The effect of acrylamide-co-vinylpyrrolidinone copolymer on the depression of talc in mixed nickel mineral flotation

Copolymers of acrylamide and vinylpyrrolidinone with varying compositions have been synthesised and employed to depress talc in a model flotation system with process plant operation conditions. Adsorption isotherms indicated that the hydrophilic acrylamide homopolymer has a very low affinity for the hydrophobic talc surface, whereas vinylpyrrolidinone homopolymer strongly adsorbs onto the talc surface. Micro-flotation experiments revealed that the copolymer system can induce stronger talc depression than the homopolymer variants, with the most effective copolymer depressant having 25-30% vinylpyrrolidinone incorporation. The copolymer system is observed to have inherited the strong talc affinity of vinylpyrrolidinone polymer and the strong hydrophilic property from polyacrylamide. This combined effect facilitates the desired strong talc depression in single mineral flotation. However, this copolymer system has similar adsorption affinity on both the talc and pentlandite, hence depressing both talc and pentlandite in the mixed mineral flotation system. This research shows that a sufficient hydrophobic balance on the polymer is necessary for the adsorption and subsequent depression for talc. However, polymer with high adsorption selectivity is required to be a successful synthetic talc depressant for mixed mineral system.     

Influence of mineralogy and ore texture on pentlandite flotation at the Nkomati nickel mine,South Africa

The influence of ore mineralogy and ore texture on flotation response was studied for 29 samples from the main mineralised zone at Pit 3 of the Nkomati Ni mine, through laboratory scale flotation testing, laboratory assay, and mineral liberation analyser examination of the ore and the concentrates. The individual sample flotation responses vary widely in terms of Ni grade, and cumulative Ni recovery. It is demonstrated that this is a complex function of ore mineralogy and ore texture. Chalcopyrite is the first sulphide to float, followed by pentlandite and finally pyrrhotite, in ore samples with dominant chalcopyrite, or where pentlandite, pyrrhotite and chalcopyrite occur in equal abundance. However in samples with a high ratio of pyrrhotite to pentlandite and chalcopyrite, pyrrhotite floats earlier than expected, reports to concentrate over the entire flotation period, and depress and extend the flotation of pentlandite over the flotation interval with no clear peak of Ni recovery during flotation. Primary silicates (e.g. olivine and pyroxene) and alteration-related minerals (talc, tremolite and chlorite) are naturally floating, and hence affect the flotation of pentlandite in a similar manner to that of pyrrhotite. The most problematic ore at Nkomati in terms of Ni recovery is characterised by fine disseminated and fine bleb- or net-texture sulphides, contain abundant olivine, pyroxene, amphibole, talc and tremolite, and include abundant metamorphism-related country rock xenoliths (with calc-silicate minerals such as diopside and tremolite).     

Pyrite and pyrrhotite open circuit potentials study: Effects on flotation

In this research, effects of chloride & sulfate ions and pH changes on the open circuit potentials of pyrite, pyrrhotite and steel and also their reduction-oxidation processes were studied. These potentials were measured at different pH values in 3 electrolytes: distilled water, sodium chloride solution and sodium sulfate solution. A pair of electrodes including reference and operative electrodes was used for measurement and control of potentials. The results showed that with pH increasing, pyrite and pyrrhotite open circuit potentials decrease whereas under the same conditions the mentioned potential for steel rises. It was also observed that use of sodium chloride and sodium sulfate solutions as electrolyte decreases the open circuit potentials, because the presence of chloride and sulfate ions speeds up the oxidation-reduction process. Furthermore, the effect of the mentioned parameters on galvanic interactions between minerals and grinding media in grinding circuits, as well as galvanic interactions between the minerals in flotation process, mineral floatability and their collector adsorption have been discussed.     

The effect of flotation and parameters for bioleaching of printed circuit boards

Waste electrical and electronic equipment (WEEE) is currently one of the fastest growing waste streams in the world. Typical for WEEE is the high content of valuable and precious metals, as well as harmful contaminants like halogens, flame retardant chemicals and plastics. Currently, WEEE treatment and metal recovery methods are imperfect, polluting and energy intensive. In this paper, novel treatment possibilities are outlined for printed circuit boards (PCB) utilizing both the flotation separation technique and acid bioleaching. Flotation, conducted after crushing and sieving of PCB, produced two fractions: metal-rich concentrate, which is more suitable for pyrometallurgical treatment than untreated PCB, and metal-poor froth suitable for acid bioleaching. It was seen that especially low pH (1.6), high initial Fe²⁺ concentration (7.8 g/l) and low PCB froth concentration in the bioleaching solution (50 g/l) were beneficial for the rapid and selective dissolution of copper. With these parameters, 99% of copper was solubilized from PCB froth in bioreactor treatment, with Cu (6.8 g/l) and Fe (7.0 g/l) being the only major metallic elements in bioleaching solution.     

