conducted. The effects of various process parameters on the leaching dynamics of metals under batch conditions were investigated. Under the best conditions (temperature 70 °C, pulp density 30%, ferric iron concentration 15.9 g/L, initial pH of the pulp 0) the recovery of copper and zinc reached 79.6% and 43.7%, respectively. It was concluded that acid leaching of base metals from old pyrite flotation tailings with pregnant leach solution for the ferric leaching of copper converter slag flotation tailings is a prospective and promising technique for the complex treatment of mining and metallurgical wastes.     

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.     

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.     

Factors affecting the floto-elutriation process efficiency of a copper sulfide mineral

Coarse mineral particles exhibit poor conventional flotation efficiency because of many factors, including the low carrying capacity of bubbles, bubble/particle adhesion problems due to cell turbulence, and low degrees of liberation (low hydrophobicity). Many attempts to improve the recovery of coarse fractions have been explored, such as floto-elutriation operating at a high solid content while dispersed in a fluidized (or expanded) bed formed with a continuous injection of compressed air and an uprising water flow. This work analyzed the comparative performances of floto-elutriation (FE) and conventional flotation (CF) on a classified copper sulfide mineral feed as an example of a difficult-to-liberate low-grade ore. Contrary to expectations, CF and FE (Hydrofloat) displayed similar particle recovery rates with feed size distributions for P80s of 130, 240 and 280 μm. However, metallurgical recoveries from classified fractions of −297+210 μm were 25% higher in FE than in CF and as expected, coarse (+297 μm) particles were not recovered in the CF, but in the FE. The recovery of fine fractions in the FE process was due to high hydraulic entrainment and surprisingly the recovery of intermediate and liberated fractions (+74−149 μm) was very low, due to its low air hold-up. However, the enhancement of the holdup in FE increased the recovery of these mid-sized fractions. Because of the hydraulic carryover caused by the bubbles and water elutriation, the metallurgical grades obtained in all cases were very low compared to conventional bench flotation. It is believed that this FE equipment works better with coarse, narrowly classified particles and high-grade feeds and that performance decreases for low-grade ores requiring high liberation. Certain features of these findings are visualized.     

Froth flotation of a Merensky Reef platinum bearing ore using mixtures of SIBX with a dithiophosphate and a dithiocarbamate

The Bushveld Igneous Complex (BIC) in northern South Africa has the largest deposit of platinum group elements (PGEs) in the world. In trace amounts, these are closely associated with base metal sulphides (BMS). Froth flotation is used as a bulk sulphide recovery to beneficiate these PGE ores. To maximise the recovery of the PGEs it is required to improve the recovery of the BMS. The chemical additives used largely determines the performance of the froth flotation process. Consequently, mixtures of collectors were used in batch froth flotation tests in an attempt to improve concentrate grades and recoveries of BMS from a Merensky Reef platinum ore. The mixtures consisted of a xanthate (SIBX) with a dithiophosphate (DTP) or a dithiocarbamate (DTC). Each mixture was tested at mole ratios of 80:20 and 60:40, with the xanthate the major component. An increase in nickel recovery was observed with all mixtures relative to pure SIBX at the expense of concentrate grade. The mixtures of DTC with SIBX increased the cumulative nickel recovery by 11%, while the mixtures with DTP increased it by 10%. Copper recovery increased by 6% with the DTP mixtures. No significant improvements in the copper recoveries and grades were observed with the mixtures of SIBX with DTC compared to pure SIBX.     

Hydrothermal modification to improve the floatability of ZnS crystals

Crystalline structure and surface properties significantly affect the floatability of metal sulphides. In this study, a novel methodology to modify zinc sulphide (ZnS) crystals was proposed to improve the floatability of the crystals. Initially, ZnS crystals, synthesised from zinc hydroxide (Zn(OH)₂) and sulphur (S) under hydrothermal conditions, were used to assess the floatability. X-ray diffraction (XRD) and transmission electron microscopy (TEM) were employed to analyse the crystalline structure and surface properties of the sulphides. Conventional flotation tests were performed to evaluate the floatability. The effects of mineraliser (KOH) concentration, precursor (Zn(OH)₂) concentration, hydrothermal temperature and holding time on the floatability of the ZnS crystals were investigated. The optimal flotation recovery of ZnS (82.53%) was obtained with a KOH concentration of 5 mol/L, a Zn(OH)₂ concentration of 10%, a holding time of 4 h and a hydrothermal temperature of 260 °C. Then, sludge containing fine and amorphous zinc compounds, which was generated during the disposal of metallurgical waste water, was employed to test the recovery of valuable metals using modified hydrothermal sulphidation. The results show that the recovery of Zn in the sludge can reach 66.3% under the optimal conditions.     

