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By B. Akwasi Osei

The lack of metrics to assess clamshell formation dredgeability limits the ability of engineers to predict the dredgeability of compact material and increases the risks associated with clamshell selection. Research is needed to develop metrics for predicting clamshell dredgeability. The objective of this paper is to evaluate possible metrics to predict the dredgeability of compact aggregates by hydraulically actuated clamshell dredges. Data on cylinder pressures and displacements, motor current, bucket tilt and cycle time (CT) were collected from a dredge. Work done and motor energy, potential metrics, were computed and their suitability in predicting dredgeability evaluated. Data analysis shows that; there is no correlation between tilt angle and work done or motor energy; a low correlation between CT and work done, and a strong correlation between CT and motor energy. Work done is a better measure of formation dredgeability, than motor energy, as it shows lower variability. However, in difficult digging conditions, work done alone does not accurately predict dredgeability. This work shows that, it is feasible to use cylinder pressures and motor signals data to predict dredgeability.

Jan 1, 2010

By Alan Boughey

The world's largest grinding mill operates at the Cadia Hill Mine in Australia. Svedala Industries supplied the 12.2-m- (40-ft-) diam, 20-MW Cadia gearless, semiautogenous grinding mill (SAG). This was a leap of more than 40% above the largest operating SAG mill. This article reviews the significant items involved in the development, design, manufacture, installation and operation of this mill. Description of the Cadia project The Cadia Hill Mine is located 25 km (15 miles) southwest of Orange, in the Central Tablelands of New South Wales, Australia. The mining lease covers about 3,100 hm2 (7,660 acres) of the Cadia Valley. The Cadia deposit is a large, homogeneous, low-grade, porphyry-type, gold-copper resource. Current reserves stand at 202 Mt (222 million st), grading 0.73 g/t (0.02 oz/st) gold and 0.17% copper. They contain 146.2 t (4.7 million oz) of gold and 347 Kt (382,500 st) of copper. The operation consists of openpit mining and ore processing to produce gold bullion and a gold/copper concentrate. The operation processes 17 Mt/a (19 millon stpy) of ore to produce 91 kt/a (100,300 stpy) of concentrate and 9.1 t/a (293,000 oz/year) of gold. The plant operates two 12-hour shifts a day and had a targeted availability of 94%. The plant facilities include a copper concentrator, o water supply system, an openpit mine, tailings pipeline and dam, and a concentrate pipeline and dewatering facility located in Blayney, about 35 km (22 miles) from the concentrator. Significant features of the process plant include o primary crusher and stockpile, a reclaim conveyor system feeding, the Cadia SAG mill (Figs. 1 and 2) and two 6.7-m- (22-ft-) diam ball mills (Fig. 3). A portion of the gold is extracted in a gravity circuit. The balance is extracted with the copper in the flotation section of the plant. Concentrate is thickened and piped about 35 km (22 miles) to the filter plant and rail loadout facility located in Blayney. Tailings are pumped to a tailings dam and process water is reclaimed from the tailings dam. The water system is designed to control and capture available runoff from local creeks, and to store these waters in dams and weirs. In addition, a major portion of the water for the process plant is treated effluent from the city of Orange and a smaller portion from Blayney Mill selection. The low-grade ore posed a dilemma for Cadia. To achieve a satisfactory economic operation, large-scale processing was required. The selection of the mills, equal to the largest designed and operated at the time, gave a plant configuration consisting of two 11.1- x 5.8-m (36- x 19-ft) SAG mills and three 6.1- x 9.1-m (20 x 30-ft) ball mills. This configuration did not provide the most satisfactory economic model for the project. A review of the single line configuration indicated that a significant reduction in the overall capital cost of the plant could be achieved by increasing the size of the mills and reducing the quantity. Increasing the size of the mills,

