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A description of:1. Air-lock at No. 14 Upcast shaft.2. Pit-top gates at Dudley winding shaft.3. Safety rider employed in sinking No. 18 shaft.1. AIR-LOCK AT No. 14 SHAFT.Some time ago it became desirable to raise and lower the McBride tunnel men to save them walking the inclined tunnel, about a mile in length, with a dip of nearly 600 ft., and a maximum gradient of 1 in 4. The main Upcast shaft, No. 14, was, therefore, equipped for winding the men. As this shaft was the main ventilating shaft for the McBride tunnel mine, it was necessary to take the greatest precautions against leakage of air. The writer therefore devised the concrete structure shown in Plate 1.The structure, as shown, surrounds the shaft inside the poppet legs. As the men are wound on double-decked cages (twelve men to each deck) provision is made at both top and bottom for the two decks to be loaded and emptied simultaneously. The doorsover the sky shaft, shown in the drawing, have been cut out and the sky shaft has been close-timbered with a double layer of T. and G. flooring boards, with building paper between, nearly up to the thimble frame, a small hole only being left for the winding rope topass through. The space between the ramps is glazed with plate glass to allow the winding-engine driver a good view of the pit top.The men are wound from two levels, 302 ft. and 611 ft.2. PIT-TOP GATES AT DUDLEY WINDING SHAFTTo minimise the shock to cage, couplings and rope when winding quickly, the writer devised the apparatus and arrangement shown in Plate n.

Jan 1, 1928

To provide suitable backfill material for the cut and fill mining method being used to extract ore from the C.S.A. mine, flotation tailings are deslimed in single stage cyclones. The sand fraction is stored on the surface in unagitated storage bins before being pumped underground, while the slime portion is thickened and pumped to a residue dam. Design and operating details of the sand preparation and storage sections are given.The initial preparation of a stope for filling, consists of laying percolation and decantation drainage pipes on the sill floor. Preparation for fllling of subsequent lifts consists of rearing up the ventilation pipes at the ends of the stope, preparing the decantation system, and building a barricade or wall. The fill is carried to, and discharged into, the stope through 4 in. diameter pipes and filling of any section is carried out 24 hr/day normally without supervision. The fill is retieulated underground to the 1200 ft level through rubber lined pipe installed in the upcast shaft. Boreholes of 3 in. diameter drilled from the1200 ft level convey fill to the 1800 ft level. The mine drainage system is basically designed to handle the water associated with hydraulic filling. A system of primary and secondary sumps has been installed to handle the percolation and decantation drainage from stapes and to allow settling of fines before the water is fed to the main pump station.A programme of investigation into the properties of the emplaced fill is in progress. Results to date indicate that the fill as presently placed and drained, will not cause any problems in the mining of vertical orebodies by cut and fill methods for lifts in excess of 300 ft.

Jan 1, 1969

Sophisticated pit optimisation software has allowed mine planning engineers to generate mining outlines that consider the numerous factors that have a major effect on the success or otherwise of a mine plan. The outlines are generally generated using a range of values for which one can assume a fixed mining cost and a variable revenue. During analysis, the expected costs and revenues are used to investigate the cashflow and sensitivities. Ultimately, the final pit and incremental outlines, or shells, are selected for a given expected price. This paper proposes an alternative method for incremental pit shell selection that considers the risk the mine be prepared to accept. The life of mine is divided into a number of Business Risk Periods (BRPs). The BRP represents the period of time over which the owner is prepared to invest capital. During each BRP, the mine will then operate at a selected investment cost which is generally equal to or greater than the minimum required to meet production targets. In this way, a robust mine plan can be produced which allows the mine to react to change without having to alter the incremental shell limits while maintaining a strategic risk profile. This paper will summarise factors to be considered in selection of the BRP and the incremental pit shells within the BRP.

