Search Documents

By David R. Maneval

The mining of coal in Alaska receives special treatment in the Surface Mining and Reclamation Act passed by the U.S. Congress in 1977 (PL95-87). Section 708 of the Act required that the National Academy of Sciences (NAS) perform a study of "Surface Coal Mining Conditions in the State of Alaska, In Order to Determine Which, If Any, of the Provisions of the Act Should be Modified with respect to Surface Coal Mining Operations in Alaska'. The special committee on Alaskan Coal Mining and Reclamation of NAS completed their study in 1980 and prepared their report (NAS, 1980). At this time, the Secretary of the Interior and the Congress have taken no special action with regard to either special regulations geared specifically to Alaska or to submitting special draft legislation to the Congress requesting amendment of PL95-87. The Alaska Department of Natural Resources (DNR) has drafted legislation which was submitted to and passed by the 1982 session of the State Legislature. The State regulatory agency has also drafted the requisite regulations necessitated by the amendments to the State's Basic Coal Mining Act. It is expected that by mid 1983 DNR will have applied for "primacy" to the office of Surface Mining, U.S. Department of the Interior (OSM). They will be submitting the amended State Law, the revised State Regulations and an operational plan as called for by OSM regulations. The presumption is that primacy will be granted to the State of Alaska under PL95-87. The Usibelli Coal Mine continues to be the only operational commercial mine for coal in the State of Alaska. It produced 7.26 x 10 kg (8 x 10 " T) of coal during 1981. The current mine was started in 1920 by Captain Lathrop, an early developer and pioneer in Alaska. It was continued by Emil Usibelli and was taken over after the death of E. Usibelli (in the 1964 earthquake) by his son Joseph Usibelli who is the current President of, the firm3 Their latest improvement is a 27.5 m (36 yd. )Bucyrus-Erie Dragline which handles most of the overburden removal. The coal is removed by a loader-truck combination. They are currently starting up the use of a unit train loading facility which handles forty cars every two hours. This loading system was a development necessitated by the Sun Eel Export contract. The Alaskan Railroad was a major cooperator in the development of the first unit train loading facility in Alaska.

Jan 1, 1982

By David S. Bolin, Bruce Kennedy, Leigh A. Readdy, Graham Mathieson

INTRODUCTION The estimation of ore reserves is a process that begins with the earliest exploration stages on a property and continues throughout any subsequent evaluation and exploitation of the deposit. During exploration and preliminary evaluation, the results of these reserve estimates constitute the basic data for prefeasibility studies and economic analysis. The decision to continue exploration and development or to abandon a prospect is often based upon these studies. During the active life of a mine, reserve computations are continuously revised to assist in development planning, cost and efficiency analyses, quality control, and improvement of extraction and processing methods. Accurate reserve estimates are also required when financing a project, purchasing or selling a property, and for accounting purposes such as depletion and tax calculation. It is important to remember that the reliability of ore reserve estimates varies progressively through time as more and more information becomes available. The lowest order of reliability of estimation of reserves exists at the time of discovery. The maximum level of certainty concerning the ore reserves within a deposit is reached when the deposit is completely mined out. Between these two extremes are variable levels of certainty as to the tonnage and grade of the resource. This is particularly true of that portion of the mineral resource that constitutes the "minable reserve," as this portion is dependent upon economic as well as geological and technological factors. In the following discussion, several of the factors affecting ore reserve computation and some of the commonly used methods of calculation are presented. The first part of the discussion is confined to classical methods of hand calculation utilizing level maps and sections. The second part of the discussion presents an overview of somewhat more sophisticated methods of geostatistics which have been developed within recent Years. Geostatistical methods frequently provide the best ore re- serve estimates for many deposits. However, a digital computer is required for such geostatistically derived estimates due to the number and complexity of the cornputations involved. For many properties the classical nongeostatistical methods are adequate, particularly in the early exploration stages. In fact, it is the author's contention that even at those properties where the geostatistical approach is employed for reserve estimation, the geological and engineering staffs should prepare an estimate by traditional means. This provides an internal audit for the reserves and at the same time requires a continuing close appraisal of geologic problems that influence the presently known geostatistically estimated reserves and production at the mine.

