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By Michael Johnston

Nautilus Minerals is exploring for high-grade gold, silver, copper, and zinc-rich submarine massive sulphide deposits (SMS) within the territorial waters of the sovereign state of Papua New Guinea. Work to date has shown that these systems contain spectacularly high metal grades. To date a total of 88 surface sulphide grab samples have been collected from Nautilus’ Solwara 1 Prospect within EL 1196 in the Manus Basin, returning average grades of 15.5 g/t gold, 191g/t silver, 10.8% copper, and 4.7% Zn. Drilling in January/February 2006 confirmed the presence of this high grade system, with sulphide intersections up to 19m below the sea floor, with a best intersection of 11.6m at 9.1 g/t Au and 13.1% Cu.

By Dave Thomas

?Uranium is a common trace metal at the earth?s surface, but these are amazing super-high grade (up to 25%) pods, veins and disseminated uraninite (UO2) near unconformities between Proterozoic siliciclastic red beds and metamorphic basement with graphitic metapelite and radiogenic granite. ?Pitchblende fills extensional features in reactivated fault zones and replaces matrix in sandstone ?One mining district in Canada ?the Athabasca Basin ->30 deposits /prospects -most in eastern ¼of basin -produces 1/3 of world?s U -employs northerners -VERY high tech -VERY high safety standards Unconformity-associated uranium depositsUnconformity-deposits

May 1, 2006

By J. B. Gayle

Numerous investigators have studied the addition of inert4 substances to coking blends. Generally, the results have yielded trends applicable only to the particular coals and inerts studied. Recently, somewhat more general results were obtained for cokes produced from blends containing added inerts, using the sole-heated oven. The present investigation represents an extension of this work to include a wider range of inerts and to determine the applicability of the various conclusions to cokes produced with the Tuscaloosa oven. The results indicated that an increase in dry-coal charge density of 1 pound per cubic foot (p.c.f.) caused a corresponding change in the hardness (1/4-inch tumbler index) of Tuscaloosa oven cokes of 0.9 percentage points. This value compares favorably with the value of 0.7 determined previously for the sole-heated oven and Bureau of Mines-American Gas Association (BM-AGA) retort.

Jan 1, 1961

By Francis Clark

WHEN 60:40 brass is heated to 825° C., given a drastic quench to obtain the beta solid solution, and reheated, various changes take place in the structure. Reheating at 200° C. causes a fine, granular, sorbitic structure to appear which has been considered alpha solid solution. These particles coalesce at higher temperatures of reheating, and at 470° C., the alpha forms twins. A further study of these transformations was considered worth investigating. EXPERIMENTAL WORK The object of the experimental work was to study the effect of mechanical work on rolled 60:40 brass, quenched and reheated, and the mechanism of twinning in the alpha reeds of a furnace-cooled specimen. It was desired to determine the effect of mechanical work on the transformations taking place on reheating the quenched material, both at 200 and at 470° C. A study of the nature of the structure resulting from the breakdown of the beta at 200° C. and a study of the mechanism of twinning above 470° C. were considered problems for investigation. The question of the twinning of the alpha reeds was also an important phase of the work. With high-power magnifications, there was hope of following these processes.

Jan 1, 1927

By Tao Long

Recently, the great amount of funds has been invested in the reconstruction of damaged infrastructure and house of subsidence area of coal mine and mine reclamation. This is a benevolent and popular work, which is highly praised. However, because of complicated collapse area problems in China, especially the less integrated control research in the collapse area of metal mine, and the technological weakness, how to effectively control the subsidence area, restore landscape and ecological environment, and utilize the abandoned mine land, are major issue related to the mining area of sustainable development. The research situation about the formation of metal mine collapse area at home and abroad is reviewed, and then the effect of the collapse area on the mine safety and ecological environment, the existing technical problems and development direction are analyzed. Lastly, the technical measures of the collapse area control and cooperative underground mining is put forward.

