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By Mech B

This paper gives a general description of mining practice at The Zinc Corporation Ltd., Broken Hill, New South Wales.The basic mining methods are those which have been common in Broken Hill for many years and have been well described in previous Proceedings. The main object of this paper is to indicate developments which have occurred in recent years and to describe changes in layout, mining methods and equipment which are taking place at present.The following headings have been adopted: History Geology and general characteristics of the orebodies, Geology, The lead lode, The zinc lode, General characteristics, Ore reserves, General mining method, Exploration and diamond drilling, Shafts and shaft sinking, Main shaft, Old main shaft, Block 5 shaft, New upcast airway, Development, Level interval, Level layout, Driving and crosscutting, Winzing, Rising, and Stope preparation.

Jan 1, 1946

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

This paper discusses support requirements for the proposed monorail system to be used in decline development. The monorail drilling and loading systems are systems that move on the rail (monorail) installed in the roof of the decline and supported by roof bolts, suspension chains and steel supports. However, due to the weight of the components of the two systems, it is imperative that the force in each roof bolt, suspension chain and steel support capable of suspending the weight of the heaviest component is determined. Numerical models that relate the weight of the monorail drilling and loading components to the required strength in the support system have been developed. Using these developed models, numerical values of the forces in each roof bolt, suspension chain and steel support, required to suspend the weight of the heaviest component of the monorail drilling and loading systems are determined.

Dec 6, 2010

Forecasting, particularly when applied to meaningful issues, is known irreverently as 'futurology'. There is quite a range of opinion ext.ant in the world when mineral resources are looked at. The opinions range from the doomsday view that there will be acute shortages of all essential commodities, to the optimistic outlook that new 'discoveries will prevent any possibility of shortages. The environment is, as yet, only studied locally and quite parochially; surprisingly there is little integration when considering the spatial atmosphere, the seas in depth and the earth's crust.

Jan 1, 1981

By T S. Giraudo

The Macraes Gold Project, operated by GRD Macraes Limited, engaged GRD Minproc Limited to assist with the evaluation of a pressure oxidation circuit, this being the key component of a recovery enhancement project. The initial study gave sufficient confidence in the process to enter into project execution. GRD Macraes Limited was the first company to license Newmont Mining Corporation's pressure oxidation technology, which aims to improve gold recovery from the carbonaceous sulfidic flotation concentrate. This paper outlines the fundamental chemistry of the process followed by a description of the plant that was ultimately installed. A summary of the commissioning is given together with an overview of the production results to date. The project ramp-up was rapid, with design temperature and throughput being achieved within days of hot commissioning. Gold recoveries have exceeded expectations, improving from around 70% prior to the installation of pressure oxidation to above 84%. The project has increased the nominal gold output at Macraes from 120,000 oz/y to 170,000 oz/y. On-going plant optimisation for process parameters is discussed, giving rise to flexibility in gold production, with key controlling parameters exceeding design.

Jan 1, 2000

By M P. Simpson

Forty one whole rock samples from the Gold Cross low-sulfidation epithermal Au-Ag deposit have d18O values that range from 4.4 to 9.3 per mil, with an average value of 7.0 per mil. Unaltered and weakly altered rocks have d18O values greater than eight per mil, and the orebody is surrounded by samples that are depleted in 18O. A strongly silicified sample adjacent to the Empire Vein System has a d18O value of 9.0 per mil, similar to previously reported analyses of vein quartz (7.0 to 11.7ë, average 9.4ë). This suggests that, in detail, Golden Cross may have a zone of 18O-enriched wall rocks in the core of the deposit, adjacent to the main underground veins. Although some workers have suggested that stable isotope geochemistry may provide useful information for epithermal mineral deposit exploration, at Golden Cross this is not the case. Alteration minerals, major elements and trace elements all define larger, less ambiguous halos than the zone of 18O-depleted wall rocks.

