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By G. H. Lewis, C. L. Krcil, N. S. Harding, D. E. Havekotte

INTRODUCTION Limestone injection work was completed in a 440-kW (1.5-million Btu per hour) pilot-scale combustor in preparation for a 50,000-kg/h (110,000-lb per hour) industrial steam boiler limestone injection test program. The main objectives were to: (1) identify the important variables that affect sulfur capture, (2) determine the maximum SO2 removal, and (3) determine the effect of limestone addition on furnace operability and heat transfer. A more detailed report is given by Lewis et al. (1985). The test program was divided into two phases. The objectives of Phase I were to determine the important variables that affect SO2 reduction and to investigate the effect of limestone addition on combustor operation and heat transfer surfaces. The objective of Phase II was to determine, using only the important variables identified during Phase I, the conditions that resulted in the greatest sulfur dioxide reduction. Phase I Test program Using one limestone, five variables were identified as having an effect on SO2 reduction. These are noted with their respective ranges: heat release rate [103 to 166 kW/m3 (10,000 to 16,000 Btu per hour-cu ft)], Ca/S molar ratio (0 to 3), flue gas oxygen concentration (3% to 5%), additive particle size [90% -150 µm (-100 mesh) and 90% -45 µm (-325 mesh)], and burner stoichiometry (SR = 0.75 and >1.15). A 2.3% sulfur northern West Virginia coal, pulverized to a nominal grind of 70% -75 µm (-200 mesh), was burned for the tests. The limestone used was obtained from a Maysville, KY, quarry and contained 51.75% (wt) CaO. An 11-run statistically designed program (plus three duplicates) was used to screen the five variables. The effect of limestone injection on the combustor heat transfer surfaces was determined by comparing the heat transfer rates through a slag panel and fouling probes during periods with and without limestone injection. Combustor operation was determined by noting any operating problems that occurred during the tests. Test procedure After furnace wall temperatures had stabilized at operating conditions, heat transfer surfaces (slag panel and fouling probes) and the convection section were blown clean, and a four-hour test was begun. The

Jan 1, 1987

Papua New Guinea has a long history of mining, with gold being produced from deposits on islands in the south east of the country over a hundred years ago. Before the Second World War, the mining and dredging of gold made the Wau-Bulolo area a

Jan 1, 1994

By Y. Sagawa

The Hishikari Gold Mine consists of epithermal vein type Au-Ag deposits, each of which has several veins standing almost in parallel each other and inclining 70 to 90 degrees. The veins have 1 to 3m widths and are extracted mainly by drifting and bench stoping with waste rock fill. The Hishikari has tried to optimise the stope dimensions by the Potvin’s Stability Graph Method, which is well known as an easy and effective tool to design appropriate stope dimensions. The method, however, has some limitations of application due to its simplicity so that it has been used only in limited cases. The authors developed a new numerical approach, based on 2D FEM solu-tion for the stability analysis. As a result, it has been possible to solve the Hishikari situation. The authors also introduce the concept of the design approach and compare the results obtained by us-ing the approach with the database of the Stability Graph Method for validation. La mine d’or Hishikari se compose de gisements d'Au-AG avec des types de filons épither-maux dont chacun a plusieurs veines qui sont presque parallèle les unes par rapport aux autres et sont inclinées de 70 à 90 degrés. Les veines ont de 1 à 3 m de largueur et elles sont principalement extraites en creusant et en abattant des bancs de roches encaissantes. Hishikari a essayé d’optimiser les dimensions du chantier d’exploitation avec la méthode de graphique de stabilité de Potvin, bien connue comme un outil facile et efficace pour avoir des dimensions de chantier d’exploitation appropriées. Toutefois, la méthode a certaines limites d’application dues à sa sim-plicité de sorte qu’elle a été utilisée seulement dans des cas limités. Les auteurs ont développé une nouvelle approche numérique, basée sur la solution 2D FEM, pour l’analyse de stabilité. Par conséquent, il a été possible d'expliquer la situation de Hishikari. Les auteurs présentent le concept de l'approche et ils comparent les résultats donnés en l'employant avec la base de données de la méthode de graphique de stabilité pour la validation. Die Hishikari-Goldmine besteht aus Au-Ag-Ablagerungen mit epithermalen Adern, die jeweils mehrere Adern aufweisen, die mit einer Neigung zwischen 70 und 90 Grad nahezu parallel zueinander verlaufen. Die Adern haben eine Breite von 1 bis 3 Metern und werden überwiegend durch Vortrieb von Stollen und Bergversatz abgebaut. Hishikari hat versucht, die Abmessungen des Abbauorts durch das Potvins Stabilitätskurvenverfahren zu optimieren, das als einfaches und effizientes Mittel für die Erzielung angemessener Abmessungen des Abbauorts bekannt ist. Das Verfahren weist jedoch aufgrund seiner Einfachheit einige Anwendungsbeschränkungen auf, so dass es nur in einigen Fällen benutzt worden ist. Vor diesem Hintergrund haben die Autoren einen neuen numerischen Ansatz entwickelt, der hinsichtlich der Stabilitätsanalyse auf der 2D-FEM-Lösung beruht. Dadurch wurde es möglich, das Problem von Hishikari zu lösen. In diesem Dokument stellen die Autoren die Grundlagen des Ansatzes vor und vergleichen die erzielten Resultate durch Benutzung der Datenbank des Stabilitätskurvenverfahrens zur Validierung.

