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By R. H. Winn

A method for the selection of the most suitable corrosion inhibitor for a particular system is given. The method involves the evaluation of surface passivity by means of copper ion displacement after the metal surface has been subjected to the corrosive atmosphere. The application of the method to the selection of suitable inhibitors for use in producing wells is described. Correlation with recognized laboratory tests has beet1 found to be good within the limitations of the test. The use of the test to evaluate inhibitor persistencc is described. INTRODUCTION Corrosion as it is associated with the industry shows wide variations in its qualitative as well as its quantitative aspects. This complicates the work of the corrosion engineer. In the past few years the efforts of the corrosion engineer have been further taxed by the large numbers of products being marketed as corrosion inhibitors by various manufacturers. These products are submitted to him together with a limited amount of data to support their use as inhibitors. The data are usually authentic and reliable yet all too often when these products are applied to an actual corrosive system the results are disappointing due to the ineffectiveness of the inhibitor and the corrosive system. The corrosion engineer must select a comparatively limited number of inhibitors for field trial from the large number available. The selection is usually made on the basis of a standardized laboratory test such as that adopted by the National Association of Corrosion Engineers or by field experience in an area; however, testing of this type is discouraging on large numbers of compounds, particularly in view of the specificity of environment. It has been an object of this work to devise a method which will allow the corrosion engineer to select the best inhibitor for a specific environment under field conditions. The test would of necessity be rapid and inexpensive. It would preferably eliminate the need for expensive and delicate instruments. STATEMENT OF THEORY The test as developed makes use of the fact that metal ions will be displaced from solution by metals having a more positive electrode potential. Thus, if an active iron surface is exposed to a solution of a copper salt, free copper will plate spontaneously on the iron surface as it is displaced from the salt solution by iron ions. If the iron surface is passified, as by coating with an insulating-type organic corrosion inhibitor, it is no longer capable of displacing the copper ions due to its own limited ionization. The result will be a change in the amount of copper plated upon exposure to a neutral copper salt solution. In practice it has been found possible to measure the magnitude of the alteration of the cell potential by determining the copper deposited on a standard iron surface under controlled conditions. This means that, as the film builds up, the resulting shift in the electrode potential tends to cause the iron surface to act more like a noble metal, resulting in a reduction in the cell potential to a point that copper will no longer plate spontaneously. Obviously, the salts of other metals such as silver, gold or platinum might be used for this operation; however, copper was selected because of the color difference of the plate, the comparatively low cost of copper salts and the ease of evaluation of the plated copper. Furthermore, its high cell potential provides a substantial driving energy. The determination of the copper is readily done colorimetrically by dissolving it in 5 per cent ammonium hydroxide solution containing 2 per cent potassium per-sulfate. The blue color of the copper ammonic complex ion so developed may be compared with standards prepared by dissolving known quantities of copper powder in the same reagent. In routine screening it is seldom desirable to determine the amount of copper plated. It is usually more important to determine the concentration of inhibitor required to prevent the plating of copper.

By G. Major-Marothy

"INTRODUCTIONA beneficiation plant to upgrade earthy iron ores is being installed by the Iron Ore Company of Canada in Sept-Iles, Quebec.The crude ore for the plant will come from the Knob Lake area, located in Northern Quebec, where mining has been in progress since 1954.Some 170 million tons of ""direct 11 ores have been mined and shipped in the past eighteen years, with an overall silica content of 7 to 8 percent. No beneficiation - except the reduction of the top size to about 2 inches - took place during this period, but extensive research work has been carried out, in order to find a feasible process.Since “washing"", gravity separation, roasting and magnetic separation experiments did not indicate positive enough results, flotation was tried and found to be feasible.The beneficiation complex being installed will consist of grinding, flotation, dewatering - i.e.: thickening, filtering and drying - balling feed preparation, and agglomeration - i.e.: balling and induration.This process will result in a product superior to the ""direct shipping"" ores and will also expand the ore reserves by allowing material with a higher silica content to be utilized.THE OREThe Knob Lake ores are grouped by mineralogical composition and by grade.BLUE ORES, representing roughly two thirds of the reserves, carry their iron content mostly in the form of massive, blue hematite and martite.YELLOW ORES, about one sixth of the reserves, have predominantly hydrated iron oxides as the iron bearing minerals.RED ORES, the other sixth of the reserves, carry the iron mostly in the form of red hematite.The gangue in all ores is mainly quartz, but aluminum oxides and hydroxides are also present, particularly in the red ores."

