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By Euler De Souza, Jason Chew, Aldrich Philip Dirige

Sillmats are structural elements upon which the safety and economics of underfill mining or sill pillar recovery in stopes of moderate width with steeply dipping ore zones, depend. The traditional stability design of cemented sillmats appears to have developed from experience and from standard static physical model tests such as uniaxial and triaxial compression. Such standard static physical model testing approach arrives at some approximate parameters, normally used to describe or predict the behaviour of sillmats in analytical or numerical models. This approach offers inherent limitations and generally results in conservative designs. Centrifuge modeling studies are combined with numerical modeling in this paper to provide insights into sillmat strength behaviour. The results of both modeling techniques suggest that, in stopes with smooth rock wall conditions, sillmat failure is driven by the fill self-weight and has minimum dependency on the fill binder content. In stopes with rough rock wall conditions, however, the wall roughness contributes significantly to the stability of the sillmat during underfill mining. This study indicates that the stability of paste fill sillmats can be designed with confidence using centrifuge physical modeling in combination with numerical modeling.

Jan 1, 2001

By Arafat Ghosheh, Yousuf Abu-Rish, Darwish Amara, Mohammed Baderkhan

The Jordan Phosphate Mines Company is one of the world's major phosphate rock and phosphatic fertilizers producers (the second largest phosphate rock exporter), with ore reserves exceeding 2 billion tonnes, mainly located in the Eshidiya area. The Eshidiya phosphate deposits, 120 km north-east of Aqaba, are of sedimentary origin (as are all phosphate ore in Jordan). Three exploitable beds (Al, A2, and A3) with variable content of P205, CaO, SiO2, etc., are processed in order to produce 68-70,73-75, and 75-77% TCP grades. The A3 bed, which is characterized as the most siliceous containing 40-48% TCP, requires physical beneficiation using wet scrubbing, screening, and desliming by hydrocylone, followed by anionic conventional flotation to achieve 75- 77% TCP grade. For the last 15 years extensive laboratory and pilot tests (including physical, chemical, and mineralogical studies) were carried out. After operating the industrial plant, certain modifications were deemed necessary to adjust the operational parameters and the final flowsheet, which was followed to obtain the targeted quantity and quality of concentrate.

Jan 1, 1999

By Michel L. Bilodeau

Mineral exploration is a sequential information gathering process. Primary exploration programs filter favorable areas through a series of geological, geochemical, and geophysical tests resulting in the discovery of mineral occurrences. The most favorable of these discoveries provides economic justification for a delineation drilling program. The purpose of delineation drilling is to provide information for estimating relevant physical characteristics of the mineral occurrence. These estimates are used to evaluate the exploration project and, eventually, the economics of mine development. Delineation drilling normally completes the exploration process. If the assessed economic characteristics of the delineated occurrence provide justification for mine development and production, an economic mineral deposit is the end result of the exploration process. Subsequently, a development drilling program is initiated to obtain the necessary data for detailed mine planning. The delineation drilling stage is in itself a sequential investment process. As delineation information accumulates, the reliability of geological estimates is improved and, thus, the uncertainty associated with the assessment of profitability criteria is reduced. Delineation drilling should continue as long as the marginal benefits, which result from improving the reliability of profitability criteria, exceed the marginal costs associated with the additional drilling.

Jan 1, 1979

By L. Alan Weakly

The purpose of this paper is to present a viable method of reinforcing rock pillars by utilizing fully grouted reinforcing steel and a cement - fly ash mortar. Although the fully bonded bolt concept is not new, as exemplified by the present widespread use of resin bolts within United States mines and by the use of cement grouted reinforcing used as back pins in the Kiruna Iron Ore Mine in northern Sweden, (1) (2) the method of stabilizing mine pillars with fully bonded reinforcing is new. This method, called the "grout-rebar system", was first incorporated at St. Joe Minerals Corporation’s Fletcher Mine in December of 1972. The system was used on an emergency basis before extensive research on the mechanics of the system could be completed. Since December of 1972, however, approximately 206 pillars have been reinforced at Fletcher and the system is currently being used on a regular basis at two other St. Joe mines and at two competitor mines in the area. Basically, this paper is a case study of the efforts, research and results of utilizing the grout-rebar system, on a selective basis, at the Fletcher Mine. In addition to information about the system, this paper will explain in detail the three rock mechanics instruments developed by St. Joe and used to measure and monitor rock movement to validate the grout- rebar system.