Jan 1, 2003

By D. Newman

"The U.S. stone industry is in a transition from surface quarries to single and multiple level underground stone mines. The factors driving this transition include surface quarries; ? reaching their economic stripping ratio, ? encroaching on the mineral reserve boundary where the remaining reserves are below the active pit, ? dust, noise, and blast vibration concerns as suburban neighborhoods and development move into formerly rural areas. Ground control is critical to the success of a safe and productive, mining operation. The primary focus of the ground control engineer in an underground stone mine is the stability of the roof or back, development and benched pillar stability, stability of the sill pillar between vertically adjacent mining levels, and floor stability. Roof stability begins with the selection of a stable back horizon based upon the presence of a thin shale parting, stylolite, or other discontinuities from which the production shot can easily ""peel"" away leaving a stable back stratum. Room width and the decision of whether to bolt the roof, the bolt length, bolt type, bolt diameter, and spacing of the support is controlled by the thickness, fracture spacing, jointing, and material properties of the back strata. Horizontal stress controls the header and crosscut azimuth as to avoid driving perpendicular to the maximum horizontal stress direction. Stable pillar dimensions depend upon the interaction of the; ? vertical stress resulting from overburden thickness and multiple level interaction of overlying level(s), ? mining height during development and the final mining height after floor shots are taken while benching, ? presence of faults, folds, or joints passing through the pillar, and ? stone lithology recognizing that the presence of oolitic bands, shale partings, and bentonite beds create zones of weakness that can destabilize an otherwise competent pillar. Floor stability is typically not a significant issue for underground stone mines because the bearing capacity of the stone exceeds the vertical pressure transferred through the pillar. However, floor heave can be a ground control issue where shale partings, moisture sensitive strata, weak stone bands, or horizontal stress is present. Roof stability and pillar stability problems can be initiated as the floor heave enables the pillar and roof to settle. The ground control issues for underground stone mines are most clearly presented through a series of case histories. The case histories presented in this paper are a mix of ground failures and successful mining. The focus is on the geotechnical data required for underground stone mine ground control for roof, pillar, and floor design and stability analysis."

Jan 1, 2016

Novel techniques were developed to obtain samples from and make measurements in the shaft of the flash furnace at the Kalgoorlie Nickel Smelter. The data obtained was used to validate mathematical models of the burners and the reaction shaft. Particle, gas and wall temperatures were measured along with the oxygen potential of the molten particles. Collected particles were characterised by sizing, chemical analysis and optical microscopy. The region of the shaft immediately beneath the burner where the rate of change was most rapid received special attention.

Jan 1, 1993

Water conservation and compliance strategy can be cause for success or failure of a new or existing mining operation. Water accounting at outset, and pit and waste dump planning, may not fully consider total inflows, or sediment loads encountered during mine life. Environmental impact assessments, and permitting often dictates the rules, and these may change, to be more stringent, during course of mine development. This paper will identify techniques to control unprojected fresh water, or increased surface water flows in surface mine operations.

Feb 23, 2014

By V. S. Swaminathan

This year, Great Britain is looking back over a century to August 14, 1850, the day when the first "Act for the Inspection of Coal Mines" was passed in that country, an act which signaled the end of over two hundred years of hitherto uncontrolled and dangerous mining practices.

Jan 11, 1950

By George S. Koch

We have devised a multivariate statistical model useful to interpret data from exploration geochemistry and also pertinent for the analysis of data from petrology, paleontology, and other geological fields. The model is applied to several simulated cases in order to compare anomaly recognition by inspection and by quadratic regression.

Jan 1, 1977

By J. H. Wernick, E. E. Thomas

THE etch-pit technique for revealing dislocations in bulk samples allows one to determine dislocation densities and to study their behavior under various mechanical and thermal treatments. We have developed a chemical polish-etch for the development of etch pits on the (1100) plane of Cd which appear to

Jan 1, 1961

The present stage of economic development of the Broken Hill lode, attained after 40 years of active exploitation, leaves little opportunity for the study of the zone of secondary sulphides.Little remains of the oxidised zone, and the observations of the secondary sulphides of copper and silver can rarely be made on specimens collected in situ. A few specimens containing secondary...