Jan 1, 1972

By K. L. C. Gonçalves, V. K. Alves

"CVRD’s Department of Development and Technology – GETEK has been developing and perfecting a general methodology for scaling up flotation and grinding circuits over several years. This methodology has been applied in several prospective CVRD mining projects. It encompasses a procedure for variability analysis based on bench scale tests, pilot scale grinding and flotation tests and modeling and simulation for scaling up the industrial flowsheet.The bench scale tests include the determination of the Work Index, simplified DWT tests, rougher flotation and locked-cycle tests with samples that are representative of the ore body variability. After completion of these studies, continuous, pilot-scale flotation tests are carried out at a Mini Pilot Plant with drill core samples representative of different stages of the mine life. Finally, studies are carried out with large volume samples using a conventional grinding and flotation pilot plant. This sample is also used for liberation characterization, grinding kinetics, flotation kinetics and industrial circuit scale up. This methodology has been successfully employed in the development of CVRD’s new mining projectINTRODUCTIONCVRD has been developing and perfecting a general methodology for scaling up flotation and grinding circuits over several years. The definition of the processing route and the sizing and scale up of the processing system are some of the products generated during the pre-feasibility of a greenfield project. This paper is focused on the scale up of grinding and flotation circuits, operations known to be of the most importance in a mineral processing plant.The methodology for the definition of the flotation and grinding circuits is divided in the following stages: bench scale ore variability studies, pilot plant operation and modeling followed by simulation. The product of this work is a calibrated simulator that can be utilized for plant operation as well as mine planning, looking to achieve optimal operational conditions.The current models used for scale up of the grinding circuits are aimed at the generating information about the size distribution of the circuit product. However, it is well known that the liberation state of these products may be of critical importance for the performance of the concentration processes downstream. Nevertheless, CVRD has been developing a methodology for liberation kinetics during grinding, aimed at optimizing grinding circuit throughput together with downstream flotation recovery."

Jan 1, 2006

By Gordon R. Sima, John L. Bennett, Roger L. King

Electric shocks are a serious problem in the mining environment. Because of extremely wide ranging use of electrically powered equipment in the mining industry, the hazard to personnel due to electrical shock and related injuries are not restricted to any single area but cover the entire spectrum of mining activity. Electrical accidents, resulting in personal injury, due to im- proper or complete lack of grounding of electrically operated mining machines, form a significant percentage of the overall electrical accident picture. The object of this paper is to describe methods of detection of unsafe electrical conditions with respect to the grounding of electrically operated mining equipment. It also includes a discussion of prototype instruments which effectively monitor the condition of the grounding safety devices currently in use on mining equipment and warn of dangerous conditions as soon as they appear. The paper is based on work accomplished under USBM Contract Number 50357115. The contract was initiated under the Pittsburgh Mining and Safety Research Center Program. It was administered under the technical direction of PM and SRC with Mr. Roger L. King as the Technical Project Officer. Mr. Daniel B. Dawkins was the Contract Administrator for the Bureau of Mines.

