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By H. Lyn Bourne

The is the 14th consecutive year for this column and the 13th in which the royalty data have been tabulated. Some holes still exist in the list of royalties that readers hopefully will fill. Construction aggregates - either sand and gravel or crushed stone - are the most common mineral commodity. In spite of their pervasive use nationwide, royalty information for these commodities has been difficult to acquire from some states. Readers can help. Share information about these missing states and next year there will be fewer states without information. Dennis Bryan sent a copy of a report, "Regional Appraisal of Mineral Materials: Carson City District, BLM." The report was authored by real-estate appraisers. It contains more than 50 references to leases of various construction materials in Nevada and nearby states. It gives a thumbnail sketch of each lease and is a good reference. However, the authors erroneously imply that the royalty rate and the in-place value are the same. Several of this year's new entries come from the report. However, the royalties for landscaping materials vary from state to state and by rock type. Some boulders used for landscaping in Nevada have a royalty reported at $1.10/t ($1/st). At the other extreme is lava from Idaho. Used for the same purpose, it has a royalty range of $11 to $18.70/t ($10 to $17/st). Chuanming Tian of Beijing, China has shared some information about the changing mineral royalty policies in China. China first implemented an oil and gas royalty in 1984. In 1992, the government expanded that to include iron ore. In 1994, the royalty policy was extended to all mineral commodities. At the same time, the royalty rate for iron ore increased tenfold.

Jan 1, 1995

By A. C. Lyons, A. J. Reed

INTRODUCTION A trend which is obvious in the United Kingdom and is thought to be evident in other parts of the Western world is the disproportionate rise in the cost of tunnel mining labour compared with other tunnel cost increases. This disproportionate increase in labour costs has been accompanied by, and may be the result of, a shortage of skilled working labour. To combat this problem there has been a steady increase in the use of machines for excavation and for the erection of tunnel linings to keep the amount of skilled labour working below ground to a minimum. The typical tunnelling cycle, where the ground needs continuous support can be excavation, temporary support, temporary lining, grouting and permanent lining. It is obviously an advantage to combine as many of these operations as possible and carry out as much as possible by machine. Machines have been developed which will, either alone or used in combination with a shield, excavate and, when necessary, temporarily support the excavation. If a segmental lining is used, it can be so designed that it will form the permanent lining and reduce the cycle time and skilled labour requirement to a minimum. In cases in which the ground can be cut accurately to a smooth profile and will stand without support for a sufficient time, the lining can be expanded against the ground and the necessity of grouting eliminated. In cases in which ground will not stand without support, the usual method is to use a bolted lining and to grout behind it preferably as the shield is shoved forward. Whichever technique is used (this is largely determined by the ground) it is an advantage to make the segments as large

Jan 1, 1974

By Warren E. Fife

This paper will discuss the application of a continuous miner using a fixed drum milling type cutting head for the underground mining of hard rock and salt ores. Continuous mining machines of the ripper chain cutting head, oscillating milling cutting head and boring arm designs have been used in rock and salt applications with various degrees of success. The large boring arm miners have been successful in potash, trona and salt, but they have not been successful in mining rock materials like iron ore. The main method used in underground mining rock materials is the conventional drilling and blasting the ore from the solid. This is particularly true for the harder, more abrasive copper and iron bearing ores. The softer salt based ores of potash, trona and rock salt can use several underground mining methods depending on seam conditions. Several of these methods use underground coal mining equipment that has been extensively redesigned to adapt it for use in mining noncoal materials. Conventional mining methods employ the use of a universal cutting machine to undercut or topcut the ore body before face drilling and blasting takes place to produce a smooth parting of the ore. The boring arm continuous miner is widely used in underground mining of potash and trona ores. It produces a tunnel with a smooth semicircular contour which results in good rib and roof stability. The ripper chain continuous miner is also widely used in mining the salt based ores of potash, trona and rock salt. They are used more frequently in Europe than in the United States and Canada. Some reasons for using the mechanical mining machines as opposed to the conventional drilling and blasting techniques for mining hard rock ore bodies are the resulting improved rib and roof control, facilitates selective mining of the ore body, and requires less man-

Jan 1, 1974

By R. Sharma

The heterogeneous distributions of nodule sizes, cover-ages and abundances and the variability of their burial und association with sediments and rock outcrops on the seafloor have emerged as crucial parameters in considering the exploitation potential of the manganese nodules in the central part of the Indian Ocean. Nodules with the long axis measuring from 20 and 40 mm (0.8 and 1.6 in.) account for nearly 60% to 70% of those evaluated from photographic and grab-sample data. In addition, 75% of the nodules (by weight) are between 20 and 60 mm (0.8 to 2.4 in.). The distribution of nodules with abundances ranging from 1 to 20 kg/m2 as well as coverages of 1 % to 90% on the seafloor are important considerations in demarcating the potential areas for nodule mining. The mining of areas with higher concentrations of nodules, such as along slopes associated with less sediment, would result in the enhanced recovery of nodules compared to areas with low nodule concentrations, such as in valleys and plains where thick sediment covers occur. The mining head will encounter substrates such as hard rock exposures, either with or without Fe-Mn encrustations, as well as soft sediment layers in the nodule fields. It must be able to collect buried nodules under sediment cover, and it must be able to negotiate uphill and downhill slopes in mining areas. Mapping prior to mining must delineate areas to he avoided and must define locations to be mined. The requirements suggest that the nodule collector must be a dynamic (active) system with sensors to detect areas of potential nodule concentrations, as well as features that are hazardous (or unfavorable) for mining operations.