Oct 1, 2010

By Carl E. Johnson

INTRODUCTION A method for predicting water-flood oil recovery was reported by H. Dykstra and R. L. Parsons' in 1950. It is now generally known as the Dykstra-Parsons method and is widely used by petroleum engineers. The method is semiempirical and consists of a correlation of four fundamental variables. These are: V, vertical permeability variation;' a, mobility ratio; Sw initial water saturation; and R, fractional recovery of oil in place at a given producing water-oil ratio. The correlation extends over a wide range in each of these variables, and may be applied to all formations with initial oil saturations of 45 per cent or greater. The method of calculation originally outlined by Dyhtra and Parsons closely follows their derivation. And this is a convenient way of illustrating the development of some rather complicated ideas. However, in actual use it is more cumbersome than necessary. The purpose of this note, therefore, is to provide a simplified method for making Dykstra-Parsons predictions. A SIMPLIFIED GRAPHICAL SOLUTION The correlation between V, a, Sw and R, corresponding to a given producing water-oil ratio, can be shown on a single graph. This is best done by plotting V against a to show lines of constant R(1 - S, . WOR-0.2). Figs. 1, 2, 3, and 4 show these plots

Jan 1, 1957

By S. S. Ibrahim

The left behind material after the combusion of fossel fuel in some power stations in Cairo; weighing tens of tons every years, was investigated. The material called "Fly Ashes" containing 7.5% Ni, 10.0% V and 9.50% Cr; could be considered as a promising source of such valiable elements. As well as the small grain size (050 70 "m) of the material, and besides the complexity of their formation history, hydrometallurgical methods appeared to be the effective technique to deal with. Yet, preconcentration is possible by means of magnetic and gravity separation methods. The end product is assaying 16.60% Ni and 20.2S!; V with weight recovery of about 60.15%.

Jan 1, 1995

By V. F. Parry

The author wishes to acknowledge the advice and criticism of A. C. Fieldner, chief, Coal Division, Bureau of Mines, United States Department of the Interior; the assistance of Durbin Van Law, consulting industrial engineer, Denver & Rio Grande Western Railroad, Denver, Colo., who furnished valuable information on trends of coal production in Utah by making available his unpublished report "Utah Coal Industry"; and the aid of J.C. Beltz, associate research chemist, Bureau of Mines, Washington, D. C., who reviewed and checked the data, and of W. S. Landers, assistant chemical engineer, Bureau of Mines, Golden, who assisted in the preparation of many of the statistical data, in this report.

Jan 1, 1942

By B. S. Butler

The principal ore deposits of Arizona are in the southern, cen-tral, and western portions of the state, which physiographically are part of the Basin and Range province, southwest of the Colo-rado Plateau (Pl. I). The Basin and Range province is characterized by numerous subparallel mountain ranges separated by plains or valleys. Most of these ranges trend northwest to north, parallel to the margin of the Colorado Plateau; but in southeastern Arizona, southern New Mexico, and northeastern Sonora, they trend northward, transverse too the edge of the Plateau. The mountains rise abruptly from plains or valleys, the margins of which in many places are pediments cut on hard rock. Some of the plains form closed basins (bolsons, playas), but most of them are drained. The Basin and Range province in Arizona is divisible into the Mountain Region and the Desert Region4 (Pls. I and II). The Mountain Region forms a belt 60 to 100 miles wide that contains most of the large ore deposits. Its longest range measures about 55 miles, the widest 20 miles, and the highest peak more than 10,000 feet above sea level or 7,000 feet above adjacent valleys or plains. Broad plain-forming valleys are exceptional, but several with maximum widths of 20 to more than 30 miles appear in the southeastern portion. In the Desert Region the mountains are relatively low, narrow, serrated, and steep, with sharply angular profiles. There are a few exceptionally large ranges of which the longest measures nearly 100 miles and the widest 18 miles. The highest peak is some 7,000 feet above sea level, and the maximum local relief amounts to about 5,400 feet. Broad and comparatively low desert plains form 50 to 75 per cent of the area.

