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By Charles O. Frush

This paper presents a summary of the current employment practices with respect to recent graduates in Mining Engineering, as observed at the Colorado school of Mines. Because our experience is limited to our own group of students it prob¬ably does not correspond exactly to that experienced at other mining schools, but because the Colorado School of Mines graduates approximately 15 per cent of the men in this field our experience is probably representative. We have chosen to limit the information presented here to that pertaining to the period since July 1, 1960. This corresponds to the experience with the class of 1961 and to the experience of the class of 1962 up to the end of the first semester of the senior year. There were 36 mining engineering degrees given in 1961 In a class numbering 43 senior. We expect to give 30 first degrees in 1962 in a class numbering 44 seniors.

Jan 1, 1962

By P. E. Wingrove

The Kayford, Boone and Nicholas Division of Bethlehem Mines Corporation operates a modern coal preparation plant in Van, West Virginia. Since the start-up of the Van Plant in 1976 until the summer of 1980, the refuse handling system proved to be a costly maintenance problem and a material handling bottleneck. During the 1980 two-week miner's vacation the original system, which included two screw conveyors and a drag conveyor, was replaced by an installation utilizing two Flexowall steep angle belt conveyors. This paper describes the technical and economic analysis that led to the new installation and reviews its first year operating performance. The Van facility is a heavy media plant designed to handle 127 mm x 0 raw metallurgical quality coal from the Powellton seam at a maximum feed rate of 590 TPH (metric tons per hour). Depending on mining conditions, the clean coal recovery rate averages between 50 and 65%. The plant is set up with three types of cleaning circuits. Coarse material (127 mm x 12 mm) is cleaned in heavy media baths. Intermediate size material (12 mm x .5 mm) is cleaned in heavy media cyclones. Fine coal (.5 mm x 0) is cleaned in a hydrocyclone1 froth flotation circuit.

Jan 1, 1981

By J. A. Procarione

Introduction Ventilation engineers have always felt somewhat uneasy when, in the normal course of their work, they are forced to consider the flow in any mine entry as "wholly rough." Under such conditions in pipes, friction factor is no longer a function of Reynolds' number. This assumption serves to simplify the problem and overcome an evident lack of information concerning the variability of K-factors. During the summer of 1985, a unique opportunity to investigate the reaction of friction factor to changing flow rates presented itself. A Utah coal mine was temporarily directing a substantial airflow through a single entry. By regulating the quantity, it would be possible to obtain a wide range of Reynolds' numbers in an entry composed of two distinct coal thicknesses. This paper describes the conducted study. The theory covering basic fluid flows in pipes that led to the questioning of constant K-factor is presented first. Experimental procedures are then described, including the use of a view camera in determining entry area and perimeter. Finally, a discussion of the observed variations in friction factor versus Reynolds' number is given. Changes of 34% and 45% in K-factor were seen in this study for the two entry heights. The implications of ignoring the relationship in this case and a call for an expanded research effort are also presented. Theory Since a gas is a fluid, the general principles of fluid mechanics are equally applicable to airflow in a mine entry as they are to the flow of water in a pipe. The Darcy equation is, therefore, equally applicable to pressure drop calculations in both situations. This famous formula was originally derived for use in circular cross sections and may be modified for any other shape by the use of hydraulic radius (defined as the area divided by the "wetted" perimeter). The Darcy equation relates the physical parameters of the pipe (diameter, length, and resistance to flow) and fluid velocity to the energy needed to create the flow. A study by Nikuradse (1933) brought to light the suspected dependency of the friction factor term in the Darcy equation upon Reynolds' number. By performing a large number of laboratory experiments, he established some basic concepts upon which other researchers built. Chief among these contributors was Colebrook (1939). Using commercial-grade pipe, he established an empirical relationship between Darcy's friction factor and Reynolds' number that is still in use today. If Colebrook's system of equations is plotted on log-log paper (as was done by Moody, 1944), the curves have some very distinct features. Of importance to the subject at hand is the relative response of friction factor to Reynolds' number at a given relative roughness. Simply stated, a pipe can appear to be either smooth or rough, depending on the flow conditions. As the Reynolds' number increases, the friction factor decreases at a decreasing rate until it becomes nearly a constant value. At this point, the flow is said to be "wholly rough." It is upon this condition that many traditional ventilation solution methods depend. For calculating pressure drops in a mine entry, Atkinson's equation is much more popular than the Darcy equation. Both of these formulas describe the same phenomenon and so have similar forms and functions. The K-factor in the Atkinson equation is equivalent to the Darcy friction factor and can be related by simply changing units, equating the pressure drops, and canceling like terms. A generalized form of this relationship is presented below as equation 1. K = yf/28,800g = [q]f/28,800 (1) where K is Atkinson's friction factor in kg/m3 (lb - min2/ft4), Y is the specific volume in Nt/m3 (lb per cu ft), p is the fluid density in kg/m3 (slugs per cu ft), g is the acceleration due to gravity (9.8 m/s2 or 32.2 fpsz), and f is Darcy's friction factor, which is unitless. The objective behind Eq. 1 is to show the strong correlation between K and f. Since these two quantities differ only by a constant (q]/28,000), the assumed independence of K-factor and Reynolds' number is suspect in light of the known response of f. Unfortunately, the formulation does not provide information concerning the actual relationship, as entry geometry likely plays a major role in the response curve. Only an indirect indication that such a relationship might exist is provided. Previous work on this subject was performed by Thomas Falkie (1958). The study involved a laboratory model composed of sections of straight, steel duct and a range of Reynolds' numbers from 15,610 to 137,200. Results from this work show a variation in friction factor with Reynolds' number in much the same way as it was found to vary in pipes. Over the range of velocities examined, a decrease of approximately 38% in K-factor was observed. This previous experience provides additional incentives to investigate the phenomenon under field conditions. Past model studies done at the University of Utah by Ozukurt (1945), Biron (1944), and Greenhalgh (1942) present similar evidence of this variability. More recently, Kharkar's (1974) study of coal mine friction factors contains similar indications that K-factor is not constant in all situations. Taken together, the data