Jan 1, 1998

By M I. Pownceby, M A. Rhamdhani, S Purohit, G Brooks

The use of concentrated solar flux for high temperature metallurgical applications is a wellresearched area with several reported literatures on solar calcination and metal oxide reduction. This paper presents two such processes with potential to reduce the carbon footprint from ironmaking through using concentrated solar flux. The first process involves solar flux incorporation to the recently developed lime-magnetite pellet (LMP) route of magnetite ore agglomeration. The second process is a hybrid solar-smelting route for iron ore-coal composite pellets that utilises solar flux to provide process heat. Both the concepts of solar agglomeration and solar smelting process were investigated using a solar simulator-hybrid reactor set-up at Swinburne University of Technology. Results from solar agglomeration experiments at 950°C suggested no change in the reaction mechanism of the LMPs under solar irradiation compared to electrical heating. The percentage and morphology of the primary phases (mainly CaFe3O5) appeared similar for both solar and electric heated LMPs. The solar-smelting of magnetite concentrate-lignite coal composite pellets at 1130°C showed a maximum metallisation of 55 per cent after 1.5 hrs of reaction. The solar-smelt process has about 25 to 35 per cent lower CO2 emission compared to the blast furnace ironmaking. Technoeconomic analysis of implementing a CST facility in an existing composite pellet process suggested a payback period of about ten years at a heliostat cost of $A100/m2 and natural gas cost of A$10/GJ. The economics of CST implementation can be further enhanced by decreasing the thermal losses from the heliostat field and reactor.

Nov 8, 2021

Both the miner and the Regional Water Board have legal obligations to ensure that all discharges of water from a mine site are free of harmful wastes, and that proper authorisation is obtained to abstract, divert, or discharge water. This requires long-term planning of water management for the whole of the mine site. Guidance is given in the paper on the issues which must be covered in water right applications.

Jan 1, 1986

A narrow (200-300m) zone of amphibolite facies metamorphic rocks underlies the Papuan ophiolite near the Musa-Kumusi divide, SW of Popondetta, Oro Province. The amphibolite rocks are bounded up-section, to the NE, by peridotite and gabbro of the ophiolite; and down-section, to the SW, by mafic greenschist, which in turn overlies blueschist facies mafic schists of the Emo Metamorphics.Amphibole compositions within the narrow contact zone show a strong chemical zonation which reflects a steep thermal gradient, with maximum temperatures >800¦C at the peridotite contact. The zonation is recorded in Ti content and A-site occupancy, relative to Al'¦

Jan 1, 1994

By S Servant

In the mine of this study, cemented hydraulic backfill is placed under the form of a slurry at a percentage of solid by weight varying from 50 to 65 per cent. This slurry settles and cures in the stope. During this process an important amount of water must be drained from the stope. Field observations in the case studied here show that the geometry of the drains system has a significant impact on the drainage of the stopes. The mode of filling produces a backfill showing an evident stratification typical of a sedimentation process. In situ observations on this material show that this stratification translates into a very strong heterogeneity in the physical properties of the cemented backfill. This heterogeneity should affect the hydraulic conductivity of the backfill in the stope. In order to study this aspect, finite element simulations were performed. This paper presents the results of these simulations, where the impact of this stratification on hydraulic conductivity is related to the performance of the backfill drainage system for various geometry. This preliminary study is aimed at providing a better understanding of the role of the drains in the dewatering process of the backfilled stope in the case of this mine.

Jan 1, 1998

By Parr JM, Scott SD, Herzig PM, Gemmell JB

PACMANUS is an actively-forming seafloor analogue of volcanic-hosted massive sulfide orebodies, situated on the crest of an andesite-dacite-rhyodacite ridge overlying rifted older arc crust in a back-arc basin. Hydrothermal deposits range from Mn-oxide crusts to Cu-Zn-rich massive sulfide mounds and chimneys containing significant Au and Ag. Preliminary data suggest a substantial magmatic component to fluids and metal sources, and we speculate that PACMANUS is the cap to a complete hydrothermal system including subvolcanic and intrusive-related mineralisation.

