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Exploring for deep shallow dipping orebodies can be very expensive and very difficult. Surface exploration is often not an option and obtaining the best location for underground drilling can cause production delays if drilling is carried out from development headings. The Otter/Juan mine has been in production for approximately 34 years. All of the production is now sourced from more than 1000 metres below the surface from an area known as the Edwards Lode. The orebody is shallow dipping and open at depth. Exploration drilling has been performed from decline stockpiles and from hangingwall drill drives. Drilling from footwall positions (eg decline stockpiles) into shallow dipping orebodies can often produce intersections which are unusually thick thus leading to optimistic resource figures. The decision to develop a dedicated hangingwall drilling position can vastly improve the quality of the data used to generate an orebody model. In 1997 a new extension to the Edwards Lode was discovered and named the æFÆ surface. In 1999 a resource calculation was performed using data obtained from drilling that was carried out from decline stockpiles. In 2002 a dedicated drill drive was developed into the hangingwall of the æFÆ surface to obtain better data to verify the previous resource calculation. For this case study obtaining more representative data reduced the resource estimate by 66 per cent. This information was obtained early enough to allow for alteration to the mine plan and without loss of capital.

Jan 1, 2003

By P. K. Chatterjee

The overall success of many strip coal mining operations depends primarily upon the efficient use of draglines to remove overburden. These machines require enormous capital investment and unless used efficiently the profitability of a venture is adversely affected. It is important that the dragline works continuously and it is imperative that the walking path and other facets of its operation be predetermined exactly so that the desired production rate is achieved with a minimum of non-productive time. Details of dragline operations which are of particular interest are walking path, prediction of best cut and positioning to maximize production, the angle of swing which produces the swiftest operation, optimum fill factor allowable for the type of material and conditions of ground, the maximum depth from which it can excavate and operating level with respect to the elevation of the high wall. The aims of the present paper are: 1. To establish a cut-off point in the cost-depth ratio. 2. Predict the most economic method of stripping in a given area using graphical plots of walking path implemented with information necessary to ensure optimum production. The optimum operating efficiency of the dragline in relation to the terrain and other conditions imposed on its operation will be evaluated. The model is based upon operating data which has been collected from site observations of a large strip coal mining operation in Central Queensland, Australia.

Jan 1, 1977

By F. O. Willhofft

The symposium on mine safety held in connection with the annual meeting of the Institute three years ago, it. was pointed out that "at present no convenient, reliable, accurate means for determining the gas content of mine air is available, for use by the ordinary mine man." In a letter addressed to Coal Age and published on Feb. 25, 1925, Thomas T. Read, then Director of the Safety Service of the Bureau of Mines, stated: "It is to be hoped that sometime in the near future some-one will invent a reliable methane indicator that is con-venient to, use and will permit the keeping of all flames of every character out of bituminous mines, through doing away with the safety lamp." The Martienssen methane detector, which has re-cently received the official approval of the Bureau of Mines, is intended to. meet the demand for such an in-strument as expressed at that meeting and on many other occasions. The characteristics which qualify it for the task are the following: (1) simplicity; (2) small bulk and light weight; (3) ruggedness; (4) accuracy sufficient for the requirements; (5) safety; and (6) dependability. The principle on which the working of the instrument is based may best be explained by giving a brief out-line of the various stages of development through which the inventor, Professor Martienssen of the University of Kiel, Germany, advanced to the final solution of the problem which he had undertaken to solve.

Jan 2, 1928

By Frank Palkovits

All technologies move through a cycle from infancy to maturity. Thickened tailings have moved into the young adulthood phase of its development. Many mining companies like the idea of reducing their water-taking and water disposal issues. This limits the risks associated with traditional surface water-covered tailings facilities and the reduced dilution possible through underground disposal of thickened or paste tailings. Just as home theater systems advertised as ?plug and play? are rarely quite that straightforward, implementing and improving a thickened-tailings system has its challenges. However, the technology has also developed in recent years so it is more accessible. Some success factors have emerged because of the new technologies. Is thickened tailings the right solution? The first step in success with thickened tailings should be whether it is, in fact, the right solution for a particular mine. And, if so, what type of thickening is best. The answer depends on several factors, including: ? Grain size: The grain size that is optimal for maximizing mill recovery is some-times not suitable for thickening. Other mineralogical factors, such as mica content, may also influence whether the tailings are suitable. ? Water issues: Available water does not always have the purity necessary for the process. Another issue to consider is that if water is in short supply, as is increasingly often the case, thickening may be a way to reduce the need to both take and dispose of water.

