The critical importance of the grinding environment on fine particle recovery in flotation

This paper examines published and new experimental evidence on the effect of the grinding environment on fine (?10 μm) value mineral recovery in flotation. Reasons for increases in fine value mineral recovery from ores with fully electrochemically inert grinding media are discussed in relation to reduced surface contamination by iron hydroxide emanating from the grinding media. The application of stirred milling technology, which allows the use of fully inert grinding media, to primary grinding applications may lead to increased fine value mineral recovery in flotation rougher applications. It is suggested that the effect of the grinding media, which is important for fine particles and progressively becomes more important as the grind size becomes finer, is principally due to the abrasion mechanism of the minerals with the grinding media in the production of fine particles. Opportunities for research and industry application are discussed.     

Comparison of energy efficiency between ball mills and stirred mills in coarse grinding

Stirred mills are primarily used for fine and ultra-fine grinding. They dominate these grinding applications because greater stress intensity can be delivered in stirred mills and they can achieve better energy efficiency than ball mills in fine and ultra-fine grinding. Investigations were conducted on whether the greater performance of stirred mills over ball mills in fine grinding can be extended to coarse grinding applications. Four different laboratory ball mills and stirred mills have been tested to grind seven ore samples with feed sizes ranging from 3.35 mm to 150 μm. A case study on full scale operations of a 2.6 MW IsaMill replacing the existing 4 MW regrind ball mill at Kumtor Gold Mine in Kyrgyzstan is also included. This paper summarizes the major findings from these investigations.     

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.     

Effects of grinding on the preg-robbing potential of quartz in an acidic chloride medium

The effect of high-energy milling on the surface properties of quartz is examined with regard to its preg-robbing behavior towards gold. A standard ring mill is used to process dry quartz samples, and the changes in the morphology of the particles, structural deformations and surface chemistry are investigated to explain the increased preg-robbing ability of quartz in acidic chloride solutions. The transition from fine grinding to mechanochemical activation of quartz can be observed from changes in the morphology of the particles, as well as the types of structural deformations. The transition occurs between 1 and 5 min of grinding in the mill used, corresponding to particle sizes around 0.55 μm. Structural studies differentiate two stages of fine grinding: particle breakage with limited structural disruption, and structural disturbance by mechanochemical alteration, which occurs after particles reach their grinding limits. Quartz keeps its structural order to some degree even after 30 min of aggressive grinding. The surface chemistry of ground quartz demonstrates generation of point defects including low valence silicon and non-bridging oxygen centers. These defect sites play an important role in the surface reactivity of the quartz, and influence the extent of gold loss during preg-robbing.     

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 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.     

Optimization of operating parameters for coarse sphalerite flotation in the HydroFloat fluidised-bed separator

Increasing the upper size limit of coarse particle flotation has been a long-standing challenge in the minerals processing industry. The HydroFloat separator, an air-assisted fluidised-bed separator, has been used in this study to float 250–1180 μm sphalerite particles in batch flotation tests and compared to results achieved utilizing a laboratory-scale conventional Denver cell. The quiescent environment within the HydroFloat cell significantly reduces the turbulent energy dissipation within the collection zone, hence decreasing the detachment of particles from bubbles during flotation. Three operating parameters including bed-level, superficial water and gas rates have been studied, and their effect on the flotation of coarse sphalerite particles is reported. It is shown that coarse sphalerite recovery increases with increasing bed-level, superficial water and gas flow rates. However, there are thresholds for each operating parameter above which recovery starts to decrease. A comparison of recovery with a conventional Denver flotation cell indicates that the HydroFloat separator vastly outperforms the conventional flotation machine for the very coarse particles (+425 μm), and this is mainly attributable to the absence of turbulence and the minimization of a froth zone, both of which are detrimental to coarse particle flotation.     