Jan 1, 1995

Gold mineralisation at the Kerimenge-Lemenge prospects is a classical example of myltiple, overprinting events, associated with intrusive-related hydrothermal systems. Repeated emplacement of high level dacitic porphyry bodies (Edie Porphyry; 2.4 - 3.8 Ma) was accompanied by a phreatomagmatic eruption which formed a diatreme-maar complex. Later hydrothermal fluids utilised permeable fault zones and diatreme contacts. Mineralisation took place during two main events. Stage I consists of early quartz - illite/sericite - carbonate phase with associated pyrite-arsenopyrite mineralisation. This phase is zoned from quartz - sericite-pyrite/arsenopyrite in the south at depth, to dominantly manganocarbonate - illitic clay - arsenopyrite- pyrite_s significant base metals in the north and at shallow levels. The majority of known gold (18.1 Mt @ 2.0 g/t Au) at Kerimenge is associated with Stage I activity. The gold occurs primary within the lattice of the sulphides and is metallurgically refractory. Fluid inclusion studies indicate early boiling and progressive cooling and dilution of hydrothermal fluids during the evolution of the Stage I event. Near paleosurface condensation of gases, which evolved off the boiling system at depth, produced abundant low pH bicarbonate fluids. During the waning phases of Stage I activity, these bicarbonate condensate fluids descended down permeable structures and formed banded manganocarbonate veins. High grade non-refractory gold mineralisation during Stage II event is restricted to favourable structures at relatively shallow levels and in the northern regions of the prospect area. Gold in this event exhibits a distinctive Au - Ag - Te -Cu geochemical signature, and is associated with hessite - tennantite - chalcopyrite mineralisation within quartz veining. Alteration and fluid inclusion studies indicate that this style of mineralisation is a result of the quenching of pulses of hot, late magmatic mineralised fluids by the cold condensate waters.

Jan 1, 1997

By T Casten

Freeport IndonesiaÆs DOZ (Deep Ore Zone) mine is scheduled to commence production in the second half of 2000. The DOZ will produce 25 000 tpd using mechanised block cave mining methods and will have a number of major improvements when compared to previous block cave mines in the district. The DOZ will have an independent and improved ventilation system, a flexible and efficient truck haulage system, and a more reliable crushing and conveying system than in previous underground mines in the district. This paper will describe the mine plan, haulage, and ventilation systems, and compare them to those used in other operations in the district.

Jan 1, 2000

The development of the lower Hunter Valley, centred around Newcastle and its suburbs is closely linked to the mining of coal and the settlements which grew to house the colliery officials, the miners and their families. The Port of Newcastle expanded and developed as a port exporting coal to many ports of the world ranging from Hawaii and Brazil to Japan and Korea. The coal which passed through the Port of Newcastle, depend- ing on the historical period in question may have been mined in Newcastle itself, or one of the nearby colliery townships such as Lambton, Wallsend, Charlestown or Hamilton, or from one of the estuarine collieries located at Carrington or Stockton or from the West Wallsend area or from East Greta, Pelaw Main or Cess- nock located on the South Maitland (Greta) coal field or from the open cut and underground mines of the Singleton Coal District.

Jan 1, 1988

Bulk mining and handling technologies have been affected in two ways by the soaring costs of oil products in recent years. Firstly, the substition of oil by coal, oil sands and oil shale demands the development of high-capacity mining and handling equipment, and secondly the rising fuel costs favour a trend away from cost-intensive heavy truck transport. To cope with the growing volumes of material in mining and handling, larger equip- ment units have to be designed. This tendency becomes apparent from the examples of the Hambach open pit mine in Germany, and the Port Kembla handling plant in Australia. A combination of the flexibility of heavy truck transport and the cost advantages of belt conveyors is achieved by the system of semi-mobile crushing installations described in this paper.

Jan 1, 1981

By H Aoki, A Nakayama, K Morita, S Yamada

The thermal conductivity of mold flux is crucial for optimising the productivity and quality of steels. This study investigates the thermal conductivity of molten oxides, including CaO-BO1.5-AlO1.5 (CBA), CaO-BO1.5-AlO1.5-SiO2 (CBAS), and CaO-MgO-BO1.5-AlO1.5 (CMBA) melts, using a hot wire method. Additionally, their structure is examined using Raman spectrometry and magic angle spinning nuclear magnetic resonance (MAS-NMR) to explore their relationship. Furthermore, first-principles calculations are employed to assess the covalency of each bond in the melts for a more quantitative analysis of their relationship. The study confirms that an increase in the CaO concentration of the CaO-BO1.5-AlO1.5 melts reduces the amount of bridging oxygen and consequently decreases thermal conductivity. The covalency of the B[4]-O bond is highest, especially when the second-nearestneighbour Al atoms transition from Al[6] to Al[4], forming the AlB3O7 structure with B[4]. The relationship between thermal conductivity and the structure of CaO-BO1.5-AlO1.5-SiO2 melts can be quantitatively assessed using the degree of polymerisation of the network structure and the variation of covalency as indicators. Furthermore, replacing CaO with MgO in the CaO-MgO-BO1.5-AlO1.5 melts reveals a mixed alkaline earth effect on thermal conductivity.