Jan 1, 2008

By Warren E. Morrison

Stimulated by 4 years of continuous economic growth without inflation, demand for energy in the United States reached an alltime high in 1964. Total value of production of goods and services (gross national product--GNP) for the year was $629 billion, an increase of 6.8 percent from 1963. Allowing for price changes based on 1958 dollars, the gain in real terms was 5 percent. In 1964 the Federal Reserve Board Index of Industrial Production stood at 132, a rise of 6.2 percent. With the impetus provided by personal income and corporate tax reductions during the year, aggregate wages, profits, and consumption reached record levels. Even the persistent balance of payments problem improved. Against this background the national consumption of commercial energy increased 4.2 percent. Energy needs continued to be met largely from domestic resources with imports, mainly crude petroleum and refined petroleum products, accounting for but 7 percent of the total. Mineral energy resources, namely gas, coal, and petroleum, contributed 96 percent of the input of energy into the economy during the year. The remaining 4 percent came from hydropower and nuclear power.

Jan 1, 1965

By D. Montagna

Large quantities of zinc-base die-cast scrap containing 3 to 4 percent aluminum are presently refined by mixing with other secondary zinc products such as galvanizer dross and distilling in a retort. The Bureau of Mines investigated alternate aluminum removal processes utilizing low-cost fluxes. Basing the investigation on the metallurgical principle that a reactive impurity can be removed from a less reactive matrix metal by melting under a flux consisting of a salt or oxide of a metal less reactive than the impurity, three separate flux mixtures were studied. In the first flux system, [FeCl3] + NaCI (the active component is underlined), part of the aluminum is' eliminated by the formation of volatile aluminum chloride and the rest by the creation of a high melting point iron-aluminum intermetallic compound that can be separated from the remaining molten zinc by filtration. Aluminum removal in the second system, [Fe203] + CaClz + NaCI, is effected by the formation of the iron-aluminum compound and possibly aluminum oxide. The third system, spent sal skimmings--essentially a mixture of [ZnCl2], NH4Cl, and ZnO, depends solely on aluminum chloride formation. Of the three flux systems investigated, the best results obtained were with system 3. Aluminum was reduced in the molten die-cast alloy from 4.10 percent to 0.01 percent when depleted sal skimmings were used.

Jan 1, 1969

COMPLYING with your frequently expressed wish and, your persistent request, I have decided to prepare a brief little book on metallic ores, based on ,the books of the ancient philosophers and on the experience of practicing miners, that will instruct you how to judge from promising indications which rock forma¬tions, which veins, stringers, or mineral matter are likely to bring. forth metallic ores so that it would be promising and gainful to work them. It explains according to the best of my ability, in separate chapters, as much as you need to :know for your, chosen profession or business, about such things as the favorable characteristics of the terrain, the strike, and dip* and outcrop of veins in general, and also about each metal in particular. Young Knappius: † Then it will be explained to me logically in. this little book and I shall be given a reasonable understanding which mines can be worked gainfully so that my investment will not be wasted but will show a profit. Daniel: Everybody should, of course, use his intelligence and should educate himself well and diligently to become able to understand, [2]‡ as far as that is. possible, from what matter, by what means, and where metals are generated by. Nature; and as a mere side issue he should not spurn the profit that ensues there from. But if his aim is solely and predominantly profit and gain rather

Jan 1, 1949

By R. Woodward Moore

A GRAPHICAL method of analyzing the data obtained from shallow earth-resistivity depth tests is presented. The method is based upon empirical results and has no theoretical basis. The usual apparent resistivity-electrode spacing curve is used together with a cumulative resistivity-electrode spacing curve plotted on the same sheet. The greatly reduced scale required for plotting the cumulative values of resistivity together with the effect of the summation of the individual resistivity values serves to minimize the effect of purely local surface anomalies and inadvertent errors of measurement. The point of intersection of tangents or straight lines drawn to intersect at zones of maximum curvature in the cumulative curve indicates the depth to the underlying material. Numerous figures are presented in which data from published reports and from recent field studies are analyzed and the results compared with actual depths established by borings or with depth values obtained by the use of theoretical methods of analysis. Smoothly rounded curves of apparent resistivity such as are often obtained in the field, and which have been a serious drawback to attempts to analyze the data empirically heretofore, appear to be susceptible to rather accurate analysis by the method described. The method is best suited to analyses involving shallow two-layer formations. It has been applied successfully, however, in analyzing the data obtained from tests made over shallow three-layer formations. As with most empirical methods, its chief advantage is its simplicity and ease of application.