Jan 1, 1977

By Ron L. Schmiermund

The Ferris-Haggarty underground mine was abandoned circa 1910, and portal discharge has since resulted in degradation of a trout fishery in the receiving waters of Haggarty Creek. Although tunnel discharge varies dramatically in volume throughout the year due to infiltration of snowmelt, compositionally it is a nearly constant low-TDS, pH-neutral, low-sulfate, copper-silicate dominated water. Between infiltration events, components of ARD accumulate in unsaturated, oxygenated environments within upper workings and are delivered to the main haulage level by the subsequent annual snowmelt flush. Mixing with non-ARD mine waters results in neutralization, dilution, and precipitation of amorphous copper carbonate and silicate phases. These mechanisms, combined with variations in the water balance, control the composition of the portal discharge. Understanding the chemical and hydrologic control mechanisms was critical to the design of a passive treatment system to decrease effluent copper concentrations.

Jan 1, 2000

By Pasi Latva-Pukkila

Underground rock drilling has gone through intense development during the last 25 years. Mechanization and hydraulics have greatly increased the efficiency of drilling and improved the environment of the drill operator. As the total economy has become more and more important, the development has resulted in greater overall economy in drilling. The latest evolution step is computerization of drilling equipment. Computerization has been applied to tunnelling jumbos, long- hole drilling rigs and bolting rigs. With today's technology computerization is the natural way to increase capacity and productivity of the drilling process. However, it is still impractical to implement total "robotic" drilling. In computerized drilling the operator's role continues to be vital in controlling the drilling process and in analyzing the effect of individual parameters on the whole excavation process. Today this is called "CAD" - Computer Aided Drilling.

Jan 1, 1989

By R. A. Jankowski, R. Gong, R. Bhaskar

The control of respirable coal dust in continuous mining sections involves careful analyses of the generation and transport of the dust, including airflow pattern analyses. Potential dust control devices must be evaluated based on their effect on both local dust transport and potential dust exposure of face personnel. The development of larger and more productive continuous miners increases the need for research on dust control and improvement of visibility in the face area. Two studies relating to underboom dust control are discussed. The first study involved the valuation of potential underboom dust control techniques to reduce operator exposure to dust and determining the relationship between the positioning of exhaust ventilation and the exposure of face personnel as a function of face and operating variables. The study involved full-scale laboratory experiments using a state- of-the-art data acquisition system. The second study related to the use of foam under a continuous miner boom and its effect on silica concentration and respirable dust. The foam study was performed in a coal mine. The laboratory studies involved the fabrication of a full-scale test facility and development of a computer based data acquisition system. This was followed by baseline data collection, smoke trace tests, detailed performance testing of various underboom dust control measures and data analyses. The underground foam test involved the design and installation of the foam generation and delivery system, testing, and data evaluation. The results of the studies are summarized in this paper.

