Search Documents

By Richard D. Begley

This paper summarizes the development of a mechanistic model to predict longwall mining induced surface subsidence. The model was originally developed from data acquired from a recent field study conducted in Northern West Virginia. Additional data has been collected from published sources and used to refine the model. The original model utilizes a negative exponential equation to predict the subsidence profile after the maximum possible subsidence is predicted based on the following input parameters: mine geometry, angle of draw and expansion factors for the deformation zones that have been identified above a longwall excavation. The model refinement involved the development of relationships between physical properties of the overburden strata and angle of draw in addition to the expansion or bulking factors. Therefore the refined model now requires only mine geometry and a typical geologic column from the site to predict the subsidence profile. This paper provides the methodology that enables the model to determine the angle of draw and expansion factors breed on physical properties of the site geology i. e. percent of strong rock and Rock Quality Designation of the overburden strata. In addition :comparisons of predicted results with measured data for over H) different sites with various geological and geometrical parameters located throughout the United States will be presented.

Jan 1, 1989

Any Australian who knows of the poverty of the early settlers in the first few years following the arrival of Captain Arthur Phillip with his "first fleet" at Botany Bay must feel proud of the progress that has been made in the 174 years that have elapsed since that momentous occasionGFor all its attractive appearance as reported by Captain Cook and Sir Joseph Banks, the shores of Botany Bay and of Port Jackson proved most ,inhospitable, and the accessible country surrounding them anything but suitable for sustaining the life of the small settlement. It is well known that during the first 15 or 20 years the hardy colonists and un-tfortunate convicts were close to starvation on several occasions and were only saved by the arrival of supply ships, which were dependent on the vagaries of wind and current, and the hazards of the uncharted seas they had to -navigate-.When Australia was little more than a miserable penal settlement, Great Britain was one of the great-powers of the world, and the U,.S.A. was an independent nation with a growing degree of prosperity that even then gave promise of its mighty awakening to become, as it has, a e.olossus of industry and commerce,-and the. leader in living standards of the world to-day.Australia, from such a shaky, and comparatively recent start, has risen to a position in which its living standards rank very high in comparison with those of most of the nations of Europe and America. Exact com-parisons are hard to fix because of the wide variety of factors involved .but the Institute of Pab?Lic Affairs .(Victoria) places Australia fifthin the world, excelled only by the U.S.A., Canada, Sweden and Switzerland. The next best after Australia are the United Kingdom, West Germany and France.How we achieved such a high standard of living is a vast subject for study but some of the economic aspects of our climb from abject poverty to a position of near affluence can readily be identified.

Jan 1, 1962

By J R. Robinson

Prior to mine development, exploration drilling within the Luise Caldera on Niolam Island encountered gold mineralisation to the northwest of the main Minifie and Lienetz orebodies. This mineralisation was referred to as the Kapit orebody, and ten diamond drill holes were subsequently completed over the area to investigate its economic potential. Several significant mineralised intercepts were returned. These were interpreted using the orebody model developed for Lihir at that time which invoked a strong horizontal control to mineralisation. Additional consideration was given to the elevated geothermal temperatures encountered, the proximity of the mineralisation to the ocean, and its perceived depth below sea level, resulting in the Kapit area being given a low priority for follow up investigation. As a result, no further work was completed over Kapit until recently. The Kapit orebody was the target of an intense diamond drilling programme during 2002/2003. Data from this latest round of drilling has significantly improved understanding of the mineralisation and associated geology. The significant orebody now delineated at Kapit substantially differs from past interpretations, and the latest data has upgraded its economic potential. A resource estimate and other critical studies completed in 2003 enabled an updated reserve statement to be issued in February 2004 including the Kapit mineralisation for the first time. Total Probable Reserves at Kapit stand at 25.2 million tonnes (Mt) at 5.28 grams of gold per tonne (g Au/t), for a total of 4.3 million ounces (Moz) of contained gold.

