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By N Stockton

Some preliminary results on the secondary leaching of previously heap leached gold ore are presented. Alternative lixiviants for gold comprising chlorine, thiourea and thiosulphate were compared with cyanide using bottle roll tests. Chlorine was subsequently selected for column leaching tests. Column tests at 1.0, 0.1 and 0.01 gL-1 Cl2 at pH 2 were conducted. The results indicated that about 23 per cent of the gold was leached over 45 days using 1 gL-1 Cl2 at pH 2. The experiments conducted proved that the chlorine/chloride system has good potential for further extraction of gold and silver from the existing cyanide heap leach residues and a process for secondary leaching of gold has been proposed. Further testwork is recommended to extend conditions, sample types, and the duration of leaching period.

Jan 1, 2000

By Davey CJ

An integrated computer software system has been developed by Fluor Australia for modelling, mine planning and production scheduling of single or multi-strata deposits. Modelling of both quantitative and qualitative variables (e.g., thickness, basal elevation, density, grade/ash content) is performed via continuous, minimum curvature surface representations. These may be developed from inverse distance weighting or geostatistical kriging interpolation techniques in conjunction with bicubic spline surface interpolation. The modelling process, rather than being purely mathematical, incorporates facilities for reflecting structural geology interpretations (e.g., faulting, pinchouts) and geological boundary conditions (e.g., subcrop and outcrop). Volumetric programs provide the facility for continuous integration over models (at any specified integration increment) for any number of specified irregular polygonal panel areas or regular rectilinearly defined block plan areas. Multi-variable integration is available, incorporating any or all operations of addition, subtraction, multiplication and division to provide multi-strata tonnages, etc., (e.g., Volume 1 * density 1 + Volume 2 density 2). These programs provide true three-dimensional integration with the facility for calculating results on an area by area, level by level, strata by strata as well as overall basis.

Jan 1, 1983

Thermodynamics of Non-Cyanide Gold Solvents The present status of the raw material base in the gold industry requires specific ways to recover gold and silver from ores. Most gold is closely associated with sulfides which require flotation followed by pressure or bacterial oxidation. The next stages of processing usually require using expensive operations for the preparation of leaching. A lot of researchers want to find a new alternative solvent for the extraction of precious metals from these products. According to the authors, one of the best ways is to use sulfur-based solvents (thiosulfate, sulfite, and their various mixtures). Thermodynamics of gold and silver thiosulfate dilution are calculated for conditions: the concentration of Au - 200 mg/L, oxygen - 3.2 mg/L, thiosulfate - 79 mg/L, OH- - 170 mg/L. the general result is the possibility of process remains under standard conditions. (erAu = = 0.035 V, ?G298 = - 29.18 kJ/g-atom ??). Silver with thiosulfate forms two strong complexes: [AgS203]- and [Ag(S203)2]3-. ?G0 = -1027 kJ/mol. The complex [AgS203]- is less stable: ?G0 = 486 kJ/mol. It should be noted that despite the high stability of gold complexes, the sodium thiosulfate may decompose to sulfite and elemental sulfur during the process, especially when heating. Gold is slightly soluble in the sodium sulfite solution, however it forms the strong complex [Au(SO3)2]3-. Depending on the concentration of free S2O32-ions, silver with sodium sulfite forms three complexes: [AgSO3]-, [Ag(SO3)2]3- and [Ag(SO3)3]5-. The dissolution speed of gold in sodium sulfite solution may be in the same range as it is during cyanidation. In this case, a practical application of sulfite process can be also used, especially for the products containing free elemental sulfur. In conclusion, sulfite-thiosulfate gold lixiviants can be used for concentrates after pressure oxidation of the cakes.

Sep 13, 2010

By C E. J de Ronde, J G. Scott

Economic gold mineralisation was discovered at Round Hill by Homestake NZ Limited in 1984 and was subsequently put into production by Macraes Mining Company Limited (MMCL) in 1990. The total mineral resource in 1990 was estimated at 41.4 Mt @ 1.42 glt at a 0.7g/t cut off, for 1.8 Moz of gold. Since project start up the mill has been upgraded twice to increase throughput and reached its current capacity of 3.2 Mtpa. At 31 December 1996, a total of 16.2 Mt of ore at average grade of 1.73 g/t has been excavated, yielding 655,000 oz of gold. MMCL began exploration along the Hyde Macraes Shear Zone in 1994. Since 1994, exploration has been successful with the discovery of over 3.5 Moz of gold. The resource now stands at 110 Mt @ 1.46 g/t gold for 5.1 Moz. This success has led to a planned expansion of the processing capacity to 9 Mtpa with the building of a new 6 Mtpa processing plant.