Jan 1, 1982

By M. N. Rubin

This report summarizes the work performed under USBM contract J0100018. The purpose of this work was to document the currently available noise control techniques that can be incorporated into new coal preparation plants during the design stage. This was accomplished by reviewing the related experience of the contractor and subcontractor (Bolt Beranek and Newman Inc.), conducting an extensive literature search, visiting operating coal preparation plants, and contacting various equipment manufacturers. The treatments addressed include equipment substitution, changes in plant layout, and retrofit treatments for existing equipment.

Jan 1, 1982

By Paul T. Alexander

After evaluating preceding studies and designs of lightweight supports, ESD selected the all-hydraulic and the hydromechanical concepts for detailed design and fabrication. Further analysis and testing showed that the all-hydraulic design was economically justified and met all the crucial engineering, manufacturing, and design criteria identified during the evaluation phase. Following Bureau of Mines approval of the detailed design, four all-hydraulic support models were fabricated. Once functional and structural testing was successfully completed, 40 additional units were constructed and distributed for in-mine testing. The units were very well-liked by all three demonstration mines. The wide- spread interest in supports for seam heights greater than 8 feet led to expansion of the design data package. ESD designed four basic models to cover the new height range, but no hardware has yet been fabricated for the new design.

Jan 1, 1982

By Welby G. Courtney

The Bureau of Mines conducted an exploratory field survey to determine the amount of airborne coal dust, the size distribution of the dust, and the rate of dust deposition in coal mine airways. The survey was taken at eight underground locations in operating coal mines. The rate of dust deposition in the airways depended upon the concentration of air-borne dust and decreased exponentially with distance from the dust source. The rate was seemingly independent of the size of the airborne dust particles. The rate of deposition, along a return, for the coal dust from a fade being mined with a continuous mining machine closely matched the rate for the rock dust from a trickle duster. A sample rock-dusting schedule for that return is presented; that is, the trickle duster should be operated for 3-112 hours after every five shifts to protect the average deposit of new float coal dust in the return, but operation after every three shifts is preferred, to protect an occasional heavier deposit.

Jan 1, 1982

By Robert W. Waytulonis

The Federal Bureau of Mines contacted 29 organizations--mine maintenance departments and equipment manufacturers--concerning service experience and construction of diesel exhaust-gas water scrubbers. Scrubbers are used primarily on mobile diesel-powered mining equipment subject to 30 CFR 36. Because the scrubbers cool exhaust gas and act as flame arresters, their failure would compromise safety in underground mines. Scrubber maintenance problems and frequency and modes of failure were identified, and construction materials and techniques were evaluated. Three failed scrubbers were randomly selected from two cooperating mines, and their modes of failure were determined. In conjunction with failure analyses, mine water and scrubber water solution samples were chemically analyzed to characterize corrosive properties. Sludge and mineral deposits and different types of corrosion and metal fatigue were identified as problems leading to scrubber failure. The corrosion process is explained; maintenance practices, construction materials, and manufacturing methods that could be used to reduce premature failures and allow longer effective scrubber life are identified; and basic changes in scrubber design are suggested. If type 304 stainless steel is used throughout the scrubber and regular maintenance is performed, long, dependable service can be expected. Other materials such as 304L, 316, or other more exotic steel alloys may be necessary under certain conditions.

Jan 1, 1982

No noticeable international economic recovery occurred during the first half of 1982. Industrial production continued to fall in the United States, Canada and Italy and to hold steady in France. Modest increases in production occurred in Japan, England and Germany, although some decline took place from March to April. The U.S. mining industry has been more severely affected by the world recession than the U.S. economy. This is not immediately obvious when comparing the U.S. mining industrial production index from 1981 to 1982, which decreased 8.5 percent, as compared to a 7.1 percent decline for total domestic industries. However, "mining" industry statistics in the U.S. national economic accounting system include petroleum and natural gas as well as minerals. When metal mining is isolated, the decrease in production from 1981 to 1982 is 19.3 percent. If metal mining production continues to perform at the same rate during the third and fourth quarters of 1982 as during the first half of the year, declines from 1981 levels will be quite severe. Aluminum will have declined 22 percent, copper 18 percent, iron ore 24 percent, lead 3 percent, steel 29 percent, titanium 2 percent and zinc 25 percent.