The French research institute, INERIS, requested the assistance of the CSIR?s Division of Mining Technology in developing further precautionary measures against explosions in their collieries. In response to this approach, an on-board active suppression system was adapted to meet their mining requirements. This system is used on tunnelling machines in Germany where the machine is positioned in the middle of the tunnel, which is therefore not much wider than the machine itself. With this configuration, the machine is never away from the tunnel face during the cutting cycle. As a result, the possible volume in front of the machine of combustible gas requiring suppression is less than 50 m3. In South Africa this technology has been adapted to suit a two-cut mining method which entails greater distances from the face and thus much larger volumes of methane mixtures requiring suppression. The particular configuration of the Explo-Stop® system being investigated required the suppression of possible methane ignition volumes of up to 180 m3of a 9% per volume methane/air mixture. The volume is a consequence of the large cross-sectional area of French underground workings and of the two-cut mining method used in France. In order to take the second cut in the coal seam, the machine is pulled back a maximum distance of 4 m, leaving an air pocket in which gas can accumulate. The machine-mounted system is required to effectively suppress any flame ignited by such a large accumulated volume of methane. The results of the protocol tests done at Kloppersbos showed that the system was capable of suppressing all simulated methane ignitions for the prescribed conditions. This was achieved with a temperature increase of less than a 100°C at the operator?s position and without any flame being detected at this position.

Jan 1, 2001

By S. Al-Hassan

The deposit has been exploited for Mn oxide and carbonate since 1916. Known reserves of MnO, however, are almost depleted and the survival of the mine will depend on the exploitation of the carbonate. Kriging, inverse distance weighting (IDW) and cross-sectional methods have been applied to the evaluation of the carbonate resources and reconciliation of the estimates obtained shows that kriging and IDW give almost identical block values. A comparison using conditional probability indicates that there is virtually no difference between the global reserve estimates derived from the cross-sectional and kriging methods. The physical causes of these similarities are highlighted and the problems of dealing with multimodal distributions are discussed together with the difficulties of semi-variogram modelling

Jun 19, 1905

By F. W. Lee

This paper purposes to outline briefly results of seismic measurements made in the vicinity of blasts in a mine and in open-quarry operations. It endeavors to examine the amplitude and frequency of such vibrations as influenced by the loading, geology, and distance from the source and includes comments on factors that may index the destructiveness of a vibration and a method for measuring the stresses produced. There seems to be no simple mathematical analysis of the transmission of seismic energy through the ground, and experimental observations are at present most feasible. SEISMOLOGIC INSTRUMENTS Instruments developed by the Bureau of Mines for recording ground vibrations from quarry blasting comprise a new type of seismometer, oscillators, and a recording oscillograph. The seismometers respond to vibrations reaching them through the earth and transmit their characteristics by electric current to the oscillograph, where they are recorded by a beam of light on a moving strip of sensitized photographic paper. The seismometers are so constructed that their sensitivity to earth vibration can be controlled, thus keeping the amplitude of the recorded wave within bounds on the paper. Each instrument has been calibrated against vibrations of known intensity and frequency so that correction factors are known for the idiosyncracies of each unit. Moreover, each seismometer is calibrated in the field for its particular sensitivity setting at each blast. By these controls the actual movement of the earth due to vibrations of one ten-thousandth of an inch can be accurately measured.

Jan 1, 1936

By Peter Radziszewski

"Typical mining operating costs can be divided between extraction (30 % to 70 %), comminution (30 % to 50 %) and separation (5% to 20% ). Cornminition process costs, representing an important contribution to mining operating costs, can themselves be divided roughly between energy (50 %) and linerkharge wear (50 %). Developing the ability of predicting mill wear will allow a more accurate estimation of wear costs. However, total media wear in a given ball mill grinding process is a product of three recognized wear mechanisms those being abrasion, corrosion and impact. Further, the contribution to total media wear of each of these wear mechanisms has not been well established. In this paper, a overview of abrasive, corrosive and impact wear mechanisms is made from which an empirical wear prediction procedure is proposed. Further, future directions in wear prediction research is outlined. INTRODUCTIONTypical mining operating costs can be divided between extraction (30 % to 70 %), comminution (30 % to 50 %) and separation (5 % to 20 %). Comminition process costs, representing an important contribution to mining operating costs, can themselves be divided roughly between energy (50 %) and lined-charge wear (50 %). With respect to wear, world wide steel consumption as grinding media is estimated at over 600 000 tomes per annum (Malghan, 1982). As such, developing the ability of predicting mill wear will allow a more accurate estimation of wear costs.Charge media wear prediction is actually estimated using the Bond abrasion test. From Bond abrasion index tests, it is possible to determine a wear rate using the following function:"