Jan 1, 1997

The Nickel Rim Depth Cu-Ni-PGE deposit, discovered by Falconbridge Limited in 1993, illustrates several trends in exploration in the Sudbury area. Recent discoveries tend to be deep, high grade, footwall hosted, close to or straddle property boundaries and made using a multi-disciplinarian exploration approach that integrates economic considerations with geology, metal geochemistry and geophysics (borehole EM). The Nickel Rim deposit occurs down dip of Inco's Victor Deep deposit (4.2 million tonnes at 2.25 per cent Ni, 7.30 per cent Cu) in the eastern margin of the Sudbury Igneous Complex (SIC) at a depth of 2650 m. Before commencing drilling an economic evaluation was done to determine the minimum size needed to be economic at the expected depths and determine the expected value of an exploration venture. The subsequent surface exploration program discovered a drill indicated, undiluted mineral inventory of 1.6 million tonnes of 1.58 per cent Ni, 10.13 per cent Cu, 4.17g/t Pt and 3.50 g/t Pd including 519 000 tonnes of 4.13 per cent Ni, 25.53 per cent Cu, 9.76 g/t Pt and 8.42 g/t Pd. The deposit is hosted by gneiss, migmatite, gabbro and pseudotachylite (Sudbury Breccia) 100 m below the base of the SIC.

Jan 1, 1996

A process was developed to produce electrolytic manganese dioxide, suitable for use in high quality dry cells, from Risdon anode mud. ImpurItIes present In the anode mud were removed prior to the electrolysis which was performed from a manganese sulphate bath using the following conditions.Temp. 95°CH2S04 50-55 g/l.Mn2+ 55-60 g/l.Anode c.d. 10·0 A/sq. ft.Operation under these conditions yielded current efficiency and power consumption figures of 95 per cent and 0·8 kWh/lb respectively. The product conformed to battery manufacturers' specifications as regards physical and chemical properties and performed satisfactorily in a battery testing programme.

Jan 1, 1968

Gold mineralisation was first noted in 1983 at the Wapolu Point area of Fergusson Island, Papua New Guinea (PNG) during a follow-up regional reconnaissance program. Subsequent exploration programs through to late 1987 completed 22,147m of costean trench sampling, 553 aircore/RC/diamond core drill holes (16,970.5m total) and metallurgical testwork, predominantly on the Wapolu grid in two mineralised areas.Subsequently in 1988, geostat-istical global and polygonal resources were calculated for the main mineral-ised zones, Didigayagaya and Dagwalala. In 1992, a 'probable total reserve' of 1.9Mf @ 2.4g/t with a 1.2g/t cut-off was calculated for these prospects.

Jan 1, 1994

Completing its historical role the opera tion of copper blast furnace of the Saganoseki Smelter was terminated in 1970, but it left some interesting data on metal lumps formed at the furnace bottom. Analysis disclosed the lumps to be speiss containing 100 to 10,000 g/t of Au and Ag and stratified to iron speiss and copper speiss. Solidifying temperatures of iron speiss and copper speiss being approximately 1,090¦C and 800¦C respectively, copper speiss precipitated at the lower side. The distribution coefficient of Au and Ag between matte and speiss (g/t Au(Ag) in matte / g/t Au(Ag) in speiss) is in the range of 1 x 10 2 to 2 x 10 ' and 1 x 10 1 to 5 respectively. The higher Cu % is in speiss, the-more An and Ag are distributed into speiss. Except for a few data, these results comfotm well to the range of distribution equilibria of Au and Ag in matte and speiss systems by Gerlach and Others.

Jan 1, 1980

By N Kojola, A Vickerfält, J Huss

The possibility of using an acidic autogenous hydrogen direct reduced iron (H-DRI) slag as a basis for vanadium extraction is investigated. For a raw material to be suitable for producing high-quality FeV, the vanadium oxide content should be sufficiently high and the phosphorus oxide content in a tolerable range. Therefore, the feasibility for vanadium extraction relies on fulfilling two demands: the predominate partitioning of phosphorus to the metal and vanadium to the slag. Correspondingly, the behaviours of phosphorus and vanadium in the Al2O3-SiO2-FeOx-VyOz-P2O5-Fe system were studied experimentally. Slag, 10 g, was put in contact with liquid iron, 30 g, for different durations (2–24 hrs) in closed alumina crucibles at 1873 K. The slag composition was varied between 33.1– 49.3 wt per cent Al2O3, 27.1–46.7 wt per cent SiO2, 16.6–30.1 wt per cent FeOx, and 2.5– 8.8 wt per cent V2O3. The equilibrium phosphorus distribution was 0.42, while the vanadium distribution ranged from 7.6 to 43.4, increasing with the Al2O3- and the FeOx-content.