Jan 1, 1999

By Rainer Kahrger, Guenter Ewald

The topic of this paper is to demonstrate the development of mobile crushing installations from throughput capacities of 200 tons per hour, to mobile and semi-mobile large- scale crushing installations with throughput capacities of 3,500 up to 10,000 tons per hour. Basis are approximately 140 mobile and semi-mobile crushing installations which have been built since 1956 or were on order in 1987. [FIGURE 1. Total Number of Mobile or Semi-mobile Crushing Plants] Sixteen of these approximate 140 mobile and semi-mobile crushing installations have a throughput higher than 3,500 metric tons per hour. Mobile crushers were developed more than 25 years ago in Europe for the quarry industry, which had special requirements - fairly low capacity, mostly softer and less abrasive material (limestone) and full mobility. So the majority of the units up to about 1980 were fully mobile, having a fairly small capacity and utilizing mainly hammermills, double roll and impact crushers. It was only when the mining industry and large quarries picked up the concept in the early 1980's that a drastic increase in capacity took place. Even though the first mobile crusher in a true mining application went into service in 1971, it took another 10 years before it proceeded in a larger scale. This trend was obviously supported by the drastic fuel price and energy cost increase in the mid 70's. Since the cost of conveying material is substantially cheaper than trucking it, the increase in energy cost, combined with increasing haulage distances provided a strong incentive. Figure 1 shows clearly the throughput distribution of all crushing installations and it also shows the sharp increase in the throughput capacities within the last few years.

Jan 1, 1989

After the war that was nearing an end In early 1945, regular consulting work for Bralorne, the Bradleys, Pennsalt Chemical Co., and a directorship and occasional consulting for Homestake Mining Co. made it almost certain that in spite of my 61 years I would have enough work to feed my family. But this routine work would not provide enough income to lay up funds for the inevitable future when I could no longer do arduous work. And it would not give me the adventure that had first lured me into mining. Therefore I was greatly relieved as well as pleased when, in the 18 months from April 1945 to November 1946, I was asked to make three of the most adventurous as well as profitable examinations that ever came my way. Pati[n]o Mines, Bolivia The first of the exciting jobs was an examination of the Pati[n]o tin mines in Bolivia. After I had unsuccessfully tried to induce Anaconda to purchase these mines In 1915, the World War I crisis that had forced the half-Indian accountant to try to sell the mines had come to an end. The European market for tin improved dramatically, and "Don Simon" was able to borrow money for expenses while the concentrates were still In Bolivia. He took advantage of the favorable conditions by buying adjacent properties and increasing development work until his Bolivian tin mines were the greatest in the world. And his daughter had marled the French Count de Boisrouvain and was cutting a swath In high International society. Pati[n]o wanted to get rid of the drudgery of running a complicated mining Industry in a remote area In Bolivia when the fleshpots of Paris were much more seductive. An added incentive was the fact that his miners had become unruly and had even threatened his 11fe.

Jan 1, 1976

By John L. Schroder

The design of an underground mining operation requires the integration of transportation, ventilation, ground control, and mining methods to form a system which provides the highest possible degree of safety for mine personnel; the lowest cost per ton of product; the most suitable quality of final product, considering its intended use; the maximum percent recovery of the reserve, consistent with cost and subsidence rights; and the minimum disturbance to the environment. The selection of mining method is generally the controlling factor in overall mine design. Some important considerations in the choice of mining method are : 1) Depth of cover over the coal seam. 2) Strength and thickness of rock strata in the overburden. 3) Characteristics of the coal seam, including thickness, nature and location of impurities or irregularities, dip (or pitch), nature and direction of cleavage, hardness, and structural strength of the coal. 4) Nature and strength of the immediate roof (top) over the coal seam. 5) Nature and strength of the floor (bottom) under the coal seam. 6) Presence and amount of explosive gas (methane). 7) Quantity of water likely to be encountered in the mining operation. 8) Any previous or concurrent mining in seams above or below the coal seam to be worked. 9) Any restrictions on recovery required for subsidence control. Topographical and geological reports and maps of the area are usually available from state and federal geological surveys, and contain excellent general information to guide more detailed investigations. Exploratory drilling, and the plotting and analysis of the information thus obtained, are essential steps in any development program, and aid in the correlation and evaluation of information gained by other means. Other good sources of information are operating coal mines that may be in the same seam in nearby properties. Such mines may indicate in advance much about roof control, drainage, ventilation, and other mining problems.

Jan 1, 1981

By Paul H. Johnson

The CH Leaching process is a new, more sophisticated, form of gold, silver, uranium and sulfide copper heap leaching wherein each layer of a multi-layered heap is divided into a number of independent cells, each of which are successively formed and leached in the top layer of the heap to their maximum economic recoveries, and then rinsed and sealed from further leaching. Triangular earth berms form the outside walls of the heap and a gridwork of rock-filled berms on top of the bottom pad form independent under- heap solution storage zones. The high degree of solution recirculation and staging that occurs in this process produces a relatively small volume of very rich pregnant solution. The cellular heap design and its mode of operation result in a uniform, intensive, long-term leach for all of the ore particles in the heap. Besides improved metal recoveries, the process appears to have several cost and operating advantages. In the CR2MR process, sulfuric acid and a soluble sulfide are added to a gold and/or silver pregnant leach liquor to precipitate most of the contained silver, mercury, copper, etc. as sulfides at a pH below 2.5. The other product of the reaction is solubilized HCN. Following a liquid-solid separation step, cyanide is regenerated in the supernatant solution by raising the pH to a level above 9.0. This process regenerates a portion of the cyanide used in the leaching step, removes contaminant ion species (silver, copper, mercury, etc.) from the leach solution and, in a subsequent step, recovers the gold in an enriched form. It is believed that substantial cost savings may be experienced by the use of this process in some gold and/or silver leach systems.