Effect of impeller rotational speed on the size dependent flotation rate of galena in full scale plant cells

The first three rougher cells in the lead circuit of the Elura concentrator (formerly Pasminco Australia Limited) were selected as the plant cells for investigation. Metallurgical surveys were performed and various hydrodynamic measurements taken, allowing the galena flotation rate constant and the bubble surface area flux (S_b) in these cells to be calculated over a wide range of gas flow rates, and at two impeller rotational speeds. It was determined that altering the impeller rotational speed did not significantly change the rate constant dependency on S_b when flotation was considered on an unsized basis.The analysis was further extended to examine the same cells parameters on a size-by-size basis. The results obtained have been used to identify differences in the flotation behaviour of the various particle size fractions, independently of surface hydrophobicity. It is shown that the physical conditions for effective flotation of fine (<9 μm) and coarse (>53 μm) particle size fractions differ substantially, suggesting that a specific hydrodynamic environment will favour a high flotation rate for fine galena, which may be detrimental to the recovery of coarse galena, and vice versa. These observations are in accord with metallurgical practice that suggest that it is difficult to improve fine particle flotation without also compromising coarse particle stability efficiency simply by modifying the cell hydrodynamics alone. A fundamental flotation model was applied to quantify differences in the flotation rate of the various particle size fractions with impeller rotational speed.     

Dual energy rapid scan radiography for geometallurgy evaluation and isolation of trace mineral particles

Procedures are described for the evaluation and isolation of trace mineral particles from sample populations at the ppm level by dual energy rapid scan radiography using an X-ray microtomography instrument (XMT). After calibration of the XMT, the samples are split into narrow size fractions, each size fraction distributed/assembled on projection plates, and then the plates placed in the sample holder of the XMT for irradiation at two energy levels (dual energy analysis). In this way, for example, more than 200,000 particles of 250 × 150 μm in size can be interrogated in less than 1 h and the composition of particles containing high density mineral phases estimated. These trace particles can be isolated for 3D analysis by high resolution X-ray microtomography and/or selected for surface characterization using XPS, TOF/SIMS, or other analytical instruments. Rapid scan radiography can be used for the examination of drill core samples, tailings samples, or any other particulate sample containing trace mineral particles.     

Effects of operating parameters on the efficiency of dry stirred milling

Stirred media milling is an industrially accepted efficient grinding method for fine and coarse particles. The stirred mills can be operated both in vertical and horizontal configurations and the selection depending on the process variables. Successful operation of horizontal stirred milling (i.e. IsaMill) in wet applications encouraged the studies in dry applications. In this study, series of dry grinding tests were performed in a prototype horizontal stirred mill (42 L) to investigate the effects of operating parameters such as stirrer speed, feed rate, media filling and ball size on grinding considering the degree of size reduction and the energy consumption. The test results have shown that the stirrer speed, the media size and the media filling are directly proportional and the feed rate is inversely proportional with the specific energy consumption. Besides, energy savings up to 27% were achieved by adjusting the milling conditions properly (suitable media size) and the size reduction values (F₅₀/P₅₀) were between 1.05 and 2.42.     

Ultrafine coal classification using 150 mm gMax cyclone circuits

A two-stage classification circuit using 150 mm diameter gMax cyclones was installed and evaluated in a coal preparation plant in an effort to achieve a clean coal product without the use of froth flotation. Particle size separations of around 37 μm were achieved while limiting ultrafine bypass to less than 10% in the circuit underflow stream. As a result, approximately 81% of the ash-bearing material in the circuit feed was rejected to the circuit overflow stream. The feed ash content was reduced from around 50% to values in the range of 22–30% in the circuit underflow stream with a mass recovery of about 30%. Further reductions in the coarse product ash content were limited due to the particle density effect and the remaining presence of a significant quantity of high-ash slime material in the coarse product. The typical D₅₀ for the coal particles was 40 μm while the estimated value for mineral matter was 17 μm. Based on the findings of the study, the use of classification to recover a low-ash, coarse fraction in the feed of a fine coal circuit is limited by the density effect regardless of the ability to eliminate ultrafine bypass.     

