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By Thomas Konstantis, Spyridon Konstantis, Panagiotis Spyridis

"Tunneling and Underground works are unequivocally subject to a diversity of inherent uncertainties associated with the geotechnical, hydro-geological and environmental regime that surrounds them. On many occasions, these uncertainties can provoke loss events of considerable consequences. The present contribution elaborates on losses that the insurance market has suffered in the recent years, following a construction failure event. These losses are being considered and analysed on a quantitative basis through evaluation of cost related data, with explicit discretization of the applied construction methodology and of the subsequent developed failure type since both of which are considered of major importance. Given the identified persistence and regularity of some loss-contributing factors, recommendations are provided on the basis of a proactive risk engineering management approach that are envisaged to reinforce the understanding of project risks and alleviate the incurred insurance cost. 1 INTRODUCTION The underground construction sector has always been a very challenging area for all involved parties, including the insurance market. Main reason for that remains the uniqueness of the uncertainties sources and the subsequent inherent risks that the tunneling and underground projects are exposed to. In many instances, these risks materialize, leading to loss events with substantially high impact on reinstatement cost and incurred delays, which depending among others on the insurance coverage will have to be bared by the Insurance market. In the recent decades and as the tunneling industry boomed and expanded with many major and significant projects undertaken around the world, the insurance market was faced with additional challenges. Increased competition during tender coupled with the scale and complexity of the projects and politically driven decision making exerted pressure on the project timelines, completion dates and budget. The pressure for on-time or even early project completion- to avoid liquidated damages or achieve bonus payments, respectively- along with cost reduction efforts and the incentive for innovative failure events. That, in turn, put additional pressure on the Insurance market which had experienced a significant amount of losses with substantial cost. One of the triggered mitigation actions was the introduction of the Code of Practice for Risk Management of Tunnel Works (ITIG, 2006 & 2012; MunichRe 2006, Wannick, 2007, Spencer, 2008, Adeyemo, 2011)."

Jan 1, 2016

By A. Yazadni-Chamzini

Maintenance strategy selection plays a significant role in mining design. The nature of maintenance strategy selection is a complex multi-criteria decision making (MCDM) problem including both tangible and intangible parameters which are often in conflicting with each other. Moreover, when decision makers are uncertain in determining and defining the ratings and the weights of alternatives and criteria respectively, fuzzy theory provides an appropriate tool to handle the existing uncertainties. In this paper, a new fuzzy MCDM method based on the concepts of TOPSIS and AHP was proposed to evaluate the feasible maintenance strategy. The linguistic terms are employed to assess the ratings and weights. Fuzzy AHP is utilized to calculate the weights of the evaluation criteria; then, the rankings of alternatives are computed based on fuzzy sets theory and TOPSIS. A real world case study is presented to illustrate a potential application of the proposed model.

Jan 1, 2012

By G. S. S. Raman, M. S. Klima

Coal refuse slurries stored in tailings impoundments can potentially lead to environmental concerns such as effluent seepage, property devaluation, and in extreme cases, dam failures. Pressure filtration offers an opportunity to reduce the impoundment requirements by providing significantly drier solids. A full-factorial design was used to study the effects of temperature and pH on dewatering a refuse slurry obtained from a coal preparation facility in Pennsylvania, USA. The experimental results indicated a strong effect of temperature on cake moisture as the cake moisture after about 2800 s was 22% at 43 °C and only 36% at 10 °C for tests conducted at pH 3. ANOVA results confirmed the statistical significance of pH, temperature, and their interaction effect. Also, the results showed no significant differences between the tests conducted near neutral and alkaline pH values at low and ambient temperatures. On the other hand, Tukey's test indicated higher reductions in cake moisture at higher temperatures for lower pH values. This study affirms the need to monitor and control the pressure filtration process as throughput could significantly vary with changes in temperature and slurry conditions.