Jan 1, 1984

By Boris K. Krstev

INTRODUCTION The flotation process, in the real industrial complexes, is characterized by the existence of the significant number of the close reciprocal interrelationship between the input and output Darameters. Traditional complexity is appeared with regard to absence of the suitable a priori information for the physico-chemical characteristics of the flotation process. It’s led to needful of taking of consideration about the undefinition which have got the place in the real conditions and application of the statistical methods. The identification of the flotation process through statistical methods is conception of the more-stages procedure, including the collection and previous analysis, the choice of the model structure and valuation method, the estimation of the model un- known parameters and the results interpretation. The essential concept for the subject of this paper belongs to the empirical models ("black-box") and deals with optimizing in constant industrial plant process. The offered method envisages the application of process operation data in order to improve the work process or constant accomplishment of better results for the target function in relation to the former achievement. The optimum process solution isn't afixed value under the influence of various factors, but "a moving point" along the response surface and deviates from the scheme solutions. The evolutive planning method constantly "investigates and approaches" the optimum solution for the selected target function. As a matter of fact the expert operates the plant without interrupting the normal work process. The essential concept of method is to attain changes in the standard conditions, but so limited that they will not effect either production or the product. This method is based on statistical concepts and availability of following the risk of error. The possible influental factors for the flotation process (m=l, 2, 3) is following: the metal content in the feed ore; production level (t/h); grinding fineness in the flotation pulp; pH-values; the consumption of collectors and modifiers The possible target functions are: the recovery of plant capacity, metal recovery in the concentration process and concentration quality. The application of the evolutive operativity method in the chalcopyrite flotation- Bucim concentrator, combines the statistical data and the operator's experience in the interpretation and making a decision which adds up to the precise automatic performance of the process. The advantages of this method in relation to the existing procedure for the flotation process are the following: • The method is unexpensive and losses are reduced to the time necessary to collect the data and record them in the computer. This method provides a constant process control and a quality decision which is better than those brought by an ex- pert with great experience ("hit-and-miss method"). • The method also provides incidental data about the effects. • The method should not be considered as a procedure for solving an incidental problem ("crash program"). • The obtained results must be interpreted by an appropriate expert in order to be considered instructive. MATHEMATICAL STATISTICS Taking into consideration the complexity of the enrichment processes it's necessary the application of "black-box" and mathematical modelling. The presence of the indefinite elements compel us towards the both theory of the probability and mathematical statistics. The mathematical statistics by the mineral processing investigation of the valuable raw materials in the last years has had greater signifance and widely spread application. The application of the mathematical statistics methods for the technological pro- cess analysis and the construction of the mathematical models is bound up by the task putting of the automatic control system de- sign (Fisher R. A. 1941, 1949). The statistical analysis in the enrichment processes is applied formerly, especially in the sampling estimation. Contemporary, in the last years is imposed the tendency of mathematical statistics application for the obtaining of quantity characteristics and quantity valuation from the individual influental factors. Above mentioned tendency in the significant level is based by the mathematical statistics development as a mathematical discipline close connected with the technical cybernetics general joining to the investigated object as a "black-box". Another essential factor in the mathematical statistics development is appeared the created mathematical theory of experiments as a new, to some extent, autonomous statistical domain. The theory of experiments has demonstrated exclusively importance for the investigation of the complex more-factorial processes in the chemistry, mineral processing, metallurgy etc. The investigation by the enrichment processes of the valuable raw materials and their management is possible by the statistical method application. In the same time, it's possible to use the wide-spread traditional methods of the statistical analysis, such as the methods of the dispersion, correlation, regression etc. Also, it may be mentioned above all the methods of the statistical planning of the extreme experiments. The more perspective are representatived some cybernetic methods which are based by the probability as an approach in the process analysis, for example, the identify of the ways and forms, the graph application as an additional instrument in the correlation analysis, the dynamic programming etc. The mathematical statistics application in the laboratory investigation scale is connected above all by the experimented material analysis and the compact representation of the obtained results. The mathematical statistics methods are more varrying and may be applied in the investigations of the wide circle of problems