Efficiency of grinding stabilizers in cement clinker processing

Stabilizers are commonly used in the production of very finely ground particulate materials. Doping with such additives is an absolutely essential process step in the production of ultra-fine and nano-fine solids; in all cases, the additives work as inhibitors that prevent the re-agglomeration of fine particles. Such additives, which are well known in both dry grinding and wet grinding processes, help to decrease the enormous energy consumption of cement production. This paper describes the essential features of stabilizers used in Portland clinker grinding processes. The results of milling experiments employing twelve commercially used and experimental additives are described. Their effectiveness in producing ultrafine (0–30 μm) particles varied widely. Novel acrylate-based grinding stabilizers showed considerable promise for the production of very fine cement clinkers.     

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.     

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.     

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.     

Improving the recovery of coarse coal particles in a Jameson cell

It has been observed in several Jameson cell installation where the source for flotation feed is deslime screens, that the recovery of coal particles greater than 0.5 mm is not as great as that of finer material. Consequently, a research project was undertaken at a CHPP in the Bowen Basin Queensland to assess the possibility of increasing the recovery of coarser particles (+0.5 mm) within the downcomer of the Jameson cell. The effect of decreasing turbulence and agitation in a commercial-scale downcomer was investigated to assess the effect on the recovery of both coarse and fine coal particles.This paper details the findings of the test work, summarising the results relating to differences in the operating parameters within the downcomer.     

Liberated 0–10 μm particles from sulphide ores,their production and separation—Recent developments and future needs

The general types of industrial behaviour of liberated valuable sulphide minerals in the 0–10 μm size fraction are described and tests to explore on-site at a concentrator the general nature of any deficient behaviour are discussed. In addition, the general types of behaviour of the gangue minerals in this size fraction are also described in terms of their mechanisms for recovery.These behaviours are also discussed in terms of the type of grinding device and grinding environment in which they were produced and the target grinding product size. Major recent improvements in the industrial capabilities of grinding devices are discussed.Present difficulties in characterization of particles in the 0–10 μm range are described and desirable improvements for the future are outlined. Further, desirable improvements for industrial processing of valuable minerals in this size range are discussed along with some remedial actions.     

Value adding limestone to filler grade through an ultra-fine grinding process in jet mill for use in plastic industries

In this study, ultra-fine grinding of limestone was carried out in jet mill using four levels of classifier rotational speed and grinding pressure. The holdup amount was determined during the grinding process, while the feed rate was kept constant at 8 kg/h. The ground product was characterized for its particle size and shape. In addition, the mechanochemical effect on the ground product was characterized through XRD. The particles size of the ground product ranged from 2.21 μm to 7.29 μm, demonstrating various particle shapes such as cubical, angular, and elongated. The degree of crystallinity of the ground product ranged from 54.5% to 93.7%. Afterwards, the ground product was incorporated as filler in polypropylene (PP), and its performance was characterized for mechanical properties. After conducting the test work, we find that the PP filled with ground limestone exhibited excellent thermal and mechanical properties. The composite flexural modulus, impact strength, tensile strength, and elongation at break were 2.1 GPa, 42 kJ/m², 22.75 MPa, and 21%, respectively, when loaded up to 20%. It likewise exhibited CTE value of 57.2 ppm/°C.     

Size-by-size analysis of dry gravity separation using a 3-in. Knelson Concentrator

Centrifugal concentrators have been widely used to recover fine gold over the past few decades, the main drawback being the large volume of water required during the operation. As water is an increasingly important commodity, investigating dry processing to reduce water usage is of great importance. This work investigated using a laboratory scale Knelson Concentrator with a dry feed, and air being used to replace water as the fluidising medium and size-by-size testing of material. A synthetic ore comprised of tungsten and quartz was used to mimic a gold ore. Response surface methodology and central composite design were used for modelling and optimisation with the experimental variables being the motor power (MP) (%), air fluidizing pressure (AFP) (psi) and solid feed rate (SFR) (g/min). The observed results of the three size classes indicated that amongst the three variables; the motor power had the greatest impact when comparing to the other two variables. Different optimum conditions had been found for −425 + 150 μm, −150 + 53 μm and −53 μm size fractions and these are 30% MP, 200 g/min SFR and 10 psi AFP, 50% MP, 160 g/min SFR and 11 psi AFP, 65% MP, 200 g/min SFR and 11 psi AFP, respectively.     

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”.