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During the drilling campaign conducted between October, 1942; and May, 1943, to prove the extent of the scheelite deposit worked by King Island Scheelite N.L. at Grassy, King Island, Tasmania, 35 holes were drilled and 606 samples recovered and assayed for their content. Of these, 238 came from outside the limits of the orebody and were for the most part barren, although a few contained some scheelite. The present statistical analysis is made only of the 368 samples taken from diamond-drill cores of the orebody itself.Drilling Pattern and Sampling MethodsFrom the purely statistical viewpoint, it is realised that methods employed in selecting sites for diamong.-drilling, and of collecting samples from drill cores, do not give a completely unbiassed set of data for analysis. Fig. 1 is a sketch plan showing the relativepositions of the drillholes, which were inclined at an angle of depression of 50 deg. The pattern does not conform to a properly randomized and' regularly-spaced grid, the positions of the later holes being determined by the results obtained from previous holes.Statistically, the bias is, therefore, favourable. The unequal spacing of the drill holes and the lack of uniformity in direction of the lines along which the holes were drilled are unsatisfactory from the mathematical viewpoint; but it is realised that these irregularitiesare difficult to avoid in determining by the most economical methods the extent of such an orebody, and are dictated by geological rather than mathematical necessity.Sampling, also, is lacking in complete uniformity. Each sample is approximately 10ft. in length. However, owing to the fact that 100% recovery of core is rarely achieved, and actual lengths of samples taken are ,primarily governed by the length of the drill runs,and secondarily by lithological changes, there is a considerable amount of variation in the sample length.AssayingWith the exception of 28 samples from three drill holes all samples dealt with in this paper were assayed in the Melbourne University Ore-Dressing Laboratory of the Commonwealth Scientific and Industrial Research Organisation. These assays were given to the nearest 0.05% WOg as it was considered that more accurate determination was not possible. This fact has been taken into consideration in the frequency analysis below. The 28 assays determined by the Tasmanian Mines Department were given to the nearest 0.01%.

Jan 1, 1950

By H. A. D. Kirsten

Shotcrete has become a strategic component of support in underground mines. The ductility of reinforced shotcrete makes it a safe and economic means of surface support under adverse static and dynamic loads. The ability of ductile shotcrete to sustain load at large deflections enables excavations in squeezing conditions to be stabilized and the safety in excavations subject to rock bursting, to be assured. A Shotcrete Working Group was formed to find the configurations of fibre reinforcement that would give shotcrete the same ductility as when reinforced with mesh. Beams and panels reinforced with various types and lengths of fibre were accordingly tested. A single mix design consisting of river sand, ordinary Portland cement, fly ash, condensed silica fume and gunite dust and segregation suppressant was used. Steel and polypropylene fibre varying in length from 25 to 50 mm were used in nominal contents of 5 % and 0.5% by mass respectively. Beams, panels and control specimens were shot into specially made formers. Suitable mixing procedures were developed for the different types of fibre. The test specimens were shot in a semi-dry mix process. The shotcrete was cured until tested 28 days after shooting. The fibre contents and uniaxial compressive and Brazilian tensile strengths of the shotcrete were determined from the control specimens. The beams and panels were tested in bending in specially manufactured test rigs. The beams were examined for simple supports and fixed ends. The panels were fixed in a manner that simulated underground conditions and were loaded statically under uniform pressure. The beams were loaded and the corresponding deflections measured up to failure. The panels were loaded up to mid-span deflections of 150 mm. The results are presented in the form of load-deflection graphs. The peak loads and ultimate deflections for the fixed end beams were considerably larger than for the simply supported beams for the various types of fibre considered. None of the beams tested, however, were able to sustain load at relatively constant value beyond the peak load. The panel tests showed that 40 mm and 50 mm long steel fibre and 30 mm, 40 mm and 50 mm long monofilament polypropylene fibre reinforcement gave shotcrete the same ductility as mesh.

Jan 1, 1998

By K J. Thompson

This paper describes the work which has been undertaken at the Ohai coalfields in Southland to restore overburden heaps during past opencast mining operations, and procedures which have been adopted to ensure that restoration is achieved concurrent with present opencast mining activities. The system implemented for the treatment of wastewater from opencast mining operati6ns is also described.

