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By Richard P. Graff

Author?s note: Diversity in practice exists in the application of Generally Accepted Accounting Principles (GAAP) for purchase price allocations with respect to the recognition of Value Beyond Proven and Probable Reserves or VBPP, as defined by the Emerging Issues Task Force (EITF) in EITF 04-3: ?Mining Assets: Impairment and Business Combinations?. This diversity appears to arise from differing interpretations of what constitutes sufficient information to value components of VBPP in business combinations. Concerned with this diversity in practice, respected members of the mining community formed a working group which prepared an issues paper that discussed this diversity and provided recommendations for recognition and valuation of VBPP in business combinations. The working group then submitted this paper to the EITF, requesting further guidance be issued on the concept of VBPP. In conjunction with the submission to the EITF, the working group also provided this paper to the United States Securities and Exchange Commission (SEC) for review and comment. The EITF discussed this issue at its May, 2008 agenda committee meeting. However, it declined to add this topic to the EITF agenda for further consideration. Absent additional guidance on VBPP from the EITF, the working group submitted a proposed definition of VBPP to the U.S. Securities and Exchange Commission for its consideration. The following article, which is consistent with the issues paper that was submitted to the EITF and the SEC, discusses the components of and valuation techniques for VBPP. The correspondence with the U.S. Securities and Exchange Commission, which includes the proposed definition of VBPP by the working group, is attached as Appendix 1 to this article.

Jan 1, 2008

SYMPOSIUM ON THE RECOVERY OF PYRITE The proceedings of the Symposium on the Recovery of Pyrite, held in Johannesburg in June 1969, may be ordered from the Secretary, S.A.I.M.M., P.O. Box 61019, Marshalltown, Transvaal, at the reduced price of R5.00. Papers and addresses presented at the Symposium were: Opening Address by Mr J. K. E. Douglas (President). "The world sulphur position," by W. C. J. van Rensburg (Member). "The South African sulphur position," by D. L. Taylor (Member). "The recovery of pyrite at Government Gold Mining Areas Limited," by R. J. Westwood (Member), G. S. Stander (Affiliate) and H. P. Carlisle (Member). "Recovery of pyrite concentrates from colliery-waste discards-A preliminary assessment," by W. A. Gilbert (Visitor) and E. A. Nugent (Member). "The role of pyrite in upgrading the uranium content of Witwatersrand conglomerate ores by means of the heavy-medium separation process," by D. J. Tutt (Associate Member). "Pyrite recovery by jigs at Durban Roodepoort Deep Limited," by E. J. Dominy (Affiliate), F. Diamond (Affiliate) and A. D. Graham (Affiliate). "The storage of pyrite concentrates," by Joan C. Campkin (Visitor) and P. J. Lloyd (Member). "The flotation plants of the Anglo-Transvaal Group," by L. A. Bushell (Member). "Pyrite flotation with amines and cyanidation of the products," by E. H. D. Carman (Member). "Gold recovery by cyanidation from mill products obtained after pyrite flotation," by P. G. Mrkusic (Graduate), M. Sciarone (Visitor) and H. E. Simonsen (Visitor). "The influence of sulphydryl and cationic flotation reagents on cyanidation of native gold," by K. G. Ashurst (Visitor) and W. P. Finkelstein (Member). "Laboratory tests on the flotation of pyrite from Witwatersrand gold ores by cationic collector," by J. Levin (Member), and M. L. Veitch (Visitor). Closing Address, by Dr R. E. Robinson (Director, National Institute for Metallurgy). INCOME TAX: POST-GRADUATE STUDY COURSE CONCESSION As a result of representations made by the Associated Scientific and Technical Societies of S.A. and the Suid-Afrikaanse Akademie vir Wetenskap en Kuns, engineers and scientists will now be allowed to deduct expenses incurred on post-graduate study courses from their taxable income. The conditions of this concession are set out in section 16 of the Income Tax Act, 1969, which reads as follows: 16. (1) The following section is hereby substituted for section 16 of the principal Act: 16. "Notwithstanding the provisions of section 23(a) and (b), there shall be allowed to be deducted from the income of any person who is a dentist or medical practitioner or practises any profession referred to in paragraph (a), so much as the Secretary may allow of any expenditure incurred by such person in the circumstances contemplated in paragraph (c), if the Dental Association of South Africa in the case of a dentist, or the Medical Association of South Africa in the case of a medical practitioner, or the "Suid-Afrikaanse Akademie vir Wetenskap en Kuns" or the Associated Scientific and Technical Societies of South Africa in any other case, certifies in such form as the Secretary may prescribe that such person- (a) if he is not a dentist or medical practitioner, is the holder of a degree in engineering or allied technology or natural sciences awarded by a recognized university on completion of a prescribed minimum period of study of not less than four academic years or the equivalent thereof and enabling the holder thereof to practise the profession of an engineer or a profession for which advanced and specialised scientific knowledge is required which relates directly to- (i) the development or exploitation of natural resources; (ii) mining or prospecting for minerals; (iii) the production of pastoral, agricultural or forest produce ; (iv) the breeding or care of farm animals; (v) any industrial process; or (vi) the training of specialists in any branch of science or technology connected with any aforementioned profession; (b) has practised his profession for not less than three years; and (c) has incurred such expenditure during the year of assessment in respect of the attendance by him of any postgraduate study course approved by such body to improve his qualifications for carrying on his profession in the Republic". (2) The amendments effected by subsection (1) shall first take effect in respect of assessments for the year of assessment ending on the twenty-eighth day of February, 1970. From this section it is clear that the concession will only be granted if the application is accompanied by a certificate, which certificate must be either signed by the Chairman of Die Akademie or the President of the Associated Societies. In terms of the authority granted such certificates must certify that: (i) the taxpayer concerned has practised his profession for not less than three years; (ii) the expenditure was incurred by the taxpayer during the year of assessment indicated in the certificate; (iii) the expenditure was incurred by the taxpayer in respect of the attendance by him of a post-graduate study course approved by the Associated Scientific and Technical Societies of South Africa or Die Suid-Afrikaanse Akademie vir Wetenskap en Kuns, as the case may be, to improve his qualifications for carrying on his profession in South Africa; (iv) the taxpayer is the holder of a degree in engineering or allied technology or natural sciences, which degree was awarded by a recognized university,