Jun 17, 2019

By H. G. Naidu, Raj K. Singhai, S. P. Singh

Introduction The economics of open-pit mining de¬pend heavily on the cost of overburden removal. It is necessary to maintain a pit slope as steep as possible. Steeper slopes lower the volume of overburden to be handled, thus lower the cost of material handling. There is, however, a limit to which the pit slopes in any given situation can be steepened. Steepening beyond this point could jeopardize the pitwall stability, ren¬dering it unsafe. It might even result in curtailment of pit production. Several factors affect the stability of pitwalls. Some are natural, such as topogra¬phy and regional and local geology. Others are man made, such as pit geometry, ground water management, equipment, drilling and blasting, and the type of excavation. Greater control can be exercised in the latter with proper planning and by following a set of procedures based on field experiences. This article deals with one such control¬lable factor - drilling and blasting. The eco¬nomics of steeper pit slopes is discussed, followed by a record of drilling and blasting practices in western Canadian surface mines. Microprocessor-based instrumentation and its application to blasthole drilling is also described. Economics of steeper pit slopes The economics of steeper pit slopes is not a new issue. It should be reinforced, though, that pit slopes must be as steep as possible to ensure maximum recovery of minerals at a minimum stripping ratio. For operating, economic, and safety reasons, different pit slopes are maintained during different phases of pit development. Figure 1 shows a hypothetical ore body, cylindrical shape, with a radius of 225 m (740 ft) and located 30 m (100 ft) beneath the surface. The mining operations are planned at two different slope angles of 25° and 40°. The ultimate pit depth was established on the basis of the criterion of "break even pit." The variation of the stripping ratio at differ¬ent mine levels is shown in Fig. 2. Table I summarizes the eco¬nomics involved for two pit slopes, 25° and 40°. Figure 3 presents the profits during the life of the mine for these two pit slopes. This study shows that steeper slopes for an open pit operation are far more advanta¬geous than flatter slopes. But, what is the optimum slope at any one point in the life of a moving pit? And how can the stable slopes be maintained over the short- and long-term? Canadian blasting practices A survey of blasting practices in western Canadian surface coal mines revealed that, at most mines, hard consolidated overburden is drilled and blasted. Blastholes are drilled with rotary crawler-¬mounted rigs using bits in the 220 to 310 mm-diam (8.5 to 12 in.-diam) range. Typi¬cal hole depths are 12 to 18 m (40 to 60 ft) to accommodate the pit bench heights of 12 to 15 m (40 to 50 ft). For a 12-m (40-ft) high bench, a typical hole will be 14.6 m (48 ft) deep including the subgrade drilling of 2.6 m (8.5 ft). Half of this hole will typically be a powder column and half will be stemmed. For a 15-m (50-ft) bench, a typical hole will be up to 18 m (66 ft) deep including the subgrade drilling. Drilling patterns are usually 9 m (29.5 ft) spacing and 9 m (29.5 ft) burden. De¬pending on local geology and hole diameter selected, drilling patterns of 8 x 8 m, 8.5 x 7.3 m, and staggered patterns of 11 x 5.5 m

Jan 1, 1989

By Sidney M. Cohen

[Portland cement is a product made from a combination of limestone and argillaceous material which would supply the necessary amounts f calcium, silica, alumina and iron to form cementious compounds when fired at high temperatures. The resulting mixture, usually around 80% limestone and 20% argillaceous, is ground nine (80% -200 mesh), blended and then fired1 (2600-2700°F) to produce a cement clinker from which the final cement product is produced by grinding. The argillaceous component can be shale, clay, slate or other sources of the individual components of silica, alumina and iron. This paper deals with a case where a clay component was under consideration. The planned Plant site was to be located at a well defined limestone deposit. When it came to the argillaceous component, the nearest uniform deposit of clay was over 50 miles away and would require purchase of land or clay from the owner plus the daily delivery costs to the plant site. This situation made the plant site economically questionable. It was decided that a local clay deposit, which was available at the plant site but was orginally not considered due to lack of chemical uniformity and its high free silica content, would be reconsidered.]

Jan 1, 1983

By J. S. McEwan

This paper describes the derivation of a set of simultaneous algebraic equations that can be solved to produce a rigorous theoretical mass balance around solution mining wells. This mass balance can then be compared with a mass balance calculated from field measurements to provide the percentage of mass balance closure. Significant deviations from the closure can be an indication that the wells are either gaining or losing circulation fluids. The equations also provide calculations for the amount of water and dissolved mineral that accumulates in the well during the mining process, as well as the net mining rate. Equations for both single mineral mining and multiple mineral mining are discussed. They can easily be formatted into spread sheets for rapid calculation of field measurement data.