Jun 19, 2024

By A Batnasan

The main techniques for removing phosphorus as an impurity element in iron ores are physical separation (eg magnetic separation, flotation) and chemical separation (acid leaching, alkaline leaching and bioleaching) processes. When phosphorus occurs in iron ores as phosphate minerals (eg apatite), physical separation techniques are available to remove the gangue minerals from iron ores. On the other hand, chemical separation techniques are suitable for reducing phosphorus levels when phosphorus is associated with iron oxides and hydroxide minerals (eg hematite and goethite) in iron ores. In this study, the dephosphorisation behaviour of high‐phosphorus iron ore was investigated using a combined alkaline roasting and water washing route. Roasting experiments were conducted in a laboratory electric muffle furnace using various concentrations of sodium hydroxide (0–50 g/kg-ore) at different temperatures (100–800°C), varying prolonged roasting times (0–2 h), and different particle size fractions (-45 – -250 μm) of iron ore. Subsequent water washing experiments were employed at various pulp densities (30–1000 g/L) and different times (30 s – 60 min) at ambient temperature. Hematite and goethite were the main minerals in the high-phosphorus iron ore identified using X-ray diffraction (XRD) analysis. The iron ore assaying 63.40 mass % Fe (total), 0.240 mass % P, 4.53 mass % SiO2, and 4.87 mass % Al2O3 were determined using X-ray fluorescence (XRF) analysis. The results showed that the dephosphorisation rate of the iron ore was reached 40 per cent under roasting with 50 g/kg-ore NaOH at 300°C for 2 hours followed by water washing for 30 s with a pulp density of 300 g/L, reducing phosphorus level from 0.240 mass % to 0.142 mass %. At the same time, the iron grade was reduced from 63.4 per cent to 59.5 per cent, caused by an iron loss. Results suggest that the phase transformation of goethite induces the dephosphorisation of high-phosphorus iron ore under alkaline roasting conditions

Aug 22, 2022

Recovery of tin and copper from concen trates containing cassiterite, stannite and chalcopyrite was studied in laboratory-scale tests designed to simulate continuous operation of the matte fuming process in which tin is recovered as SnS fume while copper accumulates in the matte. Charges made up of 300g of mixtures of the concentrate and siliceous flux were added at a uniform rate in pelletized form to crucibles containing 300g of matte and 50g of slag (both of which were recovered from previous tests), while blowing through the matte with air or oxygen-nitrogen mixtures at a rate of 3 to 5 1/min. Accurate material balances enabled the effects of process operating conditions on tin recovery and sulfur utilization to be deter- mined. When operated correctly, the matte fuming process gave tin recoveries exceeding 90 percent in spite of wide variations in the copper content of the matte and in the operat- ing conditions. Deterioration of tin recovery was associated with undersaturation of the slag with respect to silica, and with high partial pressure of sulfur dioxide (greater than approx- imately 0.2) brought about by oxygen enrichment of the blowing air. The utilization of sulfur varied widely in low grade mattes. Copper in excess of approx- imately 10% promoted formation of SO2 by the commonly accepted reaction: FeS + 1-120 2= FeO + S02 while in copper-free and lower grade mattes a substantial fraction of the oxygen was consumed in the alternative reaction FeS + i 02 = FeO + z S2.

Jan 1, 1980

By Ciminelli VST

Thiourea leaching has been applied to a refractory gold concentrate previously treated by bacterial oxidation. Following the bio-oxidation, gold has been leached in acidic thiourea solutions, under different pH(1 - 2.4), thiourea concentrations (2 - 15 g/1), ferric ion concentrations (0.04 - 3 g/1) and temperatures (298 - 333 K). Through this procedure, it was possible to increase gold extraction from 40 to 77 per cent. The increase in temperature raised gold extraction and reduced reagent consumption. Sodium sulphite addition was also very effective in lowering thiourea consumption. Staged leaching operations with interstage acid washing and carbon-in-leach experiments were tested, in an attempt to increase gold extraction. The results of thiourea leaching were compared with those obtained with cyanidation.