Jan 1, 1944

By Gooday PE, Brech GE

Ore reserve categories used by Pancontinental Mining fall into one of three fundamental groupings: Geological In Situ Reserves Mining Reserves Production Estimates The first of these describes the total reserve within stated parameters without implication as to its mineability. The second is that part of the Geological In Situ Reserve which is considered mineable under a specific mining plan. The third is the Mining Reserve adjusted with appropriate dilution and recovery factors and thus represents the best estimate of the tonnage and grade of production. This practical three category system mirrors the standard steps traditionally taken in ore reserve estimation and readily conveys the basis of the estimates. An important, but secondary classification can be added to qualify the level of confidence in these estimates using the terms: Proved (or Measured) Probable (or Indicated) Possible (or Inferred) These terms can be applied to any of the fundamental groupings, but it is generally implicit that the level of detail required to produce meaningful Mining Reserves or Production Estimates would require the basic Geological Reserve to be at a Proved or at least Probable level of confidence. The application of these classification systems by Pancontinental is illustrated by reference to the Paddington Gold Project. The methods of calculation are described including the use of geostatistics. Reconciliations of actual production with predictions from reserves are carried out monthly and annually. To date there has been an excellent correlation with Production Estimates, based on both manual and geostatistical calculations.

Jan 1, 1987

By James S. Wroth

The Act of Congress approved June 25, 1926 (Public Act 424, 69th Congress), authorized joint investigations by the U. S. Geological Survey and the U. S. Bureau of Mines to determine "the location, extent, and mode of occurrence of potash deposits in the United States." The particular deposits selected for exploration were the extensive rock-salt formations underlying large areas in western Texas and southeastern New Mexico, from which potash salts had been identified in the cuttings and sludges from oil wells drilled with cable tools. The evidence of these sludge samples was qualitative only and beyond definitely proving the presence of potash salts in the salt beds, gave little clue to the possible size, richness, or extent of the potash occurrences.

Jan 1, 1929

By A. Holland, T. Krumins, E. Olin, J. F. M. Castilleja, J. Geldart, C. Fleming, I. Jackman

A flowsheet has been developed to treat the flotation tailings from a copper mine in Mexico to recover gold and silver by cyanidation and carbon-in-pulp (CIP). The plant will treat ~15,000 t/day of a 60:40 blend of current flotation tailings and the mine’s historic tailings. The tailings also contain a significant amount of cyanide-soluble copper, and the economic viability of the project is strengthened by recovering most of the copper and cyanide associated with this process. The solution and solid phases in the CIP tailings will be separated in a counter current decantation (CCD) circuit and the CCD solution overflow will be treated by a SART Process. Copper precipitates as a salable copper sulphide product in the SART Process and the cyanide that is associated with the copper complex is regenerated as free cyanide. After SART, a portion of the SART product solution will be recycled to leach to reuse the free cyanide, while the remaining solution will be used as wash water in the CCD circuit. Incorporation of the SART Process in the flowsheet allows the CIP circuit to be operated aggressively as a gold plant, which means most of the silver in solution passes through CIP to SART. Silver co-precipitates with copper in SART and is recovered very efficiently with the copper sulphide by-product. Incorporation of SART also allows the leach and CIP circuits to be operated with relatively high levels of free cyanide, which enhances silver leach kinetics and helps keep copper from loading onto the carbon in CIP. The flowsheet was developed in the laboratory and successfully demonstrated in a fully-integrated pilot plant in 2018.

