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Four data sets are required to put constraints on the tectonic evolution of the southern rim of the Pacific Ocean basinI . data bearing on the dispersal of the major Gondwanaland continental fragments and the birth of the Southern Ocean,2. data bearing on the interaction of the Pacific Ocean basin lithosphere with the South American, Antarctic, and Australasian continental masses,3. data bearing on.the evolution of the Scotia Sea floor, and 4. data bearing on the evolution of the various continents, microcontinents. and volcanic arcs.

Jan 1, 1990

The variation of austenitic grain size with temperature for the steels XKSlSO and XK9261 was determined using the vacuum ctching technique.It is shown that the effect of temperature on the mean grain intercept is relatively small, up to approximately 950°C. Above this temperature the rate of increase of grain size rises sharply.Within the limits of the 1-4 hr vacuum etching times used the effect of time at temperature is relatively small on the ASTM grain size of the steels.

Jan 1, 1967

By K White, C Ferguson, M Harding

Understanding watercourses and river processes at a landscape scale is vital for managing risk to mining infrastructure and operations as well as minimising future environmental impacts beyond the life of the mine. Recognising this, Iluka Resources Limited engaged Alluvium Consulting Australia to identify and characterise watercourses across several neighbouring mineral tenements in the Eucla Basin, South Australia (Alluvium, 2013, 2014a, 2014b). The objective of the studies was to inform planning for risk management during mine planning and operation and appropriate landscape rehabilitation for mine closure.Mining operations in the area are by open cut extraction of mineral sands. The extracted resource is processed on-site before leaving the site for export. Several pits are to be developed in stages with simultaneous backfilling and surface reinstatement. Mining is initially taking place in the Jacinth tenement, where up to 100 km of watercourses may be directly or indirectly impacted (refer to Figure 1).The total study area (shown in Figure 1) is approximately 330 km2 and spans three smaller areas that were the focus of separate investigations during mine planning.CITATION:Ferguson, C, White, K, Harding, M and Hardie, R, 2016. Landscape-scale assessment of watercourses in remote arid South Australia to inform mine planning to mine closure, in Proceedings Life-of-Mine 2016 Conference, pp 200–204 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jun 28, 2016

By B Manda, K L. Sandvik, K Musiwa

Mining and processing of pyrite ore at Iron Duke Mine, north east of Zimbabwe, generates large quantities of acidic mine water as a result of the oxidation of iron sulfides on exposure to air and water used in the drilling and blasting operations. The acid mine drainage (AMD) contains elevated levels of dissolved Fe, Al, Mn and sulfate and has characteristic pH values ranging from about 0.50 to 3.00. Handling and disposal of such a drainage places significant impacts on the economics of the mining operation due to the corrosive nature of the drainage on infrastructure, the limitations placed on water reuse and discharge to the environment under a new law governing effluent discharge. Treated sludge at Iron Duke is chemically unstable, changing pH from ~7 to ~3 within four days after deposition in the evaporation ponds. The effluent does not meet the permitted level of neutral pH, 500 mg/l sulfate and less than 2000 mg/l total dissolved solids. Excessive acid seepage from the evaporation ponds has resulted in contamination of the ground water below the ponds area, which shows high mean annual levels of iron (368.67 mg/l), sulfate (1702.89 mg/l) and low pH (2.86) in addition to other pollutants. This poor quality seepage infiltrates into the nearby Yellow Jacket River, contaminating the public water source. Furthermore, a deterioration of the land area and vegetation adjacent to the tailings dump is also observed. Understanding the controls on the neutralisation of acidic drainage and long-term sludge stability after placement in evaporation ponds is essential to minimise costs of closure requirements and for the protection of environment and human health. This study examined the effects of the master variables pH, redox potential, acidity and alkalinity on acid neutralisation. The results indicate that ferrous hydroxides are not as stable as ferric hydroxides and that proper oxidation of iron also leads to better sludge stability, higher treatment efficiency and a manageable sludge viscosity. Metal hydrolysis and secondary mineral formation at the evaporation ponds under a wide range of pH and redox conditions is responsible for equilibrium shifts in metal reactions, resulting in high potential for metal leaching and mobility. Mechanisms of metal attenuation are also discussed.

