Use of oxidation roasting to control NiO reduction in Ni-bearing limonitic laterite

Use of limonitic laterite as an iron source in conventional ironmaking is restricted due to its gangue composition and small particle size. Even direct reduction cannot effectively produce direct reduced iron (DRI) because NiO would be reduced together with iron oxide to form Fe–Ni. A small amount of Ni (about 2 wt.%) in DRI degrades the physical properties of final steel products. The current study investigated how oxidation roasting of limonitic laterite ores affected NiO reduction, with the goal of producing Ni-free DRI and Ni-bearing slag. Ni-bearing slag can be a good secondary Ni resource. Oxidation roasting made NiO inert under H₂ reduction at 900 °C by forming Ni-olivine. Optimum roasting temperature was proposed by examining phase transformation of limonitic laterite ores during heating and by FactSage calculation of the equilibrium Ni fraction in Ni-bearing phases. Furthermore, the effect of Mg–silicate forming additives on the control of NiO reducibility was clarified to maximize the suppression of NiO reduction. Among various additives such as MgSiO₃, Mg₂SiO₄ and Fe–Ni smelting slag, Ni-free olivine-typed flux was found to be the most effective form of Ni-olivine because Ni–Mg ion exchange between Ni-bearing phase and Ni-free olivine occurs more readily than other Ni-olivine formation schemes. Finally, the mechanism of Ni-olivine formation during roasting was studied using a diffusion couple test. Calculated diffusivity values of Ni in Mg₂SiO₄ indicated that the two major routes of Ni-olivine formation while roasting limonitic laterite ore are (1) Ni partitioning within Mg–Ni silicate before crystallization and (2) Ni diffusion from spinel to Ni free olivine after crystallization.     

Agglomeration and column leaching behaviour of goethitic and saprolitic nickel laterite ores

Processability of complex, low-grade nickel (Ni) laterite ores via heap leaching is very limited due to some intractable geotechnical and hydrological challenges such as poor heap porosity/permeability and structural stability. This work presents some investigations on laboratory batch drum agglomeration and continuous column leaching behaviour of saprolitic (SAP) and goethitic (G) Ni laterite ores as part of the quest for an effective ore pre-treatment process for enhanced heap leaching. As a focus, the effect of ore mineralogy/chemistry on the agglomeration and column leaching behaviour of −2 mm (crushed from −15 mm run-of-mine) G and SAP Ni laterite ores was examined. To produce ∼5–40 mm agglomerates in <15 min, the SAP ore required a higher H₂SO₄ (30 wt.%) binder dosage compared with the G ore, although both ores displayed substantially similar, coalescence-controlled agglomeration mechanism. The resulting G agglomerates were more robust than the SAP ones based upon their compressive strength and acidic solution soak test measurements. However, over 100 days of continuous column leaching, the structural stability of the SAP agglomerate bed was slightly greater than that of G agglomerates, reflecting a lesser slump of the former. The pregnant leach solution analysis revealed greater Ni/Co extraction rates from the SAP than the G agglomerates. Whilst the total mass of acid consumed per ton dry ore processed was greater for the SAP ore, the total kg acid per kg Ni extracted was markedly lower. Incongruent leaching of gangue minerals’ constituent elements (e.g., Fe, Mn, Mg, Al and Si) occurred and contributed significantly to the overall acid consumption. The findings show the relevance of agglomeration and column leaching tests for providing useful information for plant designing and optimization of Ni laterite heap leaching operations.     

Innovative technology for processing saprolitic laterite ores by hydrochloric acid atmospheric pressure leaching

An innovative technology for processing saprolitic laterite ores from the Philippines by hydrochloric acid atmospheric leaching and spray hydrolysis is proposed. The factors that affect the hydrochloric acid atmospheric leaching of the laterite ores and spray hydrolysis of the atmospheric acid leach solution are investigated. Experimental results show that the leaching of Ni, Fe, and Mg is 98.9 wt%, 97.8 wt%, and 80.9 wt%, respectively, under optimal acid leaching conditions. The hydrolysis of Ni and Fe by the atmospheric acid leach solution approaches 100 wt% at the temperature range of 450–500 °C. Characterization results show that a serpentine mineral, nominally Mg₃Si₂O₅(OH)₄, is the major component and goethite, FeO(OH), is the minor one in the laterite ores. Treatment by hydrochloric acid atmospheric leaching breaks the mineral lattices of the laterite ores and makes amorphous silica the primary product in the atmospheric acid leach residue. The grade of Ni in the hydrolyzate increases to 4.55%. The hydrolyzate with high Ni content can be utilized for ferro-nickel production.     

