Graham Jeffrey Sparrow

Cyanide and Other Lixiviant Leaching Systems for Gold with Some Practical Applications

Selection of a leaching system for gold involves consideration of ore texture and mineralogy, chemical requirements, leaching techniques, the development of flowsheets, and environmental management. Aqueous dissolution chemistry for alkaline, neutral, and acid systems is mainly considered here. All systems require an oxidant to oxidise gold and a ligand to complex with gold in solution. Adjustment of pH is usually necessary. Alkaline lixiviant systems (pH > 10)include cyanide, ammonia-cyanide, ammonia, sulphide, nitriles, and a few other minor possibilities. Oxygen is the main oxidant. Cyanide, which is the main ligand in these systems, forms an anionic complex, “Au(CN)2”, with Au(I). Gold dissolution rates are controlled by oxygen solubility in solution. Neutral lixiviant systems (pH 5-9)include thiosulphate, halogens, sulphurous acid, and bacteria plus natural organic acids as the ligand. Oxygen is the normal oxidant and either Au(I) or Au(III)complexes are formed. Acid leaching systems (pH ⩾ 3)may cont...

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.

Extraction of platinum, palladium and gold by cyanidation of Coronation Hill ore

Abstract The effects of cyanidation conditions (pH, time, particle size of the ore, additives to the leach, and oxidative pre-treatments) at ambient temperatures on the extraction of platinum, palladium and gold from samples from the Coronation Hill deposit, Australia, were studied. Three ore lithologies were investigated: a diorite, a green tuffaceous siltstone, and a quartz-feldspar porphyry. Coarse and fine native gold, native platinum and palladium, stibiopalladinite and platinum-palladium selenides were identified in the samples. Coarse gold, but no platinum or palladium, was recovered by amalgamation. Cyanidation at pH 11.5 for 48 h of ore ground to 80% − 74 μm (standard cyanidation conditions) extracted 6–18% of the platinum, 20–50% of the palladium (the values depending on the ore type), and over 95% of the gold. Lowering the cyanidation pH to 9.5 resulted in extraction of up to 85% of the palladium, although the platinum extractions were usually less than 20%. Final grinding increased the palladium extractions but had little effect on platinum extractions, and decreased gold extractions by a few percent. The high gold extractions were obtained by 24 h cyanidation, but longer times were required for extraction of platinum and palladium. Addition of thallium salts at pH 9.5 increased the platinum extraction values to 30–35%. Addition of sodium chloride, sodium sulphite, sodium thiosulphate, lead salts and hydrogen peroxide had little effect on the extraction of the precious metals. Pre-treatment with sodium chloride/hydrogen peroxide solution at pH 1, and filtration before the subsequent cyanidation at pH 9.5, increased the platinum extraction from 20% to 37%. However, the palladium extraction was reduced to 25%.

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.

Removal of chrome spinels from Murray Basin ilmenites by low temperature roasting

Abstract Ilmenite concentrates from the Murray Basin deposits of southeastern Australia contain discrete grains of chrome spinels as a major impurity that must be removed for the ilmenite to become a satisfactory feedstock for the sulphate route to titania pigment production. Conditions were established for a magnetising roast of the ilmenite samples, followed by a magnetic separation, that produced products with (≤0·1 wt-%Cr2O3. The roasting conditions were adjusted to keep the crystallite size of any rutile formed during the roast to <10 nm. Under these conditions the rutile was appreciably soluble in sulphuric acid so that the roast products had satisfactory digestion properties. The roasting conditions are a sensitive function of the mineralogy of the ilmenite grains resulting from chemical weathering in the deposits. For mildly weathered ilmenite concentrates from the southern Murray Basin, containing intimate fine grained associations of ilmenite and pseudorutile, magnetising roasts at <650°C in slightly reducing fluidising gases gave magnetic products containing, <0·1 wt-%Cr2O3 at >80 wt-% recovery, with high solubilities in sulphuric acid. However, similar conditions were not effective for ilmenite concentrates from the central Murray Basin that contain mixtures of unweathered Mg rich ilmenite and highly weathered ilmenite. The strongly weathered ilmenite contains high levels of intra grain chromia (0·2 wt-%Cr2O3) and generates high levels of rutile on roasting, making it difficult to obtain acceptable sulphate route feedstocks from this ilmenite.