Jan 1, 1993

By Ho SE, Neil Phillips G, Houstoun SM

There are numerous individual gold deposits in the Archaean greenstone belts of Western Australia, but only 26 have productions and/or quoted reserves in excess of 10 tonnes Au. Most gold production has come from mineralization in shear zones, quartz veins or stockworks in volcanic or intrusive host rocks, particularly tholeiitic basalts or dolerites metamorphosed at sub-amphibolite facies (e.g. Kalgoorlie). Significant prod- uction has also come from BIF-hosted deposits (e.g. Hill 50), and large reserves are quoted for mafic hosts at high metamorphic grade (e.g. Big Bell). Most deposits have marked structural control, great depth extension, and are characterized by zoned alteration haloes of pyritization (with Au), K-metasomatism and carbonation. On the regional scale, most larger volcanic- or intrusive-hosted deposits occur in younger (ca 2.8 b.y.) rift-phase green- stones (Norseman-Wiluna Belt). These belts also have a high prospectivity (36 kg Au/km2 of greenstone belt). Younger (ca 3.0 b.y.) and older (ca 3.5 b.y.) platform-phase green stones of the Murchison and Sodthern Cross Provinces and the east Pilbara, respectively, are progressively less prospective (14 kg Au/km2 and 0.8 kg Au/km2, respectively) and also have progressively smaller deposits. The important regional- and local-scale characteristics of most gold deposits can be explained in terms of a metamorphic- replacement model. In this model, metamorphic H2O-CO2 fluids, focussed by faults and shear zones, infiltrated host rocks and deposit gold at 350 + 50¦C and 1-2 kb, largely in response to fluid-wallrock interaction. In many large deposits, gold precipitation was synchronous with sulphidation of Fe-rich host rocks, but the diversity and abundance of smaller deposits reflects the variety of reactions precipitating gold from ubiquitous, auriferous metamorphic fluids. Regional variations in abundance and size of deposits relate to factors such as tectonic setting, alteration styles and possibly source rocks.

Jan 1, 1986

By P C. Couper, P C. Keall, S J. Mathews, A Francis

An epithermal gold-silver deposit is presently being developed at the Golden Cross Mine in the Waitekauri Valley, 8 km northwest of Waihi, New Zealand. Highest grade mineralisation underground is mainly confined to an elongate zone of strongly banded quartz veining on the hangingwall. Footwall veining is irregular, flatter lying, and of lower grade. Hydrothermal breccias in the ore zone may have economic grades where clasts of vein quartz are present. Jigsaw breccias, shingle breccias, and layered breccias with sedimentary type structures such as layering, crossbedding, and graded bedding have been identified proximal to, and within, quartz veins. Mineralisation in the open pit occurs in and adjacent to stockwork quartz veining.   Three alteration zones are recognised: a central silicic zone enclosing Au-Ag bearing quartz veins, an argillic zone above the stockwork zone and in the footwall of the underground ore zone, and a propylitic zone which is gradational with argillic zone. Intense shearing is confined to the less competent argillic zone.   Precious metal mineralogy consists of electrum (65% Au; 35% Ag) and argentiferous tetrahedrite, polybasite, pyrargyrite, argentite, and native silver. Pyrite is the dominant sulphide mineral. Marcasite, arsenopyrite, and rare chalcopyrite and sphalerite also occur. Oxidation ranges from a few metres in the argillic western stockwork zone to 100 m in fractured rock in the footwall of the stockwork.   The open pit contains a proved and probable reserve of 2.1 million tonne grading 3.5 g/t Au and 17 g/t Ag. The underground mine will produce 2.7 million tonne grading 7.8g/t Au and 27g/t Ag.

Jan 1, 1991

By Quast K. B

The ore used in this test program was a composite sample from Waihi, New Zealand containing 2.7 g/t gold and 36 g/t silver. From the testwork it was found that a high concentration of sodium cyanide (0.1 per cent) was necessary to ensure a high recovery of both gold and silver. There appeared little chance for either differential leaching or adsorption of gold and silver. The addition of bentonite clay decreased the rates of gold and silver dissolution. The addition of 50 g/1 NaCI had little influence on the rate of silver extraction. The addition of 5 g/l carbon gave adsorptions of 70 per cent of the gold and 35 per cent of the silver, indicating the possible presence of base metal cyanides in solution. The addition of excess sodium sulphide only precipitated 70 per cent of the silver in solution, adding further evidence to the presence of base metal sulphides. The production of separate silver and gold products was not possible by either preferential leaching, preferential adsorption of gold on carbon or preferential precipitation of silver.