Jan 1, 1981

A laboratory size solvent extraction apparatus has been constructed, which incorp- orates three stages of countercurrent extraction, followed by two stages of counter- current stripping. The model uses a gravity flow system from one stage to another and air lift pumps to recirculate organic and aqueous phases. A novel pump/mixer system is used to transfer solutions from the mixer compartments to the settler compartments. The model has proven valuable in the teaching of solvent extraction principles.

Jan 1, 1999

By S C. Dominy

Various styles of gold mineralisation pose problems during sampling because of their complex gold grade distribution and presence of coarse gold. Effective sampling and sample preparation forms the basis of resource estimates. The design of a sampling and assaying protocol must be based on a thorough understanding of the mineralisation in question, which should include consideration of gold particle size and mass distribution. The proportion of coarse gold within samples has a profound effect on the requirements for sample preparation and assaying. Choice of preparation procedure and assay type is of critical importance if sampling errors are to be minimised. Traditional fire assays, using small charge sizes, often understate assays gained from large charge sizes and techniques such as screen fire assay and LeachWELL. Proper application of the Theory of Sampling reduces errors during collection, preparation and assaying. A quality control/quality assurance program must be implemented and critically assessed. A case study from the Ballarat mine is presented, which is characterised by substantial quantities of coarse-gold hosted in quartz veins. Diamond core drilling is the key input into resource delineation. A number of different sampling and assay options have been trialled over the recent project history. The current sampling protocol utilises whole drill core, using LeachWELL assays after logging and photography.CITATION:Dominy, S C, 2017. Sampling coarse gold-bearing mineralisation – developing effective protocols and a case study from the Ballarat mine, Australia, in Proceedings Eighth World Conference on Sampling and Blending , pp 71–84 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2017

By H. Jang, Y. Kawamura

Unplanned dilution and ore loss directly influence not only the productivity of underground stopes, but also the profitability of the entire mining process. Stope dilution is a result of complex interactions between a number of factors, and cannot be predicted prior to mining. In this study, unplanned dilution and ore loss prediction models were established using multiple linear and nonlinear regression analysis (MLRA and MNRA), as well as an artificial neural network (ANN) method based on 1067 datasets with ten causative factors from three underground longhole stoping mines in Western Australia. Models were established for individual mines, as well as a general model that includes all of the mine data-sets. The correlation coefficient (R) was used to evaluate the methods, and the values for MLRA, MNRA, and ANN compared with the general model were 0.419, 0.438, and 0.719, respectively. Considering that the current unplanned dilution and ore loss prediction for the mines investigated yielded an R of 0.088, the ANN model results are noteworthy. The proposed ANN model can be used directly as a practical tool to predict unplanned dilution and ore loss in mines, which will not only enhance productivity, but will also be beneficial for stope planning and design.

Jan 1, 2015

The current methodology to design stope support systems in South African gold and platinum mines is based upon the tributary area concept. The discontinuous nature of the hangingwall rock is not adequately addressed, and mechanisms leading to rock mass failures between adjacent support units are poorly understood. The work reported here aims to formulate a basis for quantifying support mechanisms and gain insights into the influence of rock discontinuities on stable hangingwall spans. Numerical models are used to qualitatively investigate the stress transfer from the support units to the discontinuous hangingwall. Stable hangingwall spans are quantified by considering two failure mechanisms, namely (i) beam buckling, and (ii) shear failure due to slip at the abutments. The output of the proposed design methodology is appropriate site specific support spacing, based on discontinuity spacing and orientation. The method is particularly suited to mines at intermediate and great depth, where typically the hangingwall is highly discontinuous due to face parallel mining induced fractures. Results of parametric studies show that with increasing in situ compressive hangingwall stress, beam thickness and friction angle, hangingwall stability is increased, leading to wider stable hangingwall spans between adjacent support units.