Jan 1, 1991

By G. J. Olson, K. S. Noah

A combined physical/biochemical process for the precombustion removal of 13 inorganic hazardous air pollutant (HAP) precursors, i.e., Sb, As, Be, Cd, Cr, Cl, Co, F, Pb, Hg, Mn, Ni and Se, from coal was tested. Biochemical processes for removal of HAP precursors from coal potentially offer advantages of deeper cleaning, more specificity and less coal loss. The slurry column is a second-generation process for the beneficiation o f fine (60-mesh x µm coal by a combination of physical separation of mineral matter and biooxidation of pyrite. Sixty-seven percent removal of pyrite from a 60-mesh Pittsburgh #8 coal was achieved at a 35% (w/w) slurry concentration and a five-day reactor residence time. Ninety percent of the heating value of the feed coal was recovered. Among the HAP precursors of most concern, over half of the Se, As and Hg were removed from the feed coal. From 40% to 70% of most HAP precursors were removed from the feed coal. Hg in the feed coal was reduced from 0.12 to 0.054 ,µg/g in the product coal, while waste coal contained 0.24 µg/g.

Jan 1, 2000

By Hans Skretting, Abraham A. Dor

This paper briefly reviews the historical development of the Rotary Kiln-Electric Furnace Process, which was first applied commercially to the treatment of qarnieritic ores in New Caledonia in the mid-fifties. It describes various technological improvements introduced, since, in industrial plants or developed in pilot facilities where a wide variety of nickel oxide ores has been treated success- fully over the years. Major categories of ores are characterized. The effect of their physical and chemical properties on their performance in the electric furnace is discussed. The relationships between ore composition, ferronickel grade and slag composition are analyzed and their effect on slag and metal tapping temperatures, nickel recovery, and electrode voltage are discussed.

Jan 1, 1979

By Vicki Seppala

Phelps Dodge Mining Company operates some of the world’s largest open pit mines. Safely mining in these deep pits requires the commitment to a proactive integrated slope management process. The slope management process also includes the continued look for improved slope design, construction, and maintenance of slope faces. This paper discusses integration of such technology as electronic programmable detonators, 3D imaging using ground-based LiDAR scans, and reflectorless slope monitoring systems in a slope management program.

Jan 1, 2005

By W. M. Husslage, M. E. Kock, T. Bakker, R. H. Heerema

Permeability in the blast furnace shaft is adversely affected by low-temperature reduction degradation of sinters. The fundamental cause of this is the expansion of the iron oxide phase resulting from the reduction of hematite to magnetite. The present study shows that expansion behavior during the hematite/magnetite reduction .step and the resulting microstructure are strongly influenced by the reduction temperature (partial oxygen pressure), the reducing agent and the nature of the hematite. The reduction rate is of minor importance. A reduction mechanism is proposed that can explain the dilation behavior and the magnetite microstructure. The results provide better insight into the reduction degradation behavior.

Jan 1, 2000

Stratabound uranium deposits consist generally of two comingled geologic and statistical populations. Barren zones, usually having radiometric assays too low for recording on digitized logs, comprise a large portion of a mineralized stratum. In between are mineralized zones; where the grades are distributed essentially lognormally. Estimation of open-pit ore reserves requires the prediction of the tonnages of selective mining units (the smallest mined volume which may be segregated as ore or waste) and their grades within panels of convenient size for open pit planning and economic analysis. The proportionate tonnage of mineralized selective mining units within each panel can be predicted by kriging an an indicator function assigned to each composite (1 if mineralized, 0 if barren). Next, using logarithms of mineralized assays, the conditional lognormal distribution of grades of mineralized selective mining units can be found. Using this distribution and the krigedestimate of the probability a selective mining unit is mineralized, the grade and tonnage of selective- mining units above any cutoff can be predicted. A case study illustrating this two- step procedure is presented for the Imouraren uranium deposit located in the Republic of Niger. Using cross validation, the estimation procedures were checked and were shown to be adequately accurate for mine planning purposes.