Jan 1, 2001

By Ward B

Central Colliery was developed to exploit large underground reserves in the German Creek lease. It was planned from the outset as a retreating longwall operation and is the first longwall mine in Queensland. Longwall mining is not new to Australia but all previous experience has been confined to the coalfields of New South Wales. There was no precedent for longwall extraction in the markedly different geological setting of the Bowen Basin in Central Queensland. Various investigations were thus carried out during the initial development of Central Colliery, prior to the installation of the longwall, to ascertain the geotechnical characteristics of the coal seam and the adjacent strata and to predict the performance of the heading and pillar design as mining proceeds to depth. Laboratory tests showed the coal to be weak with an unconfined compressive strength of 9.6 MN/m2. Site inspection and geophysical correlation revealed a shear zone and extensive intraformational shear planes in the seam. The latter were found to exert a controlling influence on rib reaction in the headings. This behaviour was predicted by finite element modelling of the development headings which indicated that mining induced stresses in the coal would be relieved by block movement of the ribs. A chain pillar width versus depth curve was established for yielding pillar design based on empirical and computational methods. This gave pillar widths from 25m to 48m over the proposed depth range of 150m to 300m. Finite element modelling of the 25m pillars chosen for the first gateroads confirmed the overall stability of the design but suggested that rib spall could be significant in the adjacent tailgate. (1) Chief Mining Engineer, Capricorn Coal Management Pty. Ltd., German Creek Mine, Queensland (2) Senior Engineering Geologist, Coal Division, The Shell Company of Australia Limited, Melbourne

Jan 1, 1986

By Y Chen, G Xu, B Albijanic

High concentrations of airborne coal particles is the main reason leading to health and safety hazards in underground coalmines. Water spray is one of the effective methods to suppress coal particles in mine sites, but the use of plain water is not effective enough to capture coal particles. By adding surfactants, coal wettability can be improved which will increase the coal dust suppression efficiency. In this study, a modified sink test was employed to evaluate coal wettability in the presence of two anionic surfactants: sodium dodecyl sulfate (SDS) and sodium dodecyl benzene sulfonate (SDBS). Sink rates were calculated by measuring sink weight over test time. Different surfactant concentrations (0.01 per cent, 0.05 per cent, 0.1 per cent, 0.2 per cent and 0.4 per cent) and two types of water (tap water and deionised water) were selected to investigate coal wettability over three size ranges (0–38, 53–75 and 75–90 µm). The results show that both SDS and SDBS surfactants can lower the water surface tension and improve coal wetting performance. Sink rates increased as the surfactant concentration increased, and surfactant solutions prepared by tap water performed better than deionised water in enhancing coal wettability. Also, coal wettability varied for different coal particle sizes; the smaller coal particles were, the worse the wettability they possessed.CITATION:Chen, Y, Xu, G, Wang, R and Albijanic, B, 2017. Study on coal wettability for dust suppression in the presence of surfactants by sink test, in Proceedings Australian Mine Vent Conference 2017, pp 105–110 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 28, 2017

Background information, formulae and a systematic set of procedures for estimating capital and operating costs for open pit mining equipment are presented in this paper. Firstly, equipment capital costs are required. These can be obtained from either suppliers, contractors or from first principles. The equipment is then divided into cost elements. Typically these are fuel and power requirements, spare parts consumption, wear items such as tyres, teeth and drill bits and finally repair and operating labour requirements. By consideration of the cost and life of each element a base cost can be derived. The base cost is then adjusted to suit specific site conditions such as the proximity to spare parts suppliers, the duty or ground conditions and the management and operator skills specific to the site. Hence, the same item of equipment can be costed to reflect different operations. Included are tables and data for typical plant items. Several examples are also given to illustrate the methodology being presented. The intention is to provide a set of procedures to allow costs to be built up from base level in a systematic and repeatable way. Applications for these costs include feasibility studies, budgetting and data for economic analyses.

Jan 1, 1990

Open pit mine plans define the complex strategy of displacement of ore and waste over the mine life. Various mixed integer linear programming (MILP) formulations have been used for production scheduling of open pit mines. The main problem with these models is the inability to solve real size mining problems. We present a mixed integer linear programming formulation for the open pit long-term production scheduling problem. To reduce the number of binary integer variables in the formulation, we aggregate blocks into larger units referred to as mining-cuts. We present a semi-supervised clustering algorithm which aggregates smaller blocks based on rock types and grade ranges into mining-cuts without introducing an error in the grade estimated values. The cluster-based aggregation algorithm is fully integrated into an MILP mine production scheduling prototype software. We have implemented, verified, and validated our models. The efficiency of the new platform is evaluated through iron ore mine case study.