Jan 1, 2000

By Qinghua Tian, Zhipeng Xu, Dong Li, Xueyi Guo, Liu Zhu

A possible process for extracting tellurium and antimony from a residue generated in alkaline sulfide leaching of telluriumbearing alkaline skimming slag was proposed. The central composite design was employed to define the optimal leaching conditions for tellurium and antimony extractions from the residue sample. Using a 23 factorial matrix, the best performance for tellurium and antimony extractions (96.36% and 99.64%) were obtained at 175 g L−1 Na2S concentration for 80 °C during 60 min. The notable degree for influencing tellurium and antimony extractions was leaching temperature, Na2S concentration, and leaching time. The analysis of variance results indicated that the regression model was significant and suitable to be used for the optimization of independent variables within the range. The model was validated under optimal operating conditions based on the experimental design results. Antimony and tellurium in the leach liquor were separated and recovered by H2O2 oxidation followed by Na2S-Na2SO3 reduction. 98.93% of antimony was precipitated selectively during H2O2 oxidation, while 95.06% of tellurium was precipitated during Na2S-Na2SO3 reduction. NaSb(OH)6 and elemental tellurium was obtained as the product.

Characterisation of Manganese Ores

Sep 13, 2010

By G. S. Maré

Tin plating operations are of a complex nature due to the many operational variables that need to be controlled in order to create a product of required quality. The work in this report lays the basics for the control of a halogen plating operation based on the DuPont tinplating set-up with regard to the chemical make-up of the plating solution, as well as operational variables such as temperature and current density. To evaluate these aspects of the process, a rotating cathode Hull-cell set-up was used. This allowed for a laboratory scale plating arrangement that simulates industry closely with regard to chemical and hydro-dynamic considerations. Typical industry accepted standards were obtained for the values of operational variables and a ?baseline? test was conducted employing these standards. All subsequent tests conducted were benchmarked against this baseline result. A large suite of results was obtained in which it was found that changes could be made to the industry accepted norm which would improve on the stability of the operation and therefore the quality of the final tin-plated product. Examples of these results are: an increase in stannous tin concentration, an increase in sodium chloride concentration and a slightly elevated plating temperature. In addition, conditions were found that must be avoided to prevent instability and poor product quality.

Jan 1, 2010

By J P. Ashby

Compared with elsewhere in the world, New Zealand quarries have tended to be mostly small. The reasons for this are a combination of the relatively high transport costs, low population density and a pioneering, self-employed attitude. The small size of the operations is reflected in low productivity, a tendency to repair and rebuild equipment rather than replace it, designed flexibility in the processing plant, minimal reserves investigation or development planning and a versatile and innovative work force who are generally not professionally trained. The Resource Management Act and the Health, Safety and Employment Act as well as a customer driven demand for higher quality and consistent products will have even greater influence upon future direction of the Industry. The trend will be towards fewer quarries operated by companies and professional managers that can adapt to the new demands.

Jan 1, 1993

By M. W. Lewis, J. Werner

Fluctuating metal prices, declining head grade, and increased operational costs are forcing many open pit mining companies to look to their balance sheets for significant cost reductions in order to remain competitive. Maintenance continues to represent a significant proportion of the total operating costs in the mining industry. Compared with the other major cost components of the mining process, very little effort has gone into maintenance optimization. The potential payoff from maintenance optimization on the cash cost of the commodity is enormous. A large percentage of these costs are incurred in unscheduled maintenance events, which typically cause serious disruptions to the production process. These maintenance-induced production losses can be a significant contributor to the overall operating costs. A structured proactive maintenance approach using the latest advances in technology can achieve dramatic cost reductions and productivity improvements. Remote condition monitoring systems are becoming increasingly commonplace and have the potential to identify problems prior to failure. Detection of equipment degradation prior to failure will enable repairs to be scheduled thereby reducing costs and interruptions to production. Remote condition monitoring technology has evolved rapidly in recent years, with a host of new tools becoming available to assist mine operators in improving their maintenance operations. These tools have capabilities such as remote real-time vital signs and performance monitoring, alarm/event handling, trend analysis, and maintenance workflow management. When these monitoring tools are used in conjunction with a mine management system, the vital signs data can be integrated and associated with maintenance history, predictive maintenance data, operational information, and geographical information. This integrated approach to the management of vital maintenance data will help shift the mobile equipment maintenance paradigm from reactive to proactive. This paper will introduce a proactive approach to the management of the maintenance process and the tools required to facilitate this process.