Jan 1, 2001

By C. J. Konya, A. Konya

"Presplitting is a form of blasting that when executed properly will create a smooth wall that will minimize secondary conditioning. In recent years, presplitting has become a popular technique to be used in both the mining and construction industry, however, presplitting fails on many occasions because of ‘geologic conditions’. This is typically because of improper explosive loads consisting of overloading and underloading blastholes. In this paper the authors introduce new factors such as a presplit factor which varies for different rock types, selection criteria for the spacing of presplits, the optimization of explosive load, as well as prediction methods for the environmental affects for the technique of Precision Presplitting. BACKGROUND Methods of controlled blasting in which a smooth, stable wall could be blasted have been an important consideration since blasting in mining and construction has begun. How can one use an explosive, with its massive energy and power, to break a non-homogenous rock without causing backbreak, poor walls, or deviations in the wall over large distances? The blasting industry has known for a long time that blasting ‘high benches’ would result in better walls and less backbreak. However, this was not a sufficiently controlled technique for critical projects, such as the Niagara Power Plant. This was one of the main challenges Hercules Powder faced when they began the project. In order to achieve this, they used a new tool in the explosive engineer’s toolkit, presplitting. The presplitting done on the Niagara Power Project utilized 1 ½” x 8” (38mm x 203mm) 40% gelatin dynamite in 2 ½” (63.5) diameter boreholes. This was done on multiple levels with benches between 12 feet to 52 feet (3.6m to 15.8m) with a total powder factor of 0.75 pounds/cubic yard (0.45 kg/m3). The spacing on these holes ranged between 18 inches to 24 inches (0.46m to 0.61m), center to center (Langfors & Kihlstrom, 1973). The strange aspect of this project was that this “lightly loaded” closely spaced row was fired before the production blast, yet this project produced some of the best walls seen at that time directly from blasting. In today’s blasting industry, presplitting is a more common way to produce smooth, stable high-walls for both the construction and the mining industry. Today, explosive loads for traditional presplitting is generally measured in pounds per foot (kilograms per meter) and typically ranges from 0.30 lb. /ft. to 0.5 lb. /ft. (0.45 kg/m to 0.75 kg/m). MECHANICS OF PRESPLITTING When analyzing a blasting process like the Niagara Project, many wonder how did that work when normally confining a charge would cause cratering and increased backbreak. Since then, presplitting mechanics have been studied extensively. It has been shown in many models that presplitting is not achieve through the collision of shockwave, but by gas pressure in the boreholes. In addition to models and finite element simulations, this has been proven full scale using Pyrodex, a propellant, to create a presplit plane with no shock wave present (Konya, Dean Barrett, & Smith Jr, Presplitting Granite Using Pyrodex, A Propellant, 1986)."

Jan 1, 2017

By Wendell R. Peters

Shallow underground voids resulting from early coal mining and other resource recovery activities over the past several decades are now being recognized as a significant cause of ground subsidence problems. Uncertain knowledge of abandoned coal mines also imposes potential hazards in current mining operations since water inundation or the release of methane gas is a serious hazard when mine excavation operations break into an abandoned mine. Requirements for detecting and mapping subversive man-made tunnels in previous work has led to the development of a surface- operated automatic earth resistivity survey system and associated computer data processing system for military applications. Field demonstration tests of this system have resulted in successful detection of tunnels at depth-to- diameter ratios greater than ten to one. Under the sponsorship of the Bureau of Mines, a similar high-resolution earth resistivity system was designed and constructed for use as a potential means of detecting coal mine workings. The hardware and software aspects of the new system are described together with applications of the method to the survey and mapping of abandoned coal mine workings. Field tests are presented showing the detection of both air-filled and water-filled mine workings.

Jan 1, 1981

By P. Lan

Gold is often associated with arsenopyrite in many refractory gold ores. The selective recovery of arsenopyrite by flotation is therefore necessary in order to maximize the recovery of gold. However, the complexity of arsenopyrite flotation often makes it difficult to obtain a high recovery from real lean gold ore. This paper presents the results of arsenopyrite flotation by ultrasonic pretreatment in micro flotation experiments and bench scale flotation tests. The tests showed that pure arsenopyrite exhibited low natural floatability due to the oxidation on the mineral surface under ambient conditions. Using ultrasonic pretreatment, the oxidation film was ruptured and the floatability was enhanced greatly. The presence of Fe (III) was found to inhibit the flotation of arsenopyrite and this depression was also removed by the ultrasonic pretreatment. This result explained why the arsenopyrite in real ore is often difficult to float and how ultrasonic pretreatment can effectively improve the arsenopyrite flotation in real gold ores. Guanidine (GMM) was found to be a superior collector for arsenopyrite in acidic medium and xanthate in alkaline medium.