Aug 21, 2024

By M I. Pownceby, J Manuel, S Suthers, A Poliakov, E Donskoi, N Ware

"The mining industry is constantly seeking methods for optimising the exploitation of mineral resources, such as iron ore, in a sustainable and eco-efficient manner for economic and social benefit. When faced with resources that have mineralogical or textural complexity, poorer grades or geological variability, it is particularly important to apply a comprehensive geometallurgical approach to the development of the mining operation, from resource definition through to processed product. Integral to this approach is the quantitative characterisation of representative samples across the whole orebody.A key factor that associates the geological characteristics of iron ores to their processing response is ore texture. While the classification of samples on the basis of chemical or physical properties alone is often not sufficient to predict their processing behaviour, classification and quantification of texture types can be effectively correlated to metallurgical performance in processes such as blending, beneficiation, pellet making, sintering and ironmaking. This paper describes some of the research conducted by CSIRO pertaining to the design and rationale of a texture-based classification system for Australian iron ores; some of the tools developed and used to apply such classification; and the application of the generated data to quantifying, modelling and predicting sintering behaviour and performance in the blast furnace.CITATION:Suthers, S, Pownceby, M I, Manuel, J, Ware, N, Donskoi, E and Poliakov, A, 2016. Geometallurgical Characterisation of Australian Iron Ores – from Ore to Processed Product, in Proceedings The Third AusIMM International Geometallurgy Conference (GeoMet) 2016, pp 351–360 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Jun 15, 2016

At Renison Limited, on the West Coast of Tasmania, modern trackless decline mining is practised to exploit the various tin bearing sulphide deposits. With footwall accesses, one of three adaptations of horizontal mechanised cut and fill, 1. unsupported, 2. pillar supported,and 3. unsupported - resuing is used to stope each defined ore block. The stoping method is basically determined by the dip and true thickness of the ore with emphasis on high ore recovery. Extreme flexibility is possible by the selective mining of stope lifts, the consumption of development waste as fill and the utilisation of the same mobile equipment in all phases of production. This has resulted in high productivity and low cost per tonne of ore.

Jan 1, 1977

This paper traces briefly the reasoning which led to the decision to extend the existing internal shaft at the Rosebery Mine and to instal underground a larger winder (No. 2 Hoist) with capacity to handle all hoisting requirements to a depth of approximately 2,500 ft. on an incline varying from 83" to 55" to the horizontal. It then covers the specifications and selection of the winder, the design of the winch chamber, rope raise and bins, the excavation, support and lining of the chambers and the transport to site and installation of the machinery. Finally a brief description is given of the mechanical and electrical parts of the winder.INTRODUCTIONThe existing shaft winder on the internal inclined shaft situated on No. 8 Level of the Rosebery Mine had reached the limit of its capacity. Hoisting 21-ton capacity skips in balance, with a 13/16-in. diameter rope, it required four layers of rope on the 3 ft. 6 in.drums to reach the loading station under No. 12 level at a depth of 864 ft. (Fig. 1). This entailed heavy wear on the multi-layered rope and heavy stresses on the cheeks of the drums. Continuous working on three shifts was necessary to handle ore, men and materials, and a hold-up from any cause whatever meant an interruption to production on the lower levels of the mine. The maximum output of the hoist whilst hoisting from the No. 12 level loading station was 50 tons per hour. It was obvious therefore, that a larger winder was required if output was to be maintained.Diamond drilling had proved the existence of sufficient ore at depths well below No. 12 to warrant exploitation and a decision had to be reached on the means of hoisting this ore.

Jan 1, 1958

By Brathwaite RL

The gold-scheelite-quartz lodes of the Wakamarina valley were the most productive of the mesothermal gold-bear in the Marlborough Schist. Most of the production came from the Empire City and Golden Bar mines which worked tug, Golden Bar lode. Only 134 m of its 1800 m strike length are accessible, in the No. 2 level of the Golden Bar Mine. A fau I has downthrown the Golden Bar Mine section by 2 47 m and displaced it left-laterally by 18 m. The lode is a typical ribbon-banded mesothermal vein, averaging c. 1.75 m wide, and enclosing narrow horses of schist country rock. It i composed of bands of milky "buck" quartz with thin laminae of schist. Scheelite occurs in short strings of grains withi quartz on the footwall side of the lode and also in small quartz-carbonate veinlets in the foot wall and hanging wall schist The lode quartz contains only localised traces of pyrite and arsenopyrite. They are more abundant in the wall-rock schist.