Jan 1, 2004

By E. J. Lintner

This paper is one of a series being prepared by the United States Bureau of Mines describing clay and shale mining and crushing methods and costs at various operations throughout the United States. These papers are designed to disseminate technical information regarding methods used. The cost tabulations represent local operating expenditures only and not total production costs. It is recognized that publication of total production costs may, in many instances, cause embarrassment to the individual operator as well as to the industry as a whole. On the other hand, it is essential to the technical discussion and study of methods employed to consider operating costs. The attention of the reader is specifically called to this differentiation in order that no misunderstanding of the scone of the cost tabulations shale ensue. The open-pit recovery of clay and shale is a comparatively simple operation in comparison with underground mining; nevertheless, it represents an appreciable factor in the cost of manufacture.

Jan 1, 1936

By V. T. Panayotov, M. I. Panayotova

Wastewater from mining and mineral processing activities is often polluted by anion forming metals, such as molybdenum and chromium. Low cost materials, including zeolites, are studied for removing heavy metals from wastewater. It is widely accepted that natural zeolites are not effective in removing metals in anionic form. The paper presents study on use of metal modified natural clinoptilolite to remove Mo (VI) and Cr (VI) ions from model soft wastewater. Modification with Pb, Fe(III), Zn and Ba increases clinoptilolite uptake capacity with respect to Mo(VI) at low pH values of the water subjected to treatment, the efficiency decreasing in the order Pb > Fe(III) > Zn > Ba. Only lead-modified clinoptilolite shows high uptake of Mo(VI) in acidic, neutral and slightly alkaline conditions. Modification with Pb, Ba and Fe(III) increases zeolite uptake capacity with respect to Cr(VI) from neutral water. The effectiveness of Ba-modified zeolite decreases in slightly acidic water. Over 90 % uptake of initially available Cr(VI) is achieved by lead-modified clinoptilolite both under neutral and slightly acidic conditions. Increasing the initial pollutant concentrations increases the removed amount both of Cr(VI) and Mo(VI) for Pb-modified clinoptilolite. The findings could be a basis for using the natural clinoptilolite to remove in consecutive steps: firstly Pb(II) and/or Fe(III) ions and then loaded zeolite to be used for Mo (VI) and/or Cr(VI) removal. The kinetics of Mo(VI) uptake by CL-Ba, CL-Fe(III), CL-Pb, and CL-Zn is best described by the pseudo-first order kinetic equation. The kinetics of Cr(VI) removal by CL-Ba, CL-Fe (II), CL-Fe (III) and CL-Zn is best described by the pseudo-first order kinetic equation. Kinetics of Cr(VI) removal by CL-Pb obeys the equation for the first order irreversible reactions. Finding way to remove anions-pollutants from the mineral processing wastewater will render this industry more environmentally friendly and socially acceptable.

Jan 1, 2016

By T. G. Hawkins, T. M. Naciuk

"The Katie Cu-Au deposit is situated in southeastern British Columbia, 425 km east of Vancouver. The deposit is a member of the alkaline group of porphyry Cu-Au deposits associated with potassically and propylitically altered andesitic volcanic rocks.The Katie porphyry system occurs in mafic to intermediate alkaline volcanic rocks of the Lower Jurassic Elise Formation of the Ross/and Group. Host metavolcanic rocks include andesitic to basaltic flows, flow breccia, massive fine tuff and crystal tuff Gabbroic intrusions alternate with the volcanic rocks in drill core. The lack of chilled margins and shear contacts, as well as their similar composition to the volcanic rocks, suggest that the gabbro was synvolcanic.Alteration exhibits a zonal pattern spatially related to zoned porphyry-style mineralization. Propylitic alteration is common and assemblages consist of epidote, chlorite, sericite and calcite. Potassic alteration zones are accompanied by elevated copper-gold grades and are composed of K-feldspar, plagioclase, quartz, biotite and chlorite. Narrow drillhole intervals of secondary magnetite, generally less than 2 m across, occur above the potassic alteration zones. A thick oxidized cap lies above the mineralized zones, although only local supergene enrichment has been identified. The lack of supergene enrichment is perhaps due to intense near-surface fracturing and leaching from a high energy groundwater environment.Sulphide mineralization consists mainly of pyrite within the propylitic halo, and pyrite and chalcopyrite with trace pyrrhotite, sphalerite, tetrahedrite, bornite and molybdenite within the potassic alteration core. Malachite, azurite and traces of chalcocite occur within the oxidized cap.Discovery and exploration techniques have included airborne geophysics, regional stream sediment sampling, prospecting, geological mapping, ground geophysical surveys, soil sampling surveys, and diamond drilling. Two primary targets of porphyry copper-gold mineralization have been identified by exploration. The Main zone strikes northwest, has an apparent true thickness ranging from 70 m to 135 m, and is a minimum of 500 m in length. The 17 zone strikes northwest, has an apparent true thickness of 90 m, and is a minimum of 300 m in length. The two zones are defined by grades that generally exceed 0.2% Cu and 0.25 g/t Au. Both zones are open along strike and to depth."

