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By A. Newman, A. Brickey

A pesar de que parte del predominio de software de programación de producción a cielo abierto se puede explicar por la preponderancia de la minería a cielo abierto en todo el mundo, hay otros factores que han dado lugar a un retraso en el software subterráneo correspondiente. Explicamos por qué matemáticamente la programación de la producción subterránea es más difícil que su contraparte a cielo abierto y daremos pautas para la investigación en el campo de la programación subterránea.

Jan 1, 2014

By C. Roos

Tradicionalmente en la industria minera, el flujo principal del proceso funcional para la evaluación de recursos y reservas procede de geólogos que elaboran el modelo geológico, de allí a los geólogos que hacen la estimación de los recursos (los modeladores) quienes estiman los grados, y a los ingenieros de mina que desarrollan el diseño de la mina y su calendario. La interfaz entre estas tres disciplinas no está siempre exenta de baches. Las presunciones hechas durante una etapa del proceso no siempre son comunicadas a los demás, a pesar que la mayoría de las empresas mineras tienen procedimientos y protocolos establecidos para promover la comunicación a lo largo de las distintas disciplinas. Estos procedimientos no siempre logran su propósito debido a una variedad de razones que incluyen restricciones de tiempo, no tener disponible a las personas apropiadas, o a una falta de entendimiento de lo que las otras disciplinas requieren, y el proceso de traspaso puede degenerar en el envío de archivos a la siguiente disciplina en el flujo de proceso. Este artículo se ocupa de la interfaz entre la estimación del recurso y las disciplinas de la ingeniería de minas y el esquema de algunas de las presunciones comunes hechas al generar un modelo del recurso y los riesgos que los ingenieros podrían correr al elevar los grados de corte, optimizar calendarios o realizar predicciones. Los casos de estudio presentados se centran en depósitos de oro a cielo abierto, pero los conceptos deben ser aplicables a toda la industria. Aquí también proveemos un resumen de preguntas clave y recomendaciones.

Jan 1, 2014

By F. A. Gates, Bengt Samuelson, James E. Edmunds, S. McCune, Neil Hario, Norman L. Weiss, J. M. Bertram, A. M. Cavaliere, F. M. Jr. Stephens, D. D. Chiang