Jan 1, 1939

By R. A. Basse, W. D. Diment, A. J. Petrina

INTRODUCTION In 1957, diamond drilling on a magnetic anomaly indicated an extensive zone of copper mineralization on what is now the Craigmont Mines property. By mid¬1958, drilling established a copper ore body. Milling commenced in September 1961 at 4536 t/d (5000 stpd) and by the end of October 1977 the mine had produced 339 662.04 t (374,363.9 st) of copper. At present, two-thirds of the mill feed is derived from underground operations and one-third from low-grade surface stockpiles. Craigmont Mines is situated 209 km (130 air miles) northeast of Vancouver (see Fig. 1), 16 km (10 miles) west of the town of Merritt, a logging, ranching, and mining community of about 7000 people. It is serviced by paved highways, Canadian Pacific Railway, British Columbia Hydro, and Inland Natural Gas Co. Water is pumped from the Nicola River, a distance of 6 km (4 miles) and a lift of 244 m (800 ft). In March 1967, the open pit mining operations at Craigmont Mines reached their economic limit and were suspended. Before this, it had been decided that a sub¬level caving method of underground mining would be used to supply ore to the concentrator after the cessation of open pit production. This chapter describes the fac¬tors influencing the choice of mining method, some of the problems encountered, mining practices, and results. GEOLOGY The ore bodies of upper Triassic age are located in a limy horizon striking east-west, closely paralleling the intrusive Guichon batholith, bounded on the south by rhyolites and on the north by graywackes, and dipping steeply to the south (Figs. 2a, b). The ore bodies are relatively narrow with a maxi¬mum width of 79 m (260 ft), a combined strike length of 853 m (2800 ft), and a vertical extent of 610 m (2000 ft). Chalcopyrite is virtually the only copper mineral, and 20% of the ore zone consists of acid solu¬ble magnetite and hematite. The area has been subjected to considerable faulting and brecciation, which is a major factor in the mining operation. Total geological reserves, at 0.7% Cu cutoff, for the deposit were 22 316 743 t (24,600,000 st) at 1.89% Cu. An additional 5 236 270 t (5,772,000 st) at 0.6% Cu were mined from the open pit. Ground Conditions The waste rocks-graywacke, andesites, and diorite -are relatively incompetent due to the high degree of fracturing and jointing, and all require varying degrees of support. The ore zones are somewhat less fractured; ground support is still required, however, although to a lesser extent than in the country rock. Ground conditions in the main ore body are better than in the smaller, nar¬rower ore bodies. Clayey fault gouge is present in most of the faults; gouge zones may be up to 6 or 9 m (20 or 30 ft) wide. The main ground problems are associated with local weakness rather than pressure. Shape of Ore Bodies (Figs. 2a, b and 3a, b) The main No. 1 ore body is approximately 244 m (800 ft) long and 46 m (150 ft) wide. It extends ver¬tically from the original top of the open pit at 4200 ele¬vation to just below the 3060 level. The No. 2 ore body is approximately 304 m (1000 ft) long, varies from stringer width at the extremities up to 79 m (260 ft) wide, and extends from 3060 level to 2400 level. Both these ore bodies have extensions re¬sulting in additional small irregular bodies. Ore bodies are mostly steep dipping, though part of the Wing ore body, an extension of No. 2 ore body, dips at 0.87 rad (50'). This ore body varies in size, but is approximately 122 m (400 ft) long, 21 m (70 ft) wide, and about 213 m (700 ft) high. No. 1 Limb ore body is a narrow extension of the No. I Main with a vertical extent of 137 m (450 ft), average width of 18 ft (60 ft), a strike length of 152 m (500 ft), and dips steeply at 1.4 rad (80°). No. 1 East is an eastern extension of the No. 1 Main with a vertical extent of 183 m (600 ft), a strike length of 91 m (300 ft), an average width of 30 m (100 ft), and dips at 1.2 to 1.4 rad (70 to 80°). No. 1 South is at the upper west end of the open pit with a vertical extent of 76 m (250 ft), a strike length

Jan 1, 1982

By Alfred L. Ransome

GOLD output in Alaska in 1950, following a 2-year decline, increased to the highest point since 1942. Tills unexpected gain in production of the Territory's greatest value commodity was largely responsible for the increase in the over-all value of mineral output in the Territory to $17,852,000 from $15,549,000 in 1949. Although coal continued to rank second to gold in value of output, production was slightly less than the record production of 1949. The mining of platinum continued to be an important factor in the mineral industry, with production of crude platinum metals exceeding that of 1949. The output of lead was three times the 1949 production, and silver gained 46 percent over the previous year's total. Copper and zinc production-entirely as byproducts from other ores-showed an increase, but remained minor. Production of tin was relatively small, but substantially above the total for 1949.

