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A good deal of experimental work has been done from time to time on the application of compressed air to the elevation of wet pulp. On account of the flat nature of most of our mill sites elevation of the pulp has to be provided for, and the various methods in vogue show considerable loss of time, in the replaceing of wearing parts. During the last two years the writer has carried out a great many experiments with a view of devising an elevator that will give continuous work with a minimum of wear.The results do not show a high efficiency for the power employed, but the lift is continuous in operation, very cheaply installed, and possesses no wearing parts. The lift was tried working in a bore hole of 8 in. diameter, but in many cases it is more convenient to sink a small well for the purpose. The result to date showed that the most efficiency was obtained when the depth of the wellwas not less than the height of the lift required. The pressure of air required in pounds per square inch was (approximately) half the number of feet to be lifted.In the majority of cases the lift required varied between 20 ft. and 50 ft. and the air pressure required between 10 lbs. and 25 lhs. per square inch. In most existing installations the air compressors in use were delivering air to the receivers at about four times that pressure, and when air so compressed was expanded to perform its work at a reduced pressure, it was apparent that the power exerted in origInally compressing the air above the working pressure required at the lift was absolutely lost.

Jan 1, 1901

Study on Selective Reduction and Magnetic Separation of Low-Grade Nickel Laterite Ore to Produce High Nickel Concentrate

Sep 13, 2010

The Significance of Internal Recycle in Froth Flotation

Sep 13, 2010

Retreating Cu-Mo Rougher Tailing by Flotation for Concentration of Copper and Molybdenum

Sep 13, 2010

By P J. Mienie

The Palaeo-Proterozoic Transvaal Supergroup in the Northern Cape Province of South Africa is host to the largest known resources of high-grade hematite ore on the Southern African continent. These ores are currently being exploited at Sishen and Beeshoek. Hematite occurances on the Sishen South exploration project (some 65 km south of Sishen mine), are at present the focus of an intensive investigation. At both the Sishen and Sishen South deposits, mineralisation is found within lithologies belonging to the Transvaal and Oliphantshoek Supergroups. The Transvaal Supergroup comprises a basal, carbonate platform sequence (the Campbell Rand Subgroup) overlain by thick units of banded iron-formation belonging to the Asbestos Hills Subgroup. The Asbestos Hills Subgroup is unconformably overlain by sedimentary rocks of the Gamagara Subgroup. The latter Subgroup has been correlated with the lowermost unit of the Oliphantshoek Supergroup. The Sishen and Sishen South deposits are located close to the western margin of the Kaapvaal craton and have been subjected to intensive, structural deformation. Both localities have undergone folding, faulting and thrusting. Early regional tectonism (2400 - 1700 Ma) played a critical role in the genesis and preservation of the ores from younger erosive events. At both deposits, four similar iron ore occurrences have been recognised. The ores have developed in specific depositional environments and hence display unique physical, chemical and metallurgical properties:The bulk of the high-grade hematite comprises laminated and massive ores, which are found in the upper parts of the Asbestos Hills Subgroup. The laminated and massive ores are best developed in basinal and pseudo-graben structures. Collapse breccia ores of medium- and lower-grade are preserved within deep palaeosinkholes. The sinkholes have developed within the dolomites of the Campbellrand Subgroup. Lower-grade conglomeratic and gritty ores constitute the remainder of the deposits. These ores are found in the overlying clastic succession of conglomerates, shales and quartzites belonging to the Gamagara Subgroup. Although, there are many stratigraphic and lithological similarities between the Sishen and Sishen South deposits, there are also important differences, which will impact on the profitable exploitation of the ores. The most striking contrast between the deposits is their size and geometry. Sishen mine is a single pit operation with an opencast mineable reserve of 895 Mt. The Sishen orebody is thicker, more continous and significantly larger than the Sishen South deposit. The Sishen South deposit has a potential opencast reserve of 228 Mt. It comprises a number of small and narrow, isolated orebodies, spread over an area of ¦64 km2 . There is a significant difference in grade between the deposits. On average the orebodies of the Sishen South deposit have a higher iron content (+0.25 per cent Fe), less potassium (- 0.090 per cent K2O) and less phosphorus (-0.019 per cent P). The Sishen South deposit has less interbedded shale in the laminated and massive ores. The footwall of these high-grade ores is a light-density, intrusive sill, compared to banded iron-formation at Sishen. With a proposed heavy-medium separation plant at Sishen South, the impact of dilution on product grades is expected to be less than that currently experienced at Sishen.