Jan 1, 1998

By J. A. Cremeens, D. D. Overton, J. C. Varnier

"Slope failures that occurred in 2008 and 2009 at the Pikeview Quarry in El Paso County, Colorado, USA were investigated using three-dimensional discrete element modeling methods. Large scale block sliding failures occurred along weak bedding planes that dip steeply into an excavation located at the toe of the slope. Several mining sequences at the toe of the slope were simulated to evaluate the effects of various degrees of extraction and backfill on slope stability. Slope stability back analysis was conducted using pre-failure and post-failure conditions to determine the joint shear strength properties. Laboratory testing was also conducted to determine intact rock properties. Three-dimensional discrete element analysis was uniquely applicable to this evaluation because it allowed representation of shear stress of the basal sliding surface and it afforded evaluation of several scenarios of extraction and backfill at the toe of the slope. INTRODUCTION The Pikeview Quarry is an aggregate quarry located approximately five miles northwest of Colorado Springs in El Paso County, Colorado, USA. The quarry has been operated by several owners since the early 1900s. The Pikeview Quarry has been a major source of construction aggregate for the Colorado Springs area for many years (MSHA 1986). In the 1950s, the site was mined by Peter Kiewit and Sons to provide limestone for the construction of the nearby Air Force Academy. A large scale slope failure occurred on December 2, 2008, which stopped mining activity at the Pikeview Quarry. A second, related slope failure occurred on September 13, 2009. The second slope failure was located immediately adjacent to the initial December 2008 failure. The objective of this publication is to discuss the three-dimensional slope stability evaluation. The causal relationships of slope instability are not within the scope of this publication. However, omitting the precursors to slope failure would limit understanding of the nature of the slope stability evaluation. The regional and local geology and the history of slope instability at the Pikeview Quarry aid in understanding these precursors. Two main failure mechanisms were hypothesized during the failure investigation to explain the triggering of the 2008 and 2009 large slope failures at the Pikeview Quarry. Both mechanisms are given equal credence in this publication. Quarry excavation was a key factor in the slope instability for both failure mechanisms. The objective of the three-dimensional model was to evaluate the effects of extraction and backfill at the toe of the slope with respect to slope stability. GEOLOGY Regional and local geology are both important factors in slope instability at the Pikeview Quarry. Geologic structure at the quarry is related to regional mountain building and basin development associated with the Laramide Orogeny and the Rocky Mountains. Lithologies at the Pikeview Quarry are also related to regional trends. Localized geologic features, primarily related to weathering and erosion, have also been major contributing factors to slope instability at the Pikeview Quarry."

Jan 1, 2017

By Rodrigo Peroni, Euler Victoria

Knowing the depth of contacts between weathering zones in phosphate mines is determinant to define the stripping ratios and the mineralized zones. Contacts are gradational and difficult to define even in the drillhole samples, so the combination with geostatistical simulation techniques helps to define the contact zone and the grade distribution, providing a way to assess the probability of the occurrence and the positioning of contacts between the barren waste and the ore zones. These inputs are used in the scheduling, anticipating risks associated with mine planning scenarios. The results from an application to a case study showed a good improvement when compared to traditional techniques.

Jun 25, 2023

By A. A. Campoli

The U. S Bureau of Mines is conducting research to develop improved guidelines for selecting the type and pattern of roof support in different geologic environments (Mucho, et al., 1995). As a part of this wider research effort, a multifaceted ground control investigation was conducted at the Diamond TB Mine of RoxCoal, Inc. The study included statistical analysis of mine-wide roof bolt performance, instrumented bolt test areas, stress damage mapping, and pillar design analysis. The mine employs state-of-the-art continuous miner technology, including mobile roof supports and continuous haulage, to retreat mine the Lower Kittanning Coalbed. A herringbone section design makes the stability of the center belt entry critical to efficient retreat mining. Roof cutters formed during section development sometimes set up unstable immediate roof conditions in advance of the belt entry mobile roof supports during retreat mining.