Jan 1, 1995

By B. Erdem, N. Çelebi

An important problem associated with tandem dragline systems is that of matching the productive capability of individual draglines in the combination. In this paper, a dependence mechanism among draglines in a tandem system is described, and possible consequential problems in a poorly-matched system are defined. The rate of linear advance and the cost of stripping are determined as two major indications for an efficient operation. The proposed decision criteria and the underlying mechanism to select a synchronous couple are introduced and used in simulating tandem dragline systems. The developed model is tested on a hypothetical strip mine where pullback stripping is applied with a tandem dragline system.

Jan 1, 1999

By F. Burat

The present research covers the results of experiments carried out with high frequency vibration screen on dewatering of coal fines. In the tests, the addition of surface active reagents was studied employing a high frequency vibrating screen. To increase the efficiency of mechanical dewatering processes, ultrasonic treatment was applied along with surfactants and optimum dewatering conditions were obtained. The moisture content of the feed material was from 82.5% to 41% by the optimization of operation parameters including the angle of the screen, vibration frequency, feed rate and screen opening size. Optimum parameters obtained in dewatering processes carried out with the vibrating screen were kept constant and the effects of surface active reagents and ultrasonic treatment on dewatering were investigated. Finally, the moisture level of the material was further reduced from 41.0% to 37.66%.

Sep 1, 2012

By J. C. Edwards

To have the capability to predict the development of localized spontaneous heating within a porous coalbed that is subjected to forced air ventilation or in an otherwise quiescent environment in which buoyancy develops, the Bureau of Mines developed three time-dependent mathematical models, which were used to calculate the temperature increase associated with chemisorption of oxygen by the coal. In each model, spontaneous heating is driven by an Arrhenius first-order reaction between the oxygen and coal. Two models are two-dimensional, and one is one-dimensional. In the first two-dimensional model, a constant-velocity forced convection airflow is specified; and in the other, buoyant flow is allowed to develop in the absence of forced convection. The third model evaluates the airflow from Darcy's law and a specification of the pressure at the surface of a one-dimensional porous coalbed. Numerical computations demonstrated how each model could be used to predict the onset of spontaneous heating when the porous coalbed was subjected to constraints of an imposed internal heat source or a high temperature airflow. The effects of particle size and coalbed compaction upon spontaneous heating were examined with the third model.