Jan 1, 2006

By Allnutt S. L, Radojkovic A. M

Century, a major shale hosted zinc deposit in the Riversleigh Fold Zone of northwest Queensland's Mount Isa Basin, was discovered by CRA Exploration in 1989. A total of 322 drill holes, completed by the end of 1992, define a resource of 118 Mt of mineralisation averaging 10.2 per cent zinc, 1.5 per cent lead and 36 grams per tonne silver. Mineralisation is developed in the upper portion of Lawn Hill Formation Unit 4 (Pmh4). The intersection of the Termite Range and Little Range Faults, major regional structures, may have been important in controlling mineralisation development. Three listric normal faults, and an unconformity between Proterozoic rocks hosting mineralisation and overlying Cambrian carbonates, divide the deposit into three distinct blocks. The deposit is essentially blind, outcropping in only two places. The deposit is characterised by an extremely simple sulphide mineral assemblage, sphalerite being the most common sulphide phase. Pyrite, the most common phase in many other shale-hosted zinc-lead deposits, accounts for only five per cent of all sulphides present. Siderite is the principal non-sulphide phase present. Both sphalerite and galena have near stoichiometric compositions. Sulphides are confined to carbonaceous shale horizons. Intervening siltstones are essentially barren. Shale and siltstone beds forming the mineralised sequence can be correlated across the entire deposit area. Systematic grade variation occurs within each mineralised shale bed. Although sulphide laminae parallel primary lithologic layering, the point of most intense mineralisation development transgresses stratigraphy. Bituminous material is intimately associated with sulphide mineralisation.

Jan 1, 1993

Historical exploration of the Ballarat East goldfield defined two primary quartz vein lode types. Tracking the intersection of these lodes with vertical lines of enrichment (or so-called æindicatorsÆ) was the basic practice for historical mining. This strategy proved highly successful in terms of identifying quartz veins associated with mineralisation. However, it provides little understanding of the timing or processes involved in the formation of these veins, or the associated emplacement of gold. Current underground mapping shows the Ballarat East gold deposit can be subdivided into four stages of vein formation. The earliest veins (V1 and V2) strike north-south and formed during bedding flexural slip and limb thrust fault propagation in parasitic chevron folds. V3 veins are subhorizontal and relate to further fold tightening involving a transition from east-west compression to north-south extension that initiated re-shear of bedded slip surfaces and additional strike-slip faults. Sporadic V4 veins overprint the shallow-dipping sections of V2 and V3 veins and are associated with weak north-south compression. The V4 veins mineralogy is distinguished by a high carbonate content and sphalerite-galena sulfide assemblage. Vein sampling indicates a significant increase in gold mineralisation where overprinting by the later deformation/veining has occurred. As a result of these observations, a new model is emerging that identifies the chronology, structural control and spatial distribution of respective quartz vein types. This is a major advance in terms of exploration targeting, resource modelling and mine planning.