Jan 1, 1996

By Lu Yang

In the last two decades, acid iron ore pellets have established their predominant position as the iron-bearing burden material for blast furnace ironmaking in North America. It is a general practice that pelletization plants are asked to produce pellets of certain chemical composition and in a narrow range of sizes. A minimum amount of fines created during handling and transportation is desirable. Such a requirement is often expressed indirectly by the minimum compression load causing a pellet to break. The behavior of pellets inside a blast furnace, of course, is very important for a smooth and efficient operation. Before its melting in the lower part of a blast furnace during this period of several hours the behavior of pellets depends on their properties after induration as well as conditions inside each furnace. Among properties of indurated pellets, the ability to maintain their physical integrity in the burden and to be reducible are most important. To maintain the integrity of pellets (i.e., minimal low temperature breakdown, swelling and softening) so that there will be adequate overall permeability for certain wind rate and local permeability for gas/solid contact for reduction. A higher reducibility will moderate the adverse effects of swelling and softening and contribute to a smooth operation and efficient use of heat and reductants in a blast furnace. An article devoted to the relationships between the quality of iron ore pellets and blast furnace practice has been published elsewhere(1); therefore, no further discussion is needed here. The mechanism of strengthening of iron ore pellets during firing has been the subject of investigation of many authors(2). There is no doubt that the bridging of the growing hematite grains, particularly starting with magnetite concentrates, is very significant in the production of strong pellets. Recently, there are many reports from blast furnace operators that they have noticed the adverse effect of LTB in their plants. It is obvious that the strength derived from bridged hematite grains may disappear when hematite is converted to magnetite with thorough breakdown of grain structures. It is suggested that some other strengthening mechanism may be required in order to improve the LTB index of acid pellets. The influence of porosity on both compression strength and reducibility follows regular patterns. It has been suggested(1) that the amount of liquid slag in pellets during firing determines to a large extent the porosity, hence, indirectly, strength and reducibility.

Jan 1, 1983

By N. Kuyucak

While Acid Rock Drainage (ARD) occurs naturally, its management poses challenges especially at the decommissioned sites. Two passive systems were designed and implemented at two different decommissioned mining sites in Canada in order to treat the seepage with ARD characteristics. One seep originating from a reclaimed tailings pond was neutral mine drainage (NMD) containing circumneutral pH and dissolved metals (e.g., nickel, zinc) in elevated concentrations. The other ARD from a tailings stack showed net acid drainage characteristics with low pH and higher metal concentrations than regulated limits. Due to the different ARD characteristics occurring at each site, different types of treatment systems were designed and implemented. Both sites experience extreme winter conditions. The NMD has been treated mainly by limestone drain and aerobic systems since 1997. The acid ARD has been treated by a system consisting of an anaerobic sulfate reducing bacteria (SRB) cell, an aerobic cell and a limestone filter since 2004. As this site has no access roads, space, or power, a significant capital cost as well as sludge disposal/storage requirements would have been required to institute a chemical (active) treatment system. Both systems have been successfully treating seepage and producing a water quality year round that could comply with regulated objectives.

Jan 1, 2010

By J. E. Urosek

The Mine safety and Health Administration (MSHA) formed the Mine Ventilation Bleeder and Gob Training Committee in 1993. This committee was charged with developing a training course designed to enhance the knowledge and understanding of practical bleeder ventilation systems. The course will probably include a discussion on the theory behind bleeder system design, an in-depth look at the strengths and weaknesses of the basic bleeder system types, case studies, and a review of the systems currently in use. A database on active underground bituminous coal mines, gob areas, and the bleeder systems presently used was developed from questionnaires distributed to nine Coal Mine Safety and Health (CMS&H) districts. Representative systems were selected for thorough investigation to develop case studies. This paper introduces some of the material to be included in the course, including a description of basic bleeder system designs, a review of a portion of the database, and facts about the gob areas and bleeder systems investigated for the development of case studies.