Dec 6, 2010

By S M. Zhang, Y X. Han, J Liu, Y J. Li

"Compared with conventional roasting and magnetic separation processes, higher reduction temperatures and stronger reducing atmospheres are used in reduction technologies to reduce iron oxides to metallic iron, which is often liberated during reduction and amenable to magnetic separation. Reduction and magnetic separation of an Australian refractory iron ore was therefore studied as a function of reduction temperature, time and coal addition. It was found that during reduction, the particle size of the reduced iron increased with reduction temperature and time. Subsequent magnetic separation tests were further conducted on the reduction products. A reduced iron powder with a Fe grade of 85.67 per cent was obtained at a Fe recovery of 87.82 per cent under optimum reduction conditions where a metallisation rate of 85.82 per cent was achieved. The mechanism involved in reduction is discussed based on XRD and SEM examinations.CITATION:Han, Y X, Liu, J, Li, Y J and Zhang, S M, 2013. Upgrading an Australian iron ore by reduction and efficient magnetic separation, in Proceedings Iron Ore 2013, pp 397-402 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Aug 12, 2013

By T R. Gibbons, I Grayand

Geotechnically massive strata may lead to a host of problems during longwall mining. These include windblast on initial goaf fall, sudden loadings on powered supports, periodic weightings, fractures that may propagate to surface, and extreme subsidence problems. It is essential to identify these problems prior to mining and to appropriately deal with them. The problem is essentially that the rock mass does not breakup evenly and when it does so it tends to produce large blocks which may suddenly move. The rock types that produce these problems tend to be massive sandstones, massive siltstones and sometimes igneous sills. The fundamental problem with these is that they do not breakup readily. This is because the ratio of stress they are subject to compared to their strength is low. Thus, the shallower mines with lower stresses are more subject to this problem than deep highly stressed ones with similar strength rock.

Nov 29, 2022

By Parker A. J, Giles D. E

Solutions of copper(II) sulphate in acidified aqueous acetonitrile leach copper from copper-lead matte derived from the drossing of lead with sulphur to produce solutions of copper(I) sulphate. The reactivity of the matte constituents towards copper(II) sulphate is Cu2S> PbS > CuS > > Pb. According to the leach conditions, part or ,all of the copper may be leached 'leaving a residue of CuS/PbS, or S/CuS/PbS04, or S/PbS04, to be recycled to the lead smelter. Complete extraction of the copper requires leaching with oxygen which occurs under mild conditions in the presence of acetonitrile and sulphuric acid in water. Copper is selectively recovered as a pure product by electrochemical or thermal disproportionation of copper(I) sulphate solutions. A notional process is proposed.

Jan 1, 1979

In gold ores, the range of value disclosed by ordinary mine sampling is very wide, and the number of 'samples required for a reliable average is correspondingly great. Even where a large number of samples is available the weighted mean value of the samplesalmost invariably exceeds the true gold content. In operating mines this disparity is overcome by arbitrary reduction of high assays until reconciliation is achieved, but it is frequently necessary, especIallyin exploration, to base an estimate or to reach a decision on a relatively small number of samples from underground openings or even on a few drill hole intersections. In these circumstances, a decisive influence may be exercised by the weight allowed to high assays and to low grade intersections, both of which events are, of course, dependent on the metal distribution in the ore. Where sufficient data are available, statistical analysis will measure the risk involved, but as this information is seldom to be had, the final choice becomes largely a matter of personal judgment.With this problem in mind, a study was made some years ago of the distribution of values in four Western Australian gold ores, with the primary purpose of learning from actual ore-bodies of differing types, something of the metal distribution in the smallquantities of ore involved in individual routine mine samples, and of the importance in each case of high assays. Adequate and suitable data were available in .four cases, in three of which high assays were of frequent occurrence due to the presence of free gold. The results give concrete expression to well known differences of metal distribution, and it is proposed, firstly to record the observed distribution; secondly to examine some of the consequences of the differences in the distributions limd thirdly, to comment, in the light of this data, on the vexed question of the treatment of high assays. Standard statistical procedure was employed in the compilation of the data. The assays were classified by value into groups, each group covering a range of value of 1 dwt. per ton. Assays of less than 1.0 dwt. per ton form the first group; assays equal to or greater than 1.0 dwt. and less than 2.0 dwt. comprise the second group; those equal to or greater than 2.0 dwt. and less than 3.0 dwt. appear in the third group; and so on. Where the number of samples is large, it can be assumed that the mid-value of the group is the mean value of all the samples in the group, thus the mean values of the first three groups are 0.5 dwt., 1.5 dwt. and 2.5 dwt. respectively.The number of samples in the group is the frequency of the group, but for convenience of plotting this frequency is converted into percentage of all the samples in the series. Plotting percentage frequency as ordinate against dwt./ton as abscissa results in a frequency curve-a curve showing the frequency...