In 1981, North American Degerstrom Contracting began gold mining on the Little Rocky Mountain Range, near Zortman, MT. In the past four years, the price of gold has fluctuated from a high of $16/g ($500 per oz) to a low of $9.60/g ($300 per oz). With that kind of a moving target, profitability presents a management challenge. Based in Spokane, WA, Degerstrom employs 85 people at the mine. It is one of the largest heap leach operations in the nation. Last year, 8 Mt (9 million st) of low-grade ore was moved to the leach pads in a two-shift, five-day operation. The current pit being worked produces 218 kt/m (240,000 stpm) of leaching ore. Ore fragmentation critical The gold bearing ore has no defined patterns or seams. It is an igneous intrusion resulting from volcanic activity. It occurs spontaneously throughout the range. To recover the gold, areas are staked out and core samples are taken and analyzed. These data determine if the ore is rich enough to mine. Once mining begins, an average of 150 holes are drilled and blasted on a two-day rotation. Five Ingersoll-Rand T4s drill on a 4- x 4.5-m (14- x 15-ft) spacing to a 7.6-m (25-ft) depth. "When we first came to this site, we drilled on a 3.6- x 3.6-m (12- x 12-ft) pattern," said Paul Baker, superintendent for N.A. Degerstrom Contracting. "We did some experimenting using a high density emulsion in the bottom of the hole and filling the rest with regular Anfo." The high density emulsion saved the company on overall drilling and blasting costs since the holes could be spaced on a wider pattern and still achieve the high degree of fragmentation needed for leaching. Blasting is important because the gold lays in the natural strata of the rock. Also, the ore is not crushed before it goes to the leach pads. So complete fragmentation is critical. Terrain dictates loading/hauling system The shot ore is worked in 6 m (20 ft) benches by two Caterpillar 245 front shovels equipped with 3 m3 (4 cu yd) buckets, a 992C wheel loader with a 12.6-m3 (16.5-cu yd) bucket, and a 988E high lift fed by a D9H. A Caterpillar D10 tractor trap dozes to the loader. According to Baker, the 6 m (20 ft) bench is an efficient lift for the front shovel and the wheel loader. In tight quarters or in a pocket of ore, Degerstrom uses the 245s. For high production areas, the wheel loader is used. The hauling fleet includes 28 Cat 773 off highway trucks. The 45 t (50 st) trucks are seven pass loaded by the 245s with a 25-second cycle time per bucket load. The 992C loads the trucks in two passes in less than 30 seconds. The cycle time is four to five minutes for a 1.2-km (4000-ft) haul on grades that average 10%. Grades, in some places, exceed 16%. "We use the 773s because they have a low weight-to-horsepower ratio - that's what you need in steep country," Baker said. The trucks haul the shot ore to the leach pad. Currently, Degerstrom is working a pad that has a capacity for 5 Mt (5.5 million st) of ore. Building the leach pads The base of the pads resemble a large drainage basin. They take about one month to construct. A 0.3-m (1-ft) layer of impervious bentonite clay is hauled in and leveled by a Cat D9H tractor. A 30-mm (1.2-in.) PVC liner is then laid in place on top of the clay base. It, in turn, is covered with tailings from an old on-site mill. Degerstrom uses the tailings to protect the liner from tears when the ore is dumped. The ore is leveled and built up in 9 m (30 ft) lifts by the D9H. The tractor rips the top layer. Then, a network of plastic irrigation pipe is put in place to distribute a cyanide solution over the surface. The leaching solution percolates through the ore and dissolves the gold. The solution drains from the pads and is pumped to a 22.7-ML (6-million gal) pregnant solution holding pond. The liquid then goes through two separate filtration units. One unit removes entrained solids and the other side adds a zinc dust to precipitate the gold. This is collected on filters. The affluent then returns to a solution holding pond for redistribution through the pipeline network. Typically, 80% of the pad's potential recovery takes place in the first 30 days of leaching. The process stops when the ambient temperature falls below the cyanide's freezing level. And each pad has a leach cycle of four to five years before recovery values decline to a point where further leaching becomes uneconomical. At Zortman, it takes 9 t (10 st) of ore to produce 31 g (1 oz) of gold. It is therefore critical to keep total production costs down and efficiency high.