Jan 1, 1988

By Sean Dessureault

The mining industry has used fleet management systems (FMSs) for decades. These systems were developed long ago in large truck-and-shovel mines, where hardware sales were the core business driver for vendors. These traditional models can be overhauled in the era of the Internet of Things (IoT), Big Data, gamification and low-cost tablet technology. Relatively low-cost tablets are ubiquitous in modern life, and they are packed with communication radios and sensors like Global Positioning System (GPS) and inertial measurement units used widely for games, fitness measures and navigation. These powerful computer devices provide the same, if not more, sensing capability than traditional fleet management hardware, all at a fraction of the cost and without complicated cables or challenging supply lines. 

By Peter F. Mataich

The phosphate mining industry is a highly competitive one, one in which survival depends on the plant's ability to produce at costs equal to, or below, the competition. The solution toward getting costs down is usually one of scaling up the plants to get the most tons through at the lowest possible cost. Control of these plants has been an art. Few of them are well instrumented because until a few years ago reliable instrumentation was non-existent. International Minerals and Chemical Corporation was a pioneer in applying computer techniques to ore reserves analysis and mine planning and the success in these areas led to the question of whether we could not extend the advantages of computer processing into the plant itself. There was a great deal of doubt on the operator's end that this could be done successfully because of the lack of instrumentation and the fact that plant control was largely an art that was learned only after many years of experience. However, in 1965 the instrumentation picture began to look up and the new generation of process computers just coming on the market were considerably lower in cost than the previous ones. A preliminary study of the installation costs and the probable benefits that could be realized from computer control convinced us that a process control computer should be seriously considered for a new phosphate mining plant which was being designed.