Jan 1, 1950

It's long been suspected that there has been an increase in the incidence of premature failure of rock bolts in underground coal mines over the past ten years. As far back as 2002 a study found a possible contributor to such premature failure in some Australian coal mines might be linked to stress corrosion cracking (SCC) (Crosky et al, 2002). That study suggested the low impact toughness could be a major influence on the susceptibility to SCC of different steel compositions. A subsequent study involving laboratory corrosion tests however failed to confirm any link with metallurgical factors (Villalba and Atrens, 2007), a finding that is at odds with field experience. Some mine sites today are still reporting prematurely failed rock bolts to the rock bolt manufacturers; the failures are being identified at SCC induced. To this end a research project has recently been commissioned at University of New South Wales (UNSW), which aims to identify the environmental factors causing SCC, through field test sites and simulating the conditions within an accelerated corrosion laboratory. This paper outlines the methodology that will be used in this study.

Nov 25, 2010

By S Winter, G Klenowski

During longwall mining in the Bowen Basin, open tension cracks formed by overburden subsidence occur to a height of about 175 m above the mining horizon. Within this zone, water in aquifers or inundated workings will flow into the mine.At the Oaky Creek mines Aquila seam, open cut pits containing pit water overlie German Creek seam longwall panels at a height of about 105 m. The highly saline water is classified as contaminated and cannot be discharged off-site. Because alternative water storage facilities were not available, innovative water management solutions were needed to allow longwall mining to proceed without underground flooding.Computer modelling and underground inflow measurements have been completed to accurately determine inflow rates into longwall mines from overlying water sources at increasing overburden heights. Research work has also included determining cost-effective blanket treatment of final voids and reshaping spoil piles to minimise downward seepage.Ponded water overlying the Aquila low, Grasstree south and A6 pits required remedial works prior to longwall undermining to prevent flooding. In the Aquila low pit it was necessary to partition the pit and remove water overlying longwall 32. Two in-pit bridges were constructed by dozing and track compacting low wall spoil. The final void above the longwall panel was kept dry during undermining. Maximum recorded underground inflow rate was 30.0 L/s.The Grasstree south pit final void, part of the ramp and much of the spoil piles were inundated prior to undermining of 308 panel. A bridge and blanket were required to prevent ponding above the longwall. Maximum recorded inflow during undermining was 23.2 L/s.The pit A6 south final void, which contained dumped, saturated rejects and inundated highwall mining entries was undermined by longwall 36 at the Oaky No 1 underground mine. Because mapping of highwall entries was not available, the maximum calculated inflow rate of 73.0 L/s was based on all design entries being fully inundated. The actual inflow rate during longwall mining ranged up to 20.0 L/s.Water management improvements that have been implemented include greater use of contaminated pit water to wash coal, use of evaporation fans, and installation of a 6 ML/day reverse osmosis plant for water purification.CITATION:Klenowski, G and Winter, S, 2017. Inflows into longwall mines from inundated overlying workings, in Proceedings 13th AusIMM Underground Operators’ Conference 2017, pp 183–190 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Oct 16, 2017

