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By Luis Fernandez-Deza, Kaveh Talebi, Seth Pollak, Soto Vardakos

INTRO PAGE

Jun 13, 2021

By Jerry C. Tien

An overcast is a form of air bridge which permits one air current to pass another without mixing. It is an essential and indispensable ventilation device for underground air distribution in horizontal mineral deposits. Overcasts are also a major source of shock pressure loss under- ground. However, a well thought out overcast plan, and a well constructed overcast, will give minimum shock loss, which in turn can save money (Tien, 1986). So far, most research on shock losses has been concentrated on splits and junctions, and bends and turns, or on underground regulators. Practically nothing could be found in the literature regarding losses caused by overcasts (see reference on shock losses). In this investigation, shock losses are quantified for an overcast with different configurations, which will give useful clues toward an aerodynamically designed overcast.

Jan 1, 1987

By Rustu S. Kalyoncu

Rainbow colors of roofing granules, adorning the roofs of buildings from single family units to largest office buildings, have their origin in the roofing shingles as burlap soaked with tar appearing as early as the 1700s. Dusting of talc and mica on the coated surfaces, to keep roofing from sticking to itself in storage, represents an early improvement to the original construction product. Sometimes, colored, natural mineral dusts were added as color pigments within the dusting material, for initial color. Prolonged exposure to sun caused the asphalt to become brittle, and it eventually cracked and leaked. Search for solutions to cracking and subsequent leaking problems eventually led to the use of crushed mineral granules on asphalt coated shingles. This provided the impetus for further growth in the asphalt roofing industry. Application of a layer of crushed mineral on the asphalt coating minimized the coating deterioration from exposure to ultraviolet light, thus significantly improving the shingle life. In the 1900s, screened waste from slate quarries was the most commonly used granule in asphalt roofing products. Records indicate the use of coarse slate screenings for asphalt roofing as early as 1906. As the asphalt roofing industry further developed, other naturally colored minerals, such as silica sand, quartz, mica, feldspar, talc, slag, granite, rhyolite, and oyster shells were used on roofing surfaces. Later on, a number of manufactured granular materials, such as crushed brick and tile, fire clay, and white porcelain were used. The first artificially colored roofing granule was produced in 1914 by running molten slag into a solution of alkaline metal silicate, which was absorbed by the porous slag. Various colors were obtained by adding pigments to the silicate coating and by subsequently heat treating the granules. In 1922, another technique was developed in which green slate granules were saturated with copper sulfate solution and heated to 760°C to form copper oxide. Since then, many other granule coloring processes have been patented, one of the most significant being the sodium silicate coatings introduced around 1975. Granules as used in construction (specifically roofing materials) are defined as natural or synthetic-mineral grains smaller than gravel and larger than sand in particle size. They are used primarily for protective and decorative purposes on the weather- exposed surface of asphalt composition roofing. Development of artificial coloring techniques has made it possible for the roofing industry to provide asphalt roofing shingles in a wide variety of colors complementing the color and style of buildings. Roofing granules, as we see today, are the products of long technological evolution.

Jan 1, 1994

By Alva L. Kuestemeyer, Randall J. Scott

INTRODUCTION Pincock, Allen & Bolt, Inc. (PAH) has developed a series of cost models for Western U.S. Gold Mines. The models indicate the economic potential of developing gold properties at various production rates and ore grades. Various combinations of mining and processing methods are also analyzed. The conclusions shown in this paper are taken from PAH's 600 page study titled, "Capital and Operating Cost Models for Western U.S. Gold Mines." PAH decided to develop the cost manual as a result of our perception of the industry's need for such a document. It was our impression that exploration development decisions were being made on the basis of limited economic information. PAH has compiled a substantial database on unit cost information. That database and the engineering capabilities of the firm allowed us to develop a reasonably detailed analysis of capital and operating costs and overall development economics. The study has developed a series of algorithms for underground and open pit mining, processing, G&A, and infrastructure costs. The capital and operating cost data served as a basis for performing after tax discounted cash flow analysis for the various development alternatives. For the base case scenarios, a 15% DCFROR was assumed and the required ore grade was back calculated using the computer. The ore grade versus production rate graphs summarize the economics of developing gold projects in the Western U.S.

