W.J. Bruckard

A review of factors that affect contact angle and implications for flotation practice.

Contact angle and the wetting behaviour of solid particles are influenced by many physical and chemical factors such as surface roughness and heterogeneity as well as particle shape and size. A significant amount of effort has been invested in order to probe the correlation between these factors and surface wettability. Some of the key investigations reported in the literature are reviewed here. It is clear from the papers reviewed that, depending on many experimental conditions such as the size of the surface heterogeneities and asperities, surface cleanliness, and the resolution of measuring equipment and data interpretation, obtaining meaningful contact angle values is extremely difficult and such values are reliant on careful experimental control. Surface wetting behaviour depends on not only surface texture (roughness and particle shape), and surface chemistry (heterogeneity) but also on hydrodynamic conditions in the preparation route. The inability to distinguish the effects of each factor may be due to the interplay and/or overlap of two or more factors in each system. From this review, it was concluded that: Surface geometry (and surface roughness of different scales) can be used to tune the contact angle; with increasing surface roughness the apparent contact angle decreases for hydrophilic materials and increases for hydrophobic materials. For non-ideal surfaces, such as mineral surfaces in the flotation process, kinetics plays a more important role than thermodynamics in dictating wettability. Particle size encountered in flotation (10-200 microm) showed no significant effect on contact angle but has a strong effect on flotation rate constant. There is a lack of a rigid quantitative correlation between factors affecting wetting, wetting behaviour and contact angle on minerals; and hence their implication for flotation process. Specifically, universal correlation of contact angle to flotation recovery is still difficult to predict from first principles. Other advanced techniques and measures complementary to contact angle will be essential to establish the link between research and practice in flotation.

Gold and silver extraction from Hellyer lead-zinc flotation middlings using pressure oxidation and thiourea leaching

Abstract A new route for the extraction of gold and silver from three different galena-sphalerite-pyrite flotation middlings (bulk scavenger tailing, zinc middling and pyrite middling) assaying 2.3, 2.6 and 1.5 g/t Au and 82, 53 and 27 g/t Ag, respectively, and containing different proportions of the sulphide minerals was investigated. Each sample was pressure oxidised for 4 h at 200°C or 220°C and 700 kPa oxygen overpressure, at an initial acidity of either 0 or 36.8 g/l H 2 SO 4 . This resulted in solid products containing, respectively, either hematite and lead jarosite, or iron hydroxy sulphates and lead sulphate. Most of the zinc and copper, together with some iron and a little silver, were dissolved and were removed by filtration. X-ray diffraction (XRD) examination of the solid products from pressure oxidation at high initial acidity confirmed that they contained mainly a basic iron sulphate (Fe(OH)SO 4 ) with some hydrated iron sulphate (Fe 2 S 2 O 9 ·5H 2 O) and lead sulphate. Thiourea leaching of these products for 4 h at natural pH (2.3–3.4) and Eh (200–350 mV vs SCE), with a thiourea addition rate of 2 g/l h −1 , extracted 83–92% of the gold and 39–96% of the silver, depending upon the sample. The lowest silver extractions were from the pyrite middling. The low silver extractions could have resulted from jarosite formation during thiourea leaching. Some of the hydrated iron sulphate reacted during the leach and there was evidence of silver re-precipitation after 3 h. Thiourea consumptions were 34–73 kg/t. XRD examination of the solid products from pressure oxidation at low initial acidity confirmed that they contained mainly hematite and lead jarosite. Thiourea leaching of these products for 4 h at pH 1.5, with an addition of 17,500 ppm iron(II), an Eh of 225–250 mV, maintained with hydrogen peroxide, and a thiourea addition rate of 2 g/l h −1 , extracted 46–88% of the gold but only 2–7% of the silver. The low gold extraction was obtained from the pyrite middling. The low silver extractions are believed to result from silver incorporation into the jarosite lattice during pressure oxidation. Thiourea consumptions were 16–70 kg/t. Gold and silver extractions by thiourea leaching were, in general, inferior to those obtained by cyanidation of similar pressure oxidation products after lime treatment. For cyanidation, gold extractions from all samples were generally around 90%, as was silver extraction from the basic iron sulphate/lead sulphate product. Silver extractions of 25–73% were obtained from the hematite/jarosite products, values significantly higher than those achieved by thiourea leaching. Total lime consumptions for pH adjustment after pressure oxidation and for cyanidation were up to 480 kg/t CaO for the basic iron sulphate/lead sulphate product and 52 kg/t CaO for the hematite/jarosite product. Cyanide consumptions were up to 5 kg/t NaCN.