Removal of titanium impurities from feldspar ores by new flotation collectors

The principal impurities in some feldspar ores are titanium and iron, which impart color and in turn degrade the quality of the ore. Mineralogical investigations on the majority of Turkish feldspar ores indicate that rutile and, scarcely, sphene are the major titanium minerals and iron mainly originates from mica minerals. Conventionally, fatty acids, and particularly sodium oleate, are extensively used to float discoloring minerals from feldspar ores with reasonable success. In this study, new collectors, oleoyl sarcosine and hydroxamate, reported for the first time in the literature, have been used to float titanium impurities. It is found that compared to fatty acids these reagents achieve superior results. The mechanism of the action is elaborated on the basis of experimental data.     

Upgrading titanium bearing Na-feldspar by flotation using sulphonates, succinamate and soaps of vegetable oils

Large reserves of feldspar ores exist in the Aegean region of Turkey. Unlike conventional ores, the main impurity in some of these deposits is titanium. In this study, the results of testwork carried out to determine flotation conditions for efficient separation of titatium bearing minerals from two different feldspar ores from the area are presented. The study commenced with mineralogical examination which indicated that the main titanium bearing mineral in the ore was rutile. Another titanium mineral, sphene, was also present in relatively small quantities. In the tests, the effects of particle size, pH, collector types and their amount were tested.The results showed that efficient separation of titanium bearing minerals could be achieved by oleate flotation at around pH 5, after grinding the ore down to 100% −180μm. Oleate provided results superior to sulphonates and succinamate. Equivalent success was obtained by the use of collectors prepared from locally available vegetable oils. Of the five oils tested, sunflower and cotton seed oils produced the most efficient separation.     

Particle shape effects in flotation. Part 1: Microscale experimental observations

There has long been speculation as to whether some particle shapes are more ‘floatable’ than others, which might be due to differences in the induction period required to achieve attachment between the particles and the air bubbles in the pulp. To resolve this, we used the Milli-Timer apparatus to directly observe the process of particle–bubble interaction and attachment by means of a magnified, high-speed video recording, thus providing a direct measure of the induction period for attachment.To assess the influence of particle shape on induction time we used two varieties of methylated borosilicate glass particles — spheres and angular ‘frit’ — in a range of tightly-sized fractions. Other factors that could affect the induction time, such as the polar angle of sliding commencement, and approach velocity, are accounted for using multiple nonlinear regression.Our results illustrate the importance of particle shape on induction period, with angular particles exhibiting induction periods that were an order of magnitude lower than those of spheres. Furthermore, the induction period was seen to decrease with increasing particle velocity, or kinetic energy on approach, but increased as the trajectory approached the limit of just grazing the bubble. These results indicate that attention should be paid to the shape of particles obtained from the grinding operation, besides particle size.     

Assessment of the floatability of chalcopyrite,molybdenite and pyrite using biosolids and their main components as collectors for greening the froth flotation of copper sulphide ores

Biosolids and representative compounds of their main components – humic acids, sugars, and proteins – have been tested as possible environment-friendly collectors and frothers for the flotation of copper sulphide ores. The floatability of chalcopyrite and molybdenite – both valuable sulphide minerals present in these ores – as well as non-valuable pyrite was assessed through Hallimond tube flotation tests. Humic acids exhibit similar collector ability for chalcopyrite and molybdenite as that of a commercial collector (Aero 6697 promoter). Biosolids show more affinity for pyrite. The copper recovery (85.9%) and copper grade (6.7%) of a rougher concentrate obtained using humic acids as main collector for the flotation of a copper sulphide ore from Chile, were very similar to those of a copper concentrate produced by froth flotation under the same conditions with a xanthate type commercial collector. This new and feasible end-use of biosolids and humic acids should be new environment-friendly organic froth flotation agents for greening the concentration of copper sulphide ore. Now, further research is needed in order to scale current laboratory assays to operational mining scales to determine efficiencies to industrial scale.     

Technique to automate measurement of water overflow rate in frother characterization testing

A novel method for measuring water overflow rate from a laboratory column in the air-water system is presented. The method is based on accumulating overflow in a tube and measuring the rate of increase in pressure. The development is a further step towards automating frother testing.     