Jan 1, 2010

By Franklin B. Woodbury

As part of its goal of helping maintain an adequate supply of minerals to meet national economic and strategic needs, the U.S. Department of the Interior's Bureau of Mines is evaluating alternative technologies to treat oxidized taconites. Since process control is an essential element in the application of these process technologies, research was performed with a prototype monitoring system. A Californium-252 neutron source and a thermal neutron capture- gamma ray spectra analysis method was utilized to measure the amount of iron and the percent solids in the process slurries. The concentrate and tailing streams in a 410-kilogram-per-hour (900- pound-per-hour) pilot plant employing a selective flocculation and cationic flotation process were monitored. The results obtained showed that the thermal neutron capture-gamma ray reaction can be used to provide process control parameters.

Jan 1, 1979

By M. Ralchenko, C. Samson, M. Svilans, C. Walsh

"Through-the-Earth (TTE) radio has been proposed as a means of emergency communications in mines. The propagation of a TTE radio signal depends chiefly on the subsurface conductivity distribution. Robust forward models of signal propagation can be computed based on a realistic conductivity estimate. These models can be used, for example, to optimize frequency, estimate signal range, identify zones of anomalously strong or weak signals, or to optimize the position of the above-ground TTE radio unit. A method was developed to determine the conductivity distribution of a mine overburden using TTE radio transmission data. At multiple locations in an underground test mine, we recorded a wideband electromagnetic signal, spanning a frequency range of approximately 500 to 8,500 Hz, transmitted from a vertical magnetic dipole at the surface. Using a finite-difference time-domain forward modeling code, a layered Earth conductivity model that accounts for topography was fitted to the observed data, which minimizes the difference between the calculated and observed attenuations for each transmitter-receiver pair, for each frequency of interest. Introduction and objectiveThrough-the-Earth (TTE) radio has been proposed for emergency underground communications. The most common type of TTE radio is based on the transmission and reception of signals by electromagnetic induction. These signals are in the extremely low to very low frequency range — in practice, between 400 and 9,000 Hz. This technology was historically developed for mine safety, although recently other applications have been suggested, such as communications in civilian infrastructure like high-rise buildings or tunnels, in the military, and even for wireless detonation and activation. In a coal mine, which is a setting of particular importance for TTE radio due to the higher risk of accidents, day-to-day communications can be done through a wired link. High-frequency line-ofsight systems, whose signals cannot penetrate rock, are also used for communications from nonfixed points. Another option is a leaky feeder system, which is a permutation of a line-ofsight system and a wired link, where a partially shielded coaxial cable allows the signal to “leak” (Delogne, 1982). In an incident such as a mine collapse or explosion, a TTE radio system can still provide a communication link, as wired systems can be severed, and line-ofsight systems cannot penetrate through collapsed portions of the mine, or to the surface (Forooshani et al., 2013; Yarkan et al., 2009)."

Jan 7, 2018

"In the conceptual stages of engineering, benchmarking and empirical ratios are commonly the only references for estimating the production rate of greenfield projects. For underground mining projects, the lack of similar cases often makes the comparison difficult due to each project’s unique geology, rock quality or other singular characteristic. This paper discusses an analytical alternative for determining the maximum achievable production rate for underground stoping mining by modeling production as a continuous process. The results show that long-term maximum capacity depends exclusively on the achievable development rate and the development intensity of the design. In operations constrained by backfill, the development rate could be limited by the backfilling rate. Extraction rate and stope size only have impacts on capacity estimates during the ramp-up period.These findings allow shaping of the maximum capacity profile of a project. They can be used as a tool for quick assessment of conceptual capacities, and sensitivities of mining rates. Benchmarking tests compared favorably with projects, mines in operation and scheduling exercises. IntroductionA key consideration in the value proposition of a mining project is its capacity or production rate (Tatman, 2001). At a conceptual level, it is common practice to refer to empirical references (Taylor, 1986; Long, 2009), rules of thumb (de la Vergne, 2014), benchmark ing studies or consultant experience to estimate or validate a target value. There is, however, a lack of analytical methodology to guide an early estimate of the maximum capacity.Araneda and Gaete (2004) presented an interesting mathematical approach to estimate the capacity of panel caving mines. Using a continuous model, they confirmed that the mine capacity is essentially determined by the rate of area incorporation, in terms of area per period, and ore column height. This simple relationship has been confirmed by multiple detailed scheduling works"