Jan 1, 1987

By Palmer C. Sweet

Mining-related wastes, which are increasing, are being utilized more and more as valuable resources in Virginia. The Environmental Protection Agency (EPA) divides mining-related wastes into two categories: those generated from mineral extraction and beneficiation and those generated from mineral processing. Materials from both categories are produced in Virginia, where waste (by-product) materials are present at a variety of mining sites and processing plants; some of which have been exploited for various uses in the past and/or present. These materials include: used concrete recycled for roadbase material; pyrite coated with magnetite for use in portland cement; minor amounts of fly ash for use in concrete block, lightweight aggregate, and flowable structural fills; granite and diabase quarry fines for use as agricultural and asphalt filler; lime kiln dust for use in neutralizing and stabilizing coal refuse; soapstone aggregate for use in terrazzo heat retention panels; kaolin for use as a filler product and as an ingredient in white cement; slate quarry fines used as landcover and combined with aggregate for crusher run and as a filler in asphalt; and coal refuse for use as fuel in a lightweight aggregate plant. There are some materials available that have never been utilized, although several ideas were investigated to turn them into usable commodities. One such material is a high-alumina scrap produced as a by-product at a ferrovanadium processing plant. However, large quantities of waste materials from metal mining, industrial mineral, and fossil fuel operations remain to be disposed of in an environmentally safe and economic manner. In 1989, almost one million tons of fly ash were produced from burning of coal at eight coal-fired power plants in Virginia. Present uses for this waste such as an ingredient in concrete block, concrete aggregate, and in structural fills cannot dispose of all the available material; however, increased use of fly ash as a soil/rock substitute for structural fills and embankments may solve the problem of its disposal. Another waste material is calcium carbonate precipitate paste produced at a lithium carbonate processing plant and presently being used for back filling an on-site inactive underground limestone mine. Flue gas desulfurization (FGD) wastes in the form of CaS04 will probably increase threefold. If this material contains a minimum of 95 percent gypsum and negligible chloride, it may have a value to a wallboard manufacturer as we move toward the year 2000. Thought must be given how this increasing waste can be disposed of in an environmentally safe manner.

Jan 1, 1995

By Thomas J. Mason, Geno Saccomanno, Victor E. Archer, Philip C. Prorok

INTRODUCTION A population of several thousand uranium miners in the Colorado Plateau region of the United States has been under surveillance since 1957 using sputum cytology and other procedures. Attempts were made to collect sputum samples annually for those miners showing normal sputum or only minor abnormalities, and on a semi-annual schedule for miners having more severe cytological abnormalities. For individuals having carcinoma in situ or malignant cells in their sputum, samples were taken more frequently. As in any human population, however, these schedules were not always adhered to, and samples were often acquired less often than scheduled. On the basis of the thousands of sputum readings that have resulted from this surveillance, one can calculate the durations or sojourn times which individuals spend in the various cytologic states into which the sputum readings are classified. This paper presents some preliminary, descriptive results concerning relationships among the cytologic state duration distributions in this population, and their associations with other variables such as radiation exposure and cigarette smoking. There are several reasons for studying cytology results in this observational setting. Of considerable interest is the opportunity to gain insight into the preclinical natural history of lung cancer. Such information would be valuable in the future design and interpretation of screening programs for the early detection and treatment of lung cancer, should this activity prove to be of benefit in experimental trials. Secondly, it is of interest to determine if some characteristic of sputum cytology readings can be utilized to determine a decision point at which to transfer a uranium miner to another job category within the industry in order to effectively reduce his probability of developing lung cancer. For example, a reading of moderately atypical cells in the sputum has been suggested as a possible decision event which would be sufficient to warrant such a transfer. In the analysis which follows, an attempt will be made to develop information relevant to these issues. The data base for our analysis, largely developed over a period of more than 20 years by Drs. Geno Saccomanno and Victor Archer, contains sputum readings classified into one of eight cytologic states: normal, metaplasia, mild atypia, moderate atypia, marked atypia, suggestive, suggestive-equivocal, and malignant. The two suggestive states will not be considered as these categories seldom contain enough observations to generate meaningful results. The duration in a given cytologic state for a given individual is calculated as the sum of the intervals between consecutive sputum readings in that state. When two consecutive readings are in two different states, half the between-readings interval is assigned to each of the two states. Among all individuals in the data base, only those having at least two sputum readings are used so that a duration can be calculated. Furthermore, the analysis is limited to individuals whose successive readings are no more than 24 months apart, so as to focus on a population which was under surveillance on a reasonably regular basis. RESULTS The overall distributions for the cytologic states are displayed as histograms in Figures 1-3.