Jan 1, 2000

By Nathan T. Lowe

This report updates an accounting of mines and prospects in the Seafoam mining district, Custer County, Idaho, begun during mineral appraisal of the Idaho Primitive Area. Thirty-two mine and prospect locations have been added to the Bureau's Mineral Industry Location Subsystem (MILS) data base as a result of initial phases of mineral appraisal of the River of No Return Wilderness; most previously known locations were field checked and the updated information added to the data base. The district's major deposits include polymetallic fissure veins of gold-silver-lead-zinc-copper hosted by granodiorite and quartz monzonite which have been intruded by dikes ranging in composition from rhyolite to dacite, and vein type replacements of silver-lead-zinc-copper hosted by calcareous xenoliths imbedded in major intrusives. These deposits may be related to the intrusion of the Rapid River pluton or the Casto pluton. The dominant northwest structural trend is parallel to an aeromagnetic high, a weak gravity salient, and the Blue Bunch Mountain dike swarm.

Jan 1, 1983

By P Paredes, T Leaño, D Gajardo, C Tramón, S Peñaloza

The definition of an ore handling system and its consequent layout design is of crucial importance in cave mining projects. It not only represents the main driver in the production capacity and reliability of the mining system, it is also a fundamental variable in the footprint’s development time and cost.In recent decades, load-haul-dump (LHD) tramming distance optimisation through the allocation of several orepasses inside the footprint has been the most common practice in ore handling system layout definition at Codelco’s operations. Today, the improvement of existing technology, such as higher-capacity LHDs and jaw-gyratory crushers, and intensive preconditioning techniques to improve fragmentation allow the conceptualisation of simple and highly productive ore handling system layouts for wide caving areas that avoid the complexity and high cost of developing numerous orepasses inside the footprint.The initial macroblocks of the Chuquicamata underground mine project (PMCHS) consider an ore handling system layout in which LHDs dump into orepasses located inside the footprint area. The ore is then transferred into crusher chambers located below the production level, where it is crushed by jaw crushers and then conveyed to surface. In an effort to reduce development time and costs, a new ore handling system layout for the PMCHS experimental macroblocks was engineered using improved technologies and techniques in a simpler mining system. This paper compares the new ore handling system layout for the project’s experimental macroblocks with the traditional LHD-orepass layout in terms of production capacity, reliability, development time and cost.CITATION:Paredes, P, Gajardo, D, Leaño, T, Peñaloza, S and Tramón, C, 2016. Simplifying the ore handling system through the use of current technology at Chuquicamata underground mine project, in Proceedings Seventh International Conference and Exhibition on Mass Mining (MassMin 2016), pp 571–584 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2016

By Andre Vervoort

Falls of ground, such as falls of roof and sidewall collapses, are a major cause of fatalities and injuries in South African collieries. Based on information obtained from the accident reports of the various Chief Inspectors of nines, a statistical analysis was conducted covering the period 1970 to 1988. This study shows that the volume of rock in a fatal fall of ground is often quite small. In intersections, fatal falls are on average larger than in roadways and at the face. It was also found that in 36 % of all fatal falls of ground no temporary or permanent support was installed at the site of the accident.