Jan 1, 1996

By S. L. Bessinger

Introduction Extensive testing in the past ten years has shown that where a uniform natural gamma background is present in the strata bordering a seam, the thickness of the boundary coal left in place after mining can be determined by measuring the attenuation of that radiation (Nelson and Bessinger, 1989). Measurements made by the authors in underground mines in Pennsylvania, West Virginia, Ohio, Illinois and Kentucky and by others in Wyoming and New Mexico have shown the presence of such a gamma background (Nelson. 1989). Natural gamma coal-thickness sensors of several configurations have been tested in mines owned and operated by the Consolidation Coal Company (Consol) in Pennsylvania and West Virginia (Nelson and Bessinger, 1988). This paper describes the installation of a natural gamma coal-thickness sensor on an operating continuous miner. Previous tests had shown that the NGB-1000 coal-thickness sensor manufactured by American Mining Electronics, Inc., of Huntsville. AL, is an accurate, mine-worthy instrument. This large gamma detector consists of a sensing head and a control panel. The sensing head contains thallium-doped, sodium iodide scintillating crystal, which is coupled to a photomultiplier tube. The control panel contains the electronic components required for calibration, count conversion and display to the operator. Methods Conditions at a Consol mine in northern West Virginia require that 10 to 15 cm (4 to 6 in.) of coal be left at the roof boundary of continuous miner development sections. This roof coal is required because the shale of the immediate roof is friable and unstable. In the past, operators have used a dirt band that is usually visible near the top of the seam as a guide in maintaining the proper cutting horizon. However, this is not always reliable. Earlier observation showed that the actual thickness of the coal left on the roof varied widely; further, it was noted that occasional, accidental excursions into the immediate roof required supplementary roof control measures, such as installation of planks or center bolts. Thus, it was concluded that operators needed a better source of guidance for control of the cutting horizon, and a roof-coal thickness sensor was scheduled for installation. The NGB-1000 sensor was installed on a Joy 12CM10 continuous miner in June 1988. The sensing head was mounted on the cutter boom of the miner, and the control panel was mounted in the operator's cab. Power for the sensor was initially derived from an intrinsically safe battery power supply. Initial measurements with the sensor showed that the calibration was the same as that used in earlier tests at two other mines, indicating the uniformity of the natural gamma background above the Pittsburgh seam. Operating personnel were initially skeptical of the instrument's accuracy, and were hesitant to use its readings as a guide in maintaining a proper cutting horizon. Because gamma attenuation, the instrument's operating principle, is somewhat abstract, attempts to demonstrate the instrument's accuracy by explaining that principle were generally ineffective. It was found, however, that an operator could usually be convinced of the usefulness of the instrument by placing a large piece of coal of fairly uniform thickness over the instrument's sensing head and allowing the operator to see that the instrument reading increased by an amount very near his estimate of the thickness of the piece. The mine was provided with seven battery power supplies and a charging station. The charging station was kept in the lampman's office, and the mechanic on each shift was instructed that he was responsible for two battery power supplies each day: a freshly charged one to be taken in at the beginning of his shift and a depleted one to be brought out at the end. This system worked well for a few weeks, but eventually some battery power supplies were left in use so long that their batteries were discharged too deeply to allow recharging. In addition, transport and recharging of the batteries represented an additional task for the mechanics, who were already very busy. Consequently, a request was filed with MSHA to allow the sensor to be powered through intrinsic safety barriers by an electronic power supply connected to machine power. The permit was granted, and the sensor was connected to machine power. After the sensor was connected to machine power, the only operating problem experienced was occasional failure of cables. A supply of the required cables was made and delivered to the mine so damaged cables could be quickly replaced. Much of the cable damage could be eliminated by slight modifications to the miner during a rebuild, so that cables could be installed in more protected locations. After the sensor had been in operation for about two months, a survey was made to determine its effect on continuous miner operations. In previous research, coal thickness measurements made in 88 locations by the natural gamma method were compared to measurements made in the same locations by observing drill cuttings and by inspections of drill holes with a borescope. That research showed that the gamma method is at least as accurate as the other two methods (Nelson and Bessinger, 1989) and is also much easier to use. The object of the survey described here was not to assess the accuracy of the natural gamma measurements. but rather to determine the effectiveness of the sensor output as a guide for the operator in maintaining control of the cutting horizon. Thus a smaller, hand-held gamma detector

Jan 1, 1992

By Charles T. Steuart

The Arkansas Novaculite of Devonian and Mississippian age is the most distinctive formation in the central Ouachita Mountains of Arkansas from the standpoint of both topography and lithology. The formation typically forms precipitous, narrow ridges (hogbacks) which reflect the steepness of dip and the resistance and thickness of the rock units. The Arkansas Novaculite varies from less than 200 feet thick in the north to locally over 800 feet thick in the south and outcrops over an east-west distance of 360 km (200 miles) in the Ouachita Mountains of Arkansas and Oklahoma. Most of the rocks of the formation are highly siliceous with novaculite predominating, and lesser quantities of chert, siliceous shale, conglomerate and rarely sandstone. Novaculite is commonly defined as a homogeneous mostly white rock, translucent on thin edges with a dull to waxy luster and composed almost entirely of microcrystalline quartz. The word novaculite comes from the Latin word novacula, meaning razor hone. This term was first used by Richard Kirwan in his mineralogy text of 1784. The origin of novaculite remains controversial but we believe it was a primary silica derived mostly from the alteration of submarine volcanic-rich materials, with numerous fine biogenic components and some minor clastics (derived mostly from foreland facies via turbidity currents) and deposited as an amorphous siliceous ooze in the deep Ouachita trough. It was subsequently converted to microcrystalline quartz during diagenesis by the compaction of the overlying rocks and the intense Late Paleozoic Ouachita Mountain tectonism. In the central and southern portions of the region three divisions of the formation have been defined: a Lower Division predominantly of light-colored novaculite; a Middle Division of interbedded dark chert and shale; and an Upper Division of massive often calcareous novaculite. In the northern exposures the formation is thinner and contains more black chert and shale, chert-shale conglom¬erate, sandstone and less novaculite. The massive novaculite of the Upper and Lower Divisions is a source of high purity silica (99+%). Indians were the first to use novaculite. They made their various tools and hunting implements from the weathered broken rock until they learned the quarrying techniques. The first white men in the area recognized the use of novaculite for sharpening their tools and knives: previously whetstones were obtained in small quantities from Europe. To this day whetstone is quarried mostly in areas which the Indians and possibly the Spanish previously worked. Some of the current mining practices have changed very little from those of the early quarryman. In areas where the novaculite of the Upper Division is poorly cemented, this friable tripolitic material has found applications in such uses as an abrasive, as a filler or extender in paints, plastics and other purposes. Some of the dense novaculite is suitable as a filler in fire bricks and also in plastics. Because some of it has the tendency to decrepitate when heated not all novaculite can be used for these applications. Several quarry operators are crushing novaculite for road, concrete, and other aggregate. This use was formerly rather limited due to the abrasiveness of the novaculite on crushing and screening equipment. Further potential uses of novaculite are speculative but include application of high purity silica in the production of silicon metal for solar cells, seed or "growing", and fusing quartz for electronics.

