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By A. T. Iannacchione

Underground limestone mines typically use the room-and-pillar method of mining in the generally flat-lying limestone formations. In some cases the dip may exceed 5° which can result in unique roof instability problems. Stability may be further exacerbated by the presence of horizontal tectonic stresses. Field studies and numerical model analyses are combined to assess the affect of the dip in various stress conditions. The results showed that where the horizontal tectonic stresses are sufficient to cause intact rock failure, increasing dip results in increased failure potential. The effect of the dip is greater when the maximum horizontal stress is parallel to the dip. Two case studies are presented and evaluated against the findings of stress analysis results. The study showed that the empirical rules of orienting headings parallel to the maximum stress are generally applicable to dipping excavations, except when mining in the up-dip direction. The importance of accurately determining the orientation of the maximum horizontal stress is emphasized because the window of favorable orientations relative to the stress is small.

By Jason Clarke, André R. Laplante

"The gravity-recoverable gold (GRG) ore characterization test was developed 15 years ago and has been applied to hundreds of ore samples. The test accurately determines the size distribution of GRG, at which grind it is recoverable, and the extent of its liberation above 150 µm. This contribution shows how the test can be simplified, what are the likely problems to occur, and what information is then lost. Two simplified versions of the test are presented, of which the simplest, grinding a 20-kg sample to a P80 of 75 µm, is retained as the best one. A database of 18 comparative tests is presented. The choice of the full GRG test vs. the simple test is discussed, as is the deportment of the GRG recovered in the standard test but not in the simplest test.INTRODUCTIONThe GRG test was developed in 1990 to characterize the progressive liberation of gold in representative samples of ores, with the primary objective of generating data for a gold gravity recovery model (Woodcock and Laplante, 1993; Laplante et al., 2000). Characterization includes the size distribution of GRG as well as the grind size at which it becomes recoverable. This is achieved in a 3-stage protocol of liberation and recovery using a laboratory centrifuge unit, a Laboratory Knelson Concentrator 3 inches in diameter (KC MD3). This protocol is shown in Figure 1. The first stage takes place at a F100 of 850 ?m (F80 of ~550 ?m), the second at F80 of 180 ?m, and the third at a P80 of 75 ?m. The rationale for the three cycles is the desire to preserve the original size distribution of the gold grains in the ore and to obtain a recovery grind size curve. The feed is a representative sample of ore that can be extracted from ore samples (e.g. drill core reject) or stream products of the grinding circuit other than circulating loads or product having undergone hydraulic classification (e.g. rod mill discharge or SAG screen undersize). Ideally the samples should weigh between 60 and 100 kg, depending on feed grade and gold top size. It is beyond the scope of this paper to fully discuss measures of representativity in the ore sample, for example Gy’s minimum sample mass model (Gy, 1979). However, further work in this regard may be an opportunity. The large initial weight facilitates representativity and minimizes weight and sulphide recovery to the concentrate (since concentrate weight is relatively constant at ~100 g), which limits the recovery of gold in nonsulphide gangue or gold carriers (i.e. sulphides). Thus, the 100 g concentrate carries principally discrete gold species, such as electrum, and native gold. The concentrate and a 600-g representative sample of gangue of each stage are screened down to 20 ?m and each size class is fire assayed (the concentrates to extinction, the tails up to 30 g per size class). The five coarsest size fractions of stage-1 concentrate are individually upgraded using a hydrosizer and the concentrates examined with an optical microscope to assess the degree of liberation of the gold phase. Scanning electron microscope work has established that the GRG has an excellent degree of liberation below 150 ?m in stage 1 and generally a good degree of liberation in stages 2 and 3 (Guerney et al., 2003); as a result the size distribution of the GRG has been found to approach closely that of gold grains (Nesset et al. 2005)."

Jan 1, 2006

By L. C. IlsLey

Explosives have been fired electrically for several decades. Mountains have been tunneled, deep shafts sunk, extensive coal and metal mine workings excavated and, in times of war, railroads, buildings, bridges, and fortifications have been destroyed by means of electrically fired explosives. Modern engineering depends so much on explosives that their safe and efficient handling and especially their detonation are of prime importance to all those connected with the use of explosive agents. Electric shot-firing is only one method of igniting explosives, but it is rapidly spreading and because of certain inherent safety features should be encouraged. Data collected by the Bureau of Mines indicate that in the United States a large percentage of the blasting in quarries, shafts, and tunnels, and from 15 to 25 per cent of the blasts in mining operations, are set off electrically. The Bureau of Mines is interested in electric shot-firing because of its use in mining, and especially because electric methods, when properly applied, are undoubtedly safer than other methods commonly used.