Jan 1, 2002

The final criterion of the payability of gold production must necessarily be the cost per ounce. While the cost per ton or per cubic yard is a convenient comparative measure, and a necessary one for operating control, it is obviously insufficient as a means of judging whether or not, in a given set of circumstances, a gold production operation can be carried out economically.. An alternative method, which brings in the ton or cubic yard as a unit of measurement, is to determine what the cost per ton or per cubic yard is equivalent to, in terms of pennyweights or grains at the ruling price of gold. But this is really only another way of stating that the cost per unit of product is the ultimate measure of profit or loss, and not the cost per unit of input material.It is accordingly of interest to determine the cost per ounce of Australian gold production and to compare it with the cost of gold production in some other parts of the world which are important gold producers. The attached Appendix, which it, is believed comprises practically all Australian mines of any consequence (except Big Bell and the de Bernales group, for which. insufficient information is available), shows this information for financial years ending between 31st December, 1938, and 31st December, 1939, the production of gold dealt with being some 1,084,000 oz., or 66% of the total Australian production for 1939. For comparative purposes the cost of producing gold at representative mines in New Guinea, New Zealand, Fiji, Canada, India, West Africa and Rhodesia, and the average cost of production at the Rand mines, has also been included.It is obviously necessary in a study of this sort that the basis of the cost should be clearly set out and should be as nearly as possible uniform. The "cost" as calculated herein comprises the following items:

Jan 1, 1940

Treatment of Gold-Bearing Oxide High-Clayish Ores The analysis of modern state of treatment technologies of gold-bearing oxide high-clayish ores showed that the main method applied both in Russia and abroad is sorption leaching and heap leaching with preliminary granulating. The use of any of these processes depends on many factors including average gold grade, site and area of a deposit, clay content and etc. The two processes are often used at the same deposit. In this case, rich ores are treated according to a plant technology, poor ores and ores with cut-off grade are treated with heap leaching. Both technologies possess their own advantages and disadvantages. A plant technology is characterised by high recovery with higher capital and operating costs at the same time. Lower recovery during heap leaching can be attributed to an unstable quality of ore granulating due to both technological and weather conditions. In case of poor high-clayish ores, the use of a combined ore treatment is possible and profitable. It includes the ore disintegration and a clayish fraction treatment by using a heap leaching. In this case, capital and operating costs reduce in comparison with a plant technology and the recovery increases in comparison with heap leaching. Sometimes it reaches a plant recovery. These ores possess special features. As a rule, gold is associated with ore fraction, so it liberates readily from the association with ore ore-forming minerals even with relatively coarse grinding explain their high solution velocity in a hydrometallurgical cycle. Clayish ores are characterised by high sorption activity to gold cyanide complexes as well as extremely low velocity of filtration and thickening of slimic pulps at the same time. The paper describes the options of oxidised high-clayish ores treatment, such as, a plant technology (PIP and CIP process), heap leaching and a combined technology (disintegration, heap leaching of a crystalline fraction and sorption cyanidation). The obtained results were used when selecting a treatment technology and during construction of a plant for oxidised high-clayish ores treatment.