Jan 1, 1990

By Sean Harvey

One of the manifestations of the dynamically induced failures in deep coal mines is the violent ejection of coal from the ribs. A possible protection against such failures is the use of a specifically designed dynamic support system with reinforcement units having the sufficient capacity to absorb the kinetic energy of the dynamic failure and a skin support that can adequately contain the disintegrated coal. With this method of mitigation in mind, a recently introduced yielding rock bolt, trade name Roofex, has been identified (Harvey and Ozbay, 2009) as a possible energy absorbing component of a dynamic support system. According to the manufacturer?s specification, the Roofex bolts can yield 300 mm (or more) displacement while holding 80 kN pull load. In order to assess the performance of this support unit in coal ribs, a series of in situ pull tests performed in two coal mines in Colorado. The results show that, under conventional pull test loading conditions, Roofex yielding bolts perform very close to their claimed specifications, thus showing a potential for their use as the main energy absorbing component of a dynamic support system. Further studies are required to confirm the in situ performance of these bolts under dynamic loading conditions and also to develop a containment component for a dynamic support system for the mitigation of dynamic coal rib failures.

Jan 1, 2009

By Christopher J. Bise

INTRODUCTION Traditionally, the purpose of mine ventilation has been defined as diluting, rendering harmless, and sweeping away hazardous gases which may accumulate in the mine workings. In addition, the ventilation sys- tem must keep the mine air as close as possible to the composition of the outside air in order to provide a healthy environment for mineworkers and to prolong the useful life of mine equipment. When plans are being developed for opening a new mine, sound engineering dictates that a calculated estimate be made of both the maximum air quantity requirement and the accept- able mine pressure (head) loss related to the under- ground layout and projections. For efficient mine ventilation, the two problem areas that demand constant attention are (1) the control of fugitive air (through short circuiting or leakage) and (2) the provision of ample airway capacity. Fugitive air is that quantity of air which enters the mine but short-circuits the mine ventilation system through improperly constructed stoppings, doors, overcasts, etc.; it represents an investment in power consumption (to move the air by the fan) while providing little, if any, return on the investment. Typically, efficient mines are those that can deliver 50% of the air handled by their fans to the working faces. Paying close attention to a new mine's ventilation system at the outset can provide numerous benefits during the later stages of mine development. It has often been noted that the main cause of trouble in many mine ventilation systems can be found within a radius of 2500 ft from the fan installation. Ample airway capacity is essential to maintain the mine's head loss within manageable limits. Similar to an industrial ventilation setting, the mine's headings are the ductwork for the handling of air. The difference between normal industrial ventilation and mine ventilation is apparent when more capacity is needed; a plant can install new or larger ductwork, but it is difficult for a mine to install new entries or enlarge existing entries. Air velocities are usually maintained under 800 fpm to limit the stirring up of dust in the intakes. With this in mind, it is usually simple to estimate the number of entries required in the ventilation system. The problem that often arises, though, is when the system's capacity is overestimated and the required head loss to achieve the desired quantity is excessive. Experience has shown that problems are often encountered when the mine's condition requires much more than three inches of water gage to ventilate the mine on a day-to-day basis. Thus, if proper attention is not paid to these parameters in the preplanning and initial development of a new mine, it may take an extremely high water gage to achieve the ultimate quantity. This chapter will not delve deeply into the theory of mine ventilation; many other books have been writ- ten on that subject. Instead, the following sections will point out the important considerations that should take place before the development of a mine ventilation system, as they relate to normal industry practice. FUNDAMENTALS OF AIRFLOW Volume Versus Water Gage The common method of measuring ventilation pressure is in terms of inches of water gage (wg); one inch of water corresponds to a pressure of 5.2 psf (pounds per square foot). Further, airflow follows a square-law relationship between volumes and pressures where, as shown in the previous section, a doubling of quantity requires a quadrupling of pressure. Table 1 shows how heads and quantities are related. Using Table 1, it is readily apparent that, for the same mining conditions, an increase in pressure from one inch wg to three inches wg results in only 34,640 cfm when the initial quantity is 20,000 cfm. Pressure Losses Pressure losses in mine ventilation systems are caused by friction between the heading walls and the air stream and by shock losses due to abrupt changes of the air stream caused by obstructions. The common pressure-loss formula for mine ventilation is: [ ] where H is the pressure loss in inches of water gage, K is the coefficient of friction, L is the length of the airway in feet, 0 is the airway perimeter in feet, V is the air velocity in feet per minute, and A is the cross-sectional area of the airway in square feet. Table 2 lists the US Bureau of Mines schedule of friction factors (K) for mine airways. Recent research indicates that the values of K are slightly less for coal mines than those presented in the USBM Schedule, and Table 3 should be consulted in this instance. It should be