Jan 1, 1995

By N. C. Miller

In situ stope leaching is an innovative mining system that reduces the surface impacts of conventional underground mining. As in any leaching operation, stope leaching requires hydrologic site characterization and control of leaching solutions. Several tracer tests, using sodium chloride (NaCl) in various concentrations, were conducted by the U. S. Bureau of Mines (USBM) in an underground experimental stope to simulate a leaching environment. The tests were conducted in fractured crystalline rock that was partially saturated. As tests progressed, the surrounding fractures became more saturated. A predominant flow direction away from the stope was found during testing. Testing procedures, preliminary results, and remediation ideas based on field observations are presented in this paper.

Jan 1, 1995

By N. J. Ferguson

An upsurge in demand for Australian coal has seen a much needed programme for a new generation of coal ports. This paper discusses design concepts in planning a large modern terminal, with the 25 Nt/a Port Kembla and the 50 Mt/a Newcastle facilities as practical examples. The need is emphasised for a studied Master Plan, in-built materials handling flexibility and staged development. Trends in bulk ship design and loading methods are noted. Features of these Australian designs include methods of dealing with poor ground conditions, functional civil engineering for environmental protection and optimal marine structure designs.

Jan 1, 1982

By H. K. Lin

The free-milling ore characteristics of the Fort Knox gold deposit are ideal for studying gold and silver behaviors in various processing circuits without serious interference of sulfide minerals and other refractory factors. A galvanic effect leads to faster silver dissolution than gold dissolution and, as a result, the Au/Ag solution concentration ratio increases through the leach circuit. In the CIP circuit, Au adsorption is more effective than Ag adsorption. The Au/Ag solution concentration ratio decreases from 10.0 to 1.0 as the slurry advances from CIP tank #1 to CIP tank #5, while Au/Ag ratio in the carbon increases from 3.2 to 15.3, as the carbon advances from CIP tank #5 to CIP tank #1. Silver is stripped more effectively than gold from the loaded carbon. The Au/Ag ratio in solution reaches 2.0 in the first 75,700 L of strip solution that pass through the loaded carbon; as the stripping process continues, the Au/Ag ratio in the strip solution increases. In the electrowinning circuit, the Au/Ag ratio in the spent electrowinning solution is significantly lower than that in the electrowinning feed, indicating that Au is more rapidly deposited onto the cathode than Ag.

Jan 1, 2010

By J. A. Haggman

In a country richly endowed in mineral wealth, Climax Arimco Mining Corporation's (CAMC) Dinkidi porphyry Au?Cu deposit in Nueva Vizcaya province, Philippines, represents the most significant mineral discovery in the country in recent times. The deposit is located within a belt of young intrusive and volcanic rocks in the Sierra Madre mountain chain of north Luzon a region previously unassociated with porphyry-style mineralization. Since 1989, drilling of the Dinkidi deposit has outlined a resource of 124 million tonnes of 1.0 g/t gold and 0.41 % copper. Regional exploration throughout CAMC's 37,000 hectare FTAA has identified a further sixteen potential porphyry-related deposits. A new mining law was passed in March 1995 but prior to that CAMC had the first Financial and Technical Assistance Agreement (FTAA) granted by the Philippine government allowing 100% foreign ownership of mining projects.