Jan 1, 1979

By Giin-Fa Fuh, Sally A. Shoop

An instrumentation program was developed based on the results of an analytical study of the geomechanical effects of longer longwall face lengths. Boundary element simulations showed that longer face lengths can cause in- creased face convergence and an increase in the shear stresses in the immediate roof. Empirical design methods predict that longer faces will cause little effect on abutment stresses as long as the roof caving is well behaved. On this basis, a longwall face instrumentation program was designed with emphasis on the interaction between face support and ground response for optimizing face conditions. At a 600-foot (183 m) face, five midface shields and two headgate shields were monitored with vary- ing hydraulic pump pressure. The instrumentation included six convergence rods with linear potentiometers to measure canopy-to-toe convergence; fourteen pressure transducers to monitor leg pressures; and two portable data loggers. Similar instrumentation will be in- stalled on a 750-foot (229 m) face for comparison. The results from the 600-foot (183 m) face indicate that a ground reaction curve, as determined from the relationship between the sup- port set pressure and the roof convergence, can be used to optimize the support pressure to control convergence, the key element to ground stability in the face area.

Jan 1, 1986

By Steven J. Page, Robert A. Jankowski, John A. Organiscak

Quartz dust is one of the most significant ongoing health concerns in coal mining today. Since initial verification of this health risk in coal mines by the Mine Safety and Health Administration (MSHA) during the early 1980’s, the Bureau of Mines has conducted numerous studies in underground mines to identify the sources of quartz dust and its fundamental characteristics. The most significant underground quartz dust source was found to be the continuous miner excavating rock. The percentage of quartz was also found to be higher in the smaller size fractions of the dust when significant amounts of rock was mined by the continuous miner. Furthermore, the quartz content in the continuous miner return remained fairly consistent at greater distances from the face. Another important factor speculated to affect quartz dust generation and its size characteristics is the morphology of the quartz bearing rock. The quartz content of dust between mines had notable differences with respect to the differences between kaolinite content in the dust. Low kaolinite content was associated with higher quartz content and increased size segregation of quartz into smaller dust particles, indicating that the free silica bonding matrix in the rock is most likely another factor responsible for quartz dust produced through mechanical comminution.

Jan 1, 1991

By Stephen G. Gemell

Gold mining has undergone a massive revival in Australia within the past decade. With very few exceptions, these operations are involved with the development of deposits which have been previously and partially worked by underground methods. In most cases, current operations have involved the open cut mining of near-surface remnants or low-grade hanging wall and footwall mineralisation. A growing number of mines are, however, progressing to the development of underground workings either as extensions of existing open cuts or as "stand alone" projects. These operations vary in scale from the Kalgoorlie operations of North Kalgurli Mines Limited and Kalgoorlie Mining Associates to the small scale "gouging" operations of individuals and partnerships. The individual production rates rarely exceed 200 000 tonnes per year, primarily because of the magnitude of reserves or the style of access available for rehabilitation. The narrow-vein mines were initially re-opened using the traditional methods, in particular shrinkage stoping and cut-and-fill stoping. As a result of experience gained in various ground conditions, certain variants have become popular. This paper investigates the fate of the traditional methods in Australian practice, and reviews some of the recently applied variants. This involves not only a perusal of current practices, but also an assessment of the influence of recent technological development on the economics of narrow-vein mining. Some consideration is also given to the effects of previously employed stoping techniques on efforts to access remnant blocks in the re-development of old mines.

Jan 1, 1989

By J. Bigu

Ventilation conditions in working environments are usually studied by anemometry or, in complex cases, b using special tracer gases such as SF6 and 85Kr. in conjunction with chromatographic and radioactivity techniques, respectively. In some working environments, such as underground uranium mines, however, use can be made of the natural radioactivity environment (e.g.. 222rn. 220rn, and their progeny) to investigate mine ventilation conditions of practical interest. Theoretical studies and experimental data in several underground uranium mines and mine locations show that certain thoron progeny to radon progeny concentration ratios are very good indicators of ventilation conditions. Furthermore, the growth and decay of radon progeny and thoron progeny provide valuable information on mine air residence and clearance times. Some specific examples have been analyzed, such as bacterially-assisted uranium leaching operations in a partially enclosed stope, and the effect on the underground environment of turning ventilation fans on and off for a period of time. Air residence times and other variables of interest, determined from the mine radioactivity environment, were compared with more traditional techniques, such as anemometry and tracer gas chromatography. These studies were also extended to a wet non-uranium mine which exhibited relatively high concentrations of radon and its progeny.