DESIGN BASIS Size of Project The size of the project generally is expressed in tons of ore milled per day, but company policies differ widely in this respect with a few adhering to this set figure but many expecting the design capacity to be liberal. To some companies that follow the latter practice a 25,000-tpd milling plant that cannot reach 30,000 or 35,000 tpd with¬out major additions is a disappointment. The engineers responsible for planning the operation must know what is intended. Other projects are sized according to the tonnage of product rather than of mine ore; examples are iron ore mills that express size in terms of tons per year of pellets or high-grade concentrates, heavy minerals plants in tons per year of rutile or ilmenite, cement plants in barrels per year, potash in tons per year of muriate, phosphate in tons per year of washer rock and flotation concentrate, and others. However, tons per day milling rate is also the common denominator in most of those instances because the major equipment has to be selected on that criterion. Equally important are operating schedules, particularly the number of operating days per week and shifts per day for ore delivery, crushing, grinding, and processing. Hours per shift also are an impor¬tant factor in design, particularly in the case of ore delivery, and the lost time expected for maintenance, cleanup, inspection, preventive maintenance shutdowns, and many others, has to be taken into ac¬count. Frequently, the ore deposit will vary in hardness and grade to such a degree that the capacity of the mill will vary widely from year to year and cannot be expressed as a fixed figure, but rather as an average over a calculable period, such as ten years, or as a specific daily capacity over the first two years, next four years, etc., or some similar projection. A good example of this kind of situation is Asarco's Mission operation 15 miles southwest of Tucson, Ariz., described by Weiss and Vincent,1 where the major rock types differed widely in hardness, so that a change in the proportion of these types in the mill feed could have imposed difficult design problems if a constant tonnage rate had been expected. Expected Life The life of a mining operation depends upon the size of the deposit and the rate of mining. The latter is a policy decision based on an economic analysis2 which, in turn, is based on many factors. From the point of view of mill planning, the estimated life of the operation determines many criteria and affects many decisions on strength of structures and degree of protection of men and materials, the quality of the materials-handling and process machinery, the type of delivery systems for ore to the mill and concentrates to the smelter or market, environmental considerations, and many others. Costly installations that may be excellent investments for a 20-year project may be waste¬ful for a 5-year operation. An example is the comparison of different methods of crushing or grinding. Take single- vs. multiple-stage grind¬ing-the former may be found more economical for a 5-year operation, and the other, more costly method more economical for a 15- to 20-year life. Mining and Ore Transportation The mining method and the means of transportation of the ore to the first crushing stage have far-reaching effects upon mill design, particularly the crushing sections, and upon site selection (see Chapter 2, Situation of Mill). Consequently, mill design cannot advance until the mine planning has reached at least tentative decisions concerning these important matters, or reach the final stages until much is known about the ore that may be expected month by month and year by year for a reasonable period. Methods of ore transportation used in over 90 underground and open-pit mines are listed briefly in Section 32 (Table 1), and more detail is found later in this section and in Section 10, Storage and Transport. These methods include surface railroad, trucks, conveyors, surface skips, underground skips, and tramming from hoist or tunnel. In surface mining, transportation of the uncrushed ore by rail or truck to the primary crushing plant-conveniently located for lowest overall cost-is most common. Rail transportation provides low perton-mile costs. Truck transportation, though in some cases less eco¬nomical, is more flexible and more adaptable to mountainous terrain; furthermore, recent improvements have reduced the operating costs to keep trucks competitive on low-grade ores. The design of the primary crushing plant will be essentially the same whether delivery is by truck or rail, but the location of the plant is subject to more rigid restrictions in the latter case because of grade limitations, and these in turn may adversely affect the selec¬tion of the mill site as well. The influence of truck haulage in reducing the stripping ratio in open-pit mines is an added reason for predicting a minor role for rail transportation in the future. The use of conveyors to transport ore from the pit to the milling plant involves (in most cases) crushing the ore to a reasonable size (10 to 20 in. maximum with today's crushers) before conveying. The Twin Buttes system3 which went into operation in late 1969, utilizes primary crushers (two) and conveyors to receive rock waste, ore, and low-grade stockpile material from trucks, then transport them to the appropriate destination. Flexibility is dependent upon the porta¬bility of the primary crushers. In the face of this development, inclined skip hoists are no longer important to the mill designers. Underground mines generally pose fewer problems of design; these will be discussed in Chapter 2. Character of Deposit The degree of homogeneity of the ore in various parts of the mine is an important factor. The characteristics that concern the millman (approximately in order of importance) are: hardness, assay, metallurgical response, size distribution,4 and moisture content. Under metallurgical response may be included degree of oxidation, mineral¬ogy, clay content, and reagent consumption, as well as process re¬sponse in general. Short-term variations-hour-to-hour, day-to-day, or even week-to-week-make operation of the plant more difficult, require greater flexibility of process design and, perhaps, also require more instrumentation and manpower. Uniformity of the ore in the characteristics noted previously and

Jan 1, 1985

By Kenneth P. Katen

Though the use of continuous mining machines consolidated the operations of cutting, drilling, blasting, and loading in one machine that would theoretically provide uninterrupted production, it was soon realized that the face haulage and roof support operations remained as major impediments to continuous production. The ability to provide not only continuous cutting and loading, but also continuous haulage and roof support (see Fig. I), was the basis for the introduction of mechanized longwall systems in this country from Europe in 1951. The early systems employed friction props for roof support, but were to become obsolete shortly after introduction of the first hydraulically powered, self-advancing supports in 1960. Since then, use of the longwall technique has grown to the point where it is currently providing more than 5% of the annual output of underground coal mines in the United States. The shortwall system had its origin in Australia, but was first introduced into this country in 1962 following the development of powered roof supports. It, too, provides for continuous cutting, loading, and roof support (see Fig. 2), but has not experienced the widespread popularity of longwall due to its inherent inability to easily provide for continuous haulage. While there are in the neighborhood of 100 longwall systems currently in operation in this country, there are only a handful of short- wall operations providing production of slightly more than 907 184.7 t/a (1,000,000 stpy) Both of these systems provide uninterrupted roof support for face operations by means of hydraulically powered steel roof supports. In each case, the coal removal consists of open-end cuts made from the narrow end of a large rectangular pillar (panel) which has been blocked out by continuous miners developing entries around its perimeter. However, this is the extent of the similarities between the two systems. It is hoped [ ] that the information presented in this chapter will enable the reader to better understand the two mechanical systems and the mining concepts involved in their employment.