Jan 1, 1997

By James J. Komadina

INTRODUCTION Newmont Gold Company's Rain operations are located in Elko County, Nevada, approximately 9 miles southeast of the Town of Carlin.The property was submitted to Newmont Exploration Limited in 1979 and underwent development drilling over the course of the next four years, culminating in the completion of a project feasibility study in October 1986. Initial project scope included development of an open pit mine, an 815,000 DSTPY heap leach, a 572,000 DSTPY mill (Mill No. 3), and associated support facilities. Approval to proceed was given by Newmont Gold Company management in June 1987. Detailed engineering began in early July, followed immediately thereafter by procurement of major equipment components through various second hand sources. PROJECT DESIGN Keep It Simple In early August, Newmont Mining Corporation, parent company of Newmont Gold Company, became the takeover target of an "investor" group headed by T. Boone Pickens. Details of the ensuing events have been well documented by the press; suffice to say their impact on this project was material and required immediate flowsheet changes. Project design throughputs were increased to 1,000,000 DSTPY and 760,000 DSTPY for the heap leach and mill, respectively, in mid- September. These changes caused major project headaches as all general arrangement drawings had been approved and structural steel as well as concrete contracts were in place. Additionally, ground breaking was to commence in two weeks time! Rain project management assembled shortly after the announced changes, to devise an engineering and construction strategy capable of handling these new business conditions. The project had been justified using a gold price of $325 per ounce with a forecasted overall cash cost per ounce of $230. Hence, strict control of capital expenditures was necessary to maintain development economics. Mill start-up had also been moved forward one month, to August 1, 1988. To meet an already ambitious schedule, construction techniques needed to be simple and proven as most of the facilities were planned for erection during winter at a site with design criteria of -40F and 90 mile per hour winds. Perhaps the most critical factor requiring consideration was that of personnel. Newmont Gold's revised business plan called for corporate-wide production increases from 587,000 ounces per year in 1987, to 1,400,000 ounces by mid- 1989. Labor requirements were forecasted to grow from 900 employees to over 2,100. This fact, in combination with the rapid growth of Nevada's gold industry, led to the expectation that turnover of operating personnel would be extremely high. It was concluded that, to meet cost, time, and operational goals, the necessary strategy could be succinctly put - "Keep It Simple." Mining/Metallurgical Factors The Rain deposit consists of three interconnecting mineable shapes with host rocks ranging from 90%+ barite to a highly silicified jasperoid. Due to a limited number of mining faces in the first two years, in-pit blending

Jan 1, 1990

By Alexander Sutulov

1. INTRODUCTION According to the classical economic approach of the open market economy, copper supply should be regulated by market forces which are (a) demand and (b) prices. In fact, demand is of primordial importance because prices are a reflection of supply-demand balance: they go up in times of shortages of supply and down when over-supply situation exists. In the past few years, this basic mechanism of market economy has been under pressure from a number of non-market forces, such as social and monetary problems of sane important copper producing countries; structural changes in copper mining companies through mergers with oil companies; and other factors which greatly contributed to distortion of the basic supply-¬demand equation. As a result, we have faced in the last three years a situation of oversupply rarely seen in the past, with a catastrophic fall in copper prices, huge losses at mines and unprecedented indebtedness of copper producers. This has deep repercussions in the copper business and its future supply outlook. This is particularly so when taking into consideration that high indebtedness and high interest rates conspire against the very survival of some traditional producers. As we write this comment (July of 1985), close to one half of U.S. copper production capacity is closed down, sane of it for good, while the same precarious situation exists in Canada, Australia, Zambia, and some other primary producers who cannot face present low prices with equally low costs. The question is then what this situation means for the medium-¬term future, until the end of this decade, Particularly when taking into consideration the fact that some of the factors influencing it are of a long-term character. 2. SUPPLY-DEMAND SITUATION IN THE LAST FIVE YEARS Since the last boon year, in 1979, when world copper consumption reached a record of 9.8 million metric tons of refined copper, demand dropped to only 9.0 million tons in 1982, when the crisis bottomed out, and then steadily recovered to the level of 9.7 million tons in 1984. Meanwhile, copper production, which in 1979 was only 9.4 million tons, in spite of the fall in demand, grew to 9.7 million tons in 1981 and 1983 and only temporarily was cut to 9.4 million tons in 1982 and to 9.5 million tons in 1984. As a result of this imbalance, hugh copper stocks were accumulated with supply systematically out¬growing demand by over 2 million tons at one time. With the Western World refined copper demand running at an average of 115,000 tons per week, this oversupply was equivalent to between 15 and 18 weeks production, i.e. twice as large as normal. This was immediately reflected in very low copper prices. Another contributing factor to relatively low copper prices was the lack of speculator interest for this commodity. Between 1981 and 1984 the world experienced a very intensive period of wealth distribution, high interest rates and intensive restructuring off world economy and its industrial sector. Therefore, much of the available money was attracted to more lucrative and attractive investments rather than to the lack-luster performance of copper, and for that matter, of the metal sector in general. The much announced and discussed recovery of the U.S. economy since 1983, and particularly in 1984, with an overall growth of the GNP by 6.8% in 1984, happened to be more oriented to the service sector than the industrial sector and therefore this recovery was not adequately reflected in copper consumption nor to metals in general. This put serious doubts about the future demand for copper. Consequently copper prices failed to improve in spite of an apparent fall in stocks and a healthy increase in the demand for copper since 1982. At the time of this writing (July, 1985), although same analysts claim that the available stocks are running at about 8 weeks of supply, which are considered to be below normal, copper prices are still very low at between 65 and 67 cents per pound of copper. Part of the answer may be the normal seasonal slow-down period in the Northern Hemisphere due to summer vacations and