Jan 1, 1985

By Andrzej Frycz

The development of the theory of reliability creates more and more possibilities of its application in different domains of technics. The methods evolved to date, both theoretical analysis of the engineering systems and investigations techniques in practice can be put in use a d translated from one science to another one [1,3,7J . It needs to consider the specific of the different sciences. But it is necessary to take account of the specific characteristic of each line of science. The reliability investigations needs, justified in technical and economical viewpoint appear too in mine ventilation technics and systems. The mine ventilation system is a complex of workings and ventilation plants, which serve to supply the networks with a sufficient quantity of fresh air from the atmosphere and then carry off the vitiated air. The condition necessary for a ventilation system to perform definite task is the correctness of all its elements, that is a capability to preserve in a expected range of tolerance, the values of basic parameters. The element of a system which ceases to work correctly is called failed. The basic parameter of all elements of a ventilation system is the characteristic curve, pressure loss - air quantity relation ?p = f(Q) (1) This curve concerning the workings and ventilation stoppings is given by the formula: ?p = R Q2 (2) thus the resistance R (of the workings or stoppings) may be taken as a characteristic parameter. The course of this characteristic f(Q) will be given in a complex form in concern with fans, ejectors, air curtains. The failure of a ventilation plant or working is therefore an event, relying on a change of its characteristic, thus for a definite time the losses of the plant's or working's capability set in for the fulfilment of its function in the ventilation system. The failures are divided in viewpoint of their nature on sudden and progressive. We are in concern with the sudden failures which are characterised by a violent change of the ventilation element parameter for example a fall of roof, or silt up of the working, opening or failure of the ventilation stopping, the stand-off of the fan or ejector etc. We distinguish two kinds of failures on the viewpoint of their action on the air flow: the "opening" failure and "closure" failure. The prevalence of factors, which causes the failures is in connection with the intersection in the mine of roads of haulage, roads of underground transport of materials and lining elements, roads of people motion with airways, as also depends on the rock pressure reaction on the workings lining and plants placed there. Other factors will yet be decisive upon supply decay of fans, ejectors and curtains. The appearance of failure of the ventilation element is therefore of random nature and determines the change of the element, when it is passing from operating state to unoperate. Qualifying the ventilation system element to one of these states is easy, when the characteristic change is clear and occurs suddenly, that was assumed. The investigations conducted by Patrushev and Karnaukh [91 in concern with the ventilation stoppings resistance change in the case of their opening, had shown the possibility of resistance value change in the range of several hundreds higher.

Jan 1, 1980

By Thomas F. Stinson

Energy-producing states see severance taxes as attractive sources of increased revenues for otherwise tight budgets. Consumers in energy-poor states, however, protest that current rates are excessive and have moved to obtain a reduction. The US Supreme Court heard two cases questioning such taxes during its 1981 term, and Congress, in each of its last two sessions, has seriously considered measures placing a ceiling on severance tax rates. Although discussions usually focus on coal severance taxes levied by Montana and Wyoming, ramifications of changes in national policy are much larger, potentially affecting mineral taxes in all states. Severance taxes yield substantial revenues in some states, as illustrated in the accompanying table. In fiscal 1980, Texas collected more than $1.5 billion from such taxes, while Louisiana, Alaska, and Oklahoma collected $525 million, $507 million, and $436 million, respectively. Oil and natural gas production were the major items taxed. While the table shows only the 10 states with the largest severance tax collections, a total of 17 states had severance tax receipts exceeding $10 million in 1980. The enormous increases in severance tax revenue between 1970 and 1980 can be seen from the table by comparing columns A and B. For the 10 states shown, revenues increased from $650 million to $3.8 billion during that period, due to a combination of higher prices, increased production, and new or increased taxes. In most states, increases in prices and production were more evident than higher tax rates. Although the revenue increase is impressive, it understates the extent of changes occurring during the past decade. Since tax collections are shown for fiscal 1980, the full impact of the decontrol of domestic oil prices is not evident. Also, taxes on natural gas reflect controlled prices. When full de-control of natural gas occurs, significantly higher revenues can be expected in gas-producing states, even if tax rates, production, and world prices remain constant. These increases in severance tax revenues, coupled with further increases expected to accompany natural gas decontrol, have produced fears that a substantial transfer of income is taking place from residents of energy-poor states to those living in energy-rich areas. Others worry that increased severance tax revenues will be used to substitute for a substantial portion of state taxes that would have otherwise been collected from firms and individuals. The resulting improvements in tax climate, it is thought, may encourage firms and workers to migrate to energy-producing states. While these issues are often portrayed as pitting energy have-nots against the haves, such is not entirely the case. Much of the original impetus for limits on coal taxes came from Texas residents, the state collecting the most severance tax revenue. One way of gaining perspective on these concerns is by examining severance taxes per capita and as a percent of state tax collections (see columns C and D in the table). When this is done, Alaska, where 1980 taxes on oil and gas production totaled more than $1300 per capita (and 35% of all state tax collections), stands out as a special case. There, the state was able to eliminate some individual taxes, place revenues in a permanent trust fund, and consider offering a rebate of some previous income