Jan 1, 1981

By Bruce D. McCellan

INTRODUCTION The gathering arm loader was patented by Joseph Joy in 1916. This unique machine motion became the standard of the industry and is used on underground loaders and continuous miners throughout the world. Figs. 1 and 2 show two types of gathering arm loaders. The gathering arm loader is a crawler-mounted ma¬chine primarily used for loading coal in a conventional mining system. Other uses also have been found for this machine in noncoal mines and other industries. MECHANICAL ASSEMBLIES The front or loading end of the machine, commonly called the head, consists of a tapered structural frame, hinged at the rear to the chassis and riding the ground at the front. Two bevel-gear cases are mounted in the head. They are connected by a timing shaft, which also acts as a conveyor drive shaft on most models. An eccentric disk attached to each gear case drives the gathering arms. The rear of the arms are attached to a swivel assembly. The motion derived from this arrange¬ment gathers loose material from the ground directly in front of the machine and loads it into the conveyor. Power for this drive is provided by two electric motors mounted under the gathering head. A speed reducer attached to each motor provides a gathering arm speed of approximately 5.2 rad/s (50 rpm). A torque limiting clutch is provided on each reducer to guard the system against sudden impact loading and continual overloading of the motors. The conveyor consists of an open box section through which a universal-type chain carries the ma¬terial to the rear of the machine, where it is discharged into the haulage vehicle. The chain is driven by a sprocket located on the timing shaft between the two gathering-head bevel-gear cases. On some large loader designs, primarily for noncoal applications, the chain is driven by a separate gear motor at the discharge end of the conveyor. However, the system driven by the timing shaft is used most widely. The conveyor frame is hinged just to the rear of the gathering head. It can be raised or lowered by hydraulic cylinders to the desired discharge height. The conveyor frame also is designed to pivot 0.79 rad (45°) either side of the center, just to the rear of the chassis. This feature increases the maneuverability of the ma¬chine and makes it possible to discharge material into the haulage system when in tight quarters such as when turning a crosscut. This function also is controlled by a hydraulic cylinder. The chassis of the loader consists of two crawler frames and cross members, holding the frames in position. The frames are either welded or cast as¬semblies. Replaceable wear runners are provided on the frames to take the wear generated by the crawler chain. An adjustable take-up idler keeps the crawler chain in proper adjustment for maximum life. The drive system for the crawler chains varies with the manufacturer, but it generally is one of two basic designs. One system provides a traction motor for each side of the machine, coupled to a speed reducer. A spur and bevel-gear train transfers power to the sprocket, which drives the crawler chain. Directional control is achieved by disk-type clutches. The other system also uses a motor for each crawler, coupled to a speed reducer. A roller chain drives the sprocket, which in