Platinum, palladium, and gold extraction from Coronation Hill ore by cyanidation at elevated temperatures

Abstract The effects of cyanidation conditions (pH, time, temperature and particle size of the ore) on the extraction of platinum, palladium, and gold from a quartz-feldspar porphyry sample from the Coronation Hill deposit, Australia, were studied. The main precious metal minerals in the sample were identified as coarse and fine native gold, native platinum and palladium, stibiopalladinite and palladium-platinum-iron alloy phases. Coarse gold, but no platinum or palladium, was recovered by amalgamation. At ambient temperature, for material 80% - 74 μm and a cyanidation time of 48 h, reducing the cyanidation pH from 11.5 to 9.5 increased the palladium extraction from 22.7 to 66.4% and the platinum extraction from 5.7 to 25.7%. Total gold extractions were not affected and were over 98%. Finer grinding increased the palladium extractions but had little effect on platinum extractions, and decreased gold extractions by a few percent. Cyanidation at elevated temperatures of up to 150°C under an overpressure of nitrogen, air, or oxygen increased the palladium and platinum extraction values. The optimum temperature range was 100–125°C. At 100°C under air, 87–92% of the palladium, 73–79% of the platinum, and 95–97% of the gold were extracted in 4–6 h at pH 9.5–11.5. Possible treatment options involving cyanidation at ambient and elevated temperatures, with or without flotation, for recovery of gold, platinum, and palladium are discussed.

Bioflotation and bioflocculation review: microorganisms relevant for mineral beneficiation

Abstract In a world of diminishing resources, current research efforts are often directed to extending the life of existing resources and developing technology to treat resources deemed uneconomic. To this end, biotechnology has been explored as a potential low cost, environmentally benign alternative to many of the current mineral processing techniques. Microorganisms and their metabolites have been successfully applied in the leaching of metals from medium and low grade sulphide minerals for many years. Recent fundamental studies have shown that selected bacteria may also assist in the beneficiation of these minerals through bioflotation and bioflocculation. The current published research demonstrates that several bacteria and their excreted proteins and polysaccharides have a significant effect on the surface chemical properties of some minerals. As such, the bacterial cells and their metabolites can be utilised as flotation collectors and modifiers enabling selective separation. To date, these studies have primarily focused on the sulphide minerals; however, there is considerable scope to expand this research for the beneficiation of oxide minerals such as those contained in iron ores. This paper provides a comprehensive review of current research on the use of microorganisms for mineral beneficiation and the potential use of genetically modified bacteria and, further, discusses the applicability of bioflotation and bioflocculation for the beneficiation of Australian iron ores.

Fundamental studies of electrochemically controlled surface oxidation and hydrophobicity of natural enargite

The surface oxidation and hydrophobicity of natural enargite (Cu(3)AsS(4)) and the formation of oxidation species at the mineral surface have been examined by a novel experimental approach that combines electrochemical techniques and atomic force microscopy (AFM). This approach allows for in-situ, synchronized electrochemical control and examination of the oxidative surface morphology of enargite. Combined with ex-situ cryo X-ray photoelectron spectroscopy surface analysis, the surface speciation of enargite surface oxidation has been obtained, comparing the newly fractured natural enargite surface with those that have been electrochemically oxidized at pHs 4 and 10. At pH 4, surface layer formations consisting of metal-deficient sulfide and elemental sulfur were identified, associated with a limited increase in root-mean-square (rms) roughness (1.228 to 3.143 nm) and apparent heterogeneous distribution of surface products as demonstrated by AFM imaging. A mechanism of initial rapid dissolution of Cu followed by diffusion-limited surface layer deposition was identified. At pH 10, a similar mechanism was identified although the differences between the initial and diffusion-limited phases were less definitive. Surface species were identified as copper sulfate and copper hydroxide. A significant increase in surface roughness was found as rms roughness increased from 0.795 to 9.723 nm. Dynamic (receding) contact angle measurements were obtained by a droplet evaporation method. No significant difference in the contact angle on a surface oxidized at pH 10 and the freshly polished surface was found. A significant difference was found between the polished surface and that oxidized at pH 4, with an increase in contact angle of about 13° (46° to 59°) after oxidation. Competing effects of hydrophilic (copper oxides and hydroxides) and hydrophobic (elemental sulfur) species on the mineral surface under oxidizing conditions at pH 4 and the change in surface roughness at pH 10 may contribute to the observed effects of electrochemically controlled oxidation on enargite hydrophobicity.