Jan 1, 1926

By Chris Doyle

From 1997-2000. PT INCO undertook the Fourth Line Expansion Project (FLEX) to increase annual production rates from 90-100 to 150 Mlbs Ni/year (68,000 MT Ni/yr). In 2003, PT INCO nickel production underwent a major sustainable step jump that finally exceeded the targeted rate. This paper presents the changes that led to this achievement. The focus of mining changed. Improving Ni grade was important but only if electric furnace feed met target chemistry so that on-specification slag could be produced. The intensity of diamond core drilling for exploration was dramatically upgraded, mine planning strategy was totally re- vamped and an ore chemistry quality assurance team was formed. In the processing plant, measurement systems were upgraded for better control. Teamwork between sections vastly improved. The final step jump in production occurred when the fifth hydro-electric power generator was added to the furnace grid to allow the power-starved furnaces to cruise in operation at 70+ MW. Keywords: Free cash flow, Ore Chemistry, Measure, Control, Improve

Jan 1, 2004

By Christopher J. Bise

INTRODUCTION The principal means for transporting men, supplies, and mined material in an underground or surface mine are rail systems, conveyor belt systems, or a combination of the two. Typically, large underground mines use rail haulage from the mining sections to the hoist, with conveyor belts being used within the sections. Small mines use conveyor belts in the mains and sub- mains as well as in the panels. The size and layout of surface mines also contribute to variations in the way these haulage systems are applied. The following sections address the typical problems confronting mining engineers in designing these haul- age systems. For rail, these include rail selection and turnout design, locomotive tractive-effort calculation, and motor sizing. Belt haulage calculations include belt width and speed specifications, tension and take- up requirements, idler spacings, and motor sizing. TRACK LAYOUTS Introduction The adequacy of mine track layouts is dependent upon such factors as rail weight (Table l), suitable ties, a well-drained roadbed, and properly designed curves, grades, and track alignment. As long as the installation conforms to the manufacturer's recommendations, good service can be expected from the rail system. Typically, mining-engineering calculations for rack layouts have dealt almost exclusively with the design of turnouts. As such, the following section provides the necessary information for computing the theoretically correct dimensions of turnouts. Turnout Design Fig. 1 shows a typical turnout with an identification of the component parts. Fig. 2 identifies the variables that are used to determine the turnout dimensions. The formulas that concern the mining engineer are those listed: [ ]

Jan 1, 1986

By Clinton C. Knox

Attention of mine operators and engineers in the Picher zinc-lead field has been drawn increasingly in recent years to marginal ore deposits lying at greater than normal depth below surface in the marked structural depressions known as the Miami and Bendelari troughs. Not only mining but prospect and development drilling is relatively costly in the bottoms of the structural depressions known as the. Miami and Bendelari troughs which cross the center of the Picher field. Nevertheless, operations by various mining companies in recent years have shown that no insuperable problems exist, and with the gradual lowering of grade cut-offs throughout the field, certain potential deep-ore areas, previously only partly explored seemed worthy of further investigation. Some such areas were called to the attention of Bureau of Mines engineers by lease holders and operators, while others were noted in the course of Bureau's systematic investigation of Tri-State ore reserves. The writer was assigned to make a study of. those possible "trough extensions," in the course of which all available nine maps, drill logs, and prior reports were assembled. Subsequently 41 churn drill holes were bored by the Bureau in the periods October 1944 to May 1945 and October 1946 to April 1947. The general area under investigation this project is indicated on figure 1, which likewise shows its relation to nearby towns, roads, and railroads. The location of individual deposits included in the report is indicated more clearly on figure 3. Minable ore was found in only two holes, and the structural information procured by drilling was less encouraging than anticipated. This report, however, presents all factual data obtained, in the belief that they may be useful in subsequent prospecting and development projects in the areas investigated. There are also set forth some of the stratigraphic data relative to the Picher field ns a whole and the Miami and Bendelari troughs which were assembled in the preliminary investigation of this project.