Jan 1, 1982

By Prescott Greene

AVAILABLE MACHINES Machines discussed in this section include only rubber tire mounted kerf cutting machines of a self¬propelled electrically powered design. Such machines handle bars up to 4.6 m (15 ft) in length, cutting kerfs of from 115 to 191 mm (4.5 to 7.5 in.). The basic design concept of all machines currently in production is very similar. The chassis features dual driving wheels on the cutter-bar end, fixed directly to the main frame. The steering wheels on the outby end of the frame are articulated in some manner to keep all four wheels on the floor over uneven bottom. Machine power for tramming, and all other functions except powering the cutting chain, is furnished from a pump motor that drives the specific pump or pumps required by a given machine. The cutter chain is driven through a mechanical drive train from a separate electric motor having no duty other than driving the chain. The general arrangement of a cutting machine is shown in Fig. 1; two types of machines, with their specifications are shown in Figs. 2 through 5. A hydraulically powered cable normally is included to eliminate the need for manual cable handling and make this a one-man machine. A bug duster, though optional, is needed for most applications. The bug duster is an hydraulically driven scroll conveyor located directly behind the cutter chain drive sprocket. Its function is to move the cuttings away from the sprocket to prevent them from packing up in this area and consuming a lot of machine power to further pulverize them. Available machines vary from models 610 mm (24 in.) high that can cut horizontally only from 203 mm (8 in.) below ground line to 229 mm (9 in.) above, to machines 1143 mm (45 in.) high that can cut in any plane from 178 mm (7 in.) below the ground line to 3.7 m (12 ft) above it. Most currently manufactured machines fall in the 17 200 to 20 400 kg (38,000 to 45,000 lb) weight range, but can be dis¬assembled for underground installation where weight and size are a problem. Maximum teardown results in all pieces having a weight of 5400 kg (12,000 lb) or less and a size allowing passage through a 1.2 x 2.1-m (4x7-ft) opening. Further reduction in either size or weight of the parts becomes a serious problem. Machines currently being manufactured are avail¬able for use with power supplies of 250 v dc; 500 v dc; 440, 550, or 950 v ac (60 Hz). Also available are machines for use on 415, 500, and 950 v 50 Hz. De¬signs to allow operation on other voltages of either 50 or 60 Hz can be made available without any serious problems. As of September 1977, a typical medium-seam ma¬chine without the extra options needed for unusual conditions, but fully equipped to operate, was selling for about $185,000. CAPABILITIES Materials That Can Be Cut The problem of deciding before a machine is put on the job whether it can work effectively is an old one that has never been completely solved and probably

Jan 1, 1982

By P. T. Bissert, J. P. Ducarme, J. D. Noll, C. C. Jobes

"There have been about 42 fatalities in underground coal mines between 1984 and 2015 where the victim was struck, pinned, or run over by a mobile machine (MM) such as a shuttle car, scoop, or battery hauler. MSHA has issued a proposed rule requiring proximity detection systems (PDS) on MMs in an effort to prevent future fatalities. Currently approved PDSs use electromagnetic (EM) technology to detect the presence of a miner and impose machine controls to prevent contact. The disadvantages of an EM-based PDS on MMs include the need for a miner-wearable component (MWC), environmental interferences, EM interferences from other devices, and low localization range and accuracy. NIOSH researchers are investigating alternate technologies that could improve the capabilities of the current PDSs. Many technologies were reviewed, and NIOSH researchers identified RFID, LIDAR, RADAR, ultrasonic detection, and computer vision as candidates for PDS development. The performance characteristics of each technology and its applicability to an underground mining environment will be discussed, along with international developments and implementations, and the concept of sensor fusion. INTRODUCTION PDSs provide safety to mine workers by detecting their presence (and in some cases location) around a machine. These PDSs provide a warning zone, which triggers an alarm, and a stop zone, which halts the machine before it contacts the miner. The Mine Safety and Health Administration (MSHA) issued a final rule requiring that PDSs be installed on all continuous mining machines (CMMs) on the working section, with the exception of full-faced machines, according to a phase-in schedule for newly manufactured and existing equipment [1]. The PDSs that were designed for CMMs have since been adapted to other mobile machines. Mobile machines include scoops, which are used as general utility vehicles, as well as coal hauling machines, such as shuttle cars, ram cars, and continuous haulage systems. MSHA has examined all accident reports from 1984 to 2014, and determined that 42 fatalities could have been prevented on mobile machines had a PDS been installed [2]. As such, MSHA issued a proposed rule that would require mobile machines on working sections (with the exception of longwall sections) to be equipped with a PDS. This rule would follow a phase-in schedule for newly manufactured and existing equipment. Currently available PDSs are based on electromagnetic (EM) technology. NIOSH researchers developed a shell-based magnetic flux density distribution model [3] for determining the position of a miner near a machine. A typical system would consist of four generators positioned around the perimeter of the machine. Each generator is a coil of insulated wire wrapped around a ferrite core. The magnetic field strength is proportional to the current running through the coil. Miners working on the section have an MWC, which is a transceiver that measures the field strength emitted by the generators and transmits a data packet over an RF link which contains the field strength reading to the PDS controller mounted on the machine. The magnetic field generators are pulsed sequentially such that the MWC can identify which generator’s field strength it is reading. This type of system utilizes the principle of magnetic flux density: the closer the MWC is to the field generators, the higher the field strength reading. Thus a miner’s presence can be determined once he gets too close to the machine, because his MWC will measure a magnetic flux density beyond a certain threshold. These thresholds are utilized on a per-generator basis to “shape” the fields for both warning and stop zones around a machine, as illustrated in Figure 1."