Jan 1, 1953

By Lester Uren

WITH the depletion of our older, and relatively shallow, oilfields and the necessity for securing new production from deeper horizons, much attention is being given to the improvement of oil-well pumps in order that they may function satisfactorily under the more difficult conditions imposed by increased depth. Until recently, practically all of the world's petroleum was secured from wells less than 3000 ft. deep, whereas horizons are now being explored to depths of from 5000 to 7000 ft. in some of the more prolific fields, and it seems probable that the greater part of the future supply must come from depths in excess of 3000 ft. While much of the early production in a new field is obtained by natural-flow methods, when the stimulus of high gas pressure fails, recourse must be had to some method of pumping. The perfection of a pump capable of economically and efficiently lifting oil from depths ranging from 3000 to 7000 ft. is therefore a matter of prime importance to the oil-producing industry. This paper is offered as an estimate of the present state of development of the oil-well plunger pump and as a review of the general principles involved in its operation in deep-well pumping. A design for a new type of oil-well pump embodying several novel features, is also described.

Jan 9, 1925

By Rex O. McLellan

The formation of a monovacancy in a metal is simulated in an elastic model by the displacement of the surface of a small spherical cavity in a large elastic continuum. The application of linear elasticity to this distortion results in a well- known formula for the energy and an expression for the concomitant entropy change due both to the shear strain in the continuum and also to the dilation of the solid resulting from the boundary conditions at the surface of the solid. Elastic data (the sliear modulus and its temperature coelficient) are used to calculate the entropy and energy of formation for many metals. Despite the simplicity of the assumptions involved, the agreement between the calculated entropies and energies and experimental values is remarkably good. In recent years there has been a large increase in measurements of the absolute concentration of mono-vacancies in metals as a function of temperature. Hence new data for both the energy and the noncon-figurational entropy of formation of monovacancies has become available. Recent measurements&apos; of the anomalous (non-Arrhenius) self-diffusion in many bcc metals has also focused interest on the prediction of the thermodynamic parameters of mono- and multi-vacancies in those metals for which no data are available. Damask and Dienes&apos; have discussed the various theoretical calculations of the energy of formation EL, of a monovacancy. These include simple models involving the breaking of atomic bonds on moving atoms from the interior of a crystal to the surface, models combining elastic calculations with surface-energy terms and detailed quantum mechanical calculations. The simler models give the correct order of magnitude of &, but tend to overestimate it by a factor of about two. The quantum mechanical calculations4"7 have been carried out for the noble and alkali metals with generally reasonably good agreement with the available Ef data. The calculation of entropy of formation Sfv14 lnvolves a fundamental calculation of the perturbation of the phonon spectrum caused by the creation of a vacancy. Huntington, Shirn. and wajda8 have given an approximate evaluation of sJV by considering an Einstein model for the localized vibrations in the immediate neighborhood of the defect and then using elastic theory to calculate the entropy associated with the shear stress field in the distorted crystal (as originally proposed by Zenerg). They also included a term due to the dilation of the crystal. They obtained a value of 1.47k for copper, in good agreement with the experimental value (1.50k). However, Nardelli and Tetta- manzi1° have recently shown that neglecting the coupling between atoms (Einstein Model) may lead to a serious error so the agreement may be somewhat fortuitous. In this work simple linear elastic theory is used to calculate the entropy and energy of formation of mono-vacancies. Despite the simplicity of some of the assumptions involved, the agreement with the available experimental data is remarkable. However. the reasonable degree of success in the application of linear elastic calculations to the excess entropy of a solute atom in a dilute solid solution1&apos; indicates that the application of elastic theory to vacancies. where the interaction of different atomic species is not involved, may not be inappropriate. THE ELASTIC MODEL The metal is assumed to be a spherical elastic continuum. A small spherical cavity of volume V = 4i;v:&apos;/3 is cut from the center. removed. and dissolved rever-sibly in the bulk of the material. TO a good approximation no net atomic bonds are broken and the material does not undergo a volume change although the externally measured volume of the body would increase by V. The radius of the sphere of metal is much larger than r Next a negative pressure is applied to the cavity causing its surface to be displaced inward by an amount simulating the relaxation of the lattice around a monovacancy. In this model the energy and entropy accompanying the distortion are taken as 4, and <. As a first approximation the equation of state for the solid is taken as: r = ro(i + *~D LiJ where K is the bulk modulus. P the hydrostatic pressure. Vo the volume of the material at 0°K and zero pressure. and d+/dT = 30. where 0 is the linear thermal expansion coefficient. The variation of entropy with hydrostatic pressure is given by the Maxwell equation: These equations give the entropy change resulting from increasing the hydrostatic pressure from 0 to P as: and since • we have: This is the entropy arising from the dilation resulting