Jan 1, 2008

By D Craw, C Smith, C Martin

The Wangaloa coal mine closed in 1989, and an attempt to grow Pinus radiata on quartz-rich overburden dumps was not successful. The site is currently being rehabilitated with re-profiling and native plant establishment. Plants are being established in a range of substrates that consist of variably disaggregated Tertiary coal measure sediments derived from the mine sequence. This material has low organic matter content (typically <10 %), originating from coal fines or carbonaceous sediment. Substrate samples were collected during December 2003 from a range of sites varying in age, slope, and parent material. Nutrient analysis includes nitrate, ammonia, total N, total P, available P and readily exchangeable base cations. Total nitrogen and phosphorus are low on areas of recently disturbed, quartz-rich overburden, with total N =1200 mg/kg, and total P =200 mg/kg. Background forested areas have total N ranging between 2800 - 3300 mg/kg and total P of 420 - 500 mg/kg. Quartz-rich overburden sites contain available nitrogen levels of less than 6 mg/kg, while background forested areas are generally >20 mg/kg. The soils and water runoff are generally acid, with pH down to three. Water samples have been collected from both ground and surface resources on a monthly basis since December 2003. The main lake on the site had pH of 4.6 in 1999, and this remained relatively unchanged when 850 tonnes of lime were added in 2002. However, the lake pH changed to 6.4 within a one month period in 2004, and has fluctuated since that time.

Jan 1, 2004

Every text book for blasting in Japan says that Hauser's law can be applied for practice. But recently it came true that there was no historical record about him and no actual datum of his own or other's to testify. According to his theory, the funnel ratio N = r/W in crater test is inversely proportional to the rock strength and proportional to the explosive amount: where r is the radius of crater and W is the burden. But experiment showed the opposite result, and revealed new information. Two kinds of experiment, crater test and boulder blasting of mortar, were carried out to investigate the effects of rock properties on blasting. The material used for test 1 and test 2 was a mortar block (60cm)3 cube, 4 vertical faces of which were used for crater test. The results from Test 1 and test 2 were summarized as follows. The funnel ratio ranges between 2.5 to 4.5 in proportion to rock strength and burden. The standard ratio of 1, reported by Hauser, was not observed in any instances. It was also noted that the funnel ratio is independent of the explosive amount. The material used for boulder blasting, test 3, was a mortar block (40cm)3 cube, in the middle of which was located a vertical borehole for the explosive charge. The results of test 3 were as follows. The necessary explosive amount for boulder blasting is in inverse proportion to the mortar strength. Fragmentation and throw are proportional to the mortar strength.

Jan 1, 1990

By Smith N. S, Rholl S. A, Otterness R. E

The Bureau of Mines is conducting research to develop blast designs for the control of fragmentation and rock displacement and the reduction of overall mining costs. This Bureau experiment, a program of 25 reduced-scale blasts conducted at the University of Missouri-Rolla's experimental quarry, determined the effects on fragmentation of changes in burden, spacing, explosive diameter, and stemming-explosive length. Combined with previous work at this experimental quarry examining air gap, delay, and decoupling effects, this experiment provided a data base of over 50 shots. The fragmented rock, 1000 to 3000 kg from each 3- to 4-hole blast in the 1 m bench of massive dolomite, was completely screened to assess fragmentation. The explosive used for all the tests was extra dynamite initiated by seismic caps fired from a sequential blasting machine with delays of 5.2- to 13-ms/m of burden. Empirical equations were developed which predicted the effect of shot design parameters on the fragmentation. These formulas predicted the average fragment size of each of the 25-shots to within 13 percent. For spacing to burden ratios of 1, 1.4 and 2, there were optimum burden to explosive diameter ratios (B/de) of 30, 26 and 20, respectively. Decreasing the B/d. ratio beyond these levels resulted in little or no reduction in the average size. For B/de ratios above these levels the average size increased but, at least a 15 percent reduction in the average size results when the shot design parameters are optimized.