Jan 1, 2017

By M. Christou

The LIFE-Environment Demonstration Project ?Treating Waste for Restoring Land Sustainability?(www.bangor.ac.uk/ies/TWIRLS/TWIRLS_home.htm), a project initiated by the University of Wales, Bangor, UK in which NAGREF is a partner, aims to demonstrate the effective use of industrial and municipal wastes for the remediation of post-industrial and degraded sites in urban and rural sites. Within the framework of this project, a schist quarry, which is located in a mountainous area near Athens and which has been exploited in the past for cement production, was chosen for bioremediation experiments in collaboration with the quarry owners, TITAN Cement Co. Three different areas within the quarry were selected representing 3 different topographies. These were a compacted, flat area with seasonal water logging (untreated control), a gently sloping, land formed area; and artificially created mounds of overburden rock material formed on the flat area. Compost produced from mixing municipal waste, green waste and digested sewage sludge at the recycling plant of the Association of Communities and Municipalities in the Attica Region (ACMAR), near Athens was used for nutrient amendment. Six replicates of three treatments were applied with varying compost application rates in order to examine the effect of the addition of organic matter in the establishment of suitable vegetation on the infertile schist material. Pine trees (Pinus halepensis Mill.), 4 per replicate, were pocket planted in a volume of 3 liters of schist or schist-compost mixture (1:2 and 2:1). Tree height, main stem diameter and branching were measured following planting and these measurements will be repeated annually to test hypotheses that compaction and water logging reduce Pine establishment by natural regeneration and that nutrient addition improves establishment of pine on schist waste. The availability of nutrients and variation in other soil physico-chemical parameters will also be studied in each treatment and related to plant growth. As a demonstration project, open days are planned on the site for the public and official bodies involved in municipal waste handling and other relevant areas. The aim of these will be to illustrate the effective re-use of waste for the sustainable restoration of post-industrial and degraded land.