Jan 1, 1981

By K. M. O’Connor

Time Domain Reflectometry (TDR) is a pulse testing technique developed by the telecommunications and power industry for use in locating breaks in cables. Variations of this principle have been utilized in applications pertinent to closure and reclamation of coal operations. This paper focuses on applications of TDR technology in the mining industry including monitoring of slope movement, abandoned mine subsidence, ground water levels and subsurface LNAPLs, landfill cover performance, and consolidation of mine tailings. TDR technology has been particularly effective for remote automated monitoring, and alarm systems have been implemented for monitoring subsidence.

Jan 1, 2003

By Wayne Charlie, Joseph P. Martin

INTRODUCTION Typically, uranium mills produce slurred tailings which are transported by pipeline to a tailings impoundment. The slurried tailings contain about 250 solids (75% liquids) by volume and, depending on the milling process, have a high or low ph. Despite decanting water from the tailings pond and losses by evaporation and seepage, Staub (1978) estimates that for lined impoundments over half of the input water is stored in the tailings. As long as the impounded tailings remain in a loose and saturated state, potential hazards exist and include potential for short and long-term seepage from the impoundments and the potential for large scale release and movement of liquified tailings should an embankment fail. The breach of an earthen dam on July 16, 1979 at Church Rock, New Mexico, and the subsequent release of 100 million gallons of water and liquified uranium tailings, underlines the potential hazard. Loose saturated tailings also have a high potential of liquefaction during earthquakes (Charlie and Wardwell, 1979; Martin, Finn and Seed, 1978; and Mittal and Morgenstern, 1976). If the tailings could be dewatered, the liner's integrity would become less critical, longterm seepage and settlements would be greatly reduced. In addition, the potential for tailings liquefaction from an embankment failure or earthquake would be greatly reduced. This paper reviews the concept, criteria, and design of various proposed interior drainage systems to dewater the tailings after deposition. ADVANTAGES OF INTERIOR DRAINAGE The primary purpose of interior drainage systems for uranium mill tailings impoundments is dewatering the tailings as rapidly and as thoroughly as possible. The expected benefits of this new technology are: • Reduction of potential seepage losses by decreasing the hydraulic head. • Reduce earthquake liquefaction potential of the tailings by inducing non-saturated conditions and/or lowering the water pressure. • Reduce the potential for large-scale release and movement of tailings should an embankment fail. • Stabilize the tailings to permit early reclamation. • Recovery and potential recycling of tailings liquid. • Inhibiting long-term seepage out of the impoundment by inducing nonsaturated conditions to provide capillary suction in the tailings.

Jan 1, 1980

By Stephen G. Hardcastle

CANMET has developed a simple yet very effective tracer gas method to test the efficiency of stench gas emergency warning systems which can easily be performed by mine personnel. This test method uses sulphur hexafluoride (SF,) as a surrogate for such stench gases as ethyl mercaptan. The test has many advantages: it is non-disruptive to production because SF, is odourless, colourless; it is safe to use as it is inert and only a small volume, insufficient to reduce oxygen levels, is necessary to dope the air with a measurable concentration; and most importantly it provides definitive results to assess the adequacy of the warning system. Specific findings of these tests have included, over doping sections of the mine and subsequent production loss due to sickness, the warning being ignored, the injection period being too short, excessive dispersion, and inadequate notification times. The simple test protocol involves releasing a known volume of the SF, under similar circumstances to those of the stench gas, and then sampling for its presence downstream. These samples are collected in air-tight syringes which are then analyzed in a laboratory with a gas chromatograph. By direct proportion, between the size of the stench and tracer gas releases, it is possible to convert the SF, concentration into equivalent stench gas concentrations profiles. From these, it is possible to define, the initial arrival time of the gas, the time it takes to reach a "target" notification concentration, whether the gas peak will exceed TWAEV, STEV and CEV exposure limits, and the duration of the peak above these levels. Since 1993, this test protocol has been employed a total of eight times on four different Canadian mine properties and the results are being used to optimize the stench gas release system with respect to volume of gas used, duration of injection, and number of injection locations.