Jan 1, 1990

The Cadjebut Mine is located 80 km southeast of Fitzroy Crossing. Cadjebut is owned 58 per cent by BHP Minerals and 42 per cent by Billiton Australia. Cadjebut is a 'Mississippi Valley Type' (MVT) deposit. Minerals present are sphalerite, galena and marcasite in a sedimentary host of dolomite and siltstone. The ore contains few deleterious impurities and no silver. The mill was commissioned in the first half of 1988 at a rate of 320 000 tonnes per annum. An expansion in the latter half of 1989 has since increased production to 550 000 tonnes per annum. The initial geological resource at the commencement of mining was 3.8 million tonnes at 13.5 per cent Zn and 4.4 per cent Pb. There is sufficient ore left to sustain operations to 1995 at a grade of 11.2 per cent Zn and 3.9 per cent Pb. The ore from underground is stockpiled on the surface via a drift conveyor and from the stockpile is fed into a grinding circuit consisting of a SAG and ball mill. Cyclone overflow at a Pgp of 90 pm gravitates to the flotation circuit. Lead and zinc are sequentially floated to produce an 80 per cent Pb grade at 90 per cent recovery and a 61 - 62 per cent Zn grade at 96 per cent recovery. The concentrates are thickened and filtered before being trucked to the port of Wyndham 600 km away. From Wyndham the concentrates are shipped to smelters in Asia, Europe and Australia.

Jan 1, 1992

By Frazer E. J, Lwin T

Measurements of coulombic efficiency (QE) for zinc electrodeposition were carried out under mass-transfer-controlled conditions using a rotating disc electrode in synthetic acidic zinc sulphate solutions. Various methods of solution preparation and purifi- cation were investigated using reagents vary- ing from analytical grade to high purity; the purest solution was obtained-by dissolving 99.9999% zinc (Koch-Light) in sulphuric acid solution (BDH Aristar). Improvements in sol- ution purity did not necessarily increase QE, especially where apparently beneficial impur- ities (e.g., Pb) were initially present. The maximum QE obtainable through variation of temperature (T), current density-(i) and elec- trode rotation rate was established using a statistical optimization technique. The QE was determined principally by T and i, with significant interaction between these factors in the highest purity solution. QE contours, which gave an overall view of solution per- formance, were constructed and predictions about the response under conditions similar to plant practice were made. Wark's Rule (the dependence of QE on zinc/acid ratio) was obeyed in the purest solution over the range of composition commonly encountered in zinc plants. High concentrations of Na2SO4 and K2SO4, which occur in some plant solutions, were found to increase QE marginally ("1.2%). The intrinsic effects of individual impurities such as Cu and Ni were evaluated by additions (up to 5 ppm) to the purest solution. Decreases in QE were found to be minimal at the impurity levels normally encountered in zinc plants.

Jan 1, 1987

By N Javati, G D. Carman, D Kikiha, B Likia, K Dala, J McPhie, R Johnstone, D R. Cooke, J L. Blackwell

The 40 Moz Ladolam gold deposit on Lihir Island, Papua New Guinea is the worldÆs largest, low-sulfidation, epithermal gold deposit, in terms of contained gold. The deposit has average grades of 2.42 g/t Au (LGL Resources Reserves Update, 2008) and gold is hosted primarily in refractory pyrite. Lihir Island is part of the 150 km long, Tabar-Lihir-Tanga-Feni, alkalic, volcanic island chain located in the New Ireland basin. Five volcanic blocks comprise Lihir; Plio-Pliestocene Huniho, Kinami and Luise volcanic edifices are built on top of remnant Pleistocene Wurtol Wedge and Londolovit Block (Wallace et al, 1983). The Ladolam gold deposit is located in the central portion of the dissected Luise volcanic edifice and is composed of four main orebodies: 1. Minifie, 2. Lienetz, 3. Coastal, and 4. Kapit. An active geothermal system is overprinted on the ore deposit, localised along north-trending lineaments, and is expressed as hot mud pools and fumaroles. The focus of open pit mapping and drill core logging, as part of ongoing PhD research, has been in the Minifie and Lienetz orebodies. The deposit geology is grouped into four main phases: 1. volcano-sedimentary host rocks, 2. porphyry-style biotite alteration and related hydrothermal breccias and veins, 3. epithermal-style breccias and veins, and 4. modern geothermal alteration. Gold is associated with epithermal-style alteration and mineralisation. Ongoing observations of the Minifie and Lienetz orebodies are highlighting differences in the geology and paragenetic histories. An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation.