Jan 1, 1995

By N Tyson

Throughout Australia and the world, there have been in recent years a number of incidents involving the reaction of ammonium nitrate based explosives with ground containing certain types of sulphides. This paper presents a comprehensive strategy designed to identify, assess and mitigate the risks involved with blasting in areas of reactive ground. Undertaking blasting in reactive or potentially reactive ground involves a degree of risk, the level of which is dependent upon the controls in place for specific sites; this in turn depends upon the degree of sound reasoning and judgement behind the implementation of these controls. Implementation of controls to minimise the risk of incidents also involves a measurable increase in production and blasting costs. To achieve a balanced solution to this matter, decision-makers need to understand how the risks eventuate, how they should be evaluated, and how control measures reduce the risk. Also required is an understanding of how the problem affects the economics of the operation, and the magnitude of this effect. It is the aim of this paper to provide a structured approach to gaining this knowledge, so that mine management, the workforce and explosive suppliers may jointly develop more effective economic and safe blasting systems for potentially reactive ground. Using a combination of fault/event tree risk assessment and semi-quantitative analysis, a structured approach to dealing with the issue will be illustrated.

Jan 1, 2001

By Wang S

Fragmentation is the basic concern in rock blasting and serves as the main measure of blasting effectiveness. A good estimation of fragment size distribution before blasting is very useful and helpful for selecting the proper equipment for loading, transporting, and crushing rocks. A fragmentation model is presented that is based on a hole by hole analysis and correlates the rock properties, burden, spacing, and explosive column with the blasted volume and fragmentation obtained. The drill bench is divided into three separate areas corresponding to different breakage mechanisms. They are: 1) a main breakage or casting area, 2) a secondary breaking area, and 3) an overbreak area. The boundary of the main breakage area is found from the magnitudes of strains in the stress waves and the tensile strength of the rock. The secondary breaking area is determined by stress wave reflections at discontinuities and interactions between colliding stress waves from adjacent boreholes. The overbreak area is that beyond the main breakage area "where the reflected tensile waves no longer break the rock. This analysis is applied on a hole by hole basis to locate the breaking points. Then connecting the breaking points on each side of a borehole forms the triangle of the main breakage area. This is repeated for the next hole and the remaining area is the secondary breakage area. After the angle and area for each hole are found the Kuz-Ram model is used to predict the fragmentation size distribution for each hole. Then summing up the size distribution for each hole gives the total fragmentation distribution. This analysis has been successfully applied to field tests performed in a limestone bench where the entire spoil pile was screened to determine the fragment size distribution. Knowing the explosive and rock properties the breaking area was found for each hole, the strain and stress values, the breaking point, and angle of breakage (104 degrees). The fragmentation size distribution was then calculated and compared to the field data with excellent agreement.

Jan 1, 1990

By Clappison D. J

Since its formation in Papua New Guinea and raising of A$2 million in 1980, Niugini Mining Limited (NML) has grown to a company capitalised on the Australian Stock Exchange at about A$350 million. In 1995 the company produced some 237 000 ounces of gold, 700 000 ounces of silver and nearly 5000 tonnes of copper from its operations at Red Dome, in North Queensland and San Cristobal in The Atacama Desert of Northern Chile. This production contrasts with Niugini Mining's first mining venture at Mt Victor in the Eastern Highlands of PNG which yielded 20 000 ounces of gold and 11 600 of silver between 1987 and 1989. Following the formation of an exploration joint venture in 1982 with Kennecott Explorations (Australia) Ltd, now a subsidiary of RTZ, Niugini Mining's early recognition of the gold potential of volcanic rocks at plate margins led to discovery of the Lihir deposit in July 1982. This deposit containing about 42 million ounces of gold in resources, is one of the largest undeveloped gold deposits in the world. Following raising of equity funds through the float of Lihir Gold Ltd and debt finance in 1995, construction of the mine at a cost of US$673 million is now underway with first full-scale production at approximately 600 000 ounces of gold per year expected in early- 1998.

Jan 1, 1996