Jan 1, 1990

By J. M. Newell, G. R. Peatfield

"The Red-Chris porphyry Cu-Au deposit is located between Ealue and Kluea lakes, southeast of the village of Iskut on the Stewart Cassiar highway in northwestern British Columbia. It was explored by Texasgulf Inc., in a joint venture with Great Plains Development Company of Canada, Ltd. and Silver Standard Mines Ltd., during the period 1973 to 1980. Exploration resumed in 1994, following acquisition of the interests originally held by Texasgulf and Great Plains Development by American Bullion Minerals Ltd.Prospectors were first drawn to the area by prominent limonitic stain zones exposed in creek gullies in the western part of the property. These zones are developed on the alteration halo at the western periphery of the deposit. The potential ore zones do not crop out and were first exposed in a bulldozer trench excavated in 1971. Subsequent exploration was by percussion and diamond drilling, guided to some extent by induced polarization surveys. In 1976, estimated resources contained within two crudely designed open pits were 41 million tonnes with an average grade of 0.56% Cu and 0.34 g/t Au, with a cutoff grade of0.25% Cu. New resource estimates based on a major drilling program completed by American Bullion in 1994 were released as this paper goes to press. These estimates were calculated using various copper cutoff grades; at the 0.4% Cu cutoff, the ""in-situ geological reserve"" is given as 100 million tonnes grading 0.58% Cu and 0.46 g! t Au (American Bullion Minerals Ltd., 1995).The deposit consists mainly of pyrite, chalcopyrite and increasing bornite at depth, in zones of quartz-vein stockwork hosted by an elongate body of altered monzonite porphyry of latest Triassic or earliest Jurassic age. The stock intruded volcanic and sedimentary rocks of probable Stuhini Group affinity and is in faulted contact with sedimentary rocks of the Jurassic Bowser Lake Group. Gypsum stockworks are developed in the higher levels of the subvolcanic porphyry system. Late-stage quartz-carbonate veins carry small amounts of galena and sphalerite. Complex faulting within the deposit makes drill-hole correlations difficult. Alteration assemblages include an early potassic phase, almost completely destroyed by later phyllic alteration, and surrounded by a propylitic assemblage dominated by ankeritic carbonates in altered sedimentary rocks.The deposit has characteristics of bath the alkalic and calc-alkalic classes of porphyry copper deposits. The development of quartzvein stockwork suggests a calc-alkalic affinity, but the hostrocks contain little or no free quartz, molybdenum is present in trace amounts only and gold values are significant. Copper grades vary between wide extremes with unusually high grades in some sections of the deposit. The weight of evidence available in the 1970s led us to place the Red-Chris deposit firmly in the alkalic family. The quartz-sericite alteration assemblage was attributed to breakdown of potassium feldspar in a near-surface alteration environment in which seawater may have had a significant role. These interpretalions are retained here, but preliminary conclusions from the most recent work suggest that they may require some modification."

Jan 1, 1995

By K. Lombard

"‘Roof support awaiting time’ (RSAT) is a term used at Goedehoop Colliery’s Simunye Shaft to describe the potential production time lost due to the continuous miner (CM) standing idle waiting for roof support to catch up. Investigations revealed that in 2013, Simunye Shaft had approximately1400 hours of RSAT, which suggests that the mine could have potentially produced an additional 280 000 t of coal. This project consisted of two parts. Firstly, the causes of the high RSAT and means to improve the situation were investigated. Secondly, as insisted by mine management, the CM cutting sequence was investigated as a possible cause of high RSAT. Machine-related challenges due to the roofbolter installing support too slowly, geological conditions (mostly hard roof conditions and slips),logistical challenges pertaining to the CM cutting sequence, man-related challenges related to operator fatigue, re-support, operator inexperience, and the absence of support targets were identified as main contributors to RSAT. Furthermore, results showed that the roofbolters in the sections at Simunye Shaft are slower than the CMs. A target of 28% reduction in RSAT was set. Experts from Kennametal and Fletcher were consulted to find solutions for the identified causes. In total, eight solutions for RSAT were identified, but the solution that contributed most significantly to reducing RSAT was to use hard roof drill bits as a standard product at Simunye Shaft. Calculations showed that by using hard roof drill bits, RSAT can be reduced by 43%, which is more than the specified 28% target. The cutting sequences at Kriel, Greenside, and Simunye Shaft, togetherwith three newly developed cutting sequences, were simulated using the UCMS (Underground Coal Mining Simulation) program. A re-aligningprinciple was incorporated into the newly developed cutting sequences toalign the cutting sequences to general support work and to reduce RSAT. A decision matrix revealed that a cutting sequence in which boxing takesplace in R3 (third road to the right of the belt road) and in which the re-aligning principle has been incorporated will be the best option for Simunye Shaft. The recommended cutting sequence will lead to a 5% increase in production.The work presented in this paper was carried out as partial fullfilment for the degree BEng (Mining Engineering)"