Jan 1, 2005

By John A. Owsiany

The effect of longwall mining on streams and shallow aquifers has been the subject of numerous research studies for more than twenty years. As a result of these studies, it has become common to consider an overburden thickness of 400 feet as the lower limit at which longwall mining methods may be used with little or no effect on hydrologic systems. This paper presents the results of a study documenting the response of a large high order stream to shallow cover undermining by longwall methods in the northern Appalachian coalfield. The field site is located in a narrow stream valley where the overburden, composed primarily of limestone and sandstone units interbedded with shales and siltstones, ranges in thickness from 340 to 285 feet above the main bench of the Pittsburgh coal seam. Streamflow monitoring and monitor well measurements indicate that no significant loss zones or water table lowering resulted from undermining the stream or the adjacent valley bottom area. Additionally, hydrographs for shallow near-stream wells exhibit fluctuations related to natural seasonal variations, which are nearly identical to those observed in a non-undermined control well in the same watershed. Environmental factors affecting streamflow were monitored through the use of a control stream in an adjacent non-undermined watershed to differentiate between natural and mining induced changes in flow.

Jan 1, 2001

By Doug Tesarik

Researchers from the National Institute for Occupational Safety and Health1 (NIOSH), together with employees from The Doe Run Company, are monitoring and evaluating the stability of a pillar retreat section at the Fletcher Mine, a lead/zinc, room-and-pillar mine near Bunker, MO. Cemented rockfill is being used to provide roof support as high-grade, remnant ore pillars are mined. Earth pressure cells and deformation meters placed in backfill panel pillars monitor stress change and vertical deformation, respectively, as nearby ore-bearing pillars are extracted. Similarly, biaxial stressmeters and borehole extensometers are used to monitor mining-induced stress change and deformation, respectively, in long-term support pillars. Data from these instruments indicate that when ore-bearing pillars were mined, stress was redistributed from support pillars to a backfill panel pillar. The data also indicate that the backfill and support pillars have remained stable for two years after pillar extraction was completed in the area. This research demonstrates the importance of using reliable instruments to document the geomechanical behavior of pillars and backfill. Gaining a better understanding of roof support and stress redistribution mechanisms is a fundamental part of NIOSH's goal to reduce and/or predict rockfalls, which are still among the leading causes of injuries in metal/nonmetal underground mines.

Jan 1, 2006

By W. Greuer

The objective of this paper is to demonstrate how site-specific slope stability analysis can be performed entirely from the desktop with satellite images, Geographic Information Systems (GIS), and slope stability software packages. This type of analysis would effectively eliminate or restrict the need for expensive field testing of geotechnical properties of a hillside, pit wall, berm, or embankment, and would be particularly useful for slope failure mitigation at remote or otherwise inaccessible locations.

Jul 17, 2000

By Peter J. Nicholson

Since the reintroduction of soil mixing to the United States over 10 years ago, the work installed by soil mixing technology has grown dramatically. From an annual turnover of less than ten million dollars per year the volume of soil mixing put in place has grown to an estimated $40 million per year in the year 2000. The majority of work done to date has involved walls, either for excavation support, permanent retaining structures or hydraulic cut-off walls. The other major use for soil mixing has been for mass stabilization of soils whether for isolation of contaminants or for retaining structures. This paper will discuss the uses of and costs of typical soil mix structures. BACKGROUND Soil mixing is defined as the in-situ combining of reagents, typically cement, with native soils, the purpose being to modify or improve the ground. This history of soil mixing covers only the use of cement slurry, aggregated with soil in-situ by mechanical mixing methods only. Soil mixing by use of high-pressure hydraulic means, also known as Jet Grouting, will not be discussed. In addition this will not cover ex-situ (above ground) mixing, or mixing dry reagents with soil. Only mixing applications for building, civil and heavy construction applications will be discussed, with only passing reference made to uses of soil mixing for environmental remediation work.