Jan 1, 1986

By E. G. King

As low-temperature heat-capacity and entropy data were available for only one sesquisulfide (cerium sesquisulfide), it appeared desirable to augment the information for this class of compounds by measuring the sulfides of antimony and indium. In addition a pressing need existed for data on indium sulfide to supplement vaporization studies in progress in the Department of Mineral Technology, University of California. This Bureau of Mines paper reports low-temperature heat-capacity measurements and entropy evaluations at 298.15° R. of antimony sesquisulfide of stoiehiometric composition (Sb2S3) and of indium sesquisulfide which was somewhat deficient in sulfur (In2S2.93). The data for the latter substance were adjusted to correspond to the stoichiome tric composition (In2S3 ) . Indium sesquisulfide has a lower entropy at 298.15' K, than either antimony sesquisulfide or cerium sesquisulfide, in line with the higher atomic density of the indium compound.

Jan 1, 1962

By E. D. Calvert

Molybdenum has been successfully cast into static and rotating molds in conventional equipment used for casting titanium and zirconium. Melting was accomplished by using an arc potential of 36 to 40 volts with a current of 10,000 amperes. Pouring efficiency, after a skull had been formed in the crucible, averaged 70 percent. Techniques are described by which four successful castings were made. Tensile properties on specimens cut from one centrifugally cast tube averaged: Tensile strength, 39,800 p.s.i.; yield strength (0.1 pct. offset), 35,300 p.s.i.; elongation, nil; reduction in area, nil; impact strength, 2 ft.-lb.; hardness DPH (diamond pyramid hardness) 180. Mold reaction was not evident in shapes cast in machined graphite. This was not true, however, for rammed graphitic-base molds.

Jan 1, 1960

By Baha Abulnaga

Slurry pipelines are used in a number of countries to convey processed ore mixed with water over long distances from remote mine sites to ports or plants for further processing. In Egypt and Sudan, numerous mineral resources are scattered in remote areas. Through the years, Egyptian geologists have identified 268 sites for metallic ores and 386 sites for industrial nonmetallic minerals. A large portion of these sites are in the Eastern Desert in an area called the Nubian Shield. Similar sites are found in the Sudan where mining contributes to less than 1 percent of the gross domestic product (GDP). There are three important components of infrastructure needed to develop these resources ? power, water and transportation. While Egypt is able to satisfy 70 percent of its needs in iron ore from the Gadideh Mine in the Western Desert, it needs to expand its other resources to reduce imports heavily sedimented. Dredging of these sediments and transporting them by slurry pipelines would enhance an industry of bricks or agricultural development based on tailings technology often used in mining. This article explores the potential for slurry technology for opening new mines in the desert or recuperating the sediments of the Nile. The technology by itself would be new to both countries and would benefit from collaboration, particularly to de-silt the Nasser-Nubia Lake.

