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A knowledge of airway friction factor (K-factor) is necessary to designing mine ventilation systems for least cost and optimum effectiveness. The friction factor used must be accurate and reliable to estimate correctly roadway resistances and pressure sources for required flows. Though the resistance to flow offered by mine airways has been an object of study over the last half century, no accurate relation has yet been established from which the prediction of the resistance of a projected airway can be made. In Kentucky, only limited and incomplete data on local K-factors are available. Therefore, there is a need for field investigations to determine these values in-situ for existing mining conditions. The author describes a study concerning K-factor that was carried out in ten coal mines, located in eastern and western Kentucky. The results of these investigations are presented in both tabular and graphical form.

Jan 1, 1991

By Amr Sallam, Ahmad Mohamed, Mohamed Alrowaimi

Renewable energy, in the form of wind and solar, has emerged as main source of energy in the modern society. The wind industry has substantially grown over the last two decades to become a key player in providing energy to public. Wind turbines are getting larger, taller, and more efficient every year. The foundation is a major component of the wind turbine system. The wind turbine foundation must provide short and long-term stability for the overall turbine to allow continuous and safe turbine operation for energy production. The turbine serviceability and stability requirements must be geotechnically and structurally satisfied. In North America, major foundation types currently utilized in the wind industry include gravity spread foundation, mono pier foundation, and pile foundation. In this case study, a Post Tensioned Pile Cap (PTPC) foundation was used to support a 3.15-3.4 Megawatt turbine with 125M Rotor supported on 87.5M-hub-height tower. Soil anchors in the form of Auger-Pressure-Grouted-Displacement (APGD) piles and Continuous Flight Auger (CFA) piles were utilized. Four distinct subsoil conditions were identified, and for which four pile designs were adapted. Due to the occasional encounter of stronger soils at 50 to 55 feet (15.2 to 16.7 m), shorter pile design was preferred. To satisfy the design uplift load, pile post grouting were utilized to increase the shaft capacity of the piles. In this paper, the pre-production pile load test program will be described. The pile load test procedure, analysis, and results will be presented. A discussion will be provided, which led to the selection of the optimum pile design to support the high tension loads taking into consideration pile length limitations.

By F. Nakhaei, M. Le Roux, Q. P. Campbell, S. Botlhoko

The results of coal sink and float tests are plotted as washability curves. Analysis of these curves indicates the most effective method by which the coal can be cleaned. Sink and float experiments are samplingdependent, destructive, and use toxic chemicals. The present research introduces an alternative method based on a 3D imaging system called RhoVol, in which the density of individual particles is determined from mass and volume measurements. A quantitative error analysis was conducted on coal sample density measurements obtained by RhoVol, and the main factors influencing the errors identified and investigated. The results show that the drawback of this technique is related to its not being able to detect hidden concavities in the particle, resulting in an overestimation of particle volume. Furthermore, it is difficult to capture the impact of porosity and cracks on the volume of coal particles obtained from 3D silhouette images. This mismatch between the RhoVol data and sink and float test results has led to attempts to apply neural network (NN) and linear regression techniques to produce a fully reliable model for correction and estimation of coal density. The NN approach offers superior predictive capability over linear regression, and the estimated density distribution is in line with the sink and float analysis. The root mean square error in estimation of density using the NN model was less than 0.05 g/cm³.

Aug 2, 2022

By Douglas Glenn Stalheim

Production of structural hot strip coils in yield strengths up to 700 MPa and higher is increasing worldwide. In addition to requiring higher strengths, structural designs require good formability, toughness and weldability. Toughness and formability come from optimization of ductility properties while weldability comes from lower carbon equivalent alloy designs. The key to the production of optimized cost effective hot strip produced structural coils is to create the correct balance of a proper alloy/processing design to create as fine and uniformly distributed cross sectional grain size as possible. In addition to the creation of the fine/uniformly distributed cross sectional grain size, the correct microstructure needs to be created. In the higher yield strength grades, not only does the grain size/distribution play a role but a proper understanding and control of microalloying precipitation kinetics is required. This paper will present various examples of hot strip alloy designs to produce cost effective hot strip structural coil in yield strengths up to 700 MPa with good ductility and weldability. In addition, steelmaking/casting parameters will be discussed to assure successful production. Optimized reheating, rolling and cooling parameters of these alloy designs will be introduced in an accompanying paper.