Jan 1, 1969

By Someswar Kesh, Gavin J. Faulkner

Introduction Powered supports in longwall mines essentially operate under two conditions: 1. Static (non-working) 2. Dynamic (working) There is a variety of static conditions which may exist in a mine. These may include a temporary stoppage of work for a few minutes or stoppage of work for days because of strikes or lack of demand for coal. Studies on pressure behavior were made during brief stoppages of work (few hours) as well as a long stoppage of work (two weeks). In dynamic conditions, the pressure variation on supports is expected to follow a certain pattern. However, in many cases that does not happen. Pressure-time characteristics of each leg of the supports were studied. This paper discusses the research findings of the studies made on powered support pressure behavior under both static and dynamic conditions. Studies on Pressure Variation in Static Conditions It has been noted (Brenner et a1.,1975) that the pressure distribution on mechanized longwall faces is random in nature and changes with time and position along the face. Figure 1 and Figure 2 show the distribution of support pressures for the front left leg of supports in two different mines. In both the cases it can be seen that the variation of the support pressures along the face is indeed absolutely random. However, the nature of randomness in both the cases is very different. The data in Figure 1 (Mine 1) is much more random than the data in Figure 2 (Mine 2). This can also be seen from the standard deviation of the data (81 26 Pa versus 4000 pa) . By comparing the results of ten other sets of data the same conclusions can be arrived at. Since it was found that in Mine 1, the pressure on support legs was extremely random, batches of ten supports each were taken and the mean pressure plotted. This mean pressure line is shown in Figure 1. It was expected that this would smooth out the variation due to individual props and reveal a better picture of the support pressure distribution. For the case cited, the standard deviation is lowered to 5043 Pa from 8126 Pa which indicates that, on an average, the supports tend to take loads more uniformly, however for individual support legs the load on the supports is very random. The significance of this is that, if the supports are selected on the basis of theoretical load calculations alone (based on the roof block concept or some other concept), only the mean value of the support load required will be known. In the case of a pressure distribution with a high standard deviation, some supports selected on the basis of mean value will fail, while others will be supporting low loads. Hence, for proper selection of support capacity, it is not only necessary to know the average load but also the expected nature of the distribution. The only reason why the supports are supporting the roof in Mine 1 well is the fact that the yield load is above the highest level of load attained. If Mine 2, which has a much higher average load, had the same sort of standard deviation, the loads on the individual legs could easily exceed the support load. For example, if the pressure distribution followed a normal curve and the mean pressure was 36,700 Pa with a standard deviation of 9000 Pa, it is possible that 67% of the supports will be within the pressure range of 36,700 plus or minus 9,000 Pa, or 22,700 and 45,700 Pa. Many of the supports could then exceed the yield load of 40,000 Pa. It is clear then that the mean load is an inadequate criterion for estimating support resistance. The nature of the problem is statistical and such an analysis should therefore be made as far as pressure data are concerned. Statistical Analysis of Pressure Data. In general, in longwall mining and other geo-mechanics studies, it is extremely difficult to specify the exact nature of the distribution of a variable. Under these circumstances, probability

Jan 1, 1986

By Ulrich Benatzky

Rheinbraun AG, Germany, is the management company of the mining and raw materials division of RWE AG. Rheinbraun's core business and major functional area is opencast lignite mining in the Rhineland. The characteristic feature of opencast mining is continuous operation using bucket-wheel excavators, belt conveyors, and spreaders. The report gives a survey of the conveyor systems used. In addition to the general performance data of these belt conveyors, this report will deal in particular with major components, such as drive stations and individual mechanical components (e.g., conveyor belts, pulleys, gears, couplings, brakes, scrapers), as well as monitoring systems. Finally, an assessment of the conveyor systems, together with future developments, will be given.

Jan 1, 2000

By M. A. Giralico

A new generation of hydraulic efficient pumper and auxiliary impeller designs is the result of an extensive research program focused on improving the overall performance of optimized SX/EW plants. Combined with the proper scale-up, these optimized designs will greatly reduce aqueous and organic entrainment losses, reduce or eliminate air entrainment and crud formation, enhance mass-transfer and maximize the overall copper yield. Currently, these new pumper and auxiliary impellers are installed full-scale in numerous plants ranging in flow rates from 90 to 15000 G.P.M. and developed heads up to 51 inches. Operational results reported by the independent metallurgist at each processing plant are reporting the lowest entrainment results in the industry. Further R&D has just been completed on a new mixing configuration that separates the systems mixing and pumping requirements. Results from the system pilot test document further improvements in reduction of both chemical entrainment and crud formation, while keeping mass transfer stage efficiencies.