Jan 1, 1927

By J. L. Arola

Since start-up of fluxed pellet production in August: 1987, the Minorca Mine has made a number of process changes to improve pellet quality and productivity. Quality improvements include: 1. Installation of a contraction t:ester. 2. Installation of on-line calcium and silica analyzers. 3. Process autocation changes. Productivity improvements include: 1. A second fluxstone storage bin and feeder. 2. A new vacuum pump and disc filter. 3. A filter aid addition system. 4. Enclosing the filter cake reclaim feeder. 5. A new grate bar cleaner.

Jan 1, 1989

By T. M. Yegulalp

One of the important factors affecting the determination of the minimum number of increments in a gross sample of coal is the variance component due to segregation. The segregation variance may be estimated by following the ASTM standard procedures. Since the estimate itself is a random variable, it is important to test if the value obtained is due to chance causes. This paper introduces a statistical test, based on the procedures published by ASTM, to test if the segregation variance is significantly greater than zero.

Jan 1, 1986

By E Topal

There are presently no suitable underground mine production scheduling optimisation models available for application to sublevel stope mining. Scheduling underground production for maximum return is therefore still a manual process, which generally involves scheduling production from the next available highest cash flow stope within the operational constraints of the project. In most mining operations, the scheduling problem is too large and complex to guarantee truly optimal results through manual scheduling. This paper reviews some general optimisation studies that have been proposed for underground hard rock mining. Furthermore, an example of a mixed integer programming model is presented for a small conceptual sublevel stoping operation. Comparative results of a schedule generated by a mixed integer programming (MIP) production scheduling model and a manually generated model on a nine stope example is presented. Results indicate that the potential benefits of the MIP production scheduling model for the purpose of maximising NPV are significant. Finally, a new constraint was formulated to limit multiple exposures of fill masses for use in an existing MIP production scheduling model. Testing of the formulation took place on the central stope of the nine stope example, which showed that the formulation performed its task without breaking other operational constraints.

Jan 1, 2007

By J. Whyatt

A package of bench design programs is being publicly released by the National Institute for Occupational Safety and Health (NIOSH). These programs are designed to provide a probabilistic estimate of the likelihood that various widths of catch bench will be retained. Program modules address the mechanisms of plane, plane step path, and wedge failure. Results are drawn from statistical and geostatistical descriptions of fracture characteristics and other input parameters, and reported as the statistical chance that a given width of catch bench will survive, that is, that a catch bench will remain wide enough to catch rolling and sliding material and prevent such material from posing hazards to miners operating below. The capabilities and application of this software are demonstrated through a case example. Comparing bench width estimates obtained using geologic data with actual widths is a key step in verifying that a program is appropriate for particular site conditions and that proper statistical descriptions of fracture parameters for that site have been developed. Such verification efforts improve confidence, and hence usefulness, of this method for designing future slope benches. Development and release of these programs appears particularly timely given recent progress in photoanalysis methods, which promise to greatly reduce the cost of collecting fracture data.

Jan 1, 2004

By B. B. Zavoico

Considerable progress has been made during the past year throughout the Russian petroleum industry. Perhaps the most remarkable achievement is the execution of the programs set out by the Central Planning Bureau of the Department of Commerce. These programs have been increased twice during the past operating year, necessitating drilling the richest reserves, instead of developing gradually as originally planned. The fulfillment of the set-out programs during the past few years makes therefore the present five-year plan of considerable importance to the international oil statisticians in approximating actual supply conditions up till 1933. The production of all Russian fields increased by 14,200,000 bbl. as compared with the 1927-1928 operating year, while the next year's program anticipates a further increase of 15,700,000 barrels. It appears therefore that production for 1932-33 will total 187,000,000 bbl., or approximately double that of the operating year just closed. It is also evident that the old fields are counted upon to produce only about 67 per cent. of the total, the new fields, not yet discovered, to furnish the balance of 33 per cent. The Soviet oil leaders therefore are hoping to have new fields discovered and developed by 1932-33. However, as in the past wildcatting

Jan 1, 1930

By Szymanski J

Although efforts towards producing a `total quality rockfill product' have allowed cemented rockfill costs to be reduced considerably, there is still tremendous potential for further cost reductions. Cemented rockfill costs range from approximately $7 to $12 (US) per tonne of fill placed. Approximately 50 per cent of the rockfill costs are associated with cement costs. A large portion of the remaining 50 per cent is associated with labour and materials handling costs. Several Canadian underground mines utilise haulage trucks and/or conventional belt conveyors to distribute the rockfill. However, due to the restrictive nature of the underground environment, these current materials handling systems have limitations (ie spillage from trucks, dust from conveyors). The Sicon conveyor system (developed in Sweden) has considerable potential for replacing these current methods. The Sicon conveyor utilises a revolutionary approach - the conveyor belt is folded over upon itself (forming a tear drop shape) so that the transported material is completely encased within the belt. The Sicon conveyor's ability to negotiate very tight comers (without transfer points) makes the conveyor very appealing for underground mining applications. An analysis comparing the pros and cons associated with the Sicon conveyor system versus conventional materials handling systems will be presented in this paper.