Extraction of lithium from micaceous waste from china clay production

The granites of South-West England are a potential source of lithium which is generally found within the mica mineral, zinnwaldite. It is mainly found in the central and western end of the St. Austell granite. When kaolin extraction occurs in these areas a mica-rich waste product is produced which is currently disposed of in tailings storage facilities. In this study a tailings sample containing 0.84% Li₂O was upgraded by a combination of froth flotation, using dodecylamine as the collector, and wet high intensity magnetic separation (WHIMS) to 2.07% Li₂O. The concentrate was then roasted with various additives, including limestone, gypsum and sodium sulphate, over a range of temperatures. The resulting products were then pulverised before being leached with water at 85 °C. Analysis of these products by XRD revealed that the water-soluble sulphates, KLiSO₄ and Li₂KNa(SO₄)₂, were produced under specific conditions. A maximum lithium extraction of approximately 84% was obtained using gypsum at 1050 °C. Sodium sulphate produced a superior lithium extraction of up to 97% at 850 °C. In all cases iron extraction was very low.Preliminary tests on the leach solution obtained by using sodium sulphate as an additive have shown that a Li₂CO₃ product with a purity of >90% could be produced by precipitation with sodium carbonate although more work is required to reach the industrial target of >99%.     

Froth transport characterization in a two-dimensional flotation cell

A study of the froth bubble transport in a two-dimensional (2D) flotation cell was performed. Experiments were developed as a 2 × 2 factorial design, in which the effect of superficial air rate (1.2–1.8 cm/s) and froth depth (2–4 cm) on the froth transport for a two phase (air–water) system was characterized.Using image analysis techniques, bubble residence times, air recovery, bubble path and bubble size increase through the froth were obtained. This information was complemented by froth surface velocity measurements using the Visiofroth system.It was found that bubbles transported from the pulp–froth interface up to the overflow, showed a minimum residence time for bubbles entering the froth near the lip wall. Also, the air-recovery significantly changes in a range of 7–20% at different operating conditions.Higher residence times promoted bubble size increase by coalescence for bubbles transported from the interface. Conversely, for lower residence times, a smaller increase in bubble size was observed.     

The dependency of the critical contact angle for flotation on particle size - Modelling the limits of fine particle flotation

The flotation behaviour of fine particles is studied in this work. Fine methylated quartz particles within the size range from 0.2 to 50 μm, and with varying contact angles, were floated in a mechanical flotation cell. Results indicate that particles of a given size need to possess a minimum critical contact angle, which increases in value as particle size decreases, for flotation to be initiated. As a consequence, a non-floating component exists within a given size fraction. This is interpreted as a fraction consisting of particles below the critical contact angle for flotation for that size. The critical contact angle for flotation is explained in terms of the existence of an energy barrier for bubble-particle attachment. The flotation results are interpreted by means of [Scheludko et al., 1976] and [Drelich and Miller, 1992] models for the floatability of fine particles. The experimental data compared very well with calculations using the Drelich and Miller equation, allowing extension to the prediction of the critical contact angle for flotation down to particle sizes well below the previous limits investigated, bridging the gap existing in the literature.     

The contributions of geometallurgy to the recovery of lithified heavy mineral resources at the Namakwa Sands mine, West Coast of South Africa

The Namakwa Sands heavy mineral deposit is located at Brand-se-Baai along the West Coast of South Africa and is a world class producer of premium quality zircon (ZrSiO₄), ilmenite (FeTiO₃) and rutile (TiO₂) concentrates from mainly aeolian sands. Superimposed on the coastal clastic Cainozoic ore-bearing sequence is a calcium-magnesium-rich pseudo-stratigraphy locally referred to as cemented hard layers that effectively lithify the mineralised sands to various degrees of hardness, rendering it unsuitable for routine treatment by wet spirals. Namakwa Sands has recently completed an expansion programme, inclusive of a SAG mill and screen installation, which allows the processing of the cemented hard layers, the first ever in the mineral sands industry. Operationally, the SAG mill and screen are performing well in tandem and are helping to improve mineral resource utilisation to levels not previously possible. Contained zircon output has increased, but at the cost of marginally lower recoveries as predicted from pilot studies. This paper reports on the systematic geometallurgical approach to improve mineral resource utilisation by successfully processing lithified ore.     

Application of hydrophobic and magnetic plastic particles for enhanced flotation recovery

The effect of hydrophobic and magnetic plastic particles having a contact angle of around 83° on flotation performance was evaluated using coal particles of varying degrees of floatability. The magnetic plastic material were recovered by a low intensity magnetic separator and recycled back to the flotation feed for re-use. Flotation rate tests conducted on coal using a conventional cell proved that combustible recovery and flotation rate were significantly enhanced with the addition of the plastic particles, especially for difficult-to-float coals, which was corroborated by flotation column tests. Carrying capacity and particle size-by-size flotation tests further showed that the magnetic plastic particles preferentially increased the recovery of coarse particles by as much as 35 absolute percentage points due to froth stabilization which reduced the selective detachment of coarse and/or weakly hydrophobic particles. The enhanced flotation recovery was attributed to the influence on liquid drainage rate in the froth zone, froth stability, bubble coalescence and flotation rates.     