Jan 1, 1982

By M. S. Cetiner

This report describes the prototype development If a medium to high voltage coupler, suitable for underground coal mining operations. The coupler has such unique features as automatic dust covers and continuous phase pin shielding. The coupler is rated at 7500V and 250A continuous operation.

Jan 1, 1982

By Mark Davis

As a follow-on to previous investigations, reduced treatment levels for the optimum fire retardant/preservative system were evaluated. An impregnation of 3.5 lbs/ft3 of Fyreprufe followed by 3.5 mils of Albi 107A plus an Albi 144 top coat provides adequate flammability and flame spread protection. However, there is a significant loss in compressive strength at this retention level, particularly parallel to the grain. A pilot-size pressure impregnation cell was installed for use in treating larger quantities of wood. Based on the optimum fire retardant/preservative treatment system - Fyreprufe/ACC impregnation followed by Albi 107A/144 intumescent paint - a model treatment pro¬cess was developed, capable of handling 3 million board feet per year. Associated annualized costs per thousand board feet (K bd. ft.) for treating 12" x 12" timber were calculated to be: Impregnation: $-63.00/K bd. ft. Coating: $112.00/K bd. ft. A literature search was conducted to identify materials and techniques suitable for a localized stench warning system for heated mine timbers. Based on this review, a number of stench warning concepts were developed. Those based on surface coating were given a preliminary laboratory evaluation.

Jan 1, 1982

By J. M. Kuchta

The Bureau of Mines investigated four simulated coal gob fires to obtain a more reliable data base for defining the state of a sealed mine fire and to evaluate the performance of various fire detectors. The fires were conducted in a multiple-entry section of the Bruceton Experimental Mine by heating 4,000 to 21,000 pounds of rubblized bituminous coal to ignition and monitoring the temperature, gas emissions, and smoke under ventilated and sealed combustion conditions. Product gas concentrations depended upon the coal bed size, coal temperature, heating rate, and oxygen concentration or ventilating condition but varied little after the fires were sealed. Expressions are given that define the temporal variation of CO and O2 after sealing when the coal temperature decreased rapidly. The CO/C02 ratio is a more sensitive fire indicator than the CO/?02 ratio, although the latter is universally used for detecting the incipient state of a coal fire. In comparison, the Jones-Trlckett ratio [ CO2 + ¾ CO ? ¼ H2 ?O2] is less reliable for this purpose. The best detector for determining the state of a mine fire after it has been sealed was a prototype submicrometer smoke particulate sensor. Limitations of the various detector systems are also discussed.

Jan 1, 1982

By Roy J. Greenfield

The overall purpose of this research was to improve the understanding of how the post disaster seismic location system works, what determines the signal size, what is the character of the seismic noise, how well the system can detect blows from a trapped miner, and how well the source of the blows can be located. Based on an improved understanding of the system, improvements were suggested in the seismic subarray configuration, and in the system deployment procedures. An important part of this research was to become familiar with the seismic system operation and to work with the Westinghouse Electric Corporation and MSHA personnel to modify and improve the system. To accomplish this we participated in the planning, execution, and data analysis for a number of seismic field tests. Several in-mine field tests were conducted. From these tests the capabilities of an in-mine seismic system were analyzed and specifications were made for designing an operational in-mine system. Several portions of the work done on this grant were covered in previous reports. In particular the capabilities or the presently operational surface based seismic location system has been thoroughly documented by Durkin and Greenfield (1981). Portions of the location procedures and effects of mine openings on seismic signals have also previously been published. These portions or the work will only be summarized in this final report.