Aug 17, 2017

By M. S. Shishvan, J. Sattarvand, R. S. H. Meij

Navigating miners during an evacuation using a smart evacuation technology can significantly decrease the evacuation time of an underground mine in case of an emergency. This paper presents a mathematical programming model to calculate the most efficient escape path for each miner as a critical component of the smart evacuation technology. In this model, the total evacuation distance of the crew is minimised and scenarios with blocked pathways, and stamina categories for the miners are simulated. It was found that all the tested scenarios are technically feasible. Using the feature that filters out blocked pathways has no downsides, as safer routes are calculated, and there is no penalty in the computation time. The paper also discusses the social and technical issues that need to be overcome before the algorithm can be implemented as an actual escape solution.

Sep 1, 2021

By Robert L. Martin, Brian W. Stump

"The Black Thunder Mine is located in the Powder River Basin of Northeastern Wyoming. Over 39.2million tons of coal were produced at the Black Thunder Mine in 1996. Just over 100 million pounds ofANFO/Emulsion blend were consumed in that same year. Currently around 30% of the overburden is cast to final with the use of explosives. With plans for a mine expansion, explosives will continue to play a very vital roll well into the future."

Jan 1, 1997

By D J Williams

Coal washery wastes comprise tailings, finer than 0.5 to 0.1 inn depending on the sophistication of the washery, and coarse reject, typically finer than 50 mm. Subaerial disposal of coal tailings slurry at about 30 per cent solids by mass results in the formation of a delta with an average slope of about 1 in 100, representing one extreme. The ravelling of coarse reject on loose-dumping, results in an angle of repose slope with an angle of about 37" (1 in 1.3), representing the other extreme. The pumped co-disposal of coal tailings and coarse reject results in the formation of an upper delta with an average slope of about to one in ten, between the two extremes. The depositional behaviour of these three waste streams is discussed and compared in this paper. The important depositional and material parameters, and a theoretical description of the processes involved, are also presented. Comparisons are made between the depositional behaviour of the various coal washery waste streams, and natural depositional systems, such as rivers, alluvial fans, and scree slopes.