Copper leaching from chalcopyrite concentrate in Cu(II)/Fe(III) chloride system

This study was conducted to develop a novel process for copper recovery from chalcopyrite by chloride leaching, simultaneous cuprous oxidation and cupric solvent extraction to transfer copper to a conventional sulfate electrowinning circuit, and hematite precipitation to reject iron. Copper leaching from chalcopyrite concentrate in ferric and cupric chloride system was investigated using a two-stage countercurrent leach circuit under a nitrogen atmosphere at 97 °C to minimize the concentrations of cupric and ferric ions in pregnant leach solution for subsequent copper solvent extraction while maintaining a maximum copper extraction. A high calcium chloride concentration (110–165 g/L) was used to maintain a high cuprous solubility and enhance copper leaching. With 3–4 h of leaching time for each stage, the copper extraction reached 99% or higher while that of iron was around 90%. With decreasing concentrate particle size from p80 of 26 to 15 μm, the copper extraction increased by about 0.2% while the iron extraction increased by about 2.0%. The concentration of Cu(II) + Fe(III) in the pregnant leach solution was able to be reduced to 0.04 M. When the cupric concentration fell below the above limiting value, the elemental sulfur present was reduced by cuprous ions to form copper sulfide, eventually stopping the leaching of copper. Under this condition, only iron was leached. A very small amount of sulfur (1.2–1.4%) was oxidized to sulfate, resulting in an increase from 3 to 9 g/L in HCl concentration. The extractions of trace metals (Cr, Pb, Ni, Ag and Zn) were 96–100%.     

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.     

Flotation of mixed copper oxide and sulphide minerals with xanthate and hydroxamate collectors

Sherwood Copper’s Minto Mine processes a high grade copper–gold deposit in Yukon, Canada. The ore mined is from a primary copper sulphide deposit with separate additional deposits of copper oxides. In conjunction with Ausmelt Chemicals, Minto is currently investigating options to recover copper oxide and sulphide minerals using flotation by blending their primary sulphide ore with oxide ores. The blend used in this laboratory scale investigation was 70% sulphide ore and 30% oxide ore on a weight basis. The copper sulphides present in the blend were bornite and chalcopyrite, while the oxides were malachite and minor azurite.From previous flotation investigations of mixed copper oxide and sulphide minerals using xanthate and hydroxamate collectors it was hard to distinguish the impact of the alkyl hydroxamate collector on sulphide recovery as the sulphide and oxide minerals occurred naturally together. In the case of the Minto operation the copper oxide and sulphide minerals occur in separate ore deposits and can be treated separately or blended together. This investigation has shown that using n-octyl hydroxamates (AM28 made by Ausmelt Limited) in conjunction with traditional sulphide collectors can successfully simultaneously recover copper sulphides and oxides by flotation from blended ore minerals. The copper sulphide recovery did not decrease when processing the blended ore compared to treating the sulphide ore independently. At a blend of 70% sulphide ore and 30% oxide ore, the rougher scavenger copper recovery was as high as 95.5%. The copper recovery from the blended ore using a mixture of collectors was shown to be superior to the recovery obtained using only xanthate after controlled potential sulphidisation.     

Copper–gold ore processing with ion exchange and SART technology

Anglo Asian Mining has developed a 50,000 oz Au/yr open pit gold mine at Gedabek in Western Azerbaijan. The deposit at Gedabek is a copper–gold porphyry, comprising both oxide and sulphide ore mineralisation, which is being mined at the rate of about 1 million tons of ore per year. Ore processing is by conventional cyanide heap leaching, which produces a pregnant leach solution (PLS) containing 1–2 ppm of gold, together with 1000 ppm or more of copper. The PLS is treated by column ion exchange, using Dow’s gold-selective MINIX resin. Loaded resin is stripped with an acidic thiourea solution, from which gold and silver are electrowon on to stainless steel mesh cathodes. Copper concentrations in the leach solutions are controlled by passing part of the PLS flow through a SART process, where the acronym stands for “Sulphidisation, Acidification, Recycling and Thickening”. The product from the SART process is a copper/silver sulphide precipitate, which is thickened, filtered and dried and then sold for copper smelting.     

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.     