Jan 3, 2010

Separation of Nickel and Cobalt from Manganese, Magnesium and Calcium by Synergistic Solvent Extraction - From Batch Tests to Pilot Plant Operation The synergistic solvent extraction system consisting of Versatic 10/LIX63/TBP was used for the separation and purification of nickel and cobalt from synthetic laterite leach solution after iron removal. Very encouraging results were obtained for separating nickel and cobalt from manganese, magnesium and calcium. In batch tests, the organic composition of the synergistic system was optimised and very large synergistic shifts for nickel and cobalt were observed. The ?pH50(Mn-Co) and ?pH50(Mn-Ni) values were as large as 1.65 and 2.40, respectively. The extraction and stripping kinetics of nickel and cobalt was fast. Semi continuous extraction tests were conducted with a pH profile of 5.5/6.1/6.5 in three stages at 40¦C. The nickel and cobalt extractions reached 99.9 per cent with only 5 mg/L Ni and <1 mg/L Co left in the raffinate. With two stages of scrubbing, only 2 mg/L Mn and less than 1 mg/L Mg and Ca were left in the scrubbed organic solution. With two stripping stages, the stripping efficiencies of Ni and Co reached over 95 per cent. A fully continuous pilot plant operation was conducted successfully for 280 hours. With an O/A ratio of about 2 and a pH profile of 5.5/5.8/6.0/6.3 for the four stages EX1/EX2/Ex3/EX4, both nickel and cobalt were almost completely extracted. Using a pH profile of 5.4/5.0 for SC1/SC2, the manganese scrubbing efficiency was over 96 per cent. Using three strip stages, over 98 per cent Ni and 99 per cent Co were stripped with only 64 mg/L Ni in the stripped organic solution. The nickel concentration in the loaded strip liquor reached 83 g/L, resulting in a ?Ni of over 28 g/L. The chemical stability of the system was also assessed over 12 weeks under proposed industrial operating conditions. Operation exclusively under strip conditions at 50¦C and 40¦C resulted in LIX63 oxime half lives of 1.1 and 2.1 years, respectively. Operation exclusively under extract conditions at 50¦C and 40¦C resulted in oxime half lives of 0.26 and 0.82 years, respectively. Losses of Versatic 10 were less than four per cent.

Sep 13, 2010

By Fredrick L. Johnson

The Bureau of Mines, under Congressional mandate, conducted a mineral survey on the Park Range WSA (Wilderness Study Area) in 1984 and 1985. The WSA, located about midway between Tonopah and Ely, encompasses 47,268 acres. No mineral resources were delineated, nor were workings found, during the investigation. There has been, however, exploration in and near the north end of the study area. ASARCO is actively exploring a block of 41 claims for hot-spring type, low-grade, large-tonnage gold deposits. Their claim block is almost entirely in the study area. The Morey mining district, 7 miles southwest of the study area, produced silver and lesser amounts of gold and lead sporadically from the mid-1860's through the mid-1960's. The WSA is underlain by Tertiary rhyolitic and latitic volcanic rocks and by Paleozoic sedimentary rocks. The Paleozoic rocks have been faulted, silicified, brecciated, and hydrothermally altered, and have anomalous concentrations of precious metals and pathfinder elements. Additional exploration in the north part of the Park Range WSA may lead to delineation of gold resources.

Jan 1, 1986

By Kent W. Hawley

An improved visibility system was developed for use on the large (100 ton+) haulage trucks typically operated in open pit and surface mines. The improved visibility system provides the truck driver' with a means of viewing hazardous situations that occur in the extensive blind areas adjacent to his true. The cost effective, reliable, visibility system consists of a closed-circuit television three unique fresnel viewers, and two improved mirrors. This report describes the visibility problems of Large trucks in the surface mining environment, shows the performance of existing mirrors, substantiates concept selection, and describes the prototype improved visibility system.

Jan 1, 1978

By Q. Cheng

This paper introduces the multi fractal interpolation method (MIM)developed for handling singularities in data analysis and for data interpolation. The MIM is a new moving average model for spatial mapping and interpolation. The model decomposes the raw data into two components: singular and nonsingular components. The former can be characterized by a localized singularity index that quantifies the scaling invariance property of measures from a multifractal point of view. The latter is a smooth component that can be estimated using ordinary kriging or other moving average models. The local singularity index characterizes the concave/convex properties of the neighbourhood values. The paper utilizes a binomial multiplicative cascade model todemonstrate the generation of one- and two-dimensional data with multi-scale singularities which can be modelled by asymmetrical multifractal distribution. It then introduces a generalized moving average mathematical model for analysing and interpolating data with singularities. Finally, it is demonstrated by a one-dimensional case study of de Wijs’ data from a profile in a zinc mine, that incorporation of spatial association and singularity can improve the interpolation result, especially for observed values with significant singularities.

Jan 1, 2015