Jan 1, 2015

By A. J. Schindler, F-P Sawyer, T. Logan, M. L. Shutey-McCann, R. M. Perry

"Test work on biooxidation pretreatment of low-grade refractory gold ores has been ongoing at Newmont Gold Company since 1988, including laboratory testing and pilot biooxidation heaps (360 to 23,000 tonne). Metallurgical results of the biooxidation process utilizing Thiobacillus ferrooxidans (T. ferrooxiddans) have culminated in the design, construction, and operation of a biooxidation demonstration facility. The facility has processed 708,000 tonnes of refractory ore, using a batched process with separate pads for biooxidation and ammonium thiosulfate gold extraction. Siliceous sulfidic refractory (SSR) and carbonaceous sulfidic refractory (CSR) ores are treated separately during biooxidation pretreatment unit operation. Following biooxidation, the ores are neutralization with lime and placed on oxide leach pads or stacked on the ammonium thiosulfate leach pad. Gold is leached with either cyanide for SSR ores or ammonium thiosulfate for preg-robbing CSR ores.This paper will only address the biooxidation pretreatment process of the SSR and CSR ores.AcknowledgmentsThe authors appreciate the cooperation and approval of Newmont Gold Company Carlin Operations management for its support and permission to publish this paper. The paper draws on efforts and assistance of the following people and several others which are to many to list who participated over the past five years in bringing this project to its fruition: Dr. James A. Brierley, Chief Microbiologist, Newmont Metallurgical Services; Dr. Rong-Yu Wan, Metallurgical Manager, Newmont Metallurgical Services."

Jan 1, 1997

By Thomas S. Jones

This report discusses processing and use of the principal forms of manganese. The focus is on trends in the United States in the 20th century and specifics of domestic manganese material flows as of about 1990. To protect proprietary data, quantitative assessments are limited to the iron and steel industry, which accounts for about 90% of consumption. Nonmetallurgical uses are considered qualitatively. Approximate disposition of a U.S. cumulative consumption during 1901-90 of an estimated 69 million st of manganese is estimated to be iron and steel slag, 69%; nonmetallurgical uses of a dissipative na¬ture, 10%; iron and steel scrap, 9%; National Defense Stockpile, 3%; unidentified and/or including items in service, 10%. In slag from iron blast furnaces and steelmaking, manganese is present as a minor, relatively im¬mobile component, which so far has been uneconomical to recover. Construction and road building are main applications for these slags. Steel slag contains the greater amount of manganese. Studies by others of global manganese fluxes and the effect of human activities on them also are discussed. One source concludes that the main effect of human activities, via deforestation and agricultural activities, has been to increase the rate of transport of manganese to the oceans.

Jan 1, 1994

During fiscal year 1966 the Bureau of Mines published 99 scientific and technical papers and engaged in 67 research projects on petroleum, natural gas, and oil shale. A basic appropriation of $3.4 million was supplemented by $1 million in cooperative funds from other Governmental agencies and industry groups. Petroleum conservation was advanced by development of scientific knowledge and improvement of technologic methods in both the extraction and utilization of petroleum, natural gas, and oil shale. Experimental work was carried out at research centers in Bartlesville, Okla., and Laramie, Wo., and in laboratories in Morgantown, W. Va., and San Francisco, Calif. ACHIEVEMENTS IN FISCAL YEAR 1966 Bureau of Mines--Petroleum Research in fiscal year 1966 (July 1, 1965- June 30, 1966) was involved in many public issues--air pollution, peaceful uses of nuclear explosives, utilization of oil-shale resources, computerization of recovery-stimulation techniques, and fuels for supersonic airplanes. Scientific contributions included compositional studies of petroleum and oil shale, determination of thermodynamic properties of petroleum compounds, development of analytical techniques, and correlations of petroleum properties with geologic occurrence. Technologic advances included rock-fracturing research, in situ combustion studies, waterflood techniques, use of radio-tracers to study stability of petroleum fuels, and hydrogenation as a tool to prepare shale-oil products.