Jan 1, 1998

By L. F. Pett

ALTHOUGH Kennecott's orebody has long been outlined, it is still necessary to define further its limits. This mine, long an advocate of churn drill methods, recently supplemented its practice by using a diamond drill. So far, the diamond drill holes have been along the main canyon floor which traverses the mine area. This zone, highly fractured, is cut by numerous friable quartz veinlets encountered within the porphyry. These veinlets, formed by filling contracted porphyry fractures with hot, siliceous, mineral-bearing solutions, are often more highly mineralized than adjacent porphyry. They pulverize on contact, yielding only slight amounts of core. For this reason neither analysis of core recovered (amounting to 37.5 pct of approximately 10,000 ft of hole drilled) nor sludge analysis considered separately would give correct assays. The following equation has been adopted: actual weight theoretical actual weight core — theoretical weight sludge Drilling is done un.der contract. Collection of core and sludge samples together with double-checked assaying is done by the company. Microscopic, specto-graphic, and X-ray examination of rock specimen saved is carried on at the University of Utah. A fund, set up by Kennecott Copper Corp. at the University, has encouraged cooperation in solution of mining research problems. The drill used is a Longyear hydraulic powered by a Waukesha 26-hp gasoline motor. A pump, Bean Royal type, is powered by a 7-hp Briggs-Stratton gasoline motor. Weight of unit is 4000 lb and is placed under tripod of 35-ft wood poles within a 25-ft sq area. The following results have been attained: Avg depth of hole drilled, ft 1215 Avg drilling depth per 8-hr shift, ft 10.11 Core Size, In. Avg depth of NC casing, ft 21 2 11/16 Avg depth of NX casing and size core, ft 178 2Ye Avg depth of BX casing and size core. ft 427 1% Avg depth of AX casing and size core, ft 585 1 5/32 Through the use of metal sludge collection boxes, 12 ft long, 18 in. wide, and varying from 12 to 24 in. in depth, sludge recovery has been maintained above 95 pct. Samples, both core and sludge, are assembled for each sample interval and averaged in 50-ft bench increments. Determinations for mineral values are made at both the mine and mill. Assay results varying in excess of 0.05 of copper assay values are rerun to determine correct assay. The organization chart for the Kennecott Utah copper mine is shown in Fig. 1. The Utah mine requires approximately four men per shovel shift to drill solid rock, properly slope banks, and trim loose material from the top of bench sections in making them safe for men and equipment working below. During recent years, while the mine has been worked day and night, powder crews have worked day shift only. Series of toe holes drilled back of advancing electric shovels comprise approximately 70 pct of all mine drilling. These holes, spaced about 21 ft apart, have a depth of 24 ft. Of the footage drilled, 25 pct is for vertical holes either 24 or 28 ft deep, spaced about 28 ft centers along the top rim of banks. This combination of toe holes, using an Ingersol-Rand DA 35 drifter mounted on a tripod and the vertical holes using an Ingersol-Rand Type X71 wagon drill is standard practice above the pit elevation. These benches are 68 ft high, with a few exceptions, while benches below the pit elevation are 50 ft high and require very little drilling other than toe holes. Five benches, varying in height from 80 to 100 ft, require a combination of toe holes plus vertical holes and bank holes or electric churn drill holes. Bank holes using the DA 35 drifter and electric churn drill using Bucyrus-Erie Model 29T, account for 5 pct of feet drilled. Until the past year, drilling of fill material on upper levels was accomplished by churn drill method. Wagon drills are now used for fills because they can be set up and drill a hole an hour. It is necessary to run an air drill 7 hr for each 8-hr shovel shift. Average drilling rates during the past year were as follows: DA 35 drifter using 2 and 3-man crews — 4.36 ft per man hr X 71 wagon drill using 2-man crews — 6.45 ft per man hr DA 35 drifter bank holes, 3-man crews — 2.53 ft per man hr Bucyrus-Erie churn drill, 2-man crews —1.9 ft per man hr Total mine avg — 4.57 ft per man hr Mine practice assigns powdermen to job locations where they perform all drilling and blasting work within the area. Almost as much labor is consumed in springing and loading holes as in drilling them. Average powder consumption is approximately as follows: Per Electric Shovel Shift, Lb To spring and chamber holes — 50 Blasting — 500 Trim banks — 50 One fifth of all powder used is Hercules hercomite bag powder while four fifths is Hercules gelamite 65 pct strength in 1 Vi in. x 8 in. sticks. Frozen material, overhanging from top banks, creates a winter hazard to shovel operators and equipment. Experiments are being conducted with a high-speed Hercules powder in an effort to eliminate this condition. Banks are trimmed by powdermen who lodge 8 or 10 bundles of powder, containing 15 or 20 sticks each, across the upper surfaces of each working face. To each bundle of powder is attached a primer with 6-ft fuse which is lighted as the powdermen, who have lowered themselves down the bank on ropes, ascend to safety. Three angle dozers, D8 caterpillars, make grades and move supplies for the average 31 drilling machines working daily. Compressed air is furnished by three Ingersol-Rand compressors, each with rated capacity of 3750 cu ft of air per min at 100 lb reservoir pressure.

Jan 1, 1952

By J. G. Kelly, C. P. Brown

"In June 1983 a new borehole hoisting system was commissioned at Sherritt Gordon's Fox Mine. Twelve hundred tons-per-day of ore from the 2600 and 2800 levels are being hoisted to the 2000 level through an 8-jt diameter bored winze. This new method replaces the diesel truck haulage system which had been in operation at the mine since 1975.A study was initiated in 1980to determine the most effective method of achieving the required production. The study examined several hoisting alternatives including: expanding the existing trucking system; converting to electrically -powered truck haulage; and deepening the skip compartments of the existing main shaft. The alternative selected as the hoisting method most suited to the requirements was borehole hoisting.Development of the new system commenced in April 1982. Underground excavations included 700 feet of 8-ft diameter borehole, a hoistroom and headframe, storage bins, and a grizzly and rockbreaker station. Equipment installations included: a 5-ft diameter double-drum Nordberg hoist; a counterweighted Kimberley-type skip; two rockbreakers; and the steel sets in the winze. These installations were completed in June 1983. Some problems were encountered after the initial commissioning but these were solved, and since August 1983, the system has proven that the designed capacity is achievable within expected costs."