Jan 1, 2008

The Martha Mine is an open cut operation situated in the midst of Waihi town which has been operating since 1987. Since then it has produced some 700,000 ounces (oz) of gold and 4,000,000 oz of silver from an epithermal vein system. It was previously worked by underground methods from 1880 to 1952 to a depth of 600 m producing 5 million ounces (Moz) of gold and 30 Moz of silver. The licensed reserves are scheduled to be mined out early in the year 2001. An additional economic resource of over 750,000 oz of gold has been identified which can be accessed by deepening the pit a further 50 m along the main Martha vein. This would extend the mine life to the year 2007. Applications were lodged on 4 July 1997 for the necessary resource consents, discharge permits and variations for the current Mining Licence conditions. The extension project requires approval from the District Council, Regional Council and the Ministry of Commerce. The consent and permitting process is extensive because of the urban setting of the mine. An approach to keep the extension as similar as possible to the current operations was adopted to give the community and authority confidence in the proposed extension. This paper summarises the extension detail and the permitting requirements it has had to address despite the extension being "more of the same". The permitting process has been exhaustive with little recognition of the standards already achieved in protecting the environment which are recognised worldwide.

Jan 1, 1997

Gold mine production in South America was 247.3 tonnes in 1995 up from 156.4 tonnes ten years earlier in 1986. This represents a 58 per cent increase and amounted to 11 per cent of total world gold mine production. The major countries involved in order of production, were: Brazil (67.4 tonnes), Peru (51.5 tonnes),. Chile (48.5 tonnes), Colombia (24.1 tonnes), Venezuela (17.1 tonnes), Bolivia (16.0 tonnes), Ecuador (9.6 tonnes), Guyana (8.8 tonnes), French Guiana (3.3 tonnes) and Argentina (one tonne). The most dramatic increase was reported in Peru where production increased from 10.9 tonnes in 1986 to 51.5 tonnes in 1995, heavily influenced by the production from Minera Yanacocha which began operations in 1993 and produced 17.2 tonnes in 1995.With the reformation of the mining laws and the introduction of investment incentives in all the South American countries in recent years, dozens of major and junior mining companies began flocking into the area in the early-1990s to conduct exploration. A recent study undertaken in Brazil shows that an average of $240 million per year will be expended in this country during the period 1993 - 2010. Twenty-five per cent of this expenditure is expected to be made on gold exploration. Similar figures for Peru are $220 million total per year with 50 per cent for gold; $170 million per year in Argentina with 25 per cent for gold and $120 million ner year in Chile also with 25 Der cent for gold.

Jan 1, 1997

Titanium tetrachloride may be reduced to trichloride with hydrogen on a hot tungsten filament and other hot: surfaces. The effect of temperature, time and size of t.he filament on the rate of reaction has been studied. A laboratory unit for preparing a few pounds of trichloride at a time is described. The yield of trichloride is 2·2 - 2·4 gm/sq. cm. of hot surface/hour at 1,000°C. and the energy requirement 3'5-4 kWh/pound of trichloride produced.'INTRODUCTIONThe lower chlorides of titanium (titanium tri- and dichlorides) are amenable to electrolysis, being electrovalent in nature. Furthermore, relative to other oxygen-free titanium salts, they are easy to prepare in a pure state and their use has been investigated to prepare titanium. The experimentaJ work described in this paper deals with the preparation of titanium trichloride by reduction of titanium tetrachloride with hydrogen on hot surfaces.REDUCING AGENTS FOR TITANIUM TETRACHLORIDEThe free energy diagram (Fig. 1) for the formation of chlorides suggests that titanium' tetrachloride may be reduced to trichloride by various reducing agents such as aluminium, zinc, arsenic, antimony, tin, mercury, silver, hydrogen, sodium, potassium and magnesium. Ruff and Neumann (1923), Thorpe (1885) , Billy and Brasseur (1935) and Seurbert and Schmidt (1892) used various metals to reduce tetrachloride to trichloride and found that the yields were invariably low and the products difficult to separate.Besides the metallic reducing agents, hydrogen may also be used to reduce titanium tetrachloride to trichloride.Goerges and Stahler (1909) obtained a deposit of violet...