Jan 1, 1949

Jan 1, 1967

By Michael D. McGill

The effect of fragmentation size distribution as it relates to front end loader productivity was evaluated by comparing loader cycle times to fragmentation size of the shot rock. At Gold Mine in Southern Utah, a gold deposit containing both limestone and sandstone formations were blasted in the same shot.

Jan 1, 1994

By Yi Luo

Bolting the mine roof is one of the noisiest underground mining operations. The main contribution to the noise is the drilling of the roof bolt holes, especially drilling in hard rock. Previous research shows that the specific energy consumed in drilling bolt holes is inversely proportional to the bite depth (i.e., penetration per drill revolution). Less specific energy at a reasonably high bite depth means less energy is wasted in the drilling process for producing heat, bit wear and noise. Therefore, proper control of the drilling operation to achieve a reasonably higher bite depth can reduce drilling noise from its sources as a proactive noise control approach. In order to explore the feasibility and develop the proactive technology for reducing drilling noise, drilling tests have been conducted in laboratory settings. The preliminary results show that significant reduction in noise dose can be achieved by drilling in medium hard rock with uniaxial compressive strength of about 276 bars (4,000 psi) and at a reasonably high bite depth of about 4.6 mm/rev (0.18 in./rev). This paper presents the findings from laboratory experimental studies.

By Felix du Breuil

Contract H0177044 included three objectives; (1) consolidate, maintain, and distribute information on STE approved lighting systems; (2) develop and present one-day seminars on underground mine lighting; (3) determine direct labor and hardware costs associated with the installation of mine illumination systems. All objectives were completed with results as follows: Task I - A file containing 1350 Statements of Test and Evaluation (STE) for all types of underground equipment was developed and publicized throughout the coal fields, A total of 40 requests for STE information was received and processed. Task II and III - A series of 50 seminars was scheduled and presented to 942 attendees throughout the coal fields. The seminars were well received and attendees indicated the information presented would be helpful in bringing their equipment into compliance. Holding a series of seminars in the coal fields, instead of a few in population centers, was favored by attendees. Task IV - Installation cost data obtained on 415 lighting systems indicate that the factors which most affect installation costs are higher hardware cost for DC powered systems and slightly higher manpower requirements for systems installed underground. Based on these data, the total cost to industry of retrofitting existing underground equipment could range up to $50 million.

Jan 1, 1979

By A. Martikainen

Field tests concerning a fog mesh method for fog removal are discussed. The first tests were performed at the Pyhäsalmi Mine, Finland. As this study gave promising results, further studies were completed at the Orivesi Mine. The results of these most recent field tests are presented and the two studies are compared. The Orivesi test set consisted of four materials. These included air filter fabric, fibrous filter fabric, a plastic greenhouse mesh, and an aluminium mosquito net. The variation of the Orivesi results was more noticeable than what was found previously. The best fog removal effect of all tests was achieved with the aluminium mosquito net, while the plastic green-house mesh performed worst. The humidity reduction by aluminium net was more profound than any of its precedents. More mesh materials will be tested in the near future.

Jan 1, 2007

By Frank E. Chase

During the past quarter century, coal mine geologists have evolved from providing a qualitative description of drill core and coal reserves into quantitative geotechnical specialists who bridge the gap between geology and engineering. In the late 1970?s and early 1980?s, hazard prediction mapping was in its infancy. At best, hazard maps identified rudimentary structures and conditions such as stream valley locations, sandstone/shale transition zones, and lineaments. A ?technical disconnect? had existed between geologists and mining engineers as each profession operated largely independent of each other. Engineers did not readily integrate geology and geologic conditions, with the exception of coal thickness, into mine planning. Geologists were relied upon for reserve estimates and did not participate in mine planning or design. The employment and market slump of the 1980?s and 1990?s combined with mining more challenging reserves mandated that this gap be bridged and forced geologists to learn to speak more ?numerically? to engineers so they could be understood. The purpose of this paper is to describe the procedures and thought processes used by coal mine geologists when determining various mineability and reserve extraction feasibility issues. The paper will also address underground geologic hazard mapping and prediction techniques. To achieve these objectives, the opinions and experiences of leading coal mine geologists, geotechnical and strata control specialists across the United States were solicited.