Jan 1, 1968

By M. P. A. Hames, D. R. Beaumont

INTRODUCTION Legend has it that when Isembard Kingdom Brunel was confronted with the cost and consequences of one of his bridges having failed, he responded cheerfully that the event had saved investors and taxpayers a fortune because it prevented him from repeating the same design at several other locations. To meet the demands of rejuvenating and expanding the transportation system in an era of industrial growth Brunel and his contemporaries introduced bold designs and many innovations. Great advances in design and construction techniques resulted, yet these too had their practical limits as the story tells. In the gold mining industry today we face the same challenge of expansion, yet designers must respond mindful of economic restraints. Is innovation the answer? It has two meanings; to bring in novelties and to make changes. Distinguishing between improvement and novelties, or simply recognizing the benefits and limitations of new ideas over proven technology is part of good design. This paper will illustrate the practical aspects of selecting appropriate solutions to a design problem by examining a variety of examples common throughout the industry. Since a variety of engineering disciplines contribute to the design of a gold processing plant, and each has some old wisdom to impart or new development to boast this paper is divided into the different areas of expertise involved.

Jan 1, 1987

By Peter N. Loncar

Most miners dream of finding a gold mine sometime during their mining careers, but not very often do such dreams come true. On the tenth day of September, 1962, our dream of finding gold came true when the results of the third drill hole were received. This was undoubtedly one of the most exciting days of our lives. Little did we realize then that this was the beginning of a major discovery destined to became one of the largest gold mines in the country. Discovering and mining gold is not entirely new, however, to Eureka. County a old Lynn Mining District in northeastern Nevada. Since 1900 there has been sporadic activity, mostly consisting of small lode mines and placers. These have produced totally about $300,000; and existence of our orebody would probably have remained unknown had it not been for the United States Geological Survey. The USGS extensively mapped the Roberts Mountain Thrust and published a paper indicating that mineral occurred along this fault and in some places was gold bearing. Utilizing this information, our Newmont Exploration geologists located an area in which: drills were brought in to test their theories. By keeping flexible in our thinning and working, we were able to bring a raw prospect to a producing mine with ore reserves of eleven million tons in a short three-year period.

Jan 1, 1965

By F. D. Everett

This report summarizes the results of a nationwide investigation completed in 1964 to determine the nonferrous base metal ore reserves and potential sources of ore and to evaluate criteria for their ultimate exploitation. Engineers of the Bureau of Mines contacted mining company officials, reviewed reports and Government records, and where necessary made personal appraisals in arriving at the results. The estimates are presented by metals in ores in geographic regions and include some of the problems affecting the reserves and potential sources. The recoverable metal in the measured ore category and in the total ore reserves was estimated as follows: [ ] At the 1963 rate of domestic mine production these total reserves would be adequate to provide copper for 62 years, lead for 121 years, and zinc for 42 years. Data also are given for gold, silver, and other coproducts. Reserve estimates greatly exceed those of previous investigations; some prior estimates of copper reserves are included for comparison (table 23). Projections of requirements for base metals indicated the United States will need approximately 45 percent more copper in 1970 and 110 percent more by 1980 than was consumed in 1960; similar projections for lead indicate increases of 20 percent by 1970 and 45 percent by 1980; zinc projections are 40 percent more by 1970 and 90 percent by 1980. Much of the production of gold and silver is a coproduct of base metals recovery and the projected need for the precious metals exceeds all the United States can produce. Some problems of the industry are mentioned. A graphic evaluation of the economic availability of the reserve and potential sources is shown.

Jan 1, 1967

By Elwin L. Fisk

The Cordero mine is located in the Nevada portion of the Opalite Mining district that straddles the Oregon-Nevada state boundary. The mercury deposits of the district occur along the margins of the broad east-west graben valley of McDermitt Creek. The only other major producers, the Opalite and Bretz mines, on the northern margin of the graben have thus far exploited only shallow deposits within 100 feet of the surface. Through December 1965, Cordero has produced 94,680 flasks of mercury, the most of which came from the large underground cinnabar deposit. Nearly all of the mercury ore deposits in this district have been found in Miocene rhyolitic intrusives, flows, tuffs, and lake beds. These rhyolites have been argillically altered at depth and extensively silicified at or near the surface. These silicified areas or "opalite" masses were the original cinnabar-bearing outcrops.