Sep 13, 2010

The prime objective of geometallurgy is to improve the profitability of mines through the use of spatial models of rock properties that have a significant impact on value. Although a key property is the grade of the component of interest, because it directly influences the revenue from saleable material, increasingly attention is being focused on other attributes. These 'non-grade' variables may have a less direct, but equally profound impact on value, either by reducing potential revenue (eg lowering recovery, deleterious elements, etc) or increasing mining or treatment costs (eg comminution, grindability, throughput, reagent use, etc). Collectively, these variables have been described as 'geometallurgical'. There are three main areas of activity required to realise the value of geometallurgy: 1. acquisition of appropriate data through sampling and measurement, 2. building spatial models of geometallurgical variables, and 3. using these models to optimise value. Each of these is built on the foundation of the preceding step. In building reliable spatial models, it is critical that the amount of error introduced to data by the sampling and measurement processes is understood and quantified, in order that the inherent variability of the component of interest can also be properly understood. Management of the quality of grade variables is now applied (with different levels of diligence and success) across nearly all sectors of the mining industry, not least because of the necessity to adhere to Codes of practice (eg JORC, 2004). The necessity for quality management and at least the broad framework of what to do in practice is reasonably understood by many relevant professionals. There are also a number of tools, including software products, statistical analyses and presentation formats, available for those managing grade data. It is also abundantly clear, that in many cases people have little understanding of why they are doing particular quality control activities, or how to sensibly interpret the results. The same is not true for many geometallurgical variables. In almost all instances the authors have seen, the magnitude of sampling and analytical/measurement error variance embedded in geometallurgical test data is untested and unknown. This is partly because until relatively recently, metallurgical data has mainly been collected for the purpose of process and mill design, with few, expensive tests carried out on 'average' samples of the orebody. Increasingly however, there is a move towards collection of much larger data sets from spatially distributed sampling, and thus a growing necessity to understand the errors inherent in both sampling and analysis of these types of variables, and how to manage them. In order to understand how the quality management practices applied to grade variables can be extended to geometallurgical variables, it is essential to understand the theoretical foundation that supports these practices. This requires an analysis of the concepts that underpin the 'principles of additivity' (Matheron, 1963), the 'theory of sampling' (Gy, 1982, Pitard, 1993), and 'statistical process control' (Shewhart, 1930). Relevant concepts are then integrated to form a foundation to develop methods to improve the quality management of geometallurgical variables. Application of this framework will significantly increase the integrity of geometallurgical data, the reliability of geometallurgical models and ultimately the value derived from implementing geometallurgy. This is an ABSTRACT ONLY no paper was prepared for this presentation.

Mar 1, 2010

By D M. Frantois-Bongarton

Variograms can be powerful tools for the analysis and control of process streams and their sampling. Their uses and proper utilisation have been described somewhat unclearly to date, and their limitations and precautions of use are too rarely mentioned. The general meaning and properties of the variogram tool and its use in geostatistics are therefore reviewed in details to derive a critical view of its proper use for process control: what to expect from it and what it cannot deliver and which uses are illusory or potentially misleading. A summary of GyÆs seminal work accompanies this essay to put the concepts into historical perspective.CITATION:Frantois-Bongarton, D M, 2012. Review of variogram properties and remarks on its use in process control, in Proceedings Sampling 2012 , pp 175-180 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 21, 2012

By M J. Andrews, P J. White, I Bogie

Active geothermal systems occur widely in Sumatra and in the Taupo Volcanic Zone (TVZ) ofNew Zealand. There are also many occurrences of epithermal gold mineralisation in Sumatra, some of which are economic. Minor sub-economic occurrences of gold have been reported in the TVZ but none have yet been mined. There are both similarities and differences in the nature of the active geothermal systems and mineral deposits in Sumatra and New Zealand, which can be systematically related to the tectonic setting. The thicker crust and consequent predominance of silicic volcanism in New Zealand leads to hydrothermal systems which are little eroded and with limited lateral outflows. The systems in Sumatra are of mixed character, with some of the characteristics of those in andesitic island arcs such as Java or the Philippines, e.g. a higher proportion of systems with identifiable acid magmatic volatiles, a greater potential for deep erosion and lateral outflows than in New Zealand; but modified by the more oblique component of subduction and thickened crust so that some are localised in volcano-tectonic depressions and consequently have high piezometric levels of primary hydrothermal fluid. The unusual occurrence of calcic zeolites in veins associated with gold mineralisation in Sumatra is thought to be indicative of æsatelliteÆ boiling zones along de-gassed hydrothermal outflows.