Jan 4, 1982

By H. E. Epstein, B. K. Parekh, W. M. Goldberger

Comminution is the single most expensive operation in mineral processing. It consumes about 50% of the energy required for mineral extraction (Agar, 1976). Current comminution technology is both energy-intensive and inefficient. A novel noncontact comminution process concept was developed in this study, whereby selective lib¬eration of minerals from an ore could be potentially achieved. The process involves application of electric and ultrasonic energy to liberate minerals from gangue particles. Introduction The mineral industry is a large user of energy in the US. Energy usage is increasing as the grade of ores processed decreases. It is es¬timated that comminution of ores uses about 32 000 kWh (115 trillion kJ), or 2% of the electric power produced in the US (NMAB-365, 1981). Among other minerals, ce¬ment, iron, and copper processing plants are the largest users of en¬ergy in comminution. Very little of this energy used in conventional grinding (about 1%) is used to gen¬erate new surface. The remainder is wasted (Table 1). Thus, there are substantial energy-saving and economic incentives to improve the efficiency of crushing and grinding techniques for mineral recovery. With an energy efficiency of only 1%, it would seem possible to devise methods to significantly improve comminution technology. This requires that breakage force be applied only where needed, not indiscriminately as in a conven¬tional ball mill (NMAB-365, 1981). Can apparatus and methods be de¬ veloped for large-scale commer¬cial use that allow energy to be fo¬cused at intergranular bounda¬ries? If this can be done, mono¬mineral grains would remain in¬tact and the grinding action would be selective and substantially more efficient. The novel comminution process concept described in this paper uses a combination of electric and ultrasonic energy. This energy breaks the ore and selectively lib¬erates minerals. The process has also been termed a two-stage or electroacoustical comminution process (Goldberger, Epstein, and Parekh, 1982). The mineral grain boundary is usually the weakest area in an ore. By applying electri¬cal energy to the ore, the rock frac¬tures mostly at grain boundaries. At the same time, the electric shock creates secondary hairline fractures in the ore. Additional application of ultrasonic energy to this ore provides further break¬age. The concept was proven on a molybdenum porphyry ore but needs additional study. Technical Background There have been substantial im¬provements in milling machinery and grinding operations, but rela¬tively few attempts have been made to develop alternatives to conventional impact milling. Sev¬eral methods, however, have been investigated that relate to this study. In the early 1970s there was considerable interest in a size re¬duction method known as the Snyder Process (Cavanaugh, 1972). It involved charging coarse ore into a pressure chamber, pressur¬izing with a gas, and activating a quick-opening (15 msec ) dis¬charge valve that connected the chamber to a discharge duct. This allowed solids to fluidize and ac¬celerate, subjecting the material to a variety of impulse phenomena that caused the desired size reduction. Other nonimpact means of achieving breakage and selective size reduction have been de¬scribed in technical and patent literature. Kanellopoulos and Ball (1975) considered using heat to induce thermal stress in quartz¬ite. This would cause cracking and, thus, increase grinding efficiency. The use of electrical energy to induce thermal stress was studied by the General Elec¬tric Co., in cooperation with the Montana School of Mines, under a research grant from the Anaconda

Jan 9, 1984

By S. P. Chong, Peter V. Avotins

The iron and steel industry continues to operate at depressed production levels due to the low level of demand for steel and foundry products. Crude steel production in the Western world dropped by 13.7% from 459.3 Gt (506 billion st) in 1981 to 396.4 Gt (437 billion st). The reduction was particularly severe in North America where the industry suffered a loss of 37.3% from 123.6 to 77.5 Gt (136 to 85 billion st). This reduced level of production severely affected iron ore and agglomerates output. Substantial curtailment was made in both sintering and pelletizing plants to meet reduced demand. The pressure on these plant operators shifted from quantity to quality. With energy cost becoming an ever increasing part of production costs, conservation was a major area of emphasis. Iron Making Sinter has continued to be the favored burden for blast furnaces overseas. As hot metal production declined and excess sintering capacity developed, many of the overseas operators that have been using pellets found it advantageous on hot metal production cost to increase the ratio of sinter at the expense of pellets. This was the case despite the depressed price level of pellets in the world market. In North America where pellets have been the predominant burden material-mainly because it is necessary to pelletize the fine grain iron ores available domestically-there was also increased interest in sinter usage. Bethlehem Steel Corporation's 13.6-m (44.5-ft) furnace at Sparrows Point experimented with a burden with more than 70% of sinter during the second half of the year and indicated some favorable results as compared to its previous 50% super flux sinter and 50% acid pellet burden. The blast furnace operators continued to press for improved burden quality. In addition to the emphasis on the importance of burden chemistry and quality, uniformity, much emphasis was placed on improving reducibilitiy and high temperature metallurgical qualities. These burden properties were recognized to hold further promise in reducing fuels rates. Sintering With reduced consumption caused by the cutback in hot metal production, many sinter plants sat idle while others operated at greatly reduced rates. The deemphasis on production in many plants allowed them to experiment with various means to reduce operating costs. The most notable action was the introduction of low cost South American pellet feed fines into the sinter ore mix. These feed fines, which are finer than 150 µm (100 mesh) ore, were previously thought to be too fine for sintering. However, when the full production rate is not required the use of up to 40% in the ore mix was reported to be successful with the aid of burnt lime and adequate premixing and granulation of the sinter mix before sintering. The emphasis on quality and tighter cost control also has resulted in a further breakthrough in sintering technology. It was recognized that good feed preparation-including selection of compatible ores, good blending, good pseudo-particle formation, additive size control increased bed depth, and closed control of the sintering bed heat pattern-could allow the production of high quality sinters with much lower heat input and FeO level. Bed depth of more than 600 mm (23.6 in.) was practiced in a number of operations. Total heat input of better than 1.3 GJ/t (290,300 kcal per st) has been achieved. Kawasaki Steel's Chiba Works has also achieved a 5.3% SiO2 in the sinter with good quality. The low heat input approach allowed production