Jan 1, 1970

By A. T. Yu, K. A. Young, Ronald M. Hays, Sanford Terry, D. R. Campbell, J. K. Lovrien, W. L. Price, C. W. Matthews, W. R. van Slyke, S. P. Cole

Storage and transport systems are essential for mineral processing. Section 10 includes systems for storage and transport of dry bulk materials, slurries, and water necessary for operation of mineral pro¬cessing plants. All aspects of these systems are presented, including material characteristics; dry material placement in and reclamation from storage stockpiles and bins; dry material conveyance; hydraulics (liquid flow and pumping); slurry characteristics; slurry storage and reclaim, slurry flow (distributors, launders, and pipes), and pumping; water (primary and secondary process, reclaimed, fire, cooling, and domestic) supply, transport, and storage; equipment selection and sizing; power transmission; controls; and health and safety. The mate¬rials considered include crude ore; process and other water supplies; initial, intermediate; and final in-plant materials; finished materials; and tailings and other waste products. Section 10 is a compilation of practical operating data for engineers, technicians, and plant opera¬tors. For the design of mineral processing plants, including selection and sizing of equipment, the design engineer should consult with equipment manufacturers, suppliers, engineering firms, and plant op¬erators with specialized experience in particular materials handling methods. When unusual conditions are encountered, actual pilot or full scale testing may be required to insure adequate design. Operation of mineral processing plants is dependent on the proper engineering and construction of storage and transport systems for dry bulk materials, slurries, and water. It is obvious that processing plants cannot produce the required quantities of semifinished and finished products unless these material can be transported adequately to, through, within, and from the plant. Costs of storage and transport systems may exceed 50% of the total capital and/or operating costs. Considerable attention must be given to the design, layout, and engineering of these systems. These systems are interrelated with the mineral processing method and it is necessary to determine the best combination of mineral processing method and storage and transport systems. Storage and transport systems perform a major supportive func¬tion to mineral processing. The final selection of these systems depends on the final selection of the processing method and related systems (environmental control, tailings and waste disposal, etc.). The process¬ing method determines the characteristics of materials to be stored and transported; distances, elevations, flow rates for transport; and storage requirements. Flowsheet After selection of the processing method and related systems, a flowsheet must be developed. This flowsheet should indicate produc¬tion rate, percent solids or moisture, and specific gravity for all initial, intermediate, and final products or product flow streams. It also may be necessary to know the size analysis, chemical analysis, abrasiveness, and other physical characteristics of various product flow streams. The flowsheet provides basic information for plant layout, storage requirements, and selection and sizing of process and transport equip¬ment. Storage and transport systems often function improperly be¬ cause of inadequate flowsheet data or process revisions, resulting in improper equipment selection and sizing. Further details on plant design are presented in Section 33. After selection and sizing of process equipment, it is necessary to select the type of material transport equipment to be used for in¬plant materials handling as well as crude ore, tailings, and final prod¬ucts. The storage and/or surge requirements, including capacity and location, then are determined. Storage and/or surge capacity is neces¬sary for leveling production rate fluctuations due to variations in material characteristics; production scheduling; equipment mismatch¬ing, failure, or breakdown; and scheduled maintenance downtime. Variations in ore grindability will cause fluctuations in throughput or production rate for grinding mills and variations in ore grade will cause fluctuations in production rate for concentrating equipment. Variations in material moisture content and percent solids can have serious effects on the capacity of material transport equipment. In some cases, a blending system may be desirable or necessary to reduce variations and to obtain uniform production and/or metallurgy. Storage capacity often is needed because of different production schedules for various phases of the mining and mineral processing complex. As an example, the mine and coarse crushing plant may operate on a one- or two-shift-per-day schedule while the fine crushing plant and concentrator on a three-shift-per-day schedule. This would require storage between the coarse and fine crushing plants. Another example would be load-out storage where unit trains are used to haul concentrate or final products. The concentrator could be operated on a three-shift-per-day schedule while only I to 2 hr may be scheduled for unit train loading. This would require adequate load-out storage so that the concentrator can be operated continuously. Process equipment with mismatched capacities requires storage or surge capacity between the mismatched equipment to achieve oper¬ating efficiency. As an example, the fine crushers may have greater capacity than the rod mills. Storage between the fine crushers and the rod mills permits both the crushers and rod mills to be operated at optimum capacity with the crushers actually operating or running a lesser number of hours per day. Mismatched capacities for process and transport equipment also may require storage or surge capacity. Mismatching of equipment capacities most often occurs because of the difficulty in equipment sizing. Thus, the design engineer must anticipate where in the flowsheet these mismatching capacities are likely to occur while making every effort to assure that equipment is selected and sized properly. Equipment breakdown can cause major operating difficulties and lost production. These breakdowns can be reduced with adequate scheduled maintenance. Lost production due to nonscheduled down¬time caused by equipment breakdown can be reduced further by standby equipment and storage and/or surge capacity ahead of essen¬tial mechanical and electrical equipment. As an example, standby pumps and conveyors often are available as part of the in-plant trans¬port system. Storage and/or surge bins and tanks often serve dual functions, such as dewatering, mixing, and smoothing production rate fluctuations.