Jan 1, 1998

By E. Siami-Irdemoosa, S. R. Dindarloo

"Pareto analysis was used to investigate the failures database of a Hitachi EX1900 hydraulic shovel over one year of continuous operation. The hydraulic system and hydraulic links, hoses and piping were identified as the most critical system and subsystem, respectively. A three-parameter gamma distribution function was identified as the best fit for a probabilistic modeling of the time between failures (TBFs). This distribution was used in a Monte Carlo reliability analysis of the hydraulic system. Autoregressive integrated moving average (ARIMA), within the context of a time series analysis of the TBFs, was proposed as a viable alternative to the traditional technique of distribution fitting. The mean absolute percentage error (MAPE) metric was used to compare the two models. The MAPE value that resulted from the ARIMA was 66 percent smaller than that from the traditional technique in predicting the TBFs. Resumen ¦ Se usó análisis Pareto para investigar la base de datos de fallas de una pala hidráulica Hitachi EX1900 durante un periodo de un año de operación continua. Se identificó al sistema hidráulico y a los enlaces hidráulicos, mangueras y tuberías como el sistema y subsistema más importantes, respectivamente. Se identificó una función de distribución de tres parámetros gamma como el más apto para un modelado de probabilidades del tiempo transcurrido entre las fallas (Tabes por sus siglas en inglés). Esta distribución fue usada en un análisis de confiabilidad Monte Carlo para un sistema hidráulico. El promedio de movimiento integrado auto-regresivo (ARIMA por sus siglas en inglés), dentro del contexto de análisis de serie temporal de los TBFs, fue propuesto como una alternativa viable a la tradicional técnica de ajuste de distribución. Se usó el porcentaje de la media absoluta de error (MAPE) para comparar los dos modelos. El valor MAPE que resultó de la ARIMA fue 66 por ciento menor que el resultado obtenido con la técnica tradicional de predicción de TBFs."

Jan 1, 2015

By R. L. Grayson, R. C. Althouse, L. A. Layne, M. J. Klishis

Introduction The researchers combine a detailed microanalysis of accident data and roof bolting field observations to develop new and meaningful risk indices for the roof bolter job classification. These indices incorporate information that has not been used in traditional measures of risk. The significant parameters that these new indices incorporate are the total number of accidents occurring within a particular work activity, the amount of time a miner usually spends performing the work activities in which the accidents occurred, and the average severity of the injuries incurred in the accidents. Being specific for a particular job, they aid comparisons among work tasks performed in the bolting cycle. The indices' specificity provide anew basis for reexamination of training and operational development of miners for specific jobs. Microanalysis The roof bolting microanalysis was performed by examining the empirical data and written summaries for all bolting accidents reported in the West Virginia Safety Information System (WVSIS) for the years 1983-1987. The WVSIS database contained more than 15,000 accident narratives. The microanalysis of bolting accidents covered 2083 reported personal injuries (RPIs). This large number of injuries, performed at an industry-wide level, provides a broad picture of the nature of roof bolting injuries. The analysis focuses on the activities associated with the cyclical routines of the roof bolting process. For this analysis, roof bolting tasks were categorized into work routines based on the sequence of work activities performed by roof bolters. Four work routines were identified within the overall bolting cycle. And these routines comprised 70% (1217) of all roof bolting accidents at the face. The routines include: face area preparation; tramming, positioning and setting the automated temporary roof support (ATRS); drilling holes; and installing bolts. The types of injuries most common within a routine and the average severity of those RPIs were also identified. Table 1 gives a breakdown of the percentage of accidents by the bolting routines. Another 19% of the roof bolting RPIs cannot be classified in a particular routine because of the vagueness of the narratives. However, it was clear that these accidents involved drilling holes or installing bolts. These results are consistent with previous research. It identified drilling and bolting, followed by tramming, as the most frequent sources of injuries (Helander, Krohn & Curtain, 1983). [Table 1-Accidents by bolting routines at the face Accidents Bolting routine Percent Number Face area preparation 18.2 222 Tram, position, ATRS 23.3 284 Drilling holes 34.2 416 Installing bolts 24.2 295 Total 1217] Roof bolter operators were involved in 82% (1706) of the 2083 accidents. The RPIs incurred by classified roof bolters occurred largely on working sections (95%) and at the face area on the working section (75%). Approximately 23% of the 1706 roof bolter RPIs occurred during tasks which were not part of the bolting cycle. These included belt utility work, providing supplies and section moves. Miners in other job classifications who were involved in roof bolting accidents accounted for the remaining 18% of the 2083 roof bolting RPIs. Among this group were section utility men and general laborers (41%), section machine operators (38%), and electricians and mechanics (11%). Another 9% were performing on-section supervisory functions. Face area preparation Face area preparation work led to a substantial number of injuries, although it accounted for the lowest number of injuries of the four routines. Several tasks were associated with the injuries incurred in this routine. Table 2 shows the percentage of injuries associated with face area preparation. [Table 2- Accidents occurring during face area preparation 26.6% Setting temporary and supplemental supports 24.8% Scaling roof and ribs 21.2% Installing ventilation devices 15.8% Handling supplies and materials 11.6% Miscellaneous tasks] The injuries commonly involved overexertion and 'struck by' falling coal and rock. Handling materials often led to overexertion injuries. Nearly 66% of the scaling-related injuries and about 33% of the injuries incurred during the setting of temporary supports involved 'strikes' by equipment, materials, top and rib. Tramming, positioning and setting the ATRS During tramming operations, workers were often struck by components of the machine and hit by loose roof and rib. Poor footing also led to slips, sometimes against the moving machinery. Table 3 gives a breakdown of accidents in this routine. Control levers were a significant problem in this routine.