Jan 1, 1997

By A. Gene Collins

The demand for lithium may increase because of its potential use id lithium batteries and in the generation of electrical power by fusion. Therefore, lithium abundance in some United States oil-field brines was surveyed. Oil-field brines from numerous locations were found to contain more than 100 mg/l lithium and several exceeded 500 mg/l. An area of 25,000 km2, with a reservoir thickness of 60 m, an effective porosity of 5 percent, and a brine with a specific gravity of 1.2 containing 100 mg/l of lithium, contains about 0.75 x 106 tonnes (1 tonne = 1 metric ton = 1,000 kg) of lithium in 7.5 km3 of brine. It is postulated that the brines were enriched in lithium by lithium-rich spring water or lithium leached from volcanic rocks followed by evaporation sequences.

Jan 1, 1978

By S. Constantinescu

The alloys mainly used for short (or segmented) chip material cutting (cast iron, porcelain, etc) are alloys of the WC-Co type; in some cases with small additions of other carbons. The same alloys but of different granulometric classes are also used for pieces exposed to wear (wire drawing, moulds) The higher performance of the WC-Co alloys in chip machining applications is accounted for by their very good heat conductivity, which is 2-3 times higher than that of fast steels. The hard-alloys producers can resort to a wide range of possibilities to achieve variation of properties of a WC-Co composition and thus they can adapt their products to the particular type of tools they may choose to manufacture. The present metallography technique allows for a correct identification and evaluation of hard sintered carbons structures. The metallography approach is an indispensable method of investigation and control in the industry.

Jan 1, 2008

By A. Ersoy, M. D. Waller

Because of a shortage of diamonds for the drilling industry in the 1970s, efforts were made to develop synthetic diamonds. The resulting development of polycrystalline diamond compacts (PDC) enabled the evolution of a new generation of rotary-drilling drag bits. Bits based on PDC cutters were very successful in drilling soft, abrasive and medium strength rocks. However, because of their thermal instability, they failed rapidly when drilling hard rocks. Thermally stable diamond (TSD) was then developed for hard-rock drilling. The rate of penetration (ROP) and bit life significantly influence the drilling cost. The ROP and the life of the TSD bits or cutters are affected by a number of complex factors. However, in laboratory testing and infield operations, the principal factors that need to be considered in predicting drilling and wear rates are bit design characteristics, bit operating parameters and the characteristics of the penetrated rock. These characteristics are comprehensively reviewed for both laboratory and field drilling results using TSD bits. Optimization of these parameters increases the drilling rate, prolongs the bit life and reduces the drilling cost.

Jan 1, 1998

By Anibal Gajardo

Clays for building bricks and roofing tiles in Chile are classified as common clays and ball clays, with deposits of sedimentary and residual origin, both Cenozoic in age. Deposits are known from north to south, through 11 of the 13 Chilean Administrative Regions, 18° to 39° South latitude, and from west to east, through three of Chile's main geomorphic features, namely the Coastal Plains, Coastal Range, and Intermediate Depression. Alluvial and lacustrine sedimentary common clay deposits are located in the northern and southern Intermediate Depression. The average chemical composition of clay deposits in I and I1 Regions, is 58 percent Si02, 16 percent A1203, and 4 percent Fe203 by weight; the primary minerals are sericite, quartz, feldspar, and albite. The average chemical composition of clays in Metropolitan, VI, and VIII Regions, is 52 percent Si02, 16 percent A1203, and 6 percent Fe203 by weight; the primary minerals are montmorillonite, quartz, and plagioclase. Residual common clay deposits occur in the Intermediate Depression and in the Coastal Range internal basins, VII, VLII, and IX Regions, and they represent surficial alteration of pumice beds and crystalline rocks. The average chemical composition of clay deposits derived from crystalline rocks is 58 percent SO2, 17 percent A1203, and 9 percent Fe203 by weight; the primary minerals are sericite, quartz, and clinochlore. Clay deposits formed from pumice beds contain 47 percent Si02, 24 percent A1203, and 10 percent Fe203 by weight; the primary minerals are kaolinite, feldspar, and cristobalite. Sedimentary ball clay deposits are located in the Coastal Plains and Coastal Range, VI, VII, VIII, and IX Regions, between 34" and 39" South latitude. The average chemical composition of the ball clay deposits is 51 percent ,902, 28 percent A1203, 5 percent Fe203 by weight; the primary minerals are kaolinite, quartz, and sericite (illite?). These clay deposits supply the Chilean building sector with over one million metric tons (mt) (1.1 million short tons [st]) of material per year. The average growth rate of clay consumption during the period 1993 through 1997 was 10 percent annually. The clays are used by both mechanized plants and handcraft facilities.