Jan 1, 2006

By Thomas Oberndorfer

INTRODUCTION Since several years the author is engaged in research activities in the field of mining method modelling (Oberndorfer, 1993 and 1994). The initial goal was to improve and assist the selection of the most appropriate mining method. However during investigations it became obvious that the most beneficial application of computer support is the fast calculation of numerical assessment rather than the consideration of all, i.e. also qualitative, criterions necessary for final decision. Numerical results serve - together with qualitative considerations based on the experience of the mining engineer - as a basis for decision. The emphasis is on "fast" calculation, as only this property makes it possible to calculate and compare many distinct variations. As long as no procedure exists, which optimizes all involved parameters, calculating many variations is of great important, because it is the only way to increase the chance to pick a alternative "close to optimum", but also to learn about the sensitivity the result will react on changes of input parameters, which are rarely known exactly. Generally speaking many trials give a better basis for final decision. It is no secrete that the most important obstacle towards quick calculations of a series of alternatives is not the processing time of the computer, but the time required for setting the input data in way that the computer can start processing, which is usually inter- active man work on the screen. Furthermore usually several distinct programs are required to cover the whole range of involved problems, with corresponding problems in automatic data transfer. The research work's goal was to overcome these problems. The approach developed was to describe a mining method more in a theoretical way (i.e. to give the computer some degree of "consciousness" on mining methods), which then can be applied under any specific conditions. E.g. a cut-and-fill operation will be "generally" the same in narrow or thick deposit areas, but the resulting keyfigures, e.g. total tonnage, productivity, dilution and loss, or finally costs per ton, will differ significantly. The approach (MMM = mining method modelling) presented reflects a deduction from the general case to a specific one (rather than the other way round). The developed model is general enough to be applicable to any mining method, and also to any degree of precision the mining method should be described, contrary to programs designed for specific mining method layouts which can be adjusted in a more or less wide range. PROGRAM DEVELOPMENT The principle correctness and feasibility of the approach was proved by a prototype program. However this program was more on a research level and far away from application for "real world" problems. The reason for this was, that it was regarded wasted time to re-invent certain tools, e.g. geometry intersection, deposit modelling or visualization. Hence it was tried to attract several software producers to invest in finalizing the model. Unfortunately this endeavor was not successful. Despite this defeat - and after some months of resignation - a final start on the project was decided. he reason for this decision was not at last the research activity in respect of low-cost underground mining methods for bulk- material as a substitute of typical quarry operations due to environmental restriction. This research work - which is in particular important for Austrian mining activities in beautiful and touristically used landscape - requires a lot of calculations in the way described. Both are still in state of progress, but - regarding mining method modelling - preliminary results can be presented. A complete re-programming of the program was required. Most obvious are the changes of programming language and operating system environment. Instead of Pascal under DOS C++ under MS-WindowsNT is now used. Several problems could be solved by this, in particular memory management, device drivers for input/output, and user friendlyness by up-to-date windows technology. However also more basic structures were changed. The idea was to utilize as many existing commercially available programs as possible to minimize programming efforts. WindowsNT helps already a lot by MFC technology, but also due to its multi-task facilities. This feature allows to run simultaneously several pro-

Jan 1, 1996

By Ross R. Bhappu

In 1989, William A. Vogely created the Mineral Economics and Management Society (MEMS). He did this after SME disbanded the Minerals Resource Management Committee, a group dedicated to technical programming of minerals economics and management issues. A few years later, SME reinstated this committee. MEMS held its fourth annual conference "Mineral resources: Managing dilemmas and opportunities," in Boulder, CO March 9-11, 1995. Some of the MEMS programs are similar to the Minerals Resource Management Committee programs. However, MEMS conferences cover more diverse issues. These include minerals policy, environmental legislation impacts on the minerals industry and minerals economics issues. The 1995 MEMS conference included six sessions and three distinguished speaker addresses. Papers topics included US and Canadian mineral policy, financial strategies for managing mining companies, Latin American mineral investment opportunities, quantitative economic modeling and the impact of environmental interest groups on the minerals sector in the United States and Canada. One session was dedicated to students who competed for $1900 in prizes for outstanding papers.

Jan 1, 1995

By Pierre M. Gy

The final purpose of exploration is to esti- mate the grade and tonnage of the various qualities of ore contained in a given mineral deposit. This estimation cannot be directly carried out on the whole of the ore body. It usually requires four steps: Core Drilling Campaign The ore body is represented by a series of cores further divided into core sections of equal or unequal length. Core Splitting The core sections are longitudinally split into two halves: the first half is kept for various examinations and physical measurements, and stored for further reference; the second half (core sample) is used for determination of the chemical analysis. Core Sample Reduction The chemical analysis cannot be directly per- formed on the core samples. It requires the constitution of assay portions usually weighing from a few decigrams to one gram (more in precious metals assaying) and passing through a 100 or 150 mesh screen.

Jan 1, 1979

By R. A. Pindred, M. P. Lam

Cyanide is a conventional lixiviant for gold extraction but its' toxicity is a major concern for the environmentalist. There are many alternative non-cyanide lixiviants, such as thiourea, thiosulphate, chlorine and hypochlorite etc., which may reduce environmental sensitivity. This paper will discuss the leaching characteristics of gold in hypochlorite leaching solutions. The dissolution of gold in a hypochlorite solution not only depends on the concentration of the hypochlorite but also on other variables, such as hydrochloric acid, sodium chloride and temperature. The dissolution rate of gold in hypochlorite solution is much faster than other lixiviants. Due to the interference of impurities, gold extraction from an ore under ambient conditions requires large quantities of reagents. However, refractory gold ores may be successfully pressure oxidized in an autoclave at reasonably low reagent levels. Gold extraction to the pressure oxidation leach solution is rapid and complete in one stage, thus eliminating the need for further cyanidation.