Jan 1, 1993

By Chee Theng

"Getting minerals out of ores is messy, but mineral processors learn to have fun with it. Celebrating the innate joys of the industry was the theme of this year’s Colorado MPD Conference. At the same time, attendees dug deeper for insights into problems that continue to dog the community, including tailings dam crises, water resource challenges, and a dwindling number of academics to teach and train the workforce (see box).Tailings dam crisesMatt Fuller of Tierra Group International Ltd. brought some of the audience to tears with his impassioned talk on “The Infamous Legacy of Upstream Tailings Dams,” which won the conference’s Lou Cope award for best paper. The talk was especially poignant coming on the heels of the January tailings dam failure in Vale S.A.’s Córrego do Feijão iron ore mine in Brazil, the latest in a series of catastrophic dam failures over the past decade that includes the Fundão dam failure in 2015.Noting that the two dams share a common characteristic — both are upstream tailing dams — Fuller explained the three basic dam construction methods and why certain upstream tailing dams are susceptible to failure (Fig. 1). The downstream, centerline and upstream methods denote the direction the centerline of the dam crest moves with each successive raise in elevation of the dam. With the downstream dam, the foundation for future raises is essentially the same as that of the starter dam, he explained. With the centerline dam, half of the dam is constructed on the same foundation as the starter dam, and with the upstream dam, each successive raise is built out over the top of the tailings.“Downstream dams are expensive to build because each successive raise is bigger, so the sustaining capex to build them keeps growing,” Fuller said. Many decades ago, mine owners realized they could construct a starter dam, put their tailings through a hydrocyclone, generate a coarse sand fraction and stack the sand over the top of the tailings to build the next raise. “Then, there is no sustaining capex, it all comes out of operations,” Fuller added. The footprint is smaller, and it is a favorable means of retaining the tailings."

Jul 1, 2019

By G. Frey, D. Scott, C. Pritchard

Controlled recirculation has been used in the metal/nonmetal mining industry for energy savings when heating and cooling air, in undersea mining and for increasing airflow to mining areas. For safe and effective use of controlled district recirculation, adequate airflow to dilute contaminants must exist prior to implementation, ventilation circuit parameters must be accurately quantified, ventilation network modeling must be up to date, emergency planning scenarios must be performed and effective monitoring and control systems must be installed and used. Safety and health issues that must be considered and may be improved through the use of controlled district recirculation include blasting fumes, dust, diesel emissions, radon and contaminants from mine fires. Controlled recirculation methods are expected to become more widely used as mines reach greater working depths, requiring that these health and safety issues be well understood. The U.S. National Institute for Occupational Safety and Health (NIOSH) conducted two controlled recirculation tests over three days at a Wyoming trona mine, utilizing an inline booster fan to improve airflow to a remote and difficult-to-ventilate development section. Test results were used to determine the effect that recirculation had on air qualities and quantities measured in that section and in other adjacent areas. Pre-test conditions, including ventilation quantities and pressures, were modeled using VnetPC. During each test, ventilation quantities and pressures were measured, as well as levels of total dust. Sulfur hexafluoride (SF6) tracer gas was used to simulate a mine contaminant to monitor recirculation wave cycles. Results showed good correlation between the model results and measured values for airflows, pressure differentials, tracer gas arrival times, mine gasses and dust levels.