May 1, 2016

By S. Natansohn

A prevalent technique for the recovery of silver from wastes generated in silver-tungsten electrical contact fabrication is electrowinning. Such scrap is placed in a permeable plastic basket which constitutes the anode compartment and this basket is suspended in a larger cell in which the cathode is mounted. A silver nitrate solution is used as the electrolyte and, upon application of an electric field, the silver migrates to the cathode where it deposits out. At times, there is tungsten migration to the cathode compartment and consequent co-deposition with the silver. This W contamination, which may amount to several thousand ppm, necessitates additional purification of the deposited Ag and its removal would therefore be advantageous. The elimination of the W impurity would also extend the operating period of the cell with corresponding increase in the Ag output and savings in labor costs.

Jan 1, 1982

By W. J. Schlitt

Introduction Heap and dump leaching are low-cost metal production techniques that are gaining in popularity among gold and copper producers. However, the flow of solution in heaps and dumps has received little attention in the literature. This is unfortunate since solution flow is one of the few parameters subject to operator control. Thus, solution management may well influence both operating costs and plant performance. Costs aside, there are two important aspects of solution flow to be considered -the metallurgical and the hydrological. Some of the metallurgical factors have recently been discussed (Jackson and Ream, 1980; Schlitt, 1984; Schlitt and Nicolai, 1987). These cover solution application rates and methods, irrigation rates, leach-rest cycles, and nonvertical solution flow. According to Caldwell and Moss (1985), however, the hydrological aspects may be more significant than the metallurgical ones. In particular, a phreatic surface will form when excessive flow leads to an accumulation of water and the associated buildup of pore pressure within the rock pile. Such internal flooding can either originate at the foundation or at some other zone of very low permeability. The latter condition gives rise to an impounded volume of solution, i.e., a perched water table (see Whiting, 1985). Caldwell and Moss point out that a rising zone of saturation is probably the most common cause of dump failures. Of course, such failures have even occurred in heaps that were carefully prepared for leaching (Milligan and Engelhardt, 1984). Thus, flooded conditions and perched water tables represent an important safety consideration as well as having an impact on metallurgical performance. The following sections describe a case history in which a perched water table developed within a copper leach dump. The description includes background information, solution flow rates, and metallurgical data. Then this situation is compared to one involving normal drainage. Description of the leach system The leach dump in the case history is located at a shutdown open pit copper mine. It was built in two lifts, with the second added some 20 years after the first. There is little detailed information available on the initial lift. It was built with rail¬hauled waste and was generally less than 30 m (100 ft) high at the crest. The area available for leaching was about 120 m (400 ft) wide and more than 380 m (1250 ft) long. The dump surface was prepared for leaching by dozing ponds approximately 12 m by 12 m by 2.5 to 3.0 m deep (40 ft by 40 ft by 8 to 10 ft). The ponds were leached by flooding with barren leach solution returned from a scrap iron cementation plant. Based on mill feed at the time, the average waste grade was probably close to 0.4% Cu. The first lift leached well. It accepted high flows, which together with the waste grade, produced a rich pregnant leach solution (PLS). Old records indicate a PLS of "50 lb Cu/ 1000 gal," or about 6 g/L. As later drilling would indicate, such a high tenor led to dissolution of considerable scrap iron that was returned to the dump. The iron then hydrolyzed and settled out on the pond bottoms. In addition, the waste settled substantially so that the unleached crest was 3.0 to 4.5 m (10 to 15 ft) above the ponded area. Eventually, the leachable copper was extracted and the dump became less permeable. Thus, the PLS tenor dropped until it became uneconomical. About ten years later, mining resumed and the decision was made to add another lift to the dump. This was done without giving much thought to a subsequent leach operation. Hence a 24-m (80-ft) lift was built on top of the original dump. The surface of the latter was not prepared in any way, e.g., by leveling and/or deep ripping, prior to over-dumping. Examination of subsequent drill cuttings indicated that the new lift contained about 0.2% Cu, with chalcopyrite and chalcocite (50:50) being the predominant copper minerals. Most of the chalcocite occurred as rimming on the abundant pyrite, with the pyrite to chalcopyrite ratio estimated at 10 to 1. The use of 100- and 120-ton trucks for haulage caused some waste compaction during emplacement. In addition, the host rock itself was relatively soft, being a porphyry intrusive material that was partially altered to clay. As a result of the initial compaction and clay swelling, the rate of water percolation from the new leach ponds was slow and the ponds often contained considerable standing water. Even frequent ripping failed to provide a sustained improvement in the percolation rate. The poor surface permeability was exacerbated by hydrolysis of iron salts which settled as a layer on the pond bottoms. Partly as a result of the permeability problems, metallurgical performance was not up to expectations. These had been based on laboratory tests which showed about 20% copper solubilization in two weeks. Continued copper extraction in the tests also suggested that a substantial percentage of the copper would eventually be recovered. However, the PLS grade in the actual operation peaked briefly at about 0.48 g/L Cu (4 lb Cu/1000 gal), then declined to a range of only 0.24 to 0.36 g/L Cu (2 to 3 lb Cu/1000 gal). The poor leaching was traced to a lack of oxidation of the sulfides. There were two principal observations supporting this conclusion. First, there was no evidence of any heat being generated within the dump. As discussed elsewhere (Schlitt and Jackson, 1981; Hiskey and Schlitt, 1982), pyrite oxidation is quite exothermic and the high pyrite content of the waste should have led to an increase in the temperature of the leach solution as it percolated through the rock pile. Second, there was no sign of any natural convective air flow through the dump.