Jan 1, 2001

By Rolf Pollak, Jan Tegtrneier, Gerhard Keller, Horst Gerhardt, Wolfgang Müller, Dieter Tetzner

INTRODUCTION Under certain geological and mining conditions mining near the surface results in an intensive destruction of the overlying rock strata. If the system of mine openings remains filled with air after having completed all mining activities because the flooding of the mines depending on the morphology of the mining area does not exceed a certain level, this circumstance might result in dangers for the public. This is particularly true where uncontrollable air- ways occur within the destroyed rock or if former uranium mining areas are concerned. In these areas radon emanation resulting in undue radon concentration should be taken into account. This circumstance can be counteracted by suitable measures of ventilation. In the following the example of a uranium deposit in the Erzgebirge Mountains (Saxon Ore Mountains) is used to illustrate how statements about the ground flow through overlying rock can be made on the basis of systematic investigations. Finally comparisons with experience gained in coal mining yield orientation data which are used as input quantities in ventilation network calculations and in project preparation of mine fans. Last not least it is proved that endangering the population by radon can be excluded by installing an appropriate fan. INITIAL SITUATION From 1946 to 1989 the uranium deposit of Schlemal-Alberode was mined starting from superficial zones down to depths of 1,800 m. The total output amounted to 80,000 t of uranium, i.e. Schlemal Alberode is one of the largest uranium deposits in the world. The mine openings comprise a cavity of 40 million m3. Al- though thin seams (0.2 to 0.3 m) and steep to semi-steep veins (60 - 70 degrees on average) were extracted using the method of overhand cut-and-fill there were also deposit sections with larger thicknesses and poor ore zones, respectively, where especially after the war until about 1955 large openings were formed due to missing filling. Due to this uncontrolled, relatively "wild" mining considerable subsidence of the surface and fractures in the border zone of the subsidence area occurred. The town of Oberschlema was destroyed and an area of 0.14 km2 was damaged by mining involving a maximum subsidence of 6 m in the centre - called deformation area. The rehabilitation work for the deposit area of Schlemal Alberode involves the flooding of the openings up to a level of the Markus-Semmler-Sohle as long-term preventive measure. Fig. 1 shows the flooding level of October 1996. The level of the -540 m drift will probably be reached in June 1997. From that point of time a general rearrangement of the ventilation has to be made because the central return air ventilation via shaft 373 will no longer be possible. This shaft has no connection with the superficia1 system of mine openings. A new return air ventilation system based on shaft 382 as return air shaft will be designed (Fig. 1). The capacity of the future mine fan system will e.g. be determined by the flow conditions in the deformation area mentioned above. For this reason investigations were carried out in a test drift 13a on the level of the Markus-Semmler-Sohle in the stretching zone of the subsidence trough for several days (Fig. 2 and 3). INFLUENCE OF THE OPERATING CONDITIONS OF THE MAIN MINE FAN ON THE RADIATION SITUATION ON SURFACE The beginning of flooding in 1991 was also the beginning of the investigations into the options of future ventilation due to the reduction of the mining area. The first analyses of radon emanation made in May 1991 showed that only about 20 % of the total radon emanation result from systematically ventilated mine workings. This means that despite of existing dams the air streams through the majority of mine openings of the deposit. The analysis of the individual radon inflows illustrates the special problematic nature of the deposit involving all possible effects on the surface. The shutdown of the main mine fan over the long period between 2.10. and 6.10.1991 showed the essential influence of the operating conditions of the fans on the whole radiation situation on surface. Fig. 4 shows the comparison of the radon concentration level and its daughters during a 'normal" weekend with the level during long-term shutdown in the old 'Casino" of the papermill of Niederschlema. This papermill as well as the 'Casino" are situated in the outcropping zone of the deposit section of Niederschlema. The -60 m level was driven as the upper level in this area, from which the panels were driven to about 20 m below the surface. Also below the "Casino" there is such a panel in the vein 'Glirck", whose roof is directly situated below the building in a depth of 39 m. Directly next to it there is the raise 2 of the panel 'Namenlos" in a depth of 27 m. The graphics clearly shows the immediate increase in radon concentration as well as in the daughter concentration after the shutdown of the fans. The elimination of the artificially generated negative pressure results in the reversal of the convection direction due to the natural ventilating pressure. The interface structure allows the mine airto pass through aero- dynamic connections to the surface. Neglecting the meteorological parameters the concentrations of radon and its daughters in- creased by about 350 % as compared to a 'normal" weekend during longterm shutdown (i.e. in 4 days). Furthermore it could be noticed that a stationary state between the radon and its daughters has not been established in the period