Jan 1, 1968

By C. Birch

"A mineral resource royalty is payment to the holder of mineral rights for the utilization of the mineral resource. In South Africa, this payment is made to the State as holder of the mineral rights (The Mineral and Petroleum Resources Royalty Act of 2008). The principle purpose of this research paper is to identify if the State is benefitting from the mineral resource royalty by considering its impact on seven individual Witwatersrand gold mines.For this study, a simple financial optimiser model was created in Microsoft Excel that links the ore flow, block listing and the cash flow (excluding or including the cost of the mineral resource royalty). Mixed integer linear programing (the Excel Solver function) is utilised to optimise either profit or NPV (at 9% and 12%) by adjusting the cut-off grade.The impact on each of the seven mines mine was different but overall R7.9 billion is estimated to be paid in mineral resource royalty over their expected remaining lives. Due to the cost of the mineral resource royalty and increasing the cut-off grades, the total revenue decreases by R10 billion. A significant portion of this lost revenue would have been paid to the State in the forms of other taxation including company income tax which decreases by R2.8 billion. It is recommended that an industry wide investigation be conducted to determine if the resource royalty is adding to the State’s revenue, or destroying value including premature job losses. IntroductionTaxation of mining companies has become a very topical subject in the context of the postapartheid South African economy. There have been increasing calls for equitable re-distribution of mineral asset wealth for the benefit of all South African citizens and not just to an elite few. As part of the Mineral and Petroleum Resources Development Act (MRDPA) of 2002, the Mineral and Petroleum Resources Royalty Act of 2008 was introduced and came into effect in 2010 (South African Revenue Services, 2008). A mineral royalty is payment to the holder of mineral rights for the utilization of the mineral resource. In the case where the holder of the mineral right is the State, then this payment is made to the State (Cawood, 1999). Studies on the potential impact of the mineral resource royalty on the minerals industry have been conducted during the formulation of the Act and subesquently. These include Cawood and Macfarlane (2003), who suggested a cap of no more than 3% (Cawood and Macfarlane, 2003), as well as Cawood (2011), who predicted rising cut-off grades and thus reducing mineral reserves. In 2013, Grobler completed a Masters in the Faculty of Economic and Management Sciences at the University of Pretoria and concluded that the impact of the mineral resource royalty on the economy is positive. She stated that the royalty has caused operating expenses to increase, but this does not appear to be significant. Furthermore, she stated that the royalty has not affected the way the mining industry operates (Grobler, 2014)."

Jan 1, 2016

By Robert E. Wallhagen

The froth foam seal topping system is part of a system for remote sealing of mine passages using flyash transported pnuematically from the surface. Making the final 5 percent of the seal with a foamed-in-place fire retardant polyurethane reduces the required flyash volume by 50 percent and provides a more reliable seal. Foam components are held at the surface under proper conditions of temperature and pressure; then conveyed down the borehole through individual hoses and mixed at the bottom with a static mixer. The polyol component is pre-mixed with fluorocarbon-12 which, when released from system pressure in the mixer, boils and causes the foam to expand. Difficulties in handling this particular foam formulation are offset by a flame spread index well below the 25 level by ASTM E-162 normally required for under-ground applications. Isocyanate, polyol (TDI) and catalyst are conveyed to the downhole package by three separate lines. Five additional lines operate the valving and convey solvent so that the mixer and package lines can be flushed immediately after foaming. The eight-hose bundle is provided in segments of several lengths so that bore-hole depths can be matched. Cast-in-place plastic cylinders hold the bundle together and transfer the weight of the hose through a special quick-release clamp to the crane's wire rope. The downhole package and hose bundle are deployed using a specially designed winch and crane system capable of borehole deployment to 1500 feet.

Jan 1, 1977