Jan 1, 1987

By J-P Strydom, J P. Z, S. Schafrik, M. Beck, Z. Agioutantis

Developing a method to accurately determine the weight and distribution of the load within a shuttle car bed eliminates the need for manual activation of the flight chain. Accordingly, this experimental study characterized the thermo-mechanical response of two unique load cells in force shunt, implemented as an engineering control to quantify load position and distribution without human intervention. The influence of environmental and installation conditions on the measurement accuracy was quantified and determined to be indirectly dependent on the characteristics of the material load cells that are mounted to.

Feb 22, 2026

By George H. K. Schenck

INTRODUCTION Boom-type miners mechanize the cutting and loading of a solid face of rock having a compressive strength as high as 124 MPa (18,000 psi). These machines differ from continuous miners in that they can cut much harder rock for any given cutter-motor power. In addi¬tion, a boom miner is more flexible, selectively cutting any cross-sectional configuration from circular to square arched faces, 1.57-rad (90°) turns, and mixed faces. The basic operational concept of these machines, utilizing a rotary cutterhead mounted on a slewing boom, gained importance in Europe during the 1960s, where, due to their specialized use in roadway development in coal mines, they were called roadheaders. Along with continuous miners, boom miners share the ability to fully mechanize the cutting and loading of rock, while eliminating the need for explosives and the shattering and overbreaking of roof and rib rock that is normal with conventional mining methods. The machines are crawler-mounted and most have a gathering-arm load¬ing system that, moves the cut rock onto an internal chain conveyor, subsequently discharging the rock into a transport device such as a shuttle car or a bridge conveyor. Light- to medium-weight boom miners are cutting softer rocks and minerals successfully, including coal, shale, rock salt, potash, phosphate, weak sandstones (e.g., those associated with uranium ore), and other softer ores such as iron. Heavy third-generation ma¬chines have extended rock cutting to much harder ores such as borate, oil shale, and almost all rock associated with coal seams, except for the harder sandstones and limestones. The mention herein of specific manufacturers, brand names, or models is made by way of example only and does not constitute endorsement by the author or The Pennsylvania State University. All costs cited herein are in 1977 US dollars. TYPES OF BOOM MINERS All boom miners utilize drag bits fixed on a rotary cutterhead that is powered by an electric motor mounted in the slewing and elevating cutter boom. The cutter¬head is either a ripping or a milling type, and the machines have evolved through three generations (Boldt, 1977). As shown in Fig. 1, only one boom miner cur¬rently produced can use both ripping and milling cutter¬heads on one machine. Table 1 lists boom-miner models built in large num¬bers and/or successfully used or marketed in North America. Table I excludes machines of which only a few have been produced and none sold in North America, such as those manufactured by Demag, Eick¬hoff, Paurat, and Thyssen-Rheinstahl. The only Ameri¬can boom miners are the F-16 and Super ROC¬MINER"', manufactured by AEC, Inc., State College, PA. Machine Generations Three generations of boom miners can be distin¬guished according to their times of introduction, their rock-cutting abilities, and their weight classifications. Fig. 2 illustrates typical best-practice production rates possible from boom miners for selected cutter-motor power sizes (other power figures can be interpolated). Productivity is related directly to the energy applied to the cutterhead and indirectly to machine weight; higher