Jan 1, 1989

By Dae S. Young

The Disjunctive Kriging (DK) procedure was developed to estimate ore recovery figures within a deposit/ domain from limited diamond drill hole (DDH) core assay data taken at the exploration stage. Local recovery figures include minable ore tonnage and ore grades within relatively small areas such as stopes or small panels on a pit bench, for a given cut-off grade. The grade and tonnage should be computed based on the true selective mining units, that is, the average grade distribution of selective mining units within a panel. Local recovery figures are critical information for economic and engineering feasibility studies before mine development, as well as for short-term mine planning for operating mines. This kind of information can be provided only by non-linear geostatistical methods including DK and the Multivariate Gaussian Approach (MG) (Journel, 2, Parker, et al, 5).

Jan 1, 1982

By Colin J. Douch

New Zealand has been a significant producer of gold for over a century. Total recorded production is over 850 tonnes. There were three major periods of production: the first and biggest was between 1860 and 1880, the second was between 1895 and 1920 and the third was between 1930 and 1945. After 1945 gold production fell continually until 1980 when the major upward adjustment in gold price led to renewed interest. There have been important new developments in each of the known gold-rich areas: Westland, Otago and Coromandel and exploration for gold in these areas and in previously unprospected regions has reached unprecedented levels. In Westland over 140 gold processing plants are each producing between two and six ounces a day from low-grade placers. A major gold dredging operation will soon be producing over a tonne of gold annually and other new dredges are proposed. There are still vast reserves of gold-bearing gravels in Westland and prospecting is continuing to reveal new resources. In the Reefton area there is renewed interest in gold-bearing quartz reefs which were the source of over a million ounces of gold between 1900 and 1951. New large opencast reserves are now defined.

Jan 1, 1988

By S. K. Mishra, S. B. Kanungo, D. Biswal

Jigging of the -2.8 +0.7-mm size fraction of a friable, low-grade manganese ore sample from Andhra Pradesh increased the Mn content of a concentrate by 7% to 8%, with an average recovery of 57%. The jig concentrate was mixed with the unjigged -0.7-mm size fraction and ground to -150 µm. The sample was then subjected to wet high-intensity magnetic separation (WHIMS). The magnetic fraction thus obtained showed an increase in Mn content of 6%, but with an overall recovery of only 50%. The results of hydrocyclone classification of -150- µm ground ore indicated that the method is not capable of upgrading the manganese content, as both ferromanganese oxides and aluminosilicate gangue minerals constitute the fine size fraction. However, when the nonmagnetic fraction obtained from the WHIMS of -150- µm ore is subjected to hydrocyclone classification and the underflow obtained is subjected to WHIMS again, an overall recovery of about 50% to 52% is obtained with at least a 7% to 8% increase in the Mn content of the product. The phosphorus content in the product is not reduced by either of the two combinations of methods. The complex association of phosphorus with all the major phases in the ore has been clearly demonstrated from the linear correlation between the distribution ratios of major constituent phases and phosphorus.

Jan 1, 2000

By Joseph J. Wojcik

INTRODUCTION A great deal more is involved in a sampling effort directed at a placer deposit than physically capturing a quantity of representative material and extracting the desired component. The distribution and dimensions of the sample sites must be chosen after consideration has been given to the type of deposit and the concentrating mechanism. Interpretation of the sampling results is likewise an extremely important part of the sampling program. SAMPLE SPACING The interpretation begins with the selection and design of the sampling program. A decision regarding genesis of the deposits and their geometry must be made in order to select a suitable technique and sampling pattern. Deposits differing genetically will have differing distributions and degrees of concentration.

Jan 1, 1992