Jan 1, 1997

By S. S. Klima, D. E. Pollock, J. A. Organiscak

"Airborne respirable coal dust capture by water sprays or wet scrubbers has been studied and developed over many decades as an engineering control to reduce dust exposure in coal mines and combat coal worker pneumoconiosis. Empirical relationships and deterministic models for particular dust capture experiments have previously been devised to show the key parameters involved in airborne coal dust capture. Many of the results from these models show that the significant parameters related to airborne dust capture are water spray pressure, water quantity, water droplet size, relative water droplet-to-dust particle velocity, and total operating air pressure of the scrubber. However, many airborne dust capture efficiency relationships and models developed for particular experiments cannot be readily applied to forecast the dust collection efficiency of different spray and scrubber design configurations, which rely on several key dimensional engineering measures. This study examines engineering measures from previous water spray and wet scrubber experiments conducted by the U.S. National Institute for Occupational Safety and Health (NIOSH) and the U.S. Bureau of Mines (USBM) to develop empirical models for wet collection of airborne dusts. A dimensionless empirical model developed for predicting airborne dust capture efficiency of water sprays and wet scrubbers is presented. IntroductionWater sprays have been one of the earliest mainstay engineering controls used to suppress respirable dust in underground coal mines. Some of the earliest reports of water spray usage on continuous mining machines indicated respirable coal mine dust reductions of 20 to 60 percent (Kobrick, 1970). Not long after the enactment of the 1969 Coal Mine Health and Safety Act, the U.S. Bureau of Mines (USBM) and others conducted extensive research into water spray and wet scrubber systems to control respirable dust in coal mines. The primary aspects of water spray dust control included prewetting for dust prevention, localized ventilation redirection and airborne dust capture, which are summarized in handbooks or best-practices guides (Kost, Yingling and Mondics, 1981; Kissell, 2003; Colinet et al., 2010). Initially, underground studies were conducted to identify the primary sources of dust during the mining processes and effective dust controls (Courtney, Jayaraman and Behum, 1978; Jankowski and Organiscak, 1983). Laboratory research systematically studied ventilation and airborne dust capture effects of water sprays and scrubbers to improve their application in underground coal mines (Tomb, Emmerling and Kellner, 1972; Divers and Janosik, 1978, 1980; Grigal et al., 1982; Ruggieri et al., 1983; Jayaraman, Jankowski and Kissell, 1985; Volkwein, Ruggieri et al., 1985; Jayaraman, Schroeder and Kissell, 1986; Divers, Jankowski and Kelly, 1987; Jones and James, 1987)."