Jan 1, 1982

By Erik Melchiorre, M. Katherine Crombie, Dante Lauretta, Chris Gholson

Rich Hill, also known as Weaver District No. 2, is among the most productive placer gold locations in Arizona. Located near Congress, AZ (Fig. 1), this district hosted major placer rushes in the 1860s and1930s. Those rushes were followed by an electronic metal detecting rush that started in the 1980s and continues today. Thousands of ounces of gold have been produced by these modern amateur and professional metal detectorists. Many of them see more gold in an afternoon than Nevada mine workers will ever see outside of the mill. Rich Hill discovery Inspired northeastward by Spanish legends, a party of 10 men left Yuma, AZ on April 1, 1863. Included in the group were Pauline Weaver (Fig. 2), Henry Wickenburg and Abraham Peeples. Wickenburg left the group near the Hassayampa River and discovered the Vulture deposit. The rest of the group arrived at the base of Rich Hill in late May. While drying antelope meat on the first day, a few of the men prospected a nearby creek bed and immediately found gold. In a couple of hours they picked up more than 2.8 kg (90 oz) of gold. The next day, four members of the party went searching for their horses and returned with news that a nearby hilltop was literally graveled with nuggets.

Jan 1, 2003

By Wayne Hilgedick

Peabody Coal Company is the nation's largest producer of coal and is headquartered in St. Louis, Missouri. In 1978 Peabody divisionalized its operations to establish more responsive and effective mine management. Peabody Coal Company's Arizona Division operates two large surface mines, the Black Mesa and Kayenta Mines. Both are located in northeastern Arizona, approximately 30 miles from Kayenta, Arizona, on the Navajo and Hopi Indian Reservations. The Black Mesa Mine produces 4 to 5 million tons of coal per year and the Kayenta Mine produces 7 to 8 million tons per year. Coal from the Black Mesa Mine is transported 274 miles by coal slurry pipeline to the Mohave Generating Plant located on the Nevada/Arizona border at Bullhead City, Arizona. This plant is jointly owned by Southern California Edison, the Los Angeles Department of Water and Power, Nevada Power Company and Salt River Project Agriculture Improvement and Power District.

Jan 1, 1983

By James L. Weeks

Introduction This paper attempts to summarize some of the major criticisms of drug testing. Eight problems with drug testing are discussed. These are: • It does not measure something employers have a reasonable need to know, • It is a distortion of priorities, • Drug use is not known to be associated with increased risk of occupational injuries, • Drug testing is proposed as a dangerously simplistic solution to poorly conceived problems, • In the absence of safeguards, it is unacceptably inaccurate, • In the absence of treatment commitments and confidentiality protection for persons dependent on drugs, it is an unethical practice of medicine, • Drug testing is an invasion of privacy, and • It will have a negative effect on treatment of persons with drug or alcohol dependencies. Employers' need to know It does not measure something employers have a reasonable need to know. It is reasonable for employers and fellow workers to want to know if workers are functionally impaired (intoxicated) while on the job. Drug testing does not measure impairment. It measures the concentration of drug metabolities in urine or blood. Unlike measurements of blood alcohol concentration (BAC), drug metabolites are not correlated with impairment or intoxication. They indicate only drug use in the one to two weeks before a urine sample. And, in some instances, they indicate chronic use. Impairment caused by drug use, on the other hand, occurs immediately and declines rapidly within hours. Drug metabolites may linger for days afterwards. Drug metabolite concentrations in urine are a function of (Morgan, 1984; Schartz and Hawks, 1985; MMWR, 1983): • the amount and purity of drug consumed before taking the test, • when the drug was consumed, • whether the person was a chronic user, in which case a certain amount of drug metabolites would be stored in body fat and slowly released, • the person's metabolic rate, and • for metabolites in urine, any dilution effects of urine. Only the first two factors has