Jan 1, 1982

By Francois E. Heuze

Today, the favored option for disposal of high-level nuclear wastes is their burial in mined caverns. As an alternative, the concept of deep disposal by rock melting (DRM) also has received some attention. DRM entails the injection of waste, in a cavity or borehole, 2 to 3 kilometers down in the earth crust. Granitic rocks are the prime candidate medium. The high thermal loading initially will melt the rock surrounding the waste. Following resolidification, a rock/waste matrix is formed, which should provide isolation for many years. The complex thermal, mechanical, and hydraulic aspects of DRM can be studied best by means of numerical models. The models must accommo¬date the coupling of the physical processes involved, and the temper¬ature dependency of the granite properties, some of which are subject to abrupt discontinuities, during [a]- ß phase transition and melting. This paper outlines a strategy for such complex modeling.

Jan 1, 1982

By Paul Soros

There has been a steady increase in shiploading capacities in the last 20 years. One of the reasons has been the general increase in ship sizes, with the largest vessels currently around 280,000 DWT in the bulk trades. Obviously, the large vessels can accept loads faster and their economic utilization justifies higher loading rates, up to a point. The high capacity facilities have loading rates far beyond this point. The economic logic of these facilities is based on the fact that for a given annual throughput the capital and operating cost of a single high capacity system is less than that of two smaller systems. Thus, where the projected throughputs are large, there is a great incentive to reduce capital costs by maximizing capacity. Tubarao, in Brazil, designed for a 100 million ton per year throughput with one finger pier and one high capacity berth is a successful demonstration of this approach. In the current state of the art, the high capacity systems have shiploading capacities in the range of 10,000 - 16,000 TPH, for annual throughputs of 20 million tons and larger per loading berth.

Jan 1, 1982

By Lorne Roberts

"This paper describes: how Falconbridge's fire protection system meets the needs of the Lockerby Mine; why the system was chosen over other alternatives; how it is unique, but can be applied to any underground situation; details of the system; how it was installed; the role of a fire protection company; and general considerations when choosing afire-suppresant system.IntroductionIn May 1981, after about three years of searching for the best fire protection system for the re-opened Lockerby nickel mine at Whitefish, near Sudbury, Ontario, Falconbridge Nickel Mines Ltd. installed an AFFF type of fire suppression system. It was designed and supplied by the Fire Protection Systems Division of Levitt -Safety Ltd.The system protects against loss of life, equipment and/or production, and it complies with legislation should fire occur in a fuel storage area and/or scooptram garage and service bay at any or all of the three levels near Shaft No.2.About 18 months after Levitt-Safety began working with Falconbridge, the system was designed and installed at a cost considerably less than that which would be incurred if fire damage caused the loss of two days of production.The fire protection system, which overcomes a limited water supply, is a self-contained balanced pressure proportioning system. It senses the flow and automatically determines the correct concentration of foam to water. It is triggered by a release mechanism which operates a bank of nitrogen cylinders, which in turn pressurize a pressure vessel to expel the foam concentrate and water from the tanks. This solution is piped to sprinkler heads in the specific area."

Jan 1, 1982

By W. J. Kennedy

Energy management in industry has recently achieved notice as a way to reduce costs without decreasing production. The same approach used in industry can be used in mining, even though the processes are different. This approach is based upon a three phase system of data collection, implementation, and monitoring, all aimed at gathering and evaluating information in such a way as to have a maximum impact on energy costs. This approach has these advantages: (1) attention is directed toward the part of each mining system where the most energy is used, (2) input is collected systematically from all persons involved in the operation of each system, and (3) a means of monitoring the progress toward established goals is created. The mining systems considered are truck haulage, conveyor haulage, blasting and crushing, and lighting. After describing the general procedure, the authors have described a specific method for improving the efficiency (and hence the energy consumption) of round-trip truck haul age systems.