Jan 1, 2019

By Oliver Bowles

A NORTH CAROLINA miner dreamed that he found high-grade mica by excavating a certain corner of his mine. The next day he sank a hole on the exact spot and found mica of excellent quality. The dream came true. May not this be as good a method as any yet known for forecasting the valuable mineral content of pegmatites? The Federal Geological Survey has found in its recent detailed studies in New England that good sheet mica occurs most abundantly in zoned pegmatites, particularly in zones lying along the wall or contact of the pegmatite mass. Although this conclusion may be a valuable guide, pegmatite mining operations must still be classed as most uncertain and hazardous enterprises. Finding of the second largest diamond ever discovered in the United States-a stone of good gem quality weighing more than 34 carats-during the course of a horseshoe pitching contest at Peterstown, W. Va., is another example of unorthodox mining methods. These should not be considered, however, as characteristic mining methods developed in 1944.

Jan 1, 1945

By S. K. Murphy

A partial shaft pillar extraction project was embarked on at a mine in the West Wits mining region in South Africa. Prior to the extraction, a rock engineering assessment was conducted with the main objective being to establish which areas of the shaft pillar could be extracted without unduly affecting the stability of major infrastructure within the shaft pillar. Standard numerical modelling methods were used to conduct this assessment and several areas were identified for extraction. However, during the extraction process, it was found that the infrastructure was affected more than was originally predicted by linear elastic modelling. In addition to this, several large events resulted in far more large aftershocks than would normally be predicted. In order to gain an understanding of this phenomenon, seismic data was integrated with numerical modelling as follows: > The seismic data in the form of displacements was integrated in 3D space within the model to assess the effect of static stress changes associated with seismic activity > Data from large events was integrated onto specific structures, where displacements associated with large events were integrated onto the plane of structure in question. This paper describes the methodology and results obtained from each of the above processes and include insights into the mining process that were previously unfathomable using standard linear elastic numerical modelling. The way forward for this rock engineering assessment of partial shaft pillar extraction is also discussed.

Jan 1, 2012

By X Li, W McKeague, M S. Powell

This paper explores the importance of operators in the overall efficiency of mineral processing plants. With the potential of future control configurations becoming more remote, it is time to review the implementation of control and utilisation of operators. This paper is based on a preliminary review of operators and control room operations and clearly shows that, in the minerals industry, operations are typically run in a reactive mode (with a focus on ensuring equipment operation), rather than in a proactive mode (with a focus on optimising the process performance). The primary reason for this misguided approach is the short sighted thinking applied to the challenges of designing technology and organisational tools capable of addressing the essential operational needs: an approach which inevitably results in a disconnect between people and technology. Contrary to the popular approach of replacing operators with expert systems, the thesis of this paper is that the shift to more complex circuits, with higher demands on processing capability and control, will require the actual upskilling of operators. Experience in mineral processing and other industries indicates that expert systems cope well with small changes but are typically unable to respond to major changes in feed material and this is where the expert operator becomes crucial. Providing operators with ongoing high level training in process understanding, and restructuring how they interact with the plant and interface with the control system is the approach being proposed here - an approach surprisingly underutilised and often absent in the minerals industry today.CITATION:Li, X, Powell, M S and McKeague, W, 2012. Unlocking processing potential by empowering our operators, in Proceedings 11th AusIMM Mill OperatorsÆ Conference, pp 333-340 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Oct 29, 2012

"Recognising that the mining industry is wedded to the break-even cut-off grade rather than a cut-off grade that optimises value, the author examines some perils and pitfalls in the use of break-even grades that lead to what the author calls ‘negatively geared ore reserves’.Negatively geared ore reserves occur when a portion of the ore reserves actually has a negative value and is effectively being subsidised by the more profitable portion of the ore reserves. This situation is generally difficult to detect with the level of information available to the public investor. The ore reserves can still be quite profitable in total, but now possess a level of risk that may not be appreciated.The risk profile (financial risk) of a negatively geared ore reserve is far greater than for a non-negatively geared ore reserve. Like a negatively geared investment property in a falling property market (for readers familiar with the Australian tax minimisation investment strategy), when the price of a commodity falls, a company that possesses a negatively geared ore reserve loses value at a magnified rate relative to a company that has a neutral or positively geared ore reserve. In the extreme case of a major commodity price reduction, as happened with the gold price in mid-2013, the negatively geared ore reserve can easily become negative value in total, not just in part.The negatively geared ore reserve is sometimes intentional – for instance in order to increase published ore reserve tonnages and meet the CEO’s key performance indicators – but is also often unintentional, resulting from various errors in the break-even calculation. Common errors in calculating a break-even grade include using fixed metallurgical recoveries, not including allowances for ideal laboratory versus real plant results, ignoring available operational data and using modal rather than mean parameter values in cost and performance calculations.This paper examines a common value distribution profile in many ore reserves and the effect of common break-even grade errors on this profile.CITATION:Poniewierski, J, 2016. Negatively geared ore reserves – a major peril of the break-even cut-off grade, in Proceedings Project Evaluation 2016, pp 236–247 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Mar 8, 2016