Jan 1, 1986

As we are all well aware, the demand for coal is increasing very rapidly. This makes the coal industry one of great opportunity for both the miner and the-supplier. Up until the middle 1960's, the price of coal had been one of the most stable of all commodities for the proceeding fifty years. In somewhat oversimplified terms this was because the inflationary tendencies had been offset by improved mine techniques and more efficient, more productive, and more reliable equipment; but, in the last seven years the price of coal has doubled. It is not the intent of this paper to discuss the many reasons for this price increase, but to show how the cost of hauling coal can be sharply decreased with a new truck concept. Not long ago a mining executive stated that all haulage trucks now available are basically the same. They look the same; they cost about the same; they have the same engine, power train, and tires; and, they have about the same operating cost. Truck size has increased which has reduced the number of drivers, but the purchase price and maintenance and repair costs have increased proportionally. Pits must be widened and loading shovels enlarged, out, all this has not reduced the cost of moving coal. Another executive said that he saw no advantage in buying new trucks because his people now know how to maintain and operate the ones they presently have. They have all the right spare parts along with a good maintenance program. Most replaceable parts of the trucks now in operation have been replaced and improved at least once and the trucks are now running better than when they were new. Besides, the trucks are no pretty well written off.

Jan 1, 1972

By W. H. Jang

Recent research has demonstrated the utility of the FTIRIIRS adsorption density equation for in-situ study of surfactant adsorption phenomena at the surfaces of reactive internal reflection elements (IREs). The FTIR/IRS adsorption density equation is given by (formula) where r is the adsorption density, A/N is the absorbance per reflection, e is the integrated molar absorptivity, 4 is the bulk concentration of the surfactant, d. is the effective thickness, and dP is the depth of penetration. Previously, the validity of the adsorption density equation had been confirmed only from solution depletion measurements with high surface area powders. Now, FTIR/IRS studies of Langmuir-Blodgett (LB) monolayer films of fatty acid transferred to the surface of a calcium fluorite IRE .clearly validate the reliability of the FTIR/IRS adsorption density equation for quantitative analysis of surfactant adsorption. In addition, it was confirmed from the observed absorption band at 1550 cm-1 that the first monolayer of the transferred fatty acid chemisorbs at the surface of fluorite.

Jan 1, 1994

As the surface reserves are being depleted, more and more stone operators are seeking underground mining options. Stability of underground openings is a major concern for the safety and productivity. This paper relates to a field study conducted at a large underground limestone mine in Central Pennsylvania. Identification of hazards leading to ground failures and assessment of roof fall risks are important in mitigating injuries. Field observations at this mine were integrated with Roof Fall Risk Index (RFRI) mapping software recently developed by West Virginia University. The system is designed to combine field study data related to geologic factors, mining-induced damages, roof profile, and ground water influx to generate RFRI maps. The mine has adopted a proactive ground control program consisting of comprehensive examination of drilling and bolting logs; precision surveying of roof and rib; accurate mapping of cutters, falls, and geologic structures; and installation of extensometers and microseismic monitoring systems. Investment of time and resources has eliminated fatalities and lost-time injuries due to ground fall. This paper presents the geologic settings, mining conditions, ground control issues along with risk management and control techniques adopted at the study mine.

Jan 1, 2009