Jan 1, 1995

At the Orieco mine quartz veins carrying sulphides were localized in parts of an extensive fault zone in the Mathinna Beds during late Middle Palaeozoic mineralization.These deposits were subsequently exposed, and oxidation of the sulphides by percolating groundwaters resulted in the formation of gossans at the surface, and replacement close to the permanent water table of the more soluble sulphides by copper sulphides (chalcocite, coveIlite). A restricted zone containing copper oxides (cuprite, tenorite), copper carbonates (malachite, azurite), iron arsenates (scorodite, pharmacosiderite) and possibly rare copper sulphates occurred above the water table.The renewed oxidation of these deposits following mine development has been facilitated by the ready circulation of air and water combined with the formation of large free surfaces ideal for evaporation of aqueous solutions. Under these new conditions of high initial sulphate concentration, low pH, and relatively high Eh, the previously formed mineral assemblage is unstable and numerous Cu (Fe,Zn)-sulphates and Fe (Cu,Pb)-arsenates are presently forming in fractures in the fault zone and on the adit and stope surfaces for some considerable distance from the Sulphide-rich ore shoots. The observed spatial sequence of the major sulphates with respect to the sulphide-rich zones (Le. chalcanthite-antleritebrochantite and langite) is consistent with decreasing sulphate concentration and increasing pH away from the oxidizing sulphides.

Jan 1, 1970

The Reefton Goldfield on the West Coast of the South Island of New Zealand was a significant gold producing area between 1860 and 1951. During this period it produced approximately 2.1 million ounces (67 tonnes) of gold. The largest single producer on the goldfield and the deepest mine in New Zealand, was the Blackwater Mine at Waiuta which produced 21 tonnes of gold. The Prolubition Shaft reached a depth of 878 metres (317 metres below sea level). The mine closed suddenly in 1951 with about 66 000 tonnes of are remaining in the stapes. Since closure there has been no attempt to determine the ore potential of the mine at depth. This paper summarises the geology, history and resource potential of the mine and sets out two options for evaluating the remaining resource. The options are either deep drilling from surface or recovery of the Prohibition Shaft and underground development and drilling. A schedule for each option is presented and discussed.

Jan 1, 1992

Process Simulation and Plant Design of a Mineral Selective Magnetising Flash Roasting Plant Using Neutrons

Sep 13, 2010

By L Forrester

The tube bundle system is an integral part of an underground coalmine’s gas monitoring system. The maintenance of these systems is vital to ensure the quality of the data produced. Integrity testing of the tube bundle lines is required monthly, as specified in Standards Australia AS/NZ 2290.3 (1990). The various forms of integrity testing are reviewed, including pressure testing and introduction of a known gas. This paper provides a recommended and standardised methodology for those techniques where none currently exists. A strengths, weaknesses, opportunities, threats (SWOT) analysis is performed. The parameters considered for the SWOT analysis include resourcing required by the mines, time for the testing to be performed, and quality of the results produced. The suitability of such integrity testing protocols to be used in a mine emergency situation is also evaluated.CITATION:Forrester, L, 2017. Tube bundle integrity testing methodologies, in Proceedings Australian Mine Vent Conference 2017, pp 129–136 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 28, 2017