By M. R. Shahan, C. E. Seaman, Mischler S. E., T. W. Beck

"Float dust deposits in coal mine return airways pose a risk in the event of a methane ignition. Controlling airborne dust prior to deposition in the return would make current rock dusting practices more effective and reduce the risk of coal-dust-fueled explosions. The goal of this U.S. National Institute for Occupational Safety and Health study is to determine the potential of open-air water sprays to reduce concentrations of airborne float coal dust, smaller than 75 µm in diameter, in longwall face airstreams. This study evaluated unconfined water sprays in a featureless tunnel ventilated at a typical longwall face velocity of 3.6 m/s (700 fpm). Experiments were conducted for two nozzle orientations and two water pressures for hollow cone, full cone, flat fan, air atomizing and hydraulic atomizing spray nozzles. Gravimetric samples show that airborne float dust removal efficiencies averaged 19.6 percent for all sprays under all conditions. The results indicate that the preferred spray nozzle should be operated at high fluid pressures to produce smaller droplets and move more air. These findings agree with past respirable dust control research, providing guidance on spray selection and spray array design in ongoing efforts to control airborne float dust over the entire longwall ventilated opening. IntroductionCoal mining processes produce airborne dust that naturally deposits on the floor, roof and ribs of mine airways. These fine dust particles, termed float coal dust when smaller than 75 µm in diameter, are deposited in the return entries of coal mines and can be reentrained by pressure waves induced by methane-air explosions (Rice et al., 1911; Nagy, 1981; U.S. National Institute for Occupational Safety and Health (NIOSH), 2006). These particles are also those most likely to participate in coal-dust-fueled explosions, which can potentiate and propagate explosions resulting in extensive damage. In order to meet compliance requirements of at least 80 percent incombustible content in the composition of material deposited in mine entries (U.S. Mine Safety and Health Administration (MSHA), 2016), coal mines apply an inerting agent, specifically, rock dust (Harris et al., 2009). Should methane be present in the mine atmosphere, the required fraction of incombustible content is further increased by 0.4 percent for each 0.1 percent methane. The removal of airborne float coal dust (AFCD) prior to deposition is an approach intended to make current rock dusting practices more effective by increasing the proportion of inert material to combustible material, resulting in a lower risk of coal-dust-fueled explosions."

Jan 1, 2018

By Marc R. Laflèche, Suzanne Paradis, A. Shirley Péloquin, Harold L. Glbson, Robin Potvin, Pierre Verpaelst