Chloride assisted leaching of chalcocite by oxygenated sulphuric acid via Cu(II)–OH–Cl

The beneficial effect of the addition of sodium chloride upon the leaching kinetics of complex iron–nickel–copper sulphides at elevated temperatures and oxygen pressures has been widely reported since the late 1970s, but the role of chloride is still being investigated or debated. Previous researchers have considered chloride as: (i) a complexing agent for cuprous ions; (ii) a surfactant that disperses the molten sulphur and thus removes passivation of the mineral surface by elemental sulphur during pressure leaching; and (iii) a reagent which increases the surface area and the porosity of the insoluble product layer on the surface. A proper understanding of the role of chloride based on the leaching of individual sulphides of known composition in the absence of host minerals at low pulp densities would be useful for the development of chloride assisted sulphate leaching processes for complex sulphide ores, concentrates, and mattes. In the present study evidence for the formation of basic salts of Cu(II) and Fe(III) during leaching are presented. The published rate data are analysed for the leaching of copper from mono-sized chalcocite particles in oxygenated sulphuric acid solutions maintained at 85 °C, a temperature lower than the melting point of sulphur. The initial leaching follows a shrinking particle (sphere) model, and the apparent rate constants are first order with respect to the concentration of dissolved oxygen and chloride. The intrinsic rate constant for the surface reaction (0.2 m s⁻¹) is two orders of magnitude larger than the calculated mass transfer coefficient of oxygen (3 × 10⁻³ m s⁻¹). The proposed reaction mechanism considers the formation of an interim Cu(II)(OH)Cl⁰ species which facilitates the leaching process.     

Iron redox-equilibria and sulphide capacity of PGM melter-type slags

New measurements have been made on the ferric to ferrous ratio as well as the sulphide capacity for platinum group metals (PGM) melter-type slags. In South Africa, these slags are produced from the smelting of low-grade copper–nickel sulphide ores, Nell [Nell, J., 2004. Melting of platinum group metal concentrates in South Africa. The South African institute of Mining and Metallurgy 104 (7), 423–428]. The typical mass compositions are 5–10% Al₂O₃, 2–15% CaO, 5–30% FeO_x, 15–25% MgO and 40–60% SiO₂ with a molar basicity defined as (CaO + MgO)/SiO₂ of 0.6–1. The industrial furnaces operate at temperatures ranging from 1450 to 1600 °C under fairly reducing conditions (typically a pO₂ close to 10⁻⁸ atm at 1500 °C). The gas–slag equilibrium was studied by subjecting a synthetic slag to controlled atmospheres in a vertical tube-furnace using Ar–CO–CO₂ (–SO₂) gas mixtures. The ratio of ferric to ferrous was determined at 1450 °C for oxygen activities, defined as pCO₂/pCO, ranging from 0.11 to 1.75 by analysing the quenched slags using the standard titration and XRF techniques. The measured Fe³⁺/Fe²⁺ ratio increased from 0.029 to 0.110 with the increasing oxygen activity. Slight non-ideal iron redox behaviour was observed, as has been reported for low alumina and low iron-containing slags. The present results are in good agreement with the trends found in the literature for similar multi-component slag systems (mostly iron bath smelting slags). Sulphide capacity was measured at partial pressures of oxygen and sulphur of approximately 10⁻⁹ and 10⁻³ atm respectively, with total-iron contents of 8.2 and 15.6 wt%, and temperature ranging from 1450 to 1525 °C. The present sulphide capacity data ranged from 10^−4.43 to 10^−3.71. The expected increase in sulphide capacity with increasing temperature was observed, and at a given temperature, the sulphide capacity increased with an increase in iron oxide content.     

Biosorptive flotation of copper ions from dilute solution using BSA-coated bubbles

Copper adsorption was carried out using the novel material known as air-filled emulsion (AFE). AFE is a stable colloidal system containing microscopic protein-coated bubbles (<10 μm) dispersed through an aqueous solution, resulting in an increased specific surface area and contact time between extractant and metal ions. Bovine serum albumen (BSA) generated emulsion concentration had a significant impact on copper removal, with maximum metal uptake obtained at 2.5 g/l of BSA-coated bubbles. It was shown that copper sorption was rapid over the first 10 min, and equilibrium conditions were reached within 40 min. Separation of the copper-loaded microcells from the aqueous solution was also investigated. Micro-flotation was employed to remove the microbubbles by means of attachment to the surface of larger air bubbles. In absence of a cationic surfactant, approximately 0.5% copper recovery was obtained at pH ranging from 5 to 8 due to the lack of hydrophobic groups on the surface of Cu-loaded BSA emulsions. Due to the fine sizes of the emulsion bubbles (<10 μm) a cationic flocculant was used to induce coagulation of the bubbles leading to easier phase separation. A combination of collector and flocculant at a concentration of 3 × 10⁻⁴ M and 0.025 g/l, respectively, led to an increase in copper recovery to nearly 35% at pH 7.     

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.     

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.