Jan 1, 2013

By J. A. Restrepo, K. D. Luxbacher

"The U.S. underground coal mining industry has recently seen the occurrences of several multifatality events that have interrupted an otherwise steadily improving safety record. These events transpired even with an unprecedented level of government regulations and safety technologies. These events may indicate that current safety practices have reached a level of diminishing returns. Risk management, a safety approach that has been successfully applied in numerous industries including mining, may provide a means to overcome this plateau. Riskgate is an Australian mining risk management program that may be successfully applied in the United States. However, fundamental differences between the Australian and U.S. coal mining industries prevent direct implementation. This paper discusses aspects of the Riskgate body of knowledge that require adaptation for its application in the United States. IntroductionThe underground coal mining industry in the United States has recently seen the occurrences of several high profile, multifatality events. These events transpired in the midst of an unprecedented level of government regulations and modern safety technologies. The recent increase in fatal events in conjunction with minimal declines in fatal and nonfatal injuries over the past decade may signify that current safety practices have reached a level of diminishing returns. In order to surpass this plateau, a modification to the manner in which the mining industry approaches worker safety may be warranted.The implementation of a risk-management- based approach may potentially provide a means for not only surmounting the current safety stagnation but also for achieving the long-desired objective of zero harm.Riskgate is an Australian risk-management- based body of knowledge that shows great potential for application in the United States. This paper discusses aspects of the Riskgate body of knowledge that require adaptation for the U.S. mining industry. The proposed adaptations are based on differences in operating practices and government regulations. This discussion will focus on the three major U.S. risk areas of explosion prevention, ground control and moving equipment."

Jan 1, 2015

By G. E. Erdman

This paper presents talc milling guided by a theme of "So You Want to Be In The Talc Business". Such a theme strikes a cautionary note to warn the uninitiated of potential pitfalls, and this paper will cover briefly some of the technical traps as well as one of the larger talc milling operations. It is important to our purpose to pre- sent some of the market and mineral complications that make the business of producing and selling industrial talcs much more intricate than an outsider would perceive. Metal markets by comparison are great worldwide pools into which production streams of generally identical metals can flow over a weir of market prices. Industrial talc markets are not as well defined and often must be developed by the producer working closely with a specific customer or groups of customers to prove out the attributes of the specific product from a new property. All the negatives of a new market seem to work against the introductory talc in the forms of industry or customer reluctance to change, sales prices, cost of delivery, mode of transportation, health considerations of the minerals and many more. Just locating the potential customers for a new industrial talc may be a substantial undertaking. Thus, an analogy of the worldwide metal market might compare starting a new talc operation to trying to gather up enough fallen rain drops to be able to start a production stream of a worthwhile size.

Jan 1, 1984

By D. Gatzemeier

The Jewett mine is a surface mine in central Texas, producing about 8,000,000 tons/year of lignite. The mine covers 46,000 acres of land. Approximately 430 people are employed by Northwestern Resources Company and through their efforts, the Jewett Mine is one of the most productive mining operations in the country. The Jewett Mine utilizes a unique structure of staff to convert Operations Research into plans which can be implemented at the "front line" level of the Operations Team. This planning not only makes the mine cost effective and efficient, but promotes creativity in mining, which promotes ongoing research to lower the cost of mining.

Jan 1, 1993

By S. M. Hossaini, M. Ataei, H. Moosavi

The abstract highlights the importance of pillar stability in underground coal mines and the influence of rock mass strength and joint orientation on pillar behavior. The study employs numerical modeling using the distinct element code (UDEC) to investigate the impact of these factors on pillar stability. The research considers different joint orientations and varying rock mass strengths, analyzing parameters like stress distribution, displacement, and failure patterns. The findings reveal that both joint orientation and rock mass strength significantly affect pillar stability, with specific combinations leading to reduced safety factors and increased deformation. The study concludes by emphasizing the importance of considering these factors in pillar design for safe and efficient mining operations.