Jan 1, 1956

A broker relies heavily on mineral valuations and independent experts' reports to provide the fundamental information, both technical and commercial, which provides the basis of estimating the sharemarket value of securities and underlying assets. It is important to appreciate the significant variance that can occur between the market value and other types of valuation at a particular point in time. In general, most experts' reports and valuations seek to provide estimates of technical/fundamental or ""fair"" value which correspond to the Corporations Law and accounting standards requirements. They do not usually attempt to estimate (stock) market values and it is usually and quite correctly the broker's task to estimate market values on the basis of a technical appraisal and consideration of all other commercial, market and financial influences. Therefore ideally, an independent experts' report or information memorandum will include material relating to commercial and management aspects of a company as well as detailed technical reviews and fundamental valuations. Wherever sensible cash flow projections can be made, cash flow-derived valuations, usually based on discounted cash flow (DCF) calculations, remain the preferred valuation methodology. Brokers in general would not argue with this preference. However, there is a paucity of examples of valuations where alternative methods are applied as a check on the reasonableness of a DCF-based estimate.

Jan 1, 1994

THE reduction of iron from the ferric to the ferrous state prior to its titration by permanganate or dichromate may give erroneous results in the presence of other reducible metals such as titanium. Stannous chloride and metallic zinc both reduce tetravalent titanium to a lower form, and so increase the consumption of the oxidizing standard: solution. Sulphuretted hydrogen and sulphurous acid do not reduce titanium, but are slow in action and troublesome to manipulate.Sulphuretted hydrogen is considered to be the best reducing agent for iron in the presence of titanium.* In the method to be described the advantages of the H 2S reduction are retained without the disadvantages of using the gas. The reduction is carried out by the action of a metallic sulphide having a solubility product of such a value that the sulphide iom, are utilized as fast as they are formed, and but little free sulphuretted hydrogen is given off.Cadmium sulphide was the first to be tried, but its solubility product (3.6 x 10 - 2~)t was apparently too low. The reduction was satisfactory, but the excess was very difficult to remove. Zinc sulphide (S.P. = 1.2 x 10 - 23) was found to be quite satisfactoryin both respects. The zinc sulphide is preferably used in the form of an emulsion prepared from pure zinc sulphate as follows: About 100 grm. of ZnS047H20 is dissolved in 2 litres of water. H2S is passed until no further precipitate is obtained, the acid formed being barely neutralized from time to time by the addition of ammonia. The precipitation of the sulphide in a slightly acid solution gives a finer emulsion than that prepared in an alkaline solution, and any iron present remains dissolved. The precipitate is allowed to settle and the clear liquid decanted. The precipitate is then shaken up with more water, acidified with a little sulphuric acid, allowed to...

Jan 1, 1922

By P Louden, C Rothnie, G Showers

Tiwest employs several methods to assess the quality of heavy mineral concentrates in ore and process samples. Mineralogy is determined using a method called MA98, developed in conjunction with JKMRC over the last three years (Zhang and Whiten, 2001). A range of other techniques are also used to characterise ore, heavy mineral concentrates and mineral products. MA98 is the culmination of efforts since 1989 to determine mineralogy from chemical analysis without recourse to heavy liquid fractionation, which has traditionally been used to discriminate between light-heavy , zircon, rutile and leucoxene. The major advance with MA98 is the use of statistical mass balancing for all chemical species, as opposed to reliance on key chemical species for each mineral. Three magnetic fractions are analysed by X-ray fluorescence spectrometry, with a minerals balance performed for each. Heavy mineral concentrates processed by this method include drill samples, process samples from the wet concentrators and Dry Mill, and mineral products. TBE can be avoided for higher heavy mineral grade samples since quartz is determined from SiO2 balance. In this sense the method is sufficiently universal and robust to cope with a wide range of sample types and compositions. The MA98 algorithm provides abundant information in addition to mineralogy, including TiO2 and Fe2O3 grades of ilmenite and leucoxene, and concentrations of weakly magnetic zircon and kyanite. Information such as this is invaluable in tracking different minerals through the processing plants. Other ore assessment procedures used by Tiwest provide for modelling of orebodies by mineral sizings, Fe in zircon, and FeO and U+Th in ilmenite. Microscopy is used to provide important qualitative information such as grain shape, surface staining and degree of cementing. These quality measurements are included with mineralogy in geological ore models and are available for mine scheduling. In using the above methods to characterise heavy mineral concentrates, Tiwest is better placed to predict and monitor plant performance and product quality from any ore zone.