Jan 1, 1985

By S. J. Oosthuizen

This paper examines innovation in relation to the availability of anew material: the metal titanium. The paper aims to highlight the need for the inclusion of entrepreneurial innovation as a necessary focus area in the development of a titanium metal value chain. Both the Department of Science and Technology (DST) and the Department of Mineral Resources (DMR) have identified the creation of titanium metals production capabilities as a key growth area for South Africa. Using historical literature as a source of data, the activities of selected innovators who used titanium metal as a central component in their success, were investigated. The origin of the process innovation behind the titanium metals industry, and two titanium product innovations: namely, medical implants and sporting goods, were detailed as case studies. It was found that individual innovators were responsible for the creation, and rapid growth, of the titanium industry and titanium product applications. There is a need to link the current research and development into titanium metals production with individuals and organizations capable of commercializing innovative processes and products.

Jan 1, 2011

By J. L. Hoover, P. W. Conrad, J. Downey, S. Parrow, S. Rosenthal

The goal of the drill bit consumption study was to develop an equation(s) based on various components that affect drill bit wear and use that equation(s) to estimate the consumption of bits during production drilling. Drill bit data from various geographic locations were compiled and used to develop linear regression equations using the Microsoft Excel software package. Multiple drill bit consumption equations were developed and statistical analysis was used to verify the accuracy of those equations. The equations will allow mining companies to more accurately forecast drill bit consumption when evaluating new mine properties. This capability becomes increasingly beneficial to forecasting as the cost of consumables increases.

Jan 1, 2012

By L Bertoss, D Carvalho

In the last decade, the use of advanced implicit modelling software/algorithms in mineral resource workflows has become a best practice in the mining industry. In most cases, these workflows can easily replicate the level of detail of explicitly modelled geometries, even for very complex deposits. These techniques are flexible and can produce fast results since software is responsible for generating 3D surfaces and solids based on geological and interpretation data. Nevertheless, depending on how unique and complex some deposits are, intense manual editing and input are required to honour the geological context and produce the desired outcome. The current geological modelling workflow at George Fisher Mine (GFM) is based on the semi-automatic built-in tool that generates stratigraphic sequences (stacked lithology layers) based on logging and mapping data. This tool (stratigraphic sequence from Leapfrog Geo™) works for the majority of areas in the mine but fails to capture observed mineralisation pinch-outs in operational areas and at extrapolated boundaries of the deposit. These pinch-outs have the potential to negatively impact grade forecasts if not resolved manually. Using built-in software tools only, the alternative for a more accurate model is to generate each individual layer and adding tens of thousands of manual interpretation points to honour pinch-outs (based on the amount of existing drilling in the mine and number of layers). This approach would require an estimated minimum of two to four years of a full-time employee (FTE) to be exhaustively completed. This study presents an automated approach to generate pinch-out points using a set of Python scripts that only take a few minutes to run. The newly generated points are then combined into the commercial software workflow as interpretation points, honouring observed pinch-outs. The remarkable time difference between approaches removes chance of manual errors and allows mine geologists to spend valuable time in validation and testing of scenarios, ensuring they are more representative of the deposit’s geology.

May 24, 2023

By D. Brady

The need for onsite gas monitoring during, or following, a major incident at an underground coal mine has long been recognised. Recommendations from inquiries following Queensland mine disasters have included ?qualified persons and equipment for rapid analysis of gas samples be available in times of emergency? and that ?this may be accomplished by either mobile laboratories or laboratories established in each mining locality.? Following the Moura No. 2 disaster, the Queensland Government allocated funds to Simtars for a mobile laboratory to respond should such a disaster occur again in the future. This vehicle was commissioned in 1997 and has been used at many mine sites under a wide range of scenarios. Simtars, armed with over ten years of operational experience with this mobile laboratory, has taken an active role in the design and construction of a modernised replacement unit. With a twenty point tube bundle system and ultrafast gas chromatograph the result is, arguably, the best equipped mobile gas analysis laboratory for mine emergency response in the world. This paper outlines the features of the new laboratory, its capabilities and the philosophy behind the design and inclusions.