Jan 1, 1986

By James H. Fletcher

Much data and experience has been accumulated, especially in the last 5 or 6 years, to show that modern roof trusses, both of the Birmingham type and the bolt-and-channel type, perform well to support certain kinds of coal mine roof. For some time now, roof trusses of the Birmingham type have been effective in controlling thick shale roof, which tends to sag and fall over a period of about eight weeks. These falls leave dame-shaped cavities which are reminiscent of the pressure arch theory of roof failure. Lately there has also been significant interest in supporting some longwall panel entries with channel and-bolt systems. These systems utilize two vertical center bolts and two 45 degree end bolts through a heavy gage channel. Both types of truss systems have long bolts, which require 8 to 9 Foot holes on a 45 degree angle, and point anchor resin bolts, which utilize a 2 or 3 foot rebar grouted in a thin annulus of polyester resin. Consequently, many coal mining people have been impressed with the very good anchorage characteristics and tensionability features of these systems. However, the same people have been proportionately dismayed to find that very often the most difficult part of introducing a trussing program to a mine is deciding on the machine specifications which determine the truss installation rate. The purpose of this paper is to indicate aspects of truss bolter design and use which are not shared by roofbolter design and its vertical pattern of bolts. Of the two types of drill units conventionally accepted, arm feed units and mast units, only one 15 judged suitable. While the arm feed unit offers low collapsed height with a long drill feed length, a great attribute for low seam use, it has not lent itself to being rolled over to a 45 degree angle. One reason for this inability is that many items which are conventionally mounted on the arm feed unit cannot be allowed to rotate with the unit. These items include drilling controls, tool Cray, lights, canopy or safety post, panic bar, and stab jack(s). If these items were independent of the arm feed drilling unit, they would have to be out of the way of the boom arms, links, and feed cylinder as the unit moves. To ease this problem, the arm feed unit can be made very short and even mounted across the machine. However, the resulting short feed length is undesirable. Several short drill steel extensions are required to drill a relatively long hole, making this procedure time-consuming as well as increasing the likelihood of drilling a crooked hole. Another reason for abandonment of the arm feed unit is that a very stiff drill guide (or "centralizer") is required to start the hole. Using the mast tilt control, the roof bolter operator will force the bit against the roof and will rotate the bit without feeding to cut sideways into the roof to form a hole face. Omission of this step often results in the skidding of the bit toward the rib. To date, no such sturdy drill guide has been designed for an arm feed drill unit. The mast style unit has been the most successful design for angle drilling. It is easily rolled over to a 45 degree angle or to 90 degrees and, on single head models, often has been made to roll through 180 degrees. The mast style is compact, in plain view and is easily fitted with heavy duty manual or hydraulically operated drill guides. It can be mounted directly on a machine chassis, or onto a swinging boom, or onto a slider arrangement in a broad variety of configurations which will accommodate controls, tool trays, stab and roof support jacks, lights, canopies, and panic bars. The only significant difficulty with mast drilling units occurs in lower mining heights when desirable low mast height is accompanied by undesirable short feed length. While the lead extension in the drill string can be as large as the mining height will allow, a short feed length will permit only the use of shorter middle extensions and drivers, and perhaps will. require the use of two different length wrenches. The associated problem is the relatively short distance from the drillhead to the drillguide. This distance is further shortened when raising the drillhead to get the bit to the roof. One might keep longer lead extensions on the machine; drill a vertical pocket for the bit, starting the angle hole at this pocket; or modify the mast. If the mast is made so that the distance from the drillhead to the drillguide remains constant while the mast extends, after which the drillhead travels up the mast, the shorter mast can be made to start the hole more easily. However, short teed length will still result in longer drilling time. It is obvious from these points that a proper choice of mast is simply the highest one which will be moved easily in the given mining height. This, of course, goes against the grain of most experienced coal mining people who realize that a relatively small decrease in mining height can create a problem correctable only with the purchase and installation of shorter masts. If the mining