Selective dissolution of uranium from Olympic Dam copper concentrates

Abstract The copper flotation concentrates produced from the Olympic Dam deposit in South Australia often contain significant amounts of uranium. Previous work indicated that uranium could be removed from a chalcocite-bornite concentrate by reaction with sulphuric acid. Reactions with djurleite, roxbyite, bornite and chalcopyrite-rich concentrates that are representative of each of the major copper sulphides in the deposit have now shown that uranium can be selectively removed from all types of copper concentrate generated from the deposit. The selective dissolution of uranium from the copper concentrates was achieved in 24 h by reaction with sulphuric acid at 30–60 ° C in an inert atmosphere. Redox potentials during most reactions were 225–250 mV vs saturated calomel electrode. It is proposed that the redox potential of the suspensions was determined by the oxidation of the copper sulphides. The addition of an oxidant (ferric alum) to increase the redox potential increased the amount of copper that dissolved but did not significantly affect the uranium dissolution rate. Oxidation of uraninite in the copper concentrates was by reaction with iron(III) from the acid dissolution of hematite, the major gangue mineral in the concentrates. It is likely that fluoride ions, from the dissolution of fluorite in the concentrates, are also involved in reactions leading to the dissolution of the uranium. Copper dissolution occurred by reaction of the copper sulphides with iron(III) and by reaction of acid with products from the aerial oxidation of the concentrates during storage.

Selective dissolution of uranium from a copper flotation concentrate

Abstract The copper flotation concentrate produced from the Olympic Dam copper— uranium gold deposit in South Australia contains uranium associated with the copper sulphide and gangue minerals. Work with a copper concentrate in which the copper mineralization was of the chalcocite-bornite type has shown that 94–97% uranium can be dissolved with sulphuric acid (> 40 g/L) at 30–60°C in 24 h. The addition of an oxidant was not necessary. Copper was leached from the concentrate along with the uranium. In the first 15 min, 5–7% copper dissolved but thereafter virtually no further copper dissolved when the reaction was carried out under nitrogen or argon (i.e., in the absence of oxygen). In air and oxygen, copper dissolution continued over the 24 h of the reaction. The initial rapid dissolution of copper was associated with oxidation of djurleite to roxbyite and dissolution of surface oxidation products. In air and oxygen, oxidation of roxbyite and bornite to blaubleibender covellite was associated with further dissolution of copper. The redox potential of the suspensions was controlled by reactions of the copper sulphide minerals. For reactions under nitrogen the redox potential of the system was 225–250 mV vs. saturated calomel electrode (SCE), while in air or oxygen the potential gradually rose (to 350 mV vs. SCE) as successive copper sulphides were oxidized. These results, and work at the Olympic Dam metallurgical pilot plant, have shown that uranium can be removed selectively from copper flotation concentrates produced from the Olympic Dam deposit.

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.

Review of processing conditions for Murray Basin ilmenite concentrates

Abstract Australia is a major world producer of heavy minerals from mineral sand deposits, but to maintain its premier position in the world market place, the development of large resources such as in the Murray Basin will be necessary. There are two types of deposits in the Murray Basin, the fine-grained (20–80 μm) WIM-type deposits and those with coarser mineralisation (90–300 μm), similar to other Australian deposits. Only the higher grade, coarse-grained deposits have been exploited, with little or no value realised from the majority ilmenite component because of its complex mineralogy. Compositional data for ilmenites from deposits across the basin are presented and potential processing options for them taking into account the grain size, composition and impurity element contamination are discussed. In particular, processes to lower chromia levels and fluidised bed and kiln-based processes to upgrade weathered ilmenites to suitable pigment feedstocks are discussed. Recommendations regarding further research required to enable commercial development of Murray Basin ilmenites are given.