Jan 5, 1983

By Shan Mafi

he Broken Aro coal mine is located at the Woodbury Wildlife Preserve 11 km (7 miles) west of Coshocton, OH on State Route 541. Underground min¬ing at the 16-hm2 (40-acre) site took place during the 1910s. During the past five years, surface mining has occurred along the western and northern perimeters of the underground region. The Ohio Department of Natural Resources (ODNR), R&F Coal and American Electric Power (AEP) are involved in a project to remine the abandoned underground complex. The program also includes installing a mine seal made of fixated flue gas desulfurization (FGD) material. Previous mining activity in this area has resulted in an acid mine drainage (AMD) problem on portions of the reserve. In particular, a prominent AMD source is present in the Simmons Run drainage. The project goal is to eliminate or reduce the AMD that is discharging into the watershed. It was initially hoped that the Simmons Run drainage area could be economically mined. Accordingly, this area was included in the mine permit. However, experience with other previously mined areas on this reserve showed that the old mine workings were more extensive than originally thought. Old mine workings present several problems. The most significant are reduced coal recovery, poor coal quality and AMD liability. All of these problems increase mining costs. And the more extensive the old workings, the more significant the increase in cost. Based on R&F's experience with old mine workings in this area, it was decided that economic recovery of the Simmons Run drainage reserve was not possible. As a result, R&F excluded the Simmons Run drainage area from its early mine plans. Some time after it was decided not to affect the Simmons Run drainage area, R&F and ODNR's Division of Mining and Reclamation discussed the possibility of removing this area from the permit and attempting to correct the AMD problem as an Abandoned Mine Lands (AML) project. Specifically, the discussion centered around mining through a portion of the old mine workings to expose the mine openings in the highwall. Once exposed, the mine openings would be sealed using FGD material from the nearby Conesville power plant operated by AEP. In addition, there was interest in reclaiming the old tipple and associated gob pile located adjacent to Coshocton County Road 17 on the opposite side of Simmons Run. R&F attempted to eliminate the AMD problem by sealing the mine openings with FGD material. Once the seals were installed, they were covered with overburden. Monitoring wells were installed to determine the effectiveness of the seal and the inundation progress in the old mine workings as well as monitoring the quality of ground water. The project was jointly funded by revenue from the sale of the coal recovered during the project and AML funds. The FGD seal is expected to cause well water levels to increase above preseal elevations. However, the additional head pressure due to increased water depths in the underground mine may cause seepage through the bottom of the mine or out through the perimeter. The in¬creased water depth in the underground mine is expected to cause lower oxygen concentrations in the underground mine water. Since acidity is produced by oxidation of iron pyrites, less oxygen should result in better quality water. Monitoring wells will be installed before seal installation so that pre- and postseal water levels and well water quality can be determined. Surface sampling locations will be identified. These will be monitored before and after seal placement.

Jan 1, 1998

By Stanislaw Siewierski

One of the major problems in the exploitation of deep copper ore deposits in the Legnicko-Glogowski Copper Field is ground control and, in particular, roof control of the headings. The depth of copper deposits ranges from 600 to 1200 m. The parameters of ore-bearing rocks and the surrounding rocks show considerable variance of strength properties along the mining thickness (Fig. 1). The roof of the openings exploitation headings consists of dolomite-calciferous rocks. The roof strata consist of layers, the thickness of which vary from several centimeters to 0.6 m (sometimes to 1.0 m). These strata differ in strength and are characterized by igneous and sedimentary discontinuities. On the basis of studies in structure, properties and geological mining conditions of the roof strata in Polish copper mines, the method of roof classification has been developed. (1) This method uses the stability index "L" of the roof, calculated from [ ] The value of "t" ranges from 0, l to 0,8. Based on the value of "L", four classes of the roof (Table 1) corresponding to different degrees of fall hazard have been developed. The mining of copper ore deposit is carried out by a single lift room-and-pillar mining system or a two-stage room-and-pillar system with roof caving. (3, 4). The single lift room-and-pillar mining system is most often used. It consists in working of the deposit of 2.5 to 4.5 m mining thickness, 150-350 m long mining line, consisting of parallel room faces 6 m wide with pillars 5.7 or 16 m wide. As the mining line advances, cross-cuts joining individual rooms are made, thus forming support pillars 3x5, 7x7, 5x16, and 7x16 m (Fig. 2). The two-stage system (Fig. 3) mines the deposit with rooms 6 m wide into blocks 25 x 35 x 25 x 45 m. During the second step these pillars are worked with rooms and cross-cuts into pillars about 3 x 4 to 4 x 5 m (retreat mining). The pillars are crushed just before the fall of the roof. The maximum opening of workings in the two-step system is about 10 m, i.e. one belt entry and one row of bearing pillars. In single lift room-and-pillar systems it is about 35 m., i.e. three belt entries and three rows of bearing pillars. The support of the headings consists of expansion bolts 1.6-1.8 m installed in a bolting pattern 1.1 x 1.5 m. In both systems the fall step is about 10 m, i.e. the width of a belt and a bearing pillar. The Indications of Rock Mass Pressure Individual Dog Heading Individual dog headings are set wide apart from each other in the unmined ore deposit area where there is no interaction of headings. The indication of rock mass pressure are, in such cases, limited to the deformations and damage of side walls. Radiometric studies of rock mass loosening showed that the maximum range of lifts in side-walls immediately after web was 2 m (4,5). In the course of time the cracked side rocks fall off, thus widening the heading. The morphology of side-walls is influenced by the location of the heading in relation to clump layer/tarry. This layer is characterized by a weak strength and "lubricating" properties. The location of the heading in