Jan 1, 1982

By Joel K. Waterland

GENERAL DESCRIPTION Open cut-and-fill stoping for many years was prob¬ably the most widely used mining method in under¬ground metal mines. Then for a time this method was largely supplanted by the blasthole stope. It again be¬came popular as many mines reached depths or grades where methods requiring large open voids to remain open for extended periods of time became unsuccessful, often as a result of excessive dilution. The open cut-and-fill method is very flexible and is readily adaptable to almost any ore body. The standard application requires that a slice of ore usually 2.4 to 3 m (8 to 10 ft) thick be removed from the back of the stope, and as the ore is taken down, the back is dressed and rockbolted. After the back is secured, the broken rock is removed through rock passes to the level below. When the rock has been removed, the rock passes are extended upward the height of the ore removed, the stope is backfilled, and another cycle is mined. This method is best employed in plunging ore bod¬ies with considerable vertical extent, ore areas that re¬quire selective mining, ore areas where weak wall con¬ditions exist, and ore bodies that have an ore value that will carry this relatively expensive mining method. Blast¬hole stoping, shrinkage stoping, and other mining meth¬ods that do not employ rock passes in a stope are not efficient in plunging or flatly dipping ore bodies because the footwall makes ore removal quite difficult. Since mining is accomplished by taking down slices of the back, only small areas of the wall rock are exposed at any one time, and these only for short periods. Due to limited back height, areas of uneconomic rock may be left in place, or they may be mined and the material gobbed in the stope. Because the miners in the stope must work under freshly blasted areas, the amount of ground control is usually great. The volume of rock that is broken during one section of mining is relatively small and the amount of nonproductive work required is high. This results in limited productivity for the scope and, be¬cause of the nonproductive work that must be done on a regular basis, the production from the stope may be quite cyclical. SUITABLE ORE BODIES The open cut-and-fill method may be adapted to al¬most any type of ore body with a relatively high vertical extent. The ore body must be accessible at both top and bottom as well as at regular intervals throughout its vertical extent. Although adaptable to most ore bodies, the method is probably best employed where the ore has poor con¬tinuity and where most types of bulk mining would pro¬duce excessive dilution. In areas of poor ore continuity, the capability of continuous and extensive sampling dur¬ing the mining of each cycle makes this method very desirable. This capability also minimizes the amount of evaluation sampling that must be done before mining is started. Because of the extractive system used, the size and shape of the stope may be as readily changed as the sampling mandates. Probably the only ore characteristic demanded is that the ore has strength enough to be sup¬ported through the use of rockbolts or cable bolts dur¬ing the mining and backfilling cycles. Good planning, systematic sampling, and careful supervision will pro¬duce a product with less dilution than any other open stoping method. PLANNING Evaluation Planning Once it has been decided that the open cut-and-fill method would be the most efficient for mining a par¬ticular ore body, the next considerations would probably be the availability of an economical backfill material and the selection of an efficient transport system for this material. Although hydraulically transported mill tail¬ings are the most widely used product, this is not always practical due to mill location or the quality of the tailings. In such cases, backfilling may be used. The type of backfill and the type of equipment used will determine if a floor or capping on the backfill is required to minimize dilution during ore removal. The early selection of rock removal equipment is im¬portant since equipment usually determines the amount of development work required to bring a stope into pro¬duction and the size of the openings needed. The size and continuity of the ore body will usually determine the type of loading equipment. The use of slushers or load¬haul-dump (LHD) equipment captive in the stopes will minimize the amount of development. If the ore con¬tinuity is such that a ramp system for extraction can be used, the cost of development will be increased but the flexibility of continuous mining will minimize the cycli¬cal nature of the production. The height of the mining section usually is deter¬mined by the strength of the wall rock and the amount of back bolting required. Once this has been decided, the appropriate drilling equipment can be chosen. The number and sizes of the rock passes employed depends upon the type and size of the extractive equipment and the type of backfill that is to be used. Since the miners must enter and leave the stope each shift, the level inter¬val is usually maintained at approximately 45 m (150 ft). Access from the level above into the stope must be main¬tained at all times. The employees perform all the work in the stope and adequate ventilation must be provided. Stope Planning Due to the flexibility of the method and the vari¬ability of the ore zones, layout is usually done on a stope basis. In areas where continuity is a problem, the size of the stope is usually determined by the boundary of the ore (with all of the ore within that boundary being removed). In areas of good continuity where ramps are to be used, the length of the stope may be determined by the length of time each of the cycles (preparation, back¬fill, mining, and ore extraction) requires. The ramp work is then laid out so that access to the various parts