Review of flotation of feldspar

Abstract This paper reviews the application of flotation for the separation of feldspar from the other minerals with which it naturally occurs (such as quartz, clay minerals, mica, ilmenite, rutile, anatase and magnetite) as well as the separation of the individual feldspar minerals themselves. Much of the published information on feldspar flotation relates to fundamental studies seeking to elucidate the mechanism of the separation and recovery of feldspar with anionic or cationic collectors and activators. However, the separation schemes currently being used industrially are largely the same as those first proposed 60 years ago because the processes used are very efficient. The general circuit for commercial separation of feldspar consists of three sequential stages of flotation, all of which are carried out in an acid medium. Usually, before flotation, the feed is deslimed, removing any fine clay minerals present. In the first flotation stage, mica is removed with an amine collector. In the second stage, titanium and iron oxide minerals, such as ilmenite and magnetite, are removed using an anionic collector. In the third stage, feldspar is activated with fluoride ions and floated with an amine. The residual product is usually high grade quartz. This procedure results in feldspar products that meet market specifications, including low iron levels. Depending on the nature of the ore being treated and the particular contaminants it contains, one or more of these stages are used. A commercial feldspar product usually contains more than one feldspar mineral, and flotation conditions for the separation of individual feldspar minerals have been developed to produce individual feldspar minerals for specific applications. Potassium feldspar can be floated from sodium feldspar in a sodium chloride solution with an amine collector at neutral or acidic pH values. To overcome environmental issues when fluoride ions are used to activate the feldspar in the third flotation stage, fluoride free flotation conditions that utilise a combination of anionic and cationic collectors have been developed.

Comparison of methods of gold and silver extraction from Hellyer pyrite and lead-zinc flotation middlings

The Hellyer massive sulphide deposit in north-west Tasmania, Australia, which is a complex fine-grained Cu-Pb-Zn-Au-Ag-pyrite-arsenopyrite orebody, has recently been brought into production. Various base metal flotation concentrates are produced for sale. However, a substantial proportion of the gold and silver report in flotation middlings or tailings and is currently not recovered. Research into a number of methods for extraction of this gold and silver was conducted on various plant products, which had different mineralogy, and which assayed 1.5–3.3 g/t Au and 27–144 g/t Ag, as well as 0.7–8.0% Zn and 1.1–11.4% Pb. Product sizings were 90% minus 38 ¼.m.

Multicomponent models of grinding and classification for scale-up from continuous small or pilot scale circuits

Abstract A multicomponent model of grinding and classification circuits as a comprehensive scaling-up method is described. Steady-state test results of an ore are presented for continuously operated ball mills ranging from 0.6 to 4.4 m diameter. Simple translation of the specific breakage rates in the breakage rate-particle size ( k-x ) plane enables an excellent estimate of the largest mill performance to be made from the smaller full scale mill, and fair to reasonable estimates from the two pilot scale mills. Ball size is the major, if not sole cause, for the observed variation in size interval for which the breakage rate was a maximum. A simple method for incorporating this effect into a scale-up procedure is presented. It is concluded that the functional form for the k-x relationship based on batch grinding tests, is inadequate for testing heterogeneous ores under grinding conditions, but otherwise the present results are consistent with the scale-up relation of Austin.

Characterisation of the Mt Weld (Western Australia) niobium ore

Abstract The aim of this paper was to characterise Mt Weld Nb–Ta ore and assess the amenability of the ore to gravity and magnetic separation, flotation, and caustic and acid leaching. The Mt Weld niobium deposit occurs in altered carbonatites in Western Australia where Sr pyrochlore and ferroniobate were the major primary Nb–Ta containing minerals. Primary apatites, pyrochlores and niobates were altered to form micrometre to submicrometre size crandallites. The fine crandallite particles were often clustered to form larger aggregates, mixed or cemented with fine niobates (up to 20 μm) and coarse (up to 200 μm) ilmenite. This study showed that the ore was not suitable for gravity concentration, magnetic upgrading and flotation due to fine particle size. Although the fine particle size was advantageous for chemical treatment and ore responded better to caustic and acid leaching, the chemical treatment alone was well short of providing a commercial grade niobium concentrate containing 60% Nb2O5+Ta2O5. It is concluded that upgrading Mt Weld ore to a commercial grade concentrate would require a combination of physical and chemical treatments where physical treatment is used to remove coarse ilmenite and chemical treatment to dissolve impurities associated with the fine ore.

Multistage leaching of metals from spent lithium ion battery waste using electrochemically generated acidic lixiviant

Lithium ion battery (LIB) waste contains significant valuable resources that could be recovered and reused to manufacture new products. This study aimed to develop an alternative process for extracting metals from LIB waste using acidic solutions generated by electrolysis for leaching. Results showed that solutions generated by electrolysis of 0.5 M NaCl at 8 V with graphite or mixed metal oxide (MMO) electrodes were weakly acidic and leach yields obtained under single stage (batch) leaching were poor (<10%). This was due to the highly acid-consuming nature of the battery waste. Multistage leaching with the graphite electrolyte solution improved leach yields overall, but the electrodes corroded over time. Though yields obtained with both electrolyte leach solutions were low when compared to the 4 M HCl control, there still remains potential to optimise the conditions for the generation of the acidic anolyte solution and the solubilisation of valuable metals from the LIB waste. A preliminary value proposition indicated that the process has the potential to be economically feasible if leach yields can be improved, especially based on the value of recoverable cobalt and lithium.