Jan 1, 1990

By S. B. Patton, T. Novak, R. L. Sanford

A comprehensive approach to the development and organization of an expert system designed to assist electricians in diagnosing electrical faults in mining machinery is discussed. The system includes diagnostic, reference, and tutorial programs. The diagnostic portion of the system is based on a causal-reasoning approach, which is used to establish a hierarchy of symptoms and the subsequent production rules for identifying component-level failures in the motor-control circuit of a shuttle car. With this approach, conclusions are developed directly from knowledge of the circuit and its behavior (in both working and failure modes) based on first principles, rather than the heuristics of a domain expert. The symptoms and their hierarchal structure are formulated by creating "what if' scenarios and performing a qualitative circuit analysis for each fault. Techniques for incorporating interactive graphical displays of the circuit schematic into the diagnostic portion of the system are also presented. Supportive computer programs that provide reference material and tutorials on the circuit are also described.

Jan 1, 1990

By Murali M. Gadde, Syd S. Peng, C. T. Holland

Cutter roof failure is one of the most common ground control problems affecting the safety and economy of an underground coal mine operation. Several attempts were made in the past to understand the processes responsible for cutter development and to devise mitigating measures. Yet, many cutter failures still happen and their advance prediction remains as elusive as before. One of the reasons for the impasse in predicting cutters lies in the limitations of the current numerical modeling technologies. The development and propagation of cutters is a progressive phenomenon and continuum models can only approximate this process. Since the discontinuum modeling is severely restricted by the lack of reliable input data, alternatives must be sought to better understand cutter failures using the continuum models. This paper proposes one approach using strain-softening material model to simulate the cutter initiation and propagation process. An attempt has also been made to provide some insights into the reasons for the change in cutter location at different parts of a development layout.

Jan 1, 2005

By U. E. Paakkinen

Progress in computer control of flotation circuits has developed parallel with advances in reliability of on-stream X-ray analysis (Cooper, 1976). Attainment of acceptable results with X-ray fluorescence analysis over the past ten years has led to application of computer control in more than 30 concentrators. Gains in applying flotation control also have given incentive to improving measurement and control devices for other essential functions in control systems. At the same time, advancing technology has made available computer systems having characteristics practical and economical for the physical environments of concentrators and ore processing plants. Benefits from flotation control have centered on increased efficiency in mineral separations, derived by stabilizing circuit operation subject to changing ore characteristics. The gains have been most evident in processing complex ores for multiple concentrate products. In many cases, particularly when coupled with grinding control, the economic results from on-line control have been substantial compared to investments in monitoring and control equipment.

Jan 1, 1979

By John S. F. Dunlop

One of the more difficult problems frequently facing geologists and mining engineers is that of estimating the ore reserves of mineral deposits. Cost conscious mining operations invariably require an accurate knowledge of tonnage and grade for both short and long term planning. A case study is presented, showing how a classical frequency distribution was fitted to a population of drillhole assays. Using these results, a grade- tonnage curve is derived using the Lasky expressions, and the results discussed. A geostatistical analysis is then presented. First, variograms are drawn, models fitted to them and analyses made of drift and anisotropy. Second, simple block kriging is carried out and the results plotted on large 1000 scale plans, suitable for overlay with routine .mine plans. Kriging results are verified using two independent techniques: random kriging on a regular grid and inverse distance weighting. Finally an attempt is made to optimize the drilling grid, and to fully tabulate the total computer running costs for the work described.