Jan 1, 2012

By Ch. D. Taylor

Past studies with blowing face ventilation systems have shown that using a machine-mounted scrubber and water sprays can reduced methane levels at the face. The current research was conducted to determine how the sprays and scrubber interact to reduce methane levels, and what spray configurations provide the best face ventilation when used with a scrubber. Testing was conducted in a ventilation test gallery designed to simulate a full-scale mine entry. In the entry, a model mining machine, equipped with a simulated dust scrubber and water spray systems, was located at the face of a 6.1 m (20 ft) box cut. Face ventilation was provided by a blowing curtain located 12.2 m (40 ft) from the face. Methane released from a manifold at the face was measured at six locations near the face to compare the ventilation effectiveness of the spray systems while operating with and without the scrubber. Airflow movement across the face was affected by the spray configuration and scrubber operation. Average face concentrations were lowest when directed sprays were used with the scrubber. Varying the direction of the scrubber exhaust affected face methane levels.

By J. Noll

Many underground metal/nonmetal mines are using enclosed cabs on mining equipment to reduce the exposure of miners to diesel particulate matter (DPM). However, some enclosed cab systems may not be meeting their expected efficiency in capturing DPM, which could be a result of inappropriate filter use, cab leaks, or insufficient pressurization. NIOSH investigated several types of filters for reducing DPM to determine which type had the best capture efficiency. The laboratory results indicated that a MERV 8 (Minimum Efficiency Reporting Value, as defined by the American Society of Heating, Refrigerating and Air- Conditioning Engineers) filter captured about 50 percent of the DPM, a MERV 16 filter collected 96-98 percent of the DPM, and two HEPA (high efficiency particulate air) grade filters captured over 99 percent of the DPM. NIOSH also performed a field study to measure the efficiency of two cabs at a limestone mine. The cab filtration systems of a loader and haul truck were found to be over 90 percent effective in removing DPM except under certain operating conditions when the efficiency was as low as 41 percent. Evidence seems to suggest that these conditions were when a window was open in the cab.

Jan 1, 2011

By Benjamin Smith Lyman

THE importance of topography to geology is so commonly underrated as to deserve to be pointed out again and again. The relation of topography to the different branches of geology may be seen best by a glance at them all. Geology is well understood to consist of the observation of facts about rocks as materials of the earth (geognosy), and the explanation of those facts (geognosy). The facts to observe are : I. GEOMETRICAL, as (1) structure, including the dip and strike at different points, the size, shape, and position of basins and saddles; (2) vertical section, that is, the thickness of rock groups and of their parts ; (3) horizontal or geographical position. 11. PHYSICAL, such as texture, whether sedimentary or crystalline ; and, if crystalline, whether granitic, porphyritic, amygdaloidal, globular, or concretionary ; whether compact or granular; coarse or fine; pervious or impervious; hard or soft; firm or tender; with cleavage or without ; massive or thin ; the microscopic structure ; also whether magnetic or not ; soluble or insoluble ; fusible or infusible ; and the constituent minerals. 111. CHEMICAL that is, the composition of the rocks and of their minerals and their affinities. IV. PALAENTOLOGICAL, as what kind of fossils are in the rocks,

By Don A. Toth

Until recently, the fire protection systems used to protect mining equipment were based on generically designed, off the- shelf concepts. These systems were reliant on single agent approaches and were generally comprised of a series of small twenty or thirty pound chemical canisters that would discharge their contents when triggered by thermal detection means. As the equipment grew in capacity, the logic used to increase the protection was to add more chemical canisters to the system. Hydraulic excavators and large-scale mining trucks grew dramatically in size into the 1990?s and beyond; with this growth came the increase in major losses due to fire. This was evidenced throughout this period, and continues to be of great concern today. During this presentation you will learn of the significant changes that have come about as a result of this history. Today, dual-agent products, infrared detection and unit-specific, engineered systems are the primary phrases used when discussing how to best protect this equipment from fire loss. This session is a must for anybody looking to optimally protect their equipment assets and for those looking to renew and/or secure new insurance for high value equipment in the future.