Jan 1, 1995

By Corbett GJ

Comparisons between individual SW Pacific Rim gold deposits have facilit-ated proposal of a new classification of porphyry-related mesothermal gold deposit types. Carbonate base-metal gold systems form at relatively high crustal levels, peripheral to porphyry intrusions. There is a spatial assoc-iation with major structures which focus fluids from degassing porphyries at depth. Milled matrix intrusive breccias and extrusive maar volcano/ diatreme breccias form as a result of pre-mineralisation phreatomagmatic eruptions. These focus degassing magmatic fluids by the provision of fracture ground preparation, typically within competent host rocks, adjacent to the breccia bodies. Local dilational structures control the form of individual deposits which may contain bonanza gold grades at sites of fluid quenching and/or repeated mineral deposition.Early quartz veining and porphyry-related alteration is common and later gold mineralisation typically occurs in association with carbonate veining and Fe and base metal sulphides as Zn > Pb > Cu. The progressive cooling by mixing of the upwelling mineralised magmatic fluid with descending bicarbonate waters, produces characteristic zonations in carbonate compositions. These vary with distance from the porphyry source, from Ca through Mg, Mn and distal Fe types. Gold mineralisation is preferentially distributed within the Mn/Mg carbon-ates.

Jan 1, 1994

By Greenwood-Smith R, Ellis W. K

A bulk analyser based on the gamma-ray resonance scattering process can be used to give a rapid and accurate measure of the percentage by weight of copper or nickel averaged over a large mass of sample, typi- cally 20 kg. It is shown that for a bulk sample containing about 3 wt.% Cu, crushing to -25 mm is the only sample preparation requited for analysis to within an error of about 0.06 wt.% Cu (errors quoted are one standard deviation). A comparison of bulk analyser and con- ventional assays on the same bulk samples can be used to estimate the sample preparation errors in the conventional methods. The squared deviations between bulk analyser and conventional assays are approximated by the sum of variances in sample preparation, conventional analysis, and bulk sample analy- sis. The sample preparation error was esti- mated as 0.26 wt.% Cu for samples crushed to -2 mm before splitting, 0.08 wt.% Cu for samples crushed to -0.35 mm before splitting, and 0.04 wt.% Cu for the average of 15 grab samples from bulk samples which had been crushed to -2 mm. The average copper grade for each group of samples was about 1.7 wt.%. Results are presented which show that

Jan 1, 1976

By P. Tarassoff

Extractive metallurgy is concerned with the winning of metals from their ores and concentrates. It is therefore not unexpected that extractive metallurgy in Canada has always been closely linked with mining. The discovery and exploitation of new orebodies has provided the impetus for the building of many of our smelters and refineries - past and present - and for the development of new metal extraction processes. The purpose of this paper is to provide a brief overview of those early metallurgical developments, and to highlight the continuing expansion and diversification of the Canadian metallurgical industry.

Jan 1, 1999

By James P. Ulery, Jon C. Volkwein

The U.S. Bureau of Mines investigated a method of using inert gas to prevent the formation of explosive gas mixtures in auger highwall mining of coal. A combination of gasoline and diesel engine exhaust gases was introduced into the auger drill hole using a short section of pipe located at the collar. Gas samples were taken and analyzed on-site with infrared detectors for oxygen, carbon dioxide, methane, and carbon monoxide. Evacuated bottle samples were also taken and analyzed by gas chromatography in the laboratory. These gas results were analyzed for explosibility. Coal samples from various depths were used to obtain the gas content of the coal using the modified direct method. Personal expo- sure to carbon monoxide was also monitored. The highest methane level observed was 9.55%. The levels of inert gas (carbon dioxide and nitrogen) were sufficiently high to prevent any ignition of the methane. Results showed that for all conditions during mining, gas concentrations were nonexplosive. The gas content of the coal was about 0.30 cm3/g. The maximum personal time-weighted average sample for carbon monoxide was 20 ppm. This system provides a safe, inexpensive, simple method for pre- venting explosions during auger mining.

Jan 1, 1993