Jan 1, 1979

By Henry O. Ash

On June 16, 1982, Rio Blanco submitted a request to the US Minerals Management Service for a suspension of operations under its federal oil shale lease. The company also wanted a five-year delay in the requirement for further minimum royalty payments and a five-year extension of the primary lease terms of 20 years. The request was granted in late July in the midst of yet another slump in oil shale activity. Rio Blanco holds one of four Federal oil shale leases that were granted under the Interior Department's Prototype Oil Shale Leasing Program in 1974. That year the DOI offered for lease six federal oil shale tracts, two each in Colorado, Utah, and Wyoming. The leases in Colorado and Utah were granted for substantial bo¬nus bids as shown in Table 1, but the Wyoming tracts had no takers. In the eight years hence, no oil shale has been produced commercially. Indeed, the prospects for successful oil shale development have fluctuated widely, continuing a 60-year- old cyclical pattern in the fledgling industry. The prototype program was developed and initiated while this infant industry struggled to begin anew, without the benefit of any federal oil shale leasing regulations, or any history of federal oil shale leasing at all. Permanent regulations were not required because the effort was limited and designed to provide the foundation for any future leasing program. Section 21 of the Mineral Leasing Act of 1920, a Nov. 30,1973 Federal Register notice, and the leases themselves constitute the legal and regulatory framework for leasing and developing prototype oil shale tracts. The prototype leases contain some unusual provisions including installment payments of the bonus bids, credit of early development costs against payments due the government, adjustment of royalty rates according to regional crude oil price fluctuations, and unusually extensive and detailed environmental stipulations. Several of the provisions were intended to provide incentives for early development and financial penalties for failure to develop. The requirement for minimum royalty payments in lieu of actual production royalties beginning in the sixth year of the lease falls in the latter category. After the sixth year, the required minimum payment is increased east year through the 15th year by an increment equal to that due in the sixth year. The minimum royalty then remains the same through the final five years of the primary lease term. Minimum royalty is calculates on a different production level for each prototype lease, based on the quantity of 125 L/t (30 gal/st) shale estimated to be recoverable from. each tract by room-and-pillar mining. That mining system was selected for reference as the only proven extraction method for oil shale at that time (early 1970s) The minimum royalty (production) rate was calculated for each tract by direct comparison of its estimated recoverable 125-L (30-gal) resource to a hypothetical tract containing 1.9 Gt (2.1 billion st) of recoverable 125-L (30-gal) shale. For the hypothetical lease tract, the minimum production for the sixth year was calculated to be 3.2 kt/d (3,500 stpd), increasing by a like amount each lease year through the 15th to a daily production rate of 31.7 kt (35,000 st). Under this formula, the hypothetical lease tract would yield $16 million in minimum royalty payments over the 20-year primary lease term if there were no production, no development credits taken, and no adjustment of the royalty rate due to fluctuation of regional crude oil prices. The production rate of 31.7 kt/d (35,000 stpd) was a somewhat arbitrary figure that was considered, in the early 1970s, to represent a minimum size for a commercial oil shale operation. A production rate of 7.9 dam3/d (50,000 bbls/day) soon supplanted that smaller mine production figure as a rule of thumb for the minimum size for a commercial facility. The proportional relation-

Jan 11, 1982

By A. Fiszer

One of the major problems with deep Polish hard coal mines are deformations of close-to-wall galleries since they are necessary and must he kept open for emergency situations and are largely responsible for work disturbances. In order to ensure safe and effective working operations, it is necessary to often rebuild galleries which, apart from the inevitable stoppage of operations, is greatly disadvantageous for the operation of a mine. The largest problems with maintaining the galleries open occur in Lower Silesian Coal Field mines where sometimes it is necessary to reconstruct the galleries several times [(Fig. 1)]. In light of what has been said above, the aim of this article is to present the reader with methods and ways of determining deformations occurring in close-to-wall galleries. Squeezing or convergence of headings is only a portion of the entire process of ground movement caused by ground disturbance due to workings. The term "squeezing of headings" is to be understood as a shortening of the distance between a point set on the roof of a heading and the corresponding one on the floor. It is very important to know the rules governing the process of squeezing in a particular mining system since it affords possibilities to define changes of the process and its parameters at any moment for a given system. This is of utmost importance because convergence of a heading is one of the most significant manifestations of ground pressure which has already been proved by studies performed by Everling (1). On the basis of those studies it has been observed that there exists an almost linear relation between the rate of convergence and loading [(Fig. 2)]. This, in turn, means that the deformations of the already exploited portion of heading are proportional to the pressure changes.

Jan 1, 1982