Jan 10, 2018

By Paul Eimon

In the world of minerals exploration, computer-oriented data storage and retrieval methods are being talked about, studied and researched. The goal is to facilitate the recording, collating, indexing, recovery, and analysis of mineral deposit-prospect data. A survey has been made of the various report forms, suggested outlines, and computer coding forms that are currently used in mineral exploration to index and/or describe mineral deposits and prospects. These forms have a number of common characteristics: location, commodities, type of deposit, host rock, etc. Such basic mineral deposit-prospect "data units" are part of most exploration programs. "Data units" may be comprehensive bound published descriptions of mines or mining districts or 3x5 file cards or elaborate computer coding forms being developed in Canada. Initial attempts to develop computer-oriented information systems have run into the same roadblock: the lack of a body of organized, processible data describing the basic qualities of mineral deposits and prospects. To solve this problem, international committees are considering coding forms, world-wide semantics, and rules or standards for data collection and description. It appears that mistrust and lack of understanding of the computer; fear of loss of proprietary information; and disagreements over terminology are distorting the picture of what has to be done if the mining industry is to grow into more meaningful information systems for mineral exploration.

Jan 1, 1969

By E.C. Jong

Resumen ¦ Un sistema de comunicaciones “a través del suelo” (TTE) comercial y disponible fue evaluado en una mina de carbón subterránea en la zona Este del Estado de Kentucky. Este estudio de campo fue realizado como parte de una evaluación multisitio mayor de sistemas TTE disponibles. La mina detallada en este estudio se encontraba paralizada al momento de la evaluación, lo que proporcionó la representación más cercana a la realidad del medio ambiente en una mina luego de una operación de cierre. Para este estudio se probaron las comunicaciones superficie a subterráneo y subterráneo a superficie. El despliegue estándar del sistema TTE usando los procedimientos recomendados por el fabricante logró obtener comunicaciones claras y nítidas entre el socavón y la superficie. Otras disposiciones de las unidades arrojaron resultados mixtos. Existen planes para examinar más los fenómenos observados en esta mina.

Apr 1, 2016

By Irwan Halim, Bhoshaga Mitrran Ravi Chandran, Kevin Huynh

Tunnel segmental lining design constitutes a significant iterative process where modeling and numerical analysis are factors in deciding the number of segments, their dimensions, gasket placements, as well as the locations of dowels and bolts. Consequently, these modifications lead to revisions in the final drawing details. However, the current tunnel engineering practices lack systematic solutions that effectively bridge the gap between the iterative design process and efficient visualization of completed drawings due to the time- consuming nature of redrawing. This paper presents a pioneering application of AI-driven solutions including using Autodesk's API, for the Northside Interceptor Tunnel (NSIT) Project. It showcases the automation of drawing design and the real-time visualization of modifications, thereby streamlining the drawing process.

Jun 23, 2024

By Dwight K. Smith

Fly ash as an additive for oil well cement was introduced to the oil industry by the Halliburton Company in 1949. Since its early usage in the Illinois area of the U. S., the total amount of fly ash pumped into wells has exceeded 30 million sacks, The advantages of using fly ash-portland cement mixtures in wells are fundamentally the same as for concrete except the conditions of placement and curing are variable and the nomenclature of properties are expressed somewhat differently. In some instances deep wells have been cemented with mixtures of fly ash and hydrated lime together with an activator which functions as a catalyst for setting. This fly ash-lime formula has many advantages for high temperature-pressure conditions found in wells.

Jan 1, 1968

By F. Calizaya

A booster fan is an underground ventilation device installed in the main airstream to handle the total quantity of air circulated to one or more working districts. It is installed in a permanent stopping and equipped with airlock doors, a monitoring system and interlocking devices between the main fan and the booster fan. When adequately installed, a booster fan can be used to reduce the main fan pressure, reduce leakage and decrease the power consumption. However, an inadequate installation can increase the likelihood of mine fies and recirculation of air contaminants. This study identifies hazards associated with the operation and maintenance of booster fans. It analyzes the accompanying risks and recommends control measures to reduce those risks to acceptable levels. The ventilation network of an existing coal mine was updated to include a booster fan. The hazards associated with the operation of this fan were identified, the risks analyzed and the esponse to each failure mode established.