The behavior of N,N-dipropyl dodecyl amine as a collector in the flotation of kaolinite and diaspore

The flotation behaviors of kaolinite and diaspore were investigated using N,N-dipropyl dodecyl amine (PN). Maximum recoveries of kaolinite and diaspore were about 90% and 50%, respectively. Both recoveries of kaolinite and diaspore increased with the increasing dosage of PN. On the basis of the zeta potential and FT-IR spectra, the ionization of surface hydroxyl and the isomorphic exchange of surface ions account for the charging mechanisms of surfaces of kaolinite and diaspore. The adsorption mechanism of PN on the surface of two minerals was mainly electrostatic. After reacting with PN, the zeta potential of kaolinite and diaspore both increased, and comparing with the original zeta potentials, potential of kaolinite increased much more than that of diaspore, this can explain why recovery of kaolinite is higher than that of diaspore.     

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.     

Cu-S flotation separation via the combination of sodium humate and lime in a low pH medium

The flotation separation of chalcopyrite and pyrite was studied in the presence of sodium humate. The results of flotation tests indicated that pyrite can be selectively depressed by sodium humate, and the activity of sodium humate was strongly affected by the pH of the pulp. At high pH values, pyrite was strongly depressed by sodium humate; however, the content of chalcopyrite was not affected. Ore flotation tests were successfully conducted in the laboratory and at the Dexing Copper Mine by applying sodium humate as a pyrite depressant. By adding 40-60 g/t of sodium humate to the pulp and adjusting the pH to 10-10.5 with CaO, a concentrate with a Cu content of 24% was obtained without reducing the Cu recovery rate. In addition, the dosage of CaO was reduced, and the recovery of Au, Ag and Mo in the copper concentrate was enhanced due to the reduced pH of the pulp. The zeta potential, adsorption of xanthate and contact angle of the mineral surface were measured, and the results from surface measurements indicated that there was a strong hydrophilic interaction between sodium humate and the surface of pyrite. Moreover, the results revealed that the interaction between sodium humate and chalcopyrite was weak. Infrared (IR) spectra of pyrite and sodium humate were obtained, and the results indicated that sodium humate was chemically adsorbed on the surface of pyrite.     

The effect of using different comminution procedures on the flotation of Platinum-Group Minerals

The use of High Pressure Grinding Rollers (HPGR) has been widely reported to have major benefits in the treatment of minerals such as iron ore and diamonds. To date there have been few investigations into its use in the treatment of ores containing Platinum-Group Minerals (PGMs). HPGRs are known to reduce energy consumption and wear costs and improve the throughput in the circuit. In the present investigation the effect of the comparative use of HPGR and conventional crushing in combination with either dry or wet rod milling on the flotation of PGMs was studied using batch flotation. Previous studies of a base metal sulphide had shown that either HPGR or conventional crushing followed by dry milling produced the highest grades and recoveries (Palm et al., 2010). However in the present study it was observed that a similar treatment of Platinum-Group Minerals produced the poorest results and the highest grades and recoveries were obtained for the case of conventional crushing in combination with wet milling. The HPGR showed no advantages in terms of flotation performance and dry milling produced particularly poor flotation results.The results were investigated further using various surface characterization techniques in order to determine the reason for the decrease in grades and recoveries of platinum when using dry milling and HPGR as opposed to the case for base metal sulphides. The feed and product samples were analysed using ToF-SIMS, XPS and MLA. The paper will propose reasons to explain the different flotation behaviour of the two ore types following the various comminution processes focusing on the surface characteristics of the ores, the particle size distribution and the pulp chemistry.     

Alkaline digestion of dunite for Mg(OH)2 production: An investigation for indirect CO2 sequestration

The sequestration of CO₂ by carbonating natural minerals has a great potential for secure reduction of net CO₂ emissions. Feedstock Mg–silicate minerals are usually converted into Mg rich solutions or Mg(OH)₂ before the carbonation process, due to the slow reaction kinetic of direct carbonation. The present work studied the alkaline digestion of Mg–silicate minerals into Mg(OH)₂ for CO₂ sequestration. Powdered dunite containing ∼73 ± 2 wt% of forsterite (Mg₂SiO₄) was dissolved using highly concentrated NaOH aqueous systems at 90 and 180 °C with varied NaOH concentration and duration of reaction. Thermal analysis and Rietveld Refinement Quantitative Phase Analysis (QPA) confirmed that an effective digestion of dunite was possible at 180 °C achieving 80 wt% of Mg(OH)₂. It was found that NaOH concentration in solution, temperature and duration of reaction significantly influence the progress of digestion.