Jan 1, 1985

By W. I. Watson

The study considered all the institutional, resource reserve, technological, and economic factors relevant to the industry&apos;s present and future status, such as markets and marketing techniques, competition from other producers both domestic and foreign, federal and state regulations affecting the industry, quarrying and processing technologies, slate reserves and resources, and industry structure and economics. Constraints to the growth of the industry were assessed and scenarios proposed to solve some of the problems.

Jan 1, 1980

By Maria I. De Rosa

The major fire and fire injury findings for all coal mining categories for 1990-1999 are reported in tables 37-38. Table 39 and figure 13 show the number of fires. fire injuries, risk rates, employees&apos; working hours, and coal production (under- ground and surface coal mines only) by time period for all coal mining categories. Table 40 shows major findings (for under- ground coal mines only) for 1978-1992. For all coal mining categories, 458 fires occurred during 1990- 1999; 157 of those fires caused 164 injuries and 2 fatalities (Ewhr = 2,070 x l0%r, Irr = 0.016: C&apos;P (for underground and surface coal mimes only) - 10,363 x l0\t, Frr = 0.044, LWD = 14,753). Twenty-nine fires and 17 injuries involved contractors. Sixty-six fires required firefighting interventions by mine rescue teams (25 times in underground mines) and fire brigades and fire departments (at least 41 times at surface coal operations). In all, 51 fires destroyed or heavily damaged equipment (including 16 pieces of mobile equipment) because of failure of other firefighting methods, late fire detection, undetected fires, or fire size. A total of 114 fires were detected late, and 42 fires were undetected. The greatest number of fires and fire injuries occurred at surface coal mines; the highest risk rate values were also calculated for this category. For all coal operations, the ignition sources that caused the greatest number of fires were flame cutting/welding spark/slag/flames (1 03 fires or 23% with 69 injuries), hydraulic fluid/fuel sprayed onto equipment hot surfaces (98 fires or 2 1 % with 29 injuries), spontaneous combustion/hot coal (62 fires or 14%), electrical short/arcing (49 fires or 11% with 18 injuries), and conveyor belt friction (3 1 fires or 7% with 6 injuries).