By Arthur A. F, MacKenzie I. H

Copper in zinc/lead sinter charged to the Imperial Smelting Furnace operated by Sulphide Corporation Pty Limited at Cockle Creek, New South Wales, concentrates in a complex lead dross which forms as the bullion is cooled to around 400°C. The 0.35 mm portion of this dross, containing approximatelY 67 per cent of the total copper, is treated via a hydrometallurgical route to produce 5000 tonne/yr of CuSO4 5H2O crystal.Around 3000 tonne/yr of copper-lead dross containing 49 per cent copper is leached at 55°C in continuous leach reactors with ammoniacal ammonium carbonate leach liquors to convert copper to soluble copper amine complexes. Leach ettlciencies of 80-85 per cent are obtained. Leach residues are repulped with ammoniacal liquid diluents from the plant and heated in a four stage steam distillation unit to reduce ammonia levels to between 0.5 and 1.0 g/Iitre. The resultant slurry is directed to an effluent treatment plant for final treatment, and separation and return of solids to the sinter plant.Copper is extracted from leach liquor by solvent extraction techniques based on the use of L1X 54 diluted 3:1 with Shellsol 2046 to give a nominal copper loading of 30 g/litre. The loaded organic is contacted with a sulfuric acid solution of pH 5 to remove the bulk of ammonia and zinc, then stripped at 55°C with a 60 g/Iitre H2SO4 95 g/Iitre copper solution to yield a 15 g/Iitre H2SO4 125 g/Iitre copper solution.This liquor is fed to an adiabatic vacuum crvstalliser operating at 35°C and 5 kPa vacuum where crystals of CuSO4 5H2O are formed. Crystals are separated from mother liquor and washed in a two stage pusher centrifuge, then dried in a hot air rotary drier. The crystal product obtained is of a high chemical purity and readily conforms to both Australian and overseas standards.

Jan 1, 1983

By A T. Job, M L. Edgar, P R. McAree

Characterising ores in-pit and in real-time presents a significant value opportunity for the mining industry. Hyperspectral imaging systems are proposed as an enabling technology to achieve this real-time characterisation. This paper considers the application of a hyperspectral imaging spectrometer mounted on a shovel. We define the performance requirements for such a spectrometer and associated subsystems during a typical dig-load cycle. The relatively high swing rotation speed of the shovel requires approximately video-rate exposure times to achieve good spatial resolution. We analyse a selection of components against these requirements and find that suitable commercial spectrometers, detectors and optics appear fit-for-purpose. Automated, real-time spectral classification systems are also available, with only relatively modest requirements for processing power. The paper argues that at near peak solar illumination, a system for field deployment at the required sensitivity is viable using only commercial off-the-shelf components.CITATION:Job, A T, Edgar, M L and McAree, P R, 2017. Real-time shovel-mounted coal or ore sensing, in Proceedings Iron Ore 2017, pp 397–406 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 24, 2017

The Stawell Goldfield, the westernmost of the main Victorian fields, has produced over a million and a quarter ounces of gold from one and three-quarter million tons of ore. The lodes are quartz reefs and sulphide bodies occurring in sediments of lower Ordovician or upper Cambrian age. The lodes are confined to a relatively narrow suite of weak rocks lying between stronger ones above and below and there is a definite relationship between fold structure, pitch and strength of folding and ore.LOCATION HISTORY AND PAST PRODUCTIONThe town of Stawell, situated 150 miles north-westerly from Melbourne on the main railway and road to Adelaide, is the most westerly of the main gold areas of Victoria. Alluvial gold was found at Pleasant Creek (now Stawell West) in 1853 and the usual "rush" followed. Numerous alluvial leads, some surface, others buried under up to 100 ft. of Tertiary sediments were worked. The alluvial gold was soon exhausted, but in 1857 gold-bearing reefs were found on the slopes of a prominent ridge on the northern edgeof the present town. Early reef mining was by small working parties mining to shallow depths only and treating their ore at public crushing plants. Some of the early returns were very high grade. The next stage was the amalgamation of claims and the formation of companies, with their own treatment plants. This stage continued till the last mine closed in 1921. These companies had, by present standards, small holdings and far too many shafts were sunk, resulting in less efficient prospecting and operation.Since 1921 there have been one or two small scale attempts to revive mining but without success. Of recent years Gold...

Jan 1, 1960

The reasons for selecting two stage pebble milling in preference to conventional methods of grinding are discussed. Commissioning of the first section was completed in October, 1969. Improvements in both the operating and engineering aspects were then incorporated in the section commiss ioned in March, 1971.

Jan 1, 1975