Selective Recovery of Platinum from Cobalt-Rich Crust

Sep 13, 2010

Stripping of Fe (III) from D2EHPA Using Different Reagents The common way to extract Fe(III) from an organic solvent is by using concentrated acidic solutions. The stripping of Fe(III) from organic solvent Di-2-ethylhexyl phosphoric acid (D2EHPA) by hydrochloric acid (HCl) solutions has been found to increase with increasing the acid concentration. However, the use of a concentrated acidic solution such as 6 M HCl can degrade the organic molecules. In this research, to increase the efficiency of Fe(III) stripping from D2EHPA, effect of some acids, salts and buffer solutions was studied. Influence of Tri-butyl phosphate (TBP) as modifier on the Fe(III) stripping operation was also investigated. According to the results, the best Fe stripping recovery was achieved in the presence of low acid concentration and the salts which was 72.4 per cent oxalic acid, 70 per cent ammonium oxalate and 75 per cent buffer of oxalic acid and ammonium oxalate. The results also showed a better recovery in the presence of the buffer rather than the salt. Finally, increasing the concentration of TBP, HCl and H2SO4 resulted in a higher stripping efficiency of Fe(III). However, in the presence of oxalic acid and its salts increasing the reagent concentration did not increase the stripping efficiency and at this condition, the recovery of iron remained constant.

Sep 13, 2010

By C A. Locke

Geophysical signatures of the Taranaki and Tongariro volcanoes reflect similarities in their age, composition and hydrothermal histories, but also reflect differences in host rock settings and edifice evolution. The Taranaki volcanoes have strong positive gravity anomalies showing that their edifices largely consist of dense andesite which reflects substantial lava production or dyke injection during construction and that large sub-edifice dyke/stock complexes occur. In contrast, the Tongariro volcanoes have strong negative gravity anomalies indicating low density edifices of volcaniclastic material; no sub-edifice intrusions can be identified, probably because of the lack of density contrast with host basement rocks. Strong positive magnetic anomalies over the younger cones can mainly be attributed to near-surface andesite lavas or dykes whereas subdued magnetic effects are associated with hydrothermally altered zones within both edifices and sub-edifice intrusions.  

Jan 1, 1995

This method determines tellurium at concentrations down to 0·01 p.p.m., with geochemical accuracy. Tellurium is volatilized from the rock at 800°C in the presence of sulphur from decomposing pyrite, condensed, dissolved, and finally determined by atomic absorption in a long tube burner. The distribution of Te is shown near some gold lodes at Kalgoorlie. Lastly, a method of synthesizing Te-free pyrite from sulphur and ferrous sulphate is described.

Jan 1, 1969

Particles of a complex nickel sulphide concentrate (containing pyrite, pyrrhotite, pentIandite, and violarite as the major mineralogical constituents) were entrained in a stream of hot gas the temperature of which was varied from 500 to -1000°C. Chemical analysis of the particulates collected from the furnace was used to detect the rapid change in sulphur content which accompanied the ignition temperature jump. X-ray diffraction analysis was used to characterize the oxidized material.These techniques showed that ignition was dependent on the mineralogical composition and occurred over the temperature range 500-900°C. Pyrite ignited first in the range 500-600°C; pyrrhotite was the least reactivemineral and required temperatures in excess of 800°C for ignition. The nickel minerals <pentIandite and violarite decomposed before ignition and ignited at temperatures intermediate to those required for the pyrite and pyrrhotite. The ignition temperature for the pyrite was found to be independent of particle size and concentration of oxygen in the atmosphere.Ignition temperatures determined by the simulated flash smelting method were compared with the results of differential thermal analysis (DTA) measurements performed on the same concentrate. Temperaturesdetermined in the DTA test as corresponding to the start of exothermic activity were shown to be at least 100°C below those determined by the simulation test.

Jan 1, 1978

Underground equipment owners and users in order to compete and succeed in being viable business propositions must continually balance costs with productivity. By reducing their hourly ownership, operation and maintenance costs and increasing their hourly production, the cost per unit of material mined will be optimized.

Jan 1, 1990