Jan 1, 1982

By Lawrence K. Nordell

This paper reviews practices used In the selection and sizing of belt conveyors and stacker systems commonly used in crushing and grinding plant facilities. Historical and modern methods of sizing this equipment are presented. The new methods Include: 1) probability theory applications for establishing component reliability, 2) "Monte Carlo" simulation techniques used in evaluating alternative system designs and in establishing service factors and corresponding availability Indices of each system, 3) quantifying the system and component operating characteristics using computer assisted modeling tools simulating rigid body and elasto-mechanic behavior for starting and stopping large and/or complex belt conveyors.

Jan 1, 1982

By J. W. Howland

A demonstration of dozer blades, designed for recontouring area-mined spoil and shown effective at Kansas and East Texas surface mine sites, was conducted at a Northeast Arizona surface coal mine to determine their effectiveness in different climate and soil conditions. The specialized tools were: 48-ft angle blade (on side-by-side tractor), 24-ft grading bar, modified 24-ft "U" blade (on side-by-side tractor), and a 13-ft deep "U" blade. Interest in the increased dirt-handling capacity of a previously built 22-ft blade, resulted in the design of a new Centerflow Blade of good versatility, able to maintain increased flow in otherwise difficult-to-move spoil. The demonstration involved 263 acres and showed a reduction of 50% in reclamation costs and 20% in fuel requirements for this mine site over conventional equipment. This savings resulted from improved technology, trained operators, and rigorous application of supervisory and monitoring techniques. Procedures and results of the demonstration are presented and analyzed here, along with guidelines for application of findings.

Jan 1, 1982

By John De La Vergne, Morri Medd

"This paper describes the sinking of a circular concrete ventilation shaft through 110feet of water-bearing overburden. The brine circulation freezing method was used to stabilize the walls during excavation. In order to take advantage of the cold winter weather, a custom-built heat exchanger (natural radiator) was installed in place of the standard mechanical refrigeration plant.IntroductionIn the late fall of 1979, Les Mines Selbaie awarded a contract to the firm of 1.S. Redpath Limited to sink a ventilation shaft on its property in northwestern Quebec. The shaft was to be sunk through 110 feet of water-bearing overburden and approximately 90 feet of rock to a final depth of 200 feet below surface. The water table was measured at 16 feet below the surface elevation.After examining the soils report and diamond drill core from the shaft area, Redpath came to the conclusion that freezing was the best way to stabilize the overburden for sinking. LocationThe mine site is 420 statute miles due north of Toronto, just east of the Ontario/Quebec border (Fig. 1). Access to the site is provided by a development road built by the Quebec government from the Matagami/lames Bay Highway. History of Ground Freezing In the Mining IndustryIn 1862, shortly after the development of mechanical refrigeration, a mine shaft was sunk through an artificially frozen water-bearing formation in the United Kingdom. The first application of ground freezing in North America was for a Michigan iron ore shaft in 1888. Since that time, the freezing method has been used to sink hundreds of shafts throughout the world. Its best known application was to sink sixteen shafts through the Blairmore Formation in Saskatchewan.The same method has been used in every case. It involves chilling a calcium chloride brine solution and circulating the brine in concentric pipes, spaced between 3 and 6 feet around the shaft perimeter and usually 4 to 5 feet from the limits of the proposed shaft excavation."

Jan 1, 1982

By Terry J. Close

The Anaconda Pintlar Wilderness has a significant mineral resource potential. Five deposits have resources totaling about 180,300 tons (163,600 t) containing silver, copper, lead, gold, and zinc. Another 11, not sufficiently exposed to allow resource estimates, may have additional mineral potential. No placer deposits are indicated. There is little potential for the production of sand, gravel, or stone; alternate sources are closer to markets. No oil, gas, uranium, or geothermal energy sources are known to be present.

Jan 1, 1982