Jan 1, 2001

By R Engle, R P. McMahan, D A. Bowman, S Schafrik, W Thabet

Preventing injuries and fatalities related to conveyor system use in the mining industry is important and increases uptime while decreasing lost time. The research presented here is funded by the National Institute for Occupational Safety and Health (NIOSH), and intended to accomplish this prevention by the use virtual environments (VEs) to improve training and help miners better understand the hazards of working around conveyor systems. Two conveyor system VEs designed to teach miners about hazards and to safely complete tasks around conveyor systems were developed for this project. The first conveyor system VE, called the instructional tour, gives the miner a semi-automated tour of a conveyor system and teaches the miner about possible hazards (eg, missing guarding) and how to safely fix conveyor problems (eg, a stuck idler) using lock-out, tag-out procedures. The second, called the virtual shift, allows the miner to freely navigate around a conveyor system to detect and avoid hazards. The miner is presented with equipment failures that need to be safely repaired during the virtual shift. If the miner fails to identify and avoid a hazard or fix a failure in a safe manner, an animation of the consequences is shown to the miner, who must start the virtual shift over. The models for the conveyor system VEs were created using 3DS Max. Desktop versions of the VEs were then created using Deep CreatorÖ, an authoring application for creating interactive 3D environments. Later, more realistic versions of the VEs were created for CAVEÖ, a room-sized visualisation system, using DIVERSE, an application programming interface. In initial evaluations of the conveyor system VEs, mining colleagues have indicated great interest in such training tools. According to their feedback, these and similar VEs would provide excellent alternatives to normal classroom training and serve as positive precursors to hands-on experience. These VEs provide new miners a chance to experience hazards associated with working around conveyor systems without being in possible danger.

Jan 1, 2008

By R Minnitt

Two methods of calibrating the parameters K and Alpha for use in Gy’s formula for the fundamental sampling error were investigated. Approximately 480 kg of broken ore from Harmony’s Target gold mine in South Africa was used to compile the replicate samples for two experiments. The duplicate series analysis (DSA) and segregation free analysis (SFA) methods differ in terms of their approach to recovering the sized materials required for the calibration procedure. In this experiment the DSA method split off six more-or-less equal masses of material from a 75 kg lot that was successively crushed to finer and finer fragment sizes. The sublots so formed were split into 32 samples and fire assayed for gold. The SFA method used approximately 250 kg of broken ore crushed to 95 per cent passing 2.50 cm, which was then passed through successively finer screens until a sublot of less than 0.0075 cm was obtained. Each of the 15 sublots obtained in this way were split using a riffle splitter into 32 samples that were fire assayed. A total of 672 fire assays were used in these experiments. The sampling parameters K and Alpha were used for the compilation of a nomogram and to calculate the gold liberation size of the ores. While there are no definitive results in terms of the actual gold grain sizes against which the calculated sizes can be compared, it appears that the DSA method produces more acceptable results.CITATION:Minnitt, R, 2014. A comparison of two methods for calculating the constants K and alpha in a broken ore, in Proceedings Sampling 2014 , pp 165–178 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 29, 2014