"The Archean Blake River Group is best known from a central volcanic complex in Rouyn-Noranda. This volcano-tectonic depression, delimited by the Hunter Creek fault to the north and the Horne Creek fault to the south, is interpreted to be a cauldron approximately 20 km in diameter. Its western and eastern boundaries are the Flavrian Plutonand the D'Alembert shear zone respectively. The complex has an exposed stratigraphic thickness of 10.5 km. Two cycles, composed of andesitic units overlain by bimodal, andesite - rhyolite units, are observed in the area, and are, in turn, overlain by an andesitic unit interpreted to be the lower part of a third cycle.The southern sector of the Blake River Group, south of the central volcanic complex, is considered to be outside the limits of the cauldron, but partly contemporaneous with it. The maximum stratigraphic thickness exposed in this area is 4 .6 km. At least two cycles occur in this sector. The oldest unit, a pre-cauldron andesitic to basaltic unit, occurs only in the southern part of the sector. In the north, two partial andesite - bimodal cycles, interpreted as contemporaneous with the cauldron volcanism, are observed.The eastern sector of the Blake River Group, east of the central volcanic complex, apparently overlies the complex. Two units are observed in this sector: a lower bimodal unit and an uper andesitic unit. The bimodal unit is coupled with the uppermost andesitic unit of the central volcanic complex to form a cycle, and the andesitic unit is considered to be the uppermost unit of the Blake River Group. The total exposed stratigraphic thickness in the sector is 2.5 km.The western sector of the Blake River Group, north and west of the central volcanic complex, is interpreted as the lateral equivalent of the complex. Three andesite - bimodal cycles are exposed in this sector, giving a total stratigraphic thickness of9 .5 km. The two lower cycles correspond directly to the central volcanic complex cycles; this correlation is facilitated by two stratigraphic marker units observed on either side of the Hunter Creek fault: the Fish-roe rhyolite flow, east of the Flavrian pluton, and a spherulitic rhyolite lapilli-tuff, west of the Flavrian pluton.Many of the volcanogenic massive sulphide deposits found in the central volcanic complex are spatially and temporally associated with synvolcanic faults, and felsic domes and feeder dykes that are the sites of felsic volcanism in the bimodal units. However, these features also occur outside the central volcanic complex, and the similarities recognized in this work between the complex and the surrounding area suggest that such deposits need not be restricted to the central volcanic complex. The presence of volcanogenic massive sulphide deposits outside the complex confirms this."

Jan 1, 1990

By J. Bongarra

This report describes the development of the instructional design and content of a training program for shaft construction miners and supervisors. The recommended training approach is based on the results of a needs, task, and competency analysis of the jobs performed by conventional shaft sinking and raise drill personnel. Learning objectives were derived from the task analysis and the appropriate training delivery system was chosen using a selection technique which employed learning algorithms. The training described is aimed at improving the safety and efficiency of constructing shafts by new hire and experienced miners and supervisors. A self-paced, self-teaching, modular approach to training, using criterion-referenced tests was proposed. Recommendation was made to divide the training into three parts: Orientation and Safety; Task Instruction; and Practical Exercise. Training requirements for conventional shaft sinking and raise drilling were described. Training program content was detailed in the form of module and lesson design approaches. In addition, the report recommends a method for developing, implementing and validating the training program. The authors conclude that such an implementation would meet the major objectives of improving the safety and efficiency of shaft construction work operations.

Jan 1, 1979

By K. B. Le Bron, T. van Aard

The empirical pillar strength formulae have been widely, and in most cases successfully, applied to design underground pillars for mines in the Bushveld Igneous Complex (BIC). This paper presents the case for combining the empirical method with numerical modelling when it comes to crush pillar stability analyses. The benefits include an understanding of potential failure mechanisms, which may then be accounted for in the designs. In addition, the impact of weak layers, such as shear zones, the impact of the dip of the orebody, differences in layouts, such as holing widths and the presence of advanced strike gulleys (ASGs) immediately adjacent to the crush pillars, and stress regime (k-ratio) may be studied using numerical modelling, as it is not accounted for in the empirical method.

Mar 5, 2025

By T Purvis

Xstrata Copper is extending the Ernest Henry Mine with a 6 Mtpa sub-level cave (SLC) mine being constructed under the pit. The SLC mine will maintain the EHM operation until 2024 after the pit is completed in August 2011. In February 2008, the first cut of the new underground mine at EHM was fired with the start of an exploration decline. Since then Xstrata Copper has continued the development and construction of a SLC mine. During this process a number of challenges arose, some of which had not been anticipated in the planning stages, that in turn has been addressed to ensure the development continued. The challenges have been many and varied and range from setting up the portal area and commencing the exploration decline development in an active large open pit, to management of underground water. Further challenges were presented by the effects of the GFC on investment recommendations, with the scale of the mine varying from a shaft mine to a smaller scale trucking only mine in mid-2009 and back to the full shaft mine in late 2009. The effect was the parallel development of two separate mining scenarios, with the challenge of maintaining commonality for as long as possible to keep both options alive until the final board approval was obtained. In 2010 the RSPT caused further impacts. This paper will describe the current progress of mine development as well as how the various challenges have been met and some key lessons learnt from these experiences.

Mar 21, 2011