Stirred milling kinetics of siliceous goethitic nickel laterite for selective comminution

The objective of this study is to determine how grinding conditions affect the breakage rate with respect to the sample mass, major elements, and minerals present in siliceous goethitic (SG) nickel laterite. This information is helpful in determining the optimal grinding conditions for selective comminution and nickel upgrade. The kinetics of batch wet grinding of nickel laterites with feed sizes of 2.38–1.68, 1.68–1.18, 1.18–0.85, 0.85–0.6, 0.6–0.42, 0.42–0.3, 0.3–0.21, and 0.21–0.15 mm were determined using a Netzsch LME4 stirred mill under the following conditions: 1000 rpm, 50% charge volume, 150.0 g of solid. The grinding behaviour of the majority of the feed samples was non-first-order due to the fast breakage rate of soft minerals and the low breakage rate of hard minerals in the feed. Therefore, an enrichment of the soft mineral was obtained in the underscreen product by selective grinding. The effect of selective grinding on Ni upgrade was evaluated by looking at grinding time, feed size, and product size. Optimum grinding time with respect to Ni upgrade was 0.25 min for SG nickel laterite samples. Generally, grinding larger particles and/or collecting finer product size yielded better Ni upgrade results. The effect of selective grinding was evaluated by the changes of the major soft and hard minerals for the selected samples. Selective grinding was also examined with respect to the major element weight ratio (e.g. Si/Ni for SG nickel laterite). With respect to Ni upgrade, the best result was achieved from the 1.18–0.85 mm feed on the −400 mesh product after grinding for 0.25 min. The Ni grade increased from 0.73% to 1.30% (upgrade 76.8%), with 14.4% Ni recovery; the Mg grade increased from 1.30% to 3.96% (upgrade 205.6%); the Si grade decreased from 28.7% to 16.2%.     

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.     

Evaluation of the response of valuable and gangue minerals on a recovery,size and liberation basis for a low-grade silver ore

Mineralogical characterisation is typically used to assist in the development of processing strategies for ores. This paper describes the application of mineralogical characterisation techniques in the development of flotation strategies for processing an ore from a low-grade silver deposit that contains a variety of rock types that have undergone hydrothermal alteration where zinc, lead and silver sulphides are typically the primary minerals of economic interest. Comprehensive mineralogical characterisation of the feed, concentrates and tailings from batch flotation tests was undertaken using both the mineral liberation analyser (MLA) and laser ablation inductively-coupled plasma mass spectroscopy (LA-ICP-MS). The results of mineralogical characterisation of the feed (head grade, 116 ppm) indicated that the majority of the silver occurred as solid solution in pyrite which assisted in the development of the flotation strategy used for this ore which resulted in approximately 87% of the total silver being recovered to rougher concentrate at a grade of 485 ppm.     

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 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.     

Atmospheric acid leaching of siliceous goethitic Ni laterite ore: Effect of solid loading and temperature

In this study, atmospheric acid leaching behaviour of siliceous goethitic nickel (Ni) laterite ore is investigated. Specifically, the effect of −200 μm feed solid loading (30 vs. 45 wt.%) and temperature (70 vs. 90 °C) on leach kinetics, acid consumption capacity and Ni and cobalt (Co) extraction was studied under isothermal, batch (4 h) leaching conditions at pH 1. Incongruent leaching was observed for constituent elements reflecting slow but steady release of value (Ni and Co) and some of gangue metals such as Fe, Mg and Al accompanied by faster and sharp release of Na and Si. Higher temperature and lower pulp solid loading, both led to a 40–50% increase in overall Ni and/or Co extraction and higher acid consumption. At 70 °C and 45 wt.% solid loading, Ni/Co extraction after 4 h was the lowest (∼14/16%) whilst the highest extraction (∼67/56%) was observed at 90 °C and 30 wt.% solid loading. Temperature appeared to have dramatic influence on Ni/Co and other impurity metals’ extractions revealing the chemical reaction controlled nature of the leaching. Higher solid loading and longer leaching time also both slowed down the leach kinetics. A two-stage chemical reactions-controlled leaching mechanism involving a faster initial leaching kinetics followed by a slower leaching at lower rate constants and higher activation energies was established for release of Ni, Co, Fe and Mg. The mechanism reflects the fast leaching of reactive host mineral phases (e.g., clays and Mg–silicates) during first 30 min followed by slow leaching of more refractory mineral phases (e.g., goethite and quartz) during the rest of leaching period. The findings provide a greater understanding for enhanced atmospheric acid leaching process of siliceous goethitic laterite ores.     

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”.     

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.     

Calcination of low-grade laterite for concentration of Ni by magnetic separation

With the continuous depletion of high-grade nickel ores such as millerite and niccolite, nickeliferous laterites have become the major source for the production of nickel metal. However, only 42% of the world’s production of nickel comes from laterites, since the concentration of Ni is relatively low (ca. 2 wt.%). In addition, other metals, such as magnesium, iron and silicon can be found in laterite, which make the concentration of nickel even more difficult.In this study, a low-grade nickeliferous laterite ore was first calcinated and then processed by using a wet magnetic separator in order to recover nickel. Since, the ore contains both Ni and Fe, the calcination of laterite is effective in altering the crystalline structure of Fe species and therefore its magnetic properties, which in turn enable the selective concentration of nickel by magnetic separation that is an easy and environmentally-friendly technique. The experimental results have indicated the importance of carefully controlling: (1) the calcination temperature; (2) the pulp density and (3) applied magnetic field strength. The main finding of this work was that magnetic separation is effective in recovering 48% of nickel from laterite, increasing the Ni grade in the recovered product from 1.5% to 2.9%, when prior to the separation the ore was calcinated at 500 °C for 1 h.     