Jan 1, 1991

By M Hassall, E Humphries

Personal risk assessment processes, also known as Take 5 or SLAM (Stop, Look, Assess, Manage) have been utilised in high hazard industries for over 30 years. However, there is limited research to understand if a personal risk assessment process improves safety and health outcomes. The author’s learnings during discussions with training participants in G3/G-MIRM (Risk Management Systems) and G2 (Risk Assessment for Supervisors) courses have provided them with significant insight into understanding if a personal risk assessment process is valued across a number of mining industry organisations. These discussions arise when the four-layer model is presented (Figure 1) where ‘Layer 4’ represents the last resort risk checking activity. In this context, training participants agree that this process has value. However, the consistent opinion is that the personal risk assessment process is not effective, that it is a ‘tick and flick exercise’ and is not achieving the desired outcomes.

Apr 16, 2024

The Porphyry Gold Mine (150km NE of Kalgoorlie) was reopened in November 1983 with production commencing in April 1984. Original reserves quoted for the open pit were 822000 tonnes at 5.2 g/t Au based on wide spaced diamond drill holes. Close spaced grade control drilling has revealed some loss of tonnage and a significantly lower grade. In the area of the pit mined to 30 June 1985 this grade control reserve estimate contained 67% of the gold in the original ore reserve with the grade being 19% lower. It is anticipated that the final grade control reserve estimate will contain around 80-85% of the gold contained in the original reserve. There has been close agreement between the Mill Head grade (4.19 g/t Au) and the grade control predicted grade (4.2 g/t Au) confirming that dilution is not a major problem. After initial poor agreement (attributed to "lock-up" in the circuit) the bullion calculated head grade steadied out at 93% of the grade control prediction. This difference is considered to represent the geostatistical bias in the drill hole/crushed rock sampling. Gold recovery has been in the range 93-95% (compared with bullion calculated grade). The net result is that the recovered grade of ore mined to 30 June 1985 is 70% of the original ore reserve grade.

Jan 1, 1986

Samples weighing O· 2 g are digested with 2 ml of 72 per cent perchloric acid for 2 hr at 180°C, and then diluted to 10 ml. with water. After settling, an aliquot of clear solution is reduced with ascorbic-citric acid, and then a zinc dithiol solution is added to form a coloured molybdenum complex which is extracted with petroleum spirit for comparison with standards.Concentrations down to 2 p.p.m. may be analysed at rates up to 150 determinations per day.

Jan 1, 1970

By Morley C

The Mount Charlotte gold mine is located 0.75 km east of the General Post Office in Hannan Street, Kalgoorlie WA and about 3.5 km north of the Fimiston Operations on the famous Golden Mile.Mechanised underground bulk-mining at Mount Charlotte has been in continuous operation since 1963. For the 30-year period ending December 1992, the mine produced 22 million tonnes (Mt) of ore for a recovery of 2.73 million ounces. During 1991/92, the mine produced 1.53 Mt of ore at a head grade of 3.15 gramshonne (g/t) for a recovery of 135 822 ounces.The principal ore source at Mount Charlotte has been the Charlotte Orebody which is now largely mined out. Since 1990, approximately 75 per cent of production has been sourced from the Reward Orebodies. The Charlotte Orebody was a relatively homogeneous quartz stockwork with simple geometry and average bulked grades in the range of 4.0 to 5.5 g/t. This is in contrast to the Reward Orebodies and the more recently defined Charlotte Deeps and Maritana Stockworks, which have complex grade distribution and geometry with lower average bulked grades (2 to 4 g/t). The lower grades are principally due to the decreased density and volume of stockwork venation and related vein halo alteration. Typically, these stockworks comprise a patchy mosaic of quartz veins and vein swarms, with irregular distribution and geometry within the quartz gabbro host rock.

Jan 1, 1993

The Que River Mine is a small under-ground base and precious metals mine located in north-western Tasmania. Current production rate is 220,000 tonne/yr from reserves grading 0.37% copper, 7.1% lead, 12.2% zinc, 195 g/tonne silver and 3.4 g/tonne gold.Ore is mined using longhole open stoping techniques. Cemented fill is placed in primary stopes to permit safe and efficient recovery of pillars. The operation is trackless with both shaft and decline access from surface.Ore produced is trucked 40 km by road to be sold to the Electrolytic Zinc Company of Australia Ltd., for treatment at their Rosebery concentrator.

Jan 1, 1985