Apr 1, 2005

By John S. Wakeman

The initial step in the pelletizing of iron ores is the production of green balls. The importance of this step cannot be overemphasized as the physical quality of the green balls has a very significant effect on the subsequent induration process, regardless of the type of the indurating equipment employed. The ability of the green balls to withstand the rigors of handling, distribution, compaction and drying will have an effect on the degree of success of any pelletizing operation. Many factors affect the physical quality of the green balls including the prior processing of the iron ore, the characteristics of the balling feed and the type and quality of the binder used (if any). When working on the development of a new project or studying an existing operation, it is often desirable to determine what effect variations in these factors will have on green ball quality. It is generally impractical to make these determinations on a commercial, or even pilot plant scale, especially when the number of variables to be studied is large. Because of this, many organizations have developed batch techniques for producing green balls to enable them to perform such investigations with a minimum of time and expense. The batch balling method is at best, only an approximation of commercial balling systems. Many practitioners employ the use of pre-formed seed pellets to initiate a batch run. Thus, the growth mechanism encountered in any experiment is essentially snowballing. Unless an instability is introduced into a balling test (such as an excessive initial feed rate or an unusual feed material characteristic) the growth mechanism of coalescence is not present to a significant extent nor is it desirable. In a continuous operation, both mechanisms are present and necessary to sustaining the system. The batch balling technique thus produces pellets in an idealized manner and cannot be used to directly predict commercial green ball qualities. A general relationship does exist however, and the procedure is fairly sensitive to variations in test parameters. Although many refinements to this technique have been introduced over the years, batch balling is still essentially an art and subject to "operator effect". Therefore, to assure consistant and reliable results, skilled technicians who carefully adhere to established practice must be utilized.

Jan 1, 1971

By J. Jorge, J. Mariño-Salguero, B. Álvarez-Rodriguez, J. M. Menéndez-Aguado, J. J. de Felipe

"This work focuses on the characterization of heat generation during dry fracture by direct impact of a tantalum ore. Moreover, an attempt was made to quantify the heat energy loss to the environment during the grinding process of a ball mill. The study was carried out using mineral samples from a Spanish mine that were characterized by measuring their density and specific heat. During the experiments, increases of temperature ranging from 5.9 to 7.3 °C after the fracturing process were observed. A numerical model was then applied to calculate temperature distributions through the mill wall. Three possible combinations of heat transfer coefficients were analyzed, from (1) Kapakyulu and Moys, (2) Duda and (3) the heat transfer laboratory of Universitat Politècnica de Catalunya. The models with the coefficients from Kapakyulu and Moys and the heat transfer laboratory of Universitat Politècnica de Catalunya were found to provide results that were in best agreement with the experimental data. IntroductionEvery year, several billion tons of metallic ores, minerals, cement and other solids used in the ceramic and chemical industries are subjected to size reduction in ball mills around the world. The U.S. Department of Energy (2005) reported that in the United States about 29 percent of the total energy used by the mining industry, or 0.4 percent of total U.S. energy consumption, is used in comminution. In Canada and Australia, comminution is responsible for about 2 percent and 1.5 percent of total energy use, respectively. In nonindustrialized countries where mining constitutes higher fractions of their industrial sectors, comminution costs represent a significant contribution to mine operating costs. The specific energy consumption values for the grinding of these materials typically range from 5 to 50 kWh/t. Thus, a significant amount of electrical energy is consumed in the ball mill grinding operation (Aguado et al., 2006; Gupta and Sharma, 2014). Grinding can represent up to 50 percent of the direct operating cost in a mineral processing plant, especially as comminution devices operate in highly inefficient ranges of 4 to 8 percent (Wang, Yang and Yu, 2012)."

Jan 1, 2017

The process of pile driving disturbs the soil around the pile. The nature and magnitude of disturbance depends upon 1) the pile size, and 2) the nature of soil. Therefore, the prediction of a single pile behavior based on soil properties of undisturbed soil samples may not match its performance. Under lateral loads, the pile may loose contact with the soil near the ground surface especially in clays. The soil pile interactions are influenced the most by pile behavior in this zone. This situation further affects the tally of prediction and performance. Piles are used in groups. The zones of overlapping stress from one pile to the other both around the pile and below their tip particularly in soft soils, contribute to group effects. Empirical solutions have been developed for prediction of the pile group behavior. The group action is more involved under vibrations. Several cases of single piles and pile groups under static and dynamic loads from published records have been analysed. It has been shown that single pile test is the best source of information on the soil-pile-soil interaction effects.

Jan 1, 1989