Jan 1, 2004

By J. A. M. Cano

Since the beginning of operations at the pellet plant at Ponta Ubu ES, Samarco Mineração SA has produced pellets for direct reduction and blast furnace processes. Of the total amount of pellets produced from 1977 through 1993, 55% are used in the blast furnace, about 45 Mt (49.6 million st). The principal components of pellet gangue are calcium oxide, magnesium oxide, silica and alumina. They should be added in adequate quantities to guarantee the mechanical resistance of the fired pellets under blast furnace conditions. Over the years, the pellets produced by Samarco had a silica content between 2.5% and 2.8% and varying binary basicity preferably between 0.8 and 0.85. On the other hand, the increasing amount of industrial waste in siderurgical plants caused by the increase of steel production has caused some countries to put into practice methods to reduce the volume of slag produced in the blast furnace. This paper’s goal is to find an alternative for decreasing the amount of slag produced in the blast furnace. It is possible to decrease pellet gangue by decreasing the silica content to about 2%, leaving the metallurgical properties and quality of pellets unaltered. For this work, Samarco pellets were used with SiO2between 2.5% and 2.8%, pellets with high silica and pellets with SiO2 between 2.0% and 2.3% low silica. Both were partially fluxed by the range of varying basicity CaO/SiO2from 0.8 to 0.95 during production.

Jan 1, 1995

By Louis H. Meece

The romance of the River lives on in the tales of Mark Twain; and many Americans still think of the Mississippi River in terms of the steamboats loaded with passengers, gamblers and cotton. Yet for many years, coal has been a major cargo on the American waterways. Currently, about 100 million tons a year of coal is handled in some part of its journey from mine to destination by barge. The bulk of this traffic has been on the Ohio River System and the Upper Mississippi River System in years past. Now we have a new entry. Over the past 10 years various movements of coal have developed down the Mississippi River to power plants at Memphis and along the Gulf Coast, and a small amount of metallurgical coal has moved down the Mississippi River for export to Japan since 1970. But in 1980 a brand new player came on the scene. 1980 was a momentous year for the American coal business. Historically, we have exported about 50 million tons a year of metallurgical coal to the world&apos;s steel mills. Just before and after World War I, we also had a fairly healthy export trade in steam coal; and again in the late forties, we supplied large amounts of coal to Europe until domestic production could recover from the ravages of World War II. But since the early fifties, the name of the export game has been metallurgical coal. Starting in the mid seventies, something began that was almost totally missed by the American coal industry; and that was the growth in the use of coal first in Europe and now in the Far East as a steam product and as an alternative to petroleum products. As these world coal markets began increasing, the U.S. at first was thought to be noncompetitive because ocean rates were still very low. For example, as recently as 1977, Australian coal could be moved all the way from Australia to the Gulf Coast for $12.50 per metric ton. Coal exporting countries were Poland, Australia and South Africa, except for the continuing exports of U.S. high quality metallurgical coal. But then came the second oil shock, and all of the sudden that $12.50 rate from Australia to the U.S. Gulf became $25.00. And then suddenly and almost unrecognized by the American industry, American coal was in fact competitive at overseas destinations because of the drastic change in trade patterns brought on by the doubling and tripling and quadrupling of bunker fuel prices for ships.