Jan 1, 2006

By Xu Youzai

Static loading teets of five nodular piles and four non-nodular piles were performed in Tianjin soft clayey soil in which measurements of axial force and movements were made. Distribution of shaft resistance and load-movment curves for pile head and pile toe are presented in this paper. The applicability of nodular piles to soft soil and their effectiveness are also discussed. Based on the comparison, it is concluded that nodular piles have an advantage over non-nodular piles mainly in contri­bution to shaft resistance. Dynamic testing was also conducted through R. P. T. for evaluating the service load of tested piles.

Jan 1, 1991

By Paul T. Alexander

After evaluating preceding studies and designs of lightweight supports, ESD selected the all-hydraulic and the hydromechanical concepts for detailed design and fabrication. Further analysis and testing showed that the all-hydraulic design was economically justified and met all the crucial engineering, manufacturing, and design criteria identified during the evaluation phase. Following Bureau of Mines approval of the detailed design, four all-hydraulic support models were fabricated. Once functional and structural testing was successfully completed, 40 additional units were constructed and distributed for in-mine testing. The units were very well-liked by all three demonstration mines. The wide- spread interest in supports for seam heights greater than 8 feet led to expansion of the design data package. ESD designed four basic models to cover the new height range, but no hardware has yet been fabricated for the new design.

Jan 1, 1982

This report is the 22d in a series summarizing Bureau of Mines research and technologic work on coal and related investigations.2 3/ A summary of work published during 1957 and of work in progress is presented. Details regarding the subjects treated in the publications may be found by referring to articles cited in the footnotes. The objective of the coal work of the Bureau of Mines is to obtain information that will lead to improved production and utilization of coal, thus conserving a most important national asset.4 5/ Though studies of mining methods and exploration of various coalfields ma; contribute directly to improved production, no less so does the work in health and safety. The many problems relating to chemical and physical properties of coal reflect the need for an understanding of these properties for the most effective use of coal in its various fields of application and for the development of new methods of processing coal. Studies of existing and proposed processes of combustion, carbonization, gasification and hydrogenation of coal are necessary for increased efficiency in its use. Engineering service provided by the Bureau is of direct value in the conservation program in preventing waste. Stations where the main coal activities of the Bureau are carried out are shown in the following list:

Jan 1, 1959

By Donald P. Richards

Is a shaft the same as a tunnel and can it be approached in the same manner for design and construction? The most obvious difference between a shaft and a tunnel is that of orientation; whereas a tunnel is commonly near horizontal or slightly dipping, a shaft is commonly vertical or steeply inclined. As a result, design and construction methods employed for tunneling, in general, cannot be directly applied to shafts, although some of the basic concepts of behavior are quite similar. Many of the shafts that have been sunk in various parts of the world are owned by private mining companies. Each company has generally accumulated years of experience in shaft sinking and mining techniques pertinent to its particular geographic location. When a new shaft is to be sunk or an existing shaft deepened, the tendency often seems to be to continue to use the methods that have been used successfully in the past.

Jan 1, 1979

By S. Vennos

As a result of Synopsis 73-16 Dynadyne responded with an unsolicited proposal offering the Bureau of Mines to develop an ultra-low volume coal mine dust personal sampler. In the ensuing technical discussions the choice of a desirable volumetric rate became a focal point. While, the initial proposal was based on a volumetric rate in the area of 50 cc/min the Government expressed interest in a variety of mass flow rates ranging from 50 cc/min to 2000 cc/min. The physical characteristics such as weight, volume and configuration of a personal sampler are to a large extent - dependent on the operating volumetric rate. It became readily apparent that the optimum size of such an instrument as a function of volumetric rate was not known and that a study of the choice of optima was essential prior to beginning the development, Hence, the Government authorized a two phase approach to this project. Phase I was to include a Technical Evaluation Study of optimum characteristics in the range between 0. 1 to 2. 0 lpm. Based on these findings and in consultation with MESA personnel, the Bureau of Mines was to authorize the second phase that is, the development. Phase II, was to include the drawing of a set of performance specifications stating among others the desired operating volumetric rate. In this report, every effort was made not only to present the actual work done over the course of this development, but to reflect also our engineering "thinking" providing answers as to the "why" we preferred certain solutions over others. We believe that this added dimension is not only useful for the Government official evaluating this report, but also that the Government is entitled to it providing the official with an educational content useful in avoiding future costly repetitions of similar situations.