Beneficiation of colemanite tailings by attrition and flotation

Boron minerals are generally concentrated using attrition methods followed by screening and classification to remove clay minerals in industrial scale. Physical concentration methods are used in Kestelek Boron Mine for the concentration of colemanite. Because of the inefficient process operation, the tailings containing about 20% B₂O₃ are discarded into the tailings pond. In this study, colemanite tailings sample taken from tailings pond was treated using scrubbing + screening followed by flotation to recover the lost boron. As a result of the experimental studies, a concentrate containing 44.5% B₂O₃ was produced with 68.4% B₂O₃ recovery.     

Two-stage fast flotation of coal tailings using reflux flotation

Low pulp density and low grade slurries in the coal and minerals industries are discharged as waste to tailings dams, incurring significant losses of valuable particles. This paper investigates the rapid processing and cleaning of hydrocyclone overflow coal slurry using two laboratory scale Reflux Flotation Cells in series as a means to economically beneficiate low quality tailings streams. The Reflux Flotation Cell incorporates a novel arrangement of inclined channels to enhance bubble-liquid segregation, enabling extremely high gas rates and liquid rates per unit of vessel area. Hence, in the first stage, fast flotation is employed to rapidly recover fine coal particles using a feed flux of 11.4 ± 0.5 cm/s, up to an order of magnitude increase in the throughput rate over conventional flotation systems. First stage product was then sent to a second stage for counter-current washing using fluidisation wash water to produce a fully deslimed product, having ash percent in agreement with the minimum ash attainable using flotation as determined through tree flotation analysis. The results demonstrate the potential for two-stage Reflux Flotation to deliver high throughput at a high separation efficiency from low quality slurry, with a fivefold reduction in the required vessel footprint, thus overcoming the principal economic deterrent of having to install banks of large-scale flotation cells.     

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.     

Acid leaching and rheological behaviour of a siliceous goethitic nickel laterite ore: Influence of particle size and temperature

This study investigates the isothermal, batch, H₂SO₄ acid leaching behaviour of siliceous goethitic (SG) nickel (Ni) laterite ore and its links to pulp rheology. Specifically, the effect of feed ore particle size (−0.2 vs −2.0 mm), leaching temperature (70 vs 95 °C) and pulp rheology on Ni and pay metal, cobalt (Co) extraction kinetics and yield was studied for 4 h on 40 wt.% solid dispersions at pH 1. The leaching behaviour was distinctly incongruent, reflecting the disproportionate proliferation of major gangue mineral’s constituent elements (e.g., Fe, Al, Mg, Na, Si) alongside Ni and Co in the pregnant leach solution. At 70 °C, Ni/Co extraction rates were notably lower (<20%) in contrast with 95 °C where a significant increase in Ni/Co extraction to 78/77% and 74/77%, respectively, for the −0.2 and −2.0 mm feeds occurred. The slurries displayed a non-Newtonian, shear thinning Bingham plastic rheological behaviour of which the viscosity and shear yield stress increased markedly in the course of 4 h leaching. The pulp viscosity and shear yield stress were greater at lower temperature than at higher temperature and they were also greater in slurries with finer than coarser feed particles. The dynamic pulp rheology, however, had no marked effect on the overall Ni/Co extraction rates. Whilst the feed ore particle size had no remarkable impact on overall Ni/Co extraction, it led to noticeably higher acid consumption and enhanced slurry rheology in the finer sized ore. The mechanism of leaching the SG ore followed a two-stage, first order chemical reaction-controlled shrinking core model, the kinetics of which gave higher rate constants and lower activation energies for the release of Ni, Co, Fe and Mg in the first stage. A faster leaching process involving more reactive minerals during the first 30 min is envisaged to be followed by leaching of the more refractory minerals.     

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.     

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.     

A Pilot Scale Study of Cenosphere Recovery and Concentration using the Inverted Reflux Classifier

Cenospheres are hollow spherical particles formed as part of the fly ash waste of coal-fired power stations. In a previous paper Kiani et al. (2015) investigated the recovery and the concentration of these particles using an Inverted Reflux Classifier (IRC) at a laboratory scale, of cross-section 0.100 m × 0.086 m, achieving a throughput advantage over a conventional fluidized bed by a factor of 54. The present paper investigated the potential to achieve scale-up, utilizing a pilot scale device with cross-section 0.3 m × 0.3 m. The product grade and recovery were examined as a function of the solids yield by varying the product volumetric rate relative to the feed volumetric rate. The performance data were compared directly with those obtained at the smaller laboratory scale. Agreement was excellent. The performance was also examined as a function of the feed slurry flux, with good agreement again evident at the laboratory and pilot scales. Overall, the separation performance was excellent, with a cenosphere recovery of about 80% achievable at a high upgrade of 19 while a recovery of 75% was achieved at an upgrade of 38. Here the feed solids flux was 4.2 t/(m² h). It is noted that much higher upgrade was achieved at a recovery of about 80% in the former study by operating at a lower solids feed flux. This paper provides the necessary basis for proceeding with a full scale implementation of this technology.     

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.