Jan 1, 2001

By K Palmer

Alluvial platinum was recovered during gold mining operations in Maud Creek, a tributary of the Howard River. Average electron probe analyses give Pt 88.75, Fe 5.27, Ir 0.70, Os 1.06, Pd 0.34, Ru 0.14, Rh 1.29, Cu 1.06. Rare grains of a Pt-Fe-Cu alloy are also present and an unusual feature of all grains is the relatively high rhodium content which finds expression in segregations of rhodium arsenide, cherepanovite. Platinum is found in the Orepuki area of Southland, where it is probably partly derived from the mafic rocks of the Longwood Complex. Rocks of the Rotoroa Complex, similar to those of the Longwood Complex, form bedrock in Maud Creek and are possibly the source of the Howard platinum. Alternatively, it might be derived by glacial transport from the Matakitaki ultramafic block 30 km to the southwest. Origin from the Rotoroa rocks is preferred as, if the Matakitaki rocks are part of the Dun Mountain Ophiolites, they might be expected to contain osmiridium rather than platinum. The ratio Pt:Ir-Os has been widely used to distinguish low Pt:Ir-Os ophiolitic from high Pt:Ir-Os gabbroic assemblages. Mafic complexes such as Longwoods, Darran, Rotoroa, and Riwaka might be expected to show similar high Pt:Ir-Os ratios and be targets for platinum exploration. Detrital platinum examined from the Parapara area, Golden Bay is mainly osmiridium, with rare grains of rutheniridosmine, very similar in composition to that found in concentrates from the Kaniere dredge, West Coast.

Jan 1, 1987

By Mohammed Y. Fattah, Wissam H. S. Al-Sudani

"During installation of pile, a soil column is formed inside the pile. For small depths, the height of the soil column typically equals to the penetration depth (full coring state). As the penetration depth increases, the developed internal shear stresses become larger and reduce the amount of soil to penetrate. At the same time, the soil in pile at the bottom is compacted. This will continue until at a given moment, the soil column is a soil plug, and does not stop moving relative to the pile. This is because the shear stresses developed in the interior of the pile are greater than the load-bearing capacity of the ground.This paper presents the results obtained from 40 model pile tests carried out to assess the effect of soil plug on the model pile capacity. The piles are embedded in sand of different densities with different lengths of piles and types of pile-ends. The model tests are divided into two parts according to the type of end of the pipe pile used in the model pile test. The first part consists of 32 model open-ended piles with different lengths of soil plug inside the pile. The second part consists of 8 model closed-ended piles.It was concluded that open-ended pipe piles behave as a closed-ended if the soil plug formed inside piles in state partial plug or full plug. Length of soil plug depends on the type of installation and relative density. The pile load carrying capacity in dense sand is several times greater than those in medium sands especially in cases of closed ended piles or open-ended since the pipe pile can produce external and internal skin friction in addition to end bearing resistance that makes the total pile capacity close to that of closed-ended pile. Whether open-ended piles are driven or pressed in the fully coring - fully unplugged mode or in the partially plugged mode, the plug does contribute to static pile base capacity."

Jan 1, 2014

Planning for the Macraes Extension project highlighted the need for additional tailings storage availability. The possible solution of developing new tailings impoundments in Northern Gully, in tum, led to concerns being expressed by the regUlatory authority over the assimilative capacity of Deepdell Creek. Philosophical differences over the appropriateness of applying a seepage model used to set standards in consents for the original project resulted in the company re-evaluating the whole monitoring system associated with tailings disposal and seepage control. Re-evaluation indicated problems arising from the existing in-stream water quality monitoring in respect of protection of the aquatic ecological system and verifiability of results. A new monitoring system was designed relocating the compliance monitoring point out of the stream and into the gully containing the tailings impoundments giving greater certainty of monitoring results to all parties and providing improved protection for the aquatic ecology of Deepdell Creek.

Jan 1, 1993

The acid plant consists of the following units: 1. Glover tower. 6. Mills-Packard type chambers. 3. Reaction towers. 6. Gay-Lussac towers.The sulphur dioxide bearing gases, used in the manufacture of the acid, are obtained from the roasting of zinc concentrate. Six Herreshoff type furnaces are available, each with an effective hearth area of 520 sq. ft.The concentrate contains approximately 30% sulphur, of which about 72 is ultimately recovered as acid. The oxides of nitrogen are added to the gases in the requisite proportions partly from nitrous vitriol and from nitric acid in the Glover tower, and partly from nitrous vitriol in some of the chambers and connecting flues.The acid leaving the Glover tower is cooled by the ordinary water-cooling coils, while the cooling of the chamber gases is effected by circulating water over the outer surface of the chambers, and by spraying nitrous vitriol and weak acid into them.Flow of gases throughout the plant is regulated and controlled by means of regulus metal fans, fitted with variable speed gears.

Jan 1, 1934