Jan 1, 1982

By M. E. Rossi

Reconciliation of long-term, block-model reserves and mine production frequently shows significant discrepancies. This reconciliation can be done comparing the long-term, block-model reserves against a grade control model (a model based on blast holes or production data). In this article, this is called an Fl factor for tonnages, grade and ounces. This reconciliation should also be done by comparing the grade control model to the mill feed (or loaded to heaps) tonnages, grades and ounces (F2 factor). The long-term model usually contains blocks with dimensions larger than the size of the selective mining unit (SMU) of the operation. The smoothing effect of an estimator, such as kriging, will generally result in a grade-tonnage curve that does not match the grade-tonnage curve of the SMUs to be recovered. This problem is partly to blame for the reconciliation discrepancies found. In addition, in-pit perfect selection is impossible. Therefore, the grade-tonnage predictions based on blastholes also need to be assessed - considering unplanned dilution and the errors of estimation of the grade control model. This article focuses on using conditional simulations to assess the uncertainty of the predicted grade-tonnage curve in terms of minable reserves and in the framework of long-term mine planning. Quantifying the uncertainty in no uncertain terms is done in conjunction with a popular pit optimizer. It yields several possible minable reserves and several possible minable tonnages and grades, per pit shell or scheduled phase. This is a step further from the usual sensitivity analyses that are performed for deposit evaluation or development planning. Specifically, this approach takes into account the volume-variance relationship, the selectivity of the mining operation and the planned and unplanned dilution. It also carries the local uncertainty implicit in grade estimation into the mine design and mine planning phases. The estimation of recoverable reserves revolves around estimating the block distribution of metal grades from the drill hole information available, typically composites of a fixed length. Ideally, this estimation should accurately predict the ore sent to the processing facilities of an operation. Here are some definitions used in this article: ? Recoverable reserves are those where an attempt has been made to estimate the actual distribution of grades (usually by a block model) that would be found in the ground at the time of mining, relative to the mine equipment size and other operational characteristics of the mine SMU. In particular, it incorporates the effect of internal dilution for each block how the grade distribution changes in variability when comparing a composite grade distribution to the "mining unit" grade distribution. In geostatistical jargon, this is known as "change of support effect," or volume-variance effect. ? Minable reserves are those recoverable reserves defined by a mine plan. In an open-pit scenario, they would incorporate pit optimization in addition to the sequencing and scheduling of production through mine life. ? All other block model estimates from drill hole data that do not explicitly attempt to estimate or measure the volume-variance effect are called in situ. It can be argued that these models may be inappropriate tools for predicting the grade-tonnage curve of the material to be sent for processing (minable reserves). The purpose of estimating recoverable reserves can be thought of as estimating the appropriate amount of dilution that should be incorporated into the model so that it adequately predicts (long-term and on average) the grade-tonnage curve of the ore sent to the processing facilities (mean grades and ton-

Jan 1, 1999

By Prescott Greene

AVAILABLE MACHINES Machines discussed in this section include only rubber tire mounted kerf cutting machines of a self¬propelled electrically powered design. Such machines handle bars up to 4.6 m (15 ft) in length, cutting kerfs of from 115 to 191 mm (4.5 to 7.5 in.). The basic design concept of all machines currently in production is very similar. The chassis features dual driving wheels on the cutter-bar end, fixed directly to the main frame. The steering wheels on the outby end of the frame are articulated in some manner to keep all four wheels on the floor over uneven bottom. Machine power for tramming, and all other functions except powering the cutting chain, is furnished from a pump motor that drives the specific pump or pumps required by a given machine. The cutter chain is driven through a mechanical drive train from a separate electric motor having no duty other than driving the chain. The general arrangement of a cutting machine is shown in Fig. 1; two types of machines, with their specifications are shown in Figs. 2 through 5. A hydraulically powered cable normally is included to eliminate the need for manual cable handling and make this a one-man machine. A bug duster, though optional, is needed for most applications. The bug duster is an hydraulically driven scroll conveyor located directly behind the cutter chain drive sprocket. Its function is to move the cuttings away from the sprocket to prevent them from packing up in this area and consuming a lot of machine power to further pulverize them. Available machines vary from models 610 mm (24 in.) high that can cut horizontally only from 203 mm (8 in.) below ground line to 229 mm (9 in.) above, to machines 1143 mm (45 in.) high that can cut in any plane from 178 mm (7 in.) below the ground line to 3.7 m (12 ft) above it. Most currently manufactured machines fall in the 17 200 to 20 400 kg (38,000 to 45,000 lb) weight range, but can be dis¬assembled for underground installation where weight and size are a problem. Maximum teardown results in all pieces having a weight of 5400 kg (12,000 lb) or less and a size allowing passage through a 1.2 x 2.1-m (4x7-ft) opening. Further reduction in either size or weight of the parts becomes a serious problem. Machines currently being manufactured are avail¬able for use with power supplies of 250 v dc; 500 v dc; 440, 550, or 950 v ac (60 Hz). Also available are machines for use on 415, 500, and 950 v 50 Hz. De¬signs to allow operation on other voltages of either 50 or 60 Hz can be made available without any serious problems. As of September 1977, a typical medium-seam ma¬chine without the extra options needed for unusual conditions, but fully equipped to operate, was selling for about $185,000. CAPABILITIES Materials That Can Be Cut The problem of deciding before a machine is put on the job whether it can work effectively is an old one that has never been completely solved and probably

Jan 1, 1982

By Stephen H. Stow

Economically important heavy minerals have been studied in the sediments of offshore Alabama and Mississippi. Sandy samples have been collected from a 200 square mile area from Dauphin to Horn Islands. The minerals concentrate in sediments under 10-35 feet of water and may reach 2.4% by weight. The distribution of total heavies is related to the hydrodynamics of offshore Alabama-Mississippi; the greatest concentrations are in the Gulf where the longshore current is diverted from its westerly flow. Preliminary observations indicate that the distribution of individual minerals can be correlated with wave energy, particle size, specific gravity, and morphology of the mineral grains. The coarser, lighter minerals--kyanite, sillimanite and staurolite--are concentrated in relatively low energy areas in the Mississippi Sound and at the end of Dauphin Island in the Gulf; maximum concentration for each is 34.5%, 6.0% and 43.2%, respectively, of the heavy minerals. The smaller, more denseminerals--ilmenite, rutile, leucoxene, zircon and monazite--are concentrated in high energy areas in the Gulf; maximum concentration for each, respectively, is 42.0%, 5.9%, 9.652, 10.7% and 1.3% of the heavy minerals. On the basis of a minimum concentration of 1% heavy minerals, two areas have been identified and the total amounts of different minerals have been calculated to a depth of 25 feet of sediment. The data are preliminary and in no way indicate recoverable reserves, but they do show that on the order of 20 million tons of heavies exist. These deposits should be termed "indicated resources".

Jan 1, 1975