Jan 1, 1975

By K. Barr, F. Grenier, J. Hettinger

"To reach the Hemlo Gold Inc. objective of becoming the lowest cost, safest Canadian gold producer while causing no environmental damage, management embarked from its inception on a strategy to build an organization which would foster employee involvement wherever possible. This paper describes the organizational structure and how various innovative concepts have been implemented.BACKGROUNDThe Golden Giant Mine in the Hemlo area of North Western Ontario is owned by Hemlo Gold Mines Inc. 50.1% is controlled by Noranda. The mine commenced operations in March 1985 at a modest production rate. After it reaches full production by the end of 1988, with completion of underground development work, the mine is expected to operate at 3000 t/d and produce in excess of 300,000 ounces of gold per annum. This makes it one of the three largest producing gold mines in North America.Two other companies, Lac Minerals and Teck-Corona, operate mines on either side of the Golden Giant Mine. Noranda Geco Division is located 55 kilometers to the north at Manitouwadge. The majority of Hemlo Gold's 270 employees live in Manitouwadge where a mix of accommodation was constructed and made available for purchase by employees. Bus transportation is provided to and from the workplace with modern coaches. The housing and transportation policies were implemented prior to start-up and designed to help attract and maintain good employees.As shown in Figure 1, the operation is divided into five operating and support departments with each reflecting a high degree of departmental control and responsibility. For example, the mill department is responsible for all aspects directly relating to the milling complex including safety, operation, mechanical maintenance, assaying, instrumentation, and environmental control. Similarly, the mine department is responsible for all phases of operation, mine engineering, geology, and maintenance of underground mobile equipment. The plant department is responsible for maintenance of hoisting plants and conveyances, underground stationary equipment, all site electrical services, surface buildings, housing and yard maintenance. Efforts are made at every level of the organization, from departmental organization to individual job assignments, to ensure responsibilities are clear and that they encompass complete functions."

Jan 1, 1988

By Sean Gallagher, John R. Bartels, Dean H. Ambrose

Remote operation of continuous miners has enhanced the health and safety of underground miners in many respects; however, numerous fatal and non-fatal continuous miner struck-by accidents have occurred when using remote controls. In an effort to prevent these injuries, NIOSH researchers at Pittsburgh Research Laboratory examined the workplace relationships between continuous miner operators and various tramming modes of the equipment using motion captured data, predicted operator response times, and field- of- view data to determine causes of operator-machine struck-by events in a virtual mine environment. Factors studied included machine speed, direction of escape, operator facing orientation relative to the machine, work posture, distance from machine, and operator anthropometry. Close proximity to the machine, high machine tramming speeds, a right-facing orientation and operator positioning near the tail all resulted in high risk of being struck. It is hoped that this data will provide an improved rationale for operator positioning for remotely operated continuous miners.

By U. Schoepf

Economic and at the same time environmentally acceptable mining of small vertical ore bodies and veins to great depths has always been a challenge for miners. The Bauer cutter system, widely used in civil engineering, has been further developed for applications in exploration and mining, to a sustainable system for selective mining with a focus on small vertical ore bodies and veins, as well as residual ore in open pit operations. Current situation of conventional mining of small ore bodies and vertical veins Small ore bodies and vertical or near-vertical ore veins often pose a challenge to the mining industry. The relatively small size of the ore deposit over great depths or lengths, with only limited exposure to the surface makes economical mining difficult. Open pit mining of such ore bodies must either be terminated early because of the high stripping ratio and the enormous environmental impact or is unprofitable from the very beginning. The required massive stripping of waste rock leads to high material hauling times and costs, intensive roadway and bench construction and maintenance, possible safety hazards due to heavy vehicle movement, pit wall stability risks, intensive surface water management and massive waste dump requirements. In addition, it can be assumed that increasingly extensive rehabilitation measures will be required in the future after mine closure. Particularly in vertically oriented magmatic and hydrothermal deposits the technical, economic and environmental limits for further pushbacks, to extend the open pit depth, can lead to an early end of mine life even if a significant quantity of residual ore is left behind. Going underground is only a viable option in a limited number of mines with very high ore values and volumes. The high development cost of an underground mine, as well as the higher underground mining costs, often prevent further mining to greater depths.

Feb 1, 2023