Mar 1, 2014

"A commercially available through-the-earth (TTE) communications system was evaluated at an underground coal mine in eastern Kentucky. This field study was conducted as part of a larger multisite evaluation of available TTE systems. The mine detailed in this paper was idled at the time of the study, which provided the nearest practical representation of a postevent mine shutdown environment. For this study, surface-to-underground and undergroundto- surface communications were tested. Standard deployments of the TTE system using the manufacturer’s recommended procedures were able to achieve clear communications between underground and surface locations. Other arrangements of the units yielded mixed results. Future experimentation is planned to further examine the phenomena observed at this mine IntroductionThe ability to establish effective communication between surface and underground personnel after a mine event is essential for both evacuation and rescue efforts. Current wireless underground communication technologies are reliant on intermediary physical infrastructure, such as antennas and repeaters, to enable underground communications. As such, communications are easily disabled during a mine event because of the inherent vulnerabilities to physical damage (Damiano, 2012; National Academy of Engineers, 1970; Yenchek et al., 2011). A true emergency communications system would function regardless of a mine’s physical state. Through-the-earth (TTE) wireless communication technology only requires the use of a single transmitter and a single receiver to achieve point-to-point communication through solid strata. With this capability, TTE systems have the potential to serve as a true emergency communications platform."

Apr 1, 2016

By V. P. Petrov, V. S. Znamensky

There are a great number of asbestos deposits in the USSR, and some are quite large. The main deposits of chrysotile asbestos are shown in Fig. 1. The Bazhenovo deposit, which is located in the Sverdlovsk region, is the largest and has been mined since 1889. Systematic studies of asbestos deposits were begun from the very be- ginning of Soviet government and are most often connected with the names K.E. Tarasov, P.M. Tatarinov, B.J. Merenkov, N.D. Sobolev, F.V. Syromjatnikov, V.N. Lodochnikov, V.P. Petrov, V.P. Eremeev, L.A. Sokolova, V.R. Artemov, and V.F. Dybkov. Investigations by these people and others helped the USSR be one of the most important producers of asbestos in the world. In this chapter, the main asbestos deposits of the USSR are explained in a general outline from the point of view of their genesis, common features, and differences that must be taken into account for forecasting, exploration, and the mining of asbestos. All chrysotile asbestos deposits of the USSR were formed from magnesium rock as a result of comparatively low grade temperature, hydrothermal alteration, and free crystallization of fibrous chrysotile in open cavities that served as catalysts. Chrysotile, a widely distributed rock-forming mineral, is not in general an important mineral for industry as it is a long fibrous and main vein variety. However, it must serve for geologists as an important prospecting criteria. And, the geologist should search for the mineral not only in places of possible chrysotile generation, but in places where asbestos may be crystallized in its long fibers. The Soviet geologists distinguish chrysotile asbestos deposits in two groups: (1) ultrabasic rocks and (2) sedimentary dolomite terranes, taking into account how the maternal rock was affected by hydrothermal solutions. The largest deposits are related to ultrabasite. In dolomite, the largest deposits are small and their asbestos contains almost no iron, which makes it a better filler for insulating plastics. There are a few types of asbestos deposits within the limits of the ultrabasic group, the largest being a Bazhenovo type to which the deposits of the Urals are attributed. This Bazhenovo asbestos is characterized by the most complete zonality of disposition for various kinds of asbestization. These types of deposits are, for the most part, the only ones presently working in the USSR. The Bazhenovo deposit is connected to the gabbro-peridotite intrusive belt that is about 180 km long and is broken into a number of fault blocks as shown in Fig. 2. Granitoid veins of intruded rocks are found along the deepest faults, and were sources of hydrothermal solutions that affected the maternal rock. The nearest granitoid vein, ultrabasite, is substituted by listwanite that consists of quartz, talc, carbonate, and some other essentially talcous rock. The next zone of hydrothermal alteration is composed of serpentine. The boundry between serpentinite and unaltered ultrabasite is as irregular, diffuse, and rough as the boundaries between all other zones. Nearer the vein of granitic rock, ultrabasite is entirely serpentinized into massif serpentinization that penetrates through cracks and fissures. There is an unaltered ultrabasite space between fissures, and the biggest block is preserved unaltered in the central part of the massif. The Soviet geologists call this ultrabasic sopka. Preservation of the unchanged blocks is of great importance for deposit formation as volume changes are impossible during the process of asbestization and serpentinization. It may be confidently said that asbestization in Bazhenovo type deposits is possible only when alteration of ultrabasite has not completely occurred. Under conditions of massif volume, a very big mass of substance must be carried out of the ultrabasite for the progressive stage of serpentinization. However, in the regressive stage,

Jan 1, 1986