Jan 11, 1987

By I. B. Tulu, M. M. Sears

In 2015, 40 longwall mines provided nearly 60 percent of U.S. underground coal production. During this time, about 80 percent of ground-fall-related fatalities occurred in areas where the mine roof was supported. In an attempt to better understand the status quo of current longwall support practices in the United States, during a period between 2010 and 2017, researchers from the U.S. National Institute for Occupational Safety and Health (NIOSH) visited a sample of 25 longwall mines representing nearly 50 percent of the currently active longwall mines operating in all of the major U.S. longwall producing regions. The data presented here indicate that the types and amount of roof support being installed in modern U.S. longwall mines have reduced ground-fall accident and injury rates to levels never seen before. However, the complete elimination of unintended roof falls in U.S. longwall mines has remained elusive. The data in this paper can be used by mine operators and engineers to determine where their chosen type and density of roof support occur on the spectrum of observations contained in this database. Additionally, this data can be used to enhance, supplement and support additional research on longwall gateroad support systems. 

By John D. Price

In the constant search for ways and means of improving the quality of coke and thus making its use at the blast-fornace more satisfactory and economical, it has become increasingly evident that the present laboratory tests do not show the full picture of a coke's value as a blast-furnace fuel. This fact was recognized by Dr. A. C. Pieldner thirty-five years ago: (Ref. 1) here is perhaps no phase of fuel technicology which has aroused more discussion or regarding which there is more difference of opinion than the properties of coke which make it most suitable for blast-furnaces and other uses." Dr. Fieldner followed this assertion with a statement to the effect that men who write on this subject today (1923) have, because of conflicting experimental evidence brought forth in recent years, far less positive opinions on the matter than did writers of some years past. One is led to wonder if these statements are not just as true now as they were when first made.

Jan 1, 1959

By Virginia T. McLemore, Devlon R. Shaver

Clinkers are the result of a seam of coal catching fire and burning the surrounding rocks. The fires caused the surrounding rocks to be baked at high temperatures, removing organic material and partially forming glass. These rocks become hard, orange, and brick like, forming clinker. The purpose of the study is to determine their potential to contain economic levels of critical minerals, including Rare Earth Elements. Some clinkers contain over 200 ppm total REE.

Feb 1, 2024

By Michael J. Beus, Jan G. Patricio

INTRODUCTION The U.S. Bureau of Mines, Spokane Mining Research Center, has been conducting and documenting rock mechanics research in the Coeur d’Alene mining district for the past ten years (1, 2, 3, 4, 14, 15). A current project combines the information gathered from rock mechanics instrumentation, geologic structure, and analytic techniques to improve shaft and support design for the deep mines of the district (3). This project required knowledge of the anticipated loads and the physical properties of the rock mass to provide accurate input data for computer design models. Several techniques have been utilized to determine the physical parameters including the USGS (U.S. Geological Survey) solid inclusion probe, the CSIR (Council for Scientific and Industrial Research) biaxial and triaxial strain cells, and the CSM (Colorado School of Mines) dilatometer. This paper describes the determination of the in situ modulus of deformation in three mines over a period of three years and the evaluation of these results as they applied to this project. The need for in-situ rock mass properties such as the modulus of deformation has stimulated research and development of a large number of instrumentation techniques. One area of continuing interest involves various methods of radial borehole dilation. This concept was pursued by Menard (9) with an apparatus known as the pressure meter, which is used frequently in soil mechanics. Other developments in rock mechanics include the CPC (cylindrical pressure cell) by Panek (11) and the dilatometer by Rocha et.al. (13) in Portugal. The CSM dilatometer, which also utilizes this technique, was chosen for this project because it required a small amount of background study, was inexpensive, and could easily be combined with other experiments being planned for the test locations. Although no previous field applications to hard rock (compressive strengths greater than 20,000 psi) has been documented, the instrument was selected since good results have been achieved in granite, oil shale, and several coal measure rocks (11, 12).

Jan 1, 1977