Jan 1, 1982

By Bobby C. Carson

INTRODUCTION The National Park Service is entering the seventh year of monitoring caves for the presence of radon and radon daughter products. The purpose of this paper is to summarize the radiation monitoring program at Mammoth Cave National Park, and to present some of the results of this program. Mammoth Cave National Park completed five years of collecting data on May 1, 1981: although Mammoth Cave encompasses approximately 361 km of underground passageways, this paper will concentrate on only a 2.2 km section of the cave known as the Historic Tour. Included in this paper is a discussion of the methods the Nations Park Service uses to protect employees from exposure to alpha radiation. MONITORING METHODS The National Park Service monitors cave atmospheres utilizing the procedures provided by the Mine Safety and Health Administration in their Radiation Monitoring Training Manual (Anon., 1976). This procedure is described as the Kusnetz Method (Kusnetz, 1956) of radon daughter monitoring. Due to the length of the tours at Mammoth Cave, it has been determined to be the most practical procedure. The Historic Tour is a 2.2 km (1.4 mile) loop through passageways ranging in size from 18 m high by 12 m wide, to 0.9 m high by 0.6 m wide. Seven five minute walking samples were taken for this cave tour by drawing at least 10 1 of air through a 25 mm fiberglass filter utilizing a Monitaire Sampler Pump. The radon daughter concentration levels were determined using an alpha scintillation counter to measure the alpha activity on the filter paper. The Monitaire Sampler Pump was calibrated each day prior to monitoring the cave tour and the scintillation counter was calibrated by procedures described by the Mine Safety and Health Administration (Beckman, 1975) at six month intervals. Guidelines established by the National Park Service and approved by the Mine Safety and Health Administration require weekly sampling when the average working level exceeds 0.30 (NPS-14, 1980). A working level is an atmospheric concentration of radon (Rn-222) daughters which will deliver 1.3 x 10 5 MeV of alpha energy per liter of air in decaying through Ra C&apos; (Po-214). The Historic Tour has continually exceeded the 0.30 working level average and has been monitored weekly. Generally, only radon daughter working level data has been collected on the Historic Tour due to limited personnel. However, other special measurements of the uncombined fractions of radon daughters with wire screens, tsivoglou method for radon daughter sampling (Thomas modification, 1970), and thoron daughter monitoring. These special measurements have not been routine due to time limitations involved in radon daughter sampling of other occupied portions of the cave. SUMMARY OF DATA The Historic Tour has been the most consistantly monitored tour since elevated levels of alpha radiation were found to exist at Mammoth. Cave. It is also the only natural entrance to the main sections of the cave and provided an opportunity to study man made actions upon the natural entrance. For these reasons the Historic Tour was isolated for study. Beginning October 10, 1977, and ending November 20, 1977, a pilot project was undertaken involving the Historic Tour and the practice of covering the natural entrance to this tour with sheet metal in the winter months. The purpose was to study radiation levels on the Historice Tour while the covers were on and off the natural entrance. In this pilot project, comparisons were made with incast air with covers on and off the entrance, and outcast air with covers on and off the entrance. TABLE 1 Incast air Mean W.L. Covers on . . . . 1.46 W.L. Increased 54% Covers off. . . . 0.67 W.L. when covers on Outcast air Mean W.L. Covers on . . . . 1.33 W.L. Decreased 5% Covers off. . . . 1.40 W.L. when covers on The natural entrance was artificially covered in the winter months (Yarborough, 1978) to protect the visitor from the extremely cold incast air, in the first four years of monitoring. The data in Table 1, illustrated in Figures 1 and 2, shows that this action increased the radon daughter working levels on the Historic Tour by 54% when the covers were on the entrance and the airflow was incast. While the air flow was outcast at the natural entrance, it made little difference as to whether the entrance was closed or open. Some interesting findings were observed when

Jan 1, 1981

By F. J. Klingler, R. Hausmann, K. M. Swaffar, J. C. Neyer

This paper presents the results of a study of case histories of projects on which sinking caissons were used to construct shafts for tunnel construction and other underground works. Design methodologies are reviewed with respect to the impact of soil and groundwater conditions on the performance and constructability of sinking caissons. Factors which have led to success or failure are reviewed and evaluated, and recommendations for future investigation, design and construction activities are presented. Cost data for five recent projects are reviewed and compared to the cost of shaft installation by other means.

Jan 1, 1997