Apr 4, 2022

By Anthony Iannacchione, Tim Miller, Nicole Evanek

The Subtropolis Mine is a room-and-pillar mine extracting the Vanport Limestone near Petersburg, OH. In February 2018, mine management began implementing a new mine layout to better control the negative effects of excessive levels of horizontal stress. Almost immediately, the conditions in the headings improved. Conversely, and as expected, stress-related damage was concentrated within crosscuts. Over the last 18 months, the mine operator has diligently experimented with different techniques/methods to lessen the impact of the instabilities in the outby crosscuts. The range of controls used by the mine operator include angled crosscuts, crosscut offsets, increased distance between crosscuts, arched crosscuts, cable-bolted crosscuts, altered blasting pattern, and windows. A window is used to resist roof deformation by leaving a strong brow of roof rock within the crosscuts.

By J. P. Gliddon

Since 1980 a number of significant changes have been introduced at Palabora in an effort to off-set the effects of reduced product price and increased unit cost of inputs. These changes and improvements include major equipment rationalization programs, the introduction of trolley assisted truck haulage, adoption of a sophisticated computerized truck control system, pit design enhancements, improved drilling practices, use of emulsion explosives and the installation of an in-pit crushing and conveying system. The various changes, their effect on the operation and their contribution to cost reduction and improved productivity at Palabora are considered.

Jan 1, 1987

By Verya Nasria, William Bent, William Hogan

"The South Hartford Conveyance and Storage Tunnel (SHCST) project is a significant component of the Hartford Metropolitan District’s (MDC) long-term control plan (LTCP), which is overseen by the Connecticut Department of Energy and Environmental Protection (CTDEEP). This project will address a portion of the MDC’s Clean Water Project (CWP), which will reduce combined sewer overflows (CSOs), eliminate sanitary sewer overflows (SSOs), and reduce nitrogen released into the Connecticut River. The purpose of the SHCST project is to eliminate West Hartford and Newington SSOs, eliminate Franklin area CSOs discharging to Wethersfield Cove and minimize CSO discharges to the South Branch Park River. The locations of each overflow are shown in Fig. 1. In 2010, the district prepared a preliminary design report (PDR) for the SHCST project, which included relief of the Folly Brook trunk sewer and proposed to keep the tunnel boring machine (TBM) retrieval shaft within the city limits of Hartford. Figure 1 shows the 2010 PDR recommended tunnel route. Subsequent to the PDR, the objectives of the SHCST have slightly shifted. In accordance with ongoing revisions to the LTCP, relief of the Folly Brook Trunk Sewer is no longer necessary. Additionally, the MDC has decided to perform less sewer separation in the Franklin Avenue drainage area. To replace the sewer separation, new relief points are proposed within the Franklin area and will be diverted to the SHCST. Figure 1 also shows the current recommended tunnel route (Alignment F). During 2012, the MDC conducted an evaluation of the potential of connecting the proposed North Tunnel (originally proposed as an independent tunnel with its own pump station) into the South Hartford Conveyance and Storage Tunnel. This evaluation concluded that the two proposed tunnels could reasonably be connected together and operated as a single tunnel system utilizing the tunnel pump station at the eastern terminus of the South Hartford Tunnel (Fig. 2). It also was concluded that this alternative was less costly than two independent tunnel systems."

Jan 1, 2014

This one day field trip is sponsored by the Denver Region Exploration Geologists? Society, May 14th 2009. The trip highlights major geologic features and some notable uranium occurrences along the north-south mountain-basin boundary southwest of Denver and continuing north to Fort Collins, Colorado. Uranium deposits visited include sandstone?hosted roll front type and polymetallic veins hosted in metamorphic rocks. A replacement ?episyenite? type occurrence is considered via description and hand samples but is not visited owing to logistical constraints. Recovery methods addressed include underground mining and in-situ leach recovery. Samples of Precambrian basement rocks, deposit host rocks, uranium ore samples, and ore polished sections viewed by reflected light microscopy will be available. A road map showing the locations of the stops serves as the front cover of this field guide. The Robinson episyenite-hosted occurrence (not visited) is labeled as Stop 6. A stratigraphic column is provided which shows the position of the host rocks at each stop (Figure 1-1).

Jan 1, 2009