Jan 1, 1982

By L. Thomassen, N. Spooner, J. M. Thomas

NI-CR and some Ni-Cr-Fc alloys, when used as electrical resistance heating elements in reducing atmospheres, at times suffer rapid breakdowns due to so-called "green rot." These reducing atmospheres are most frequently used in bright annealing and heat-treating furnaces which are kept for long periods of time at 1500' to 1800°F. The green rot is a preferential oxidation of the chromium in the alloy to such an extent that the remaining metal frequently becomes ferromagnetic. The Curie point (magnetic transition point) at room temperature for a pure Ni-Cr alloy is known to be at about 7 pct Cr.¹ This represents a severe loss in chromium for one popular grade of resistance alloy, which is nominally 80 pct Ni-20 pct Cr. The formation of oxides along the grain boundaries makes the ribbon brittle and gives the fracture a green, earthy appearance—hence the name. The authors have studied this phenomenon, simulating the industrial conditions by heating resistance ribbons in various reducing atmospheres, such as moist hydrogen or the atmosphere specified in the ASTM test.² This atmosphere contains 16 pct H2, 10 pct CO, 4 pct C02, 1 pct CH1, 1.5 to 2.5 pct H2O, balance N2. Little success was had in producing severe corrosion with the commercial 80 pct Ni-20 pct Cr alloys. It had been noticed, however, that 90 pct Ni-10 pct Cr alloy wires deteriorated very rapidly when placed in the bottom of narrow thermocouple protection tubes. The bottoms of the tubes with the wires were heated in air. This took place both in porcelain tubes and in Inconel* tubes either with the cold end open or loosely stoppered with asbestos. On examination the wires showed microstructures which would be classified as green-rot corrosion. Subsequent investigations with various alloy wires confirmed the observation that corrosion occurred in some wires, including 80 pct Ni-20 pct Cr alloys, much more rapidly in such tubes than had occurred previously in test atmospheres mentioned above. When 90 pct Ni-10 pct Cr alloys are oxidized with an abundant air supply, the oxide coat consists of both nickel and chromium oxides. In this case thermodynamic equilibrium between the oxides is not established. However, when oxidized 90 pct Ni-10 pct Cr alloy specimens were heated to 1500" to 1800°F inside small diameter tubes, the oxide coat transformed to a bright metallic outer layer of nickel or nickel-rich alloy, beneath which appeared a green oxide, followed by intergranularly attacked metal. The obvious explanation is that in the narrow tubes, thermodynamic equilibrium is being established according to the equation NiO + Cr (in alloy) ? Ni + Cr2O3 This reaction is favored due to the fact that the oxygen pressure over nickel oxide is more than 10"' times the oxygen pressure over the green chromium oxide. The initial oxygen in the tubes is depleted very rapidly by oxidation of the wire and of the tube, if it is metallic. A confirmation of the occurrence of the reaction, reduction of NiO by the chromium in the alloy, was obtained by putting oxidized nickel foil into an evacuated quartz tube along with a piece of bright 90 pct Ni-10 pct Cr alloy ribbon, and heating the tube at 1820°F for a number of hours. The nickel oxide on the foil was completely reduced to pure nickel, leaving a bright foil. The 90 pct Ni-10 pct Cr alloy became strongly magnetic and showed the typical green-rot structure. Weights of the samples before and after testing showed the weight loss of the foil to approximately equal the weight gain of the ribbon. Evidently the bottom of long, narrow protection tubes, stoppered or not, may under certain conditions act just as the sealed quartz tubes in respect to being a confined space in which oxygen is not freely replenished. This condition of oxygen depletion can be prevented by introducing a small amount of air into the bottom of the tube. Many conditions becloud the green-rot phenomenon, such as carburization, sulphidization, and the presence of other corrosive agents. However, these experiments have shown that the basic reaction is a case of internal oxidation which can occur simply by adjusting the oxygen pressure to a low enough level so that the atmosphere will leave the nickel intact and oxidize the chromium. The action of the other agents in commercial atmospheres can then be taken up as individual cases. From these and numerous other tests completed to date or still in progress, it is believed that a more critical evaluation of electrical resistance alloys for their resistance to green-rot corrosion can be made. A much better understanding of the mechanism for this type of corrosion should also result. A more complete report on these experiments will be published later. In the May 1953 issue: TP 35213 Discussion: Solubility of Carbon and Oxygen in Molybdenum by W. E. Few and G. K. Manning, discussion by N. A. Gokcen: p. 747, first column, the fifth, sixth, and seventh lines should be the first, second, and third lines, followed by the present first, second, third, fourth, eighth, ninth, and tenth lines.

Jan 1, 1954

By G. R. Speich

The interlarnmelm spacing, growth rate, and degree of segregation that accompany cellular precipitation in four Fe-Zn alloys containing 9.7, 15.2, 23.5, and 30.5 at. pct Zn have been determined in the temperature range 400" to 600°C. The chemical free-energy change for the reaction was calculated from the available thermodynamic data and the known compositions of the phases. The fraction of the chemical free-energy change for equilibrium segregation that is converted into interfacial free energy decreases from 0.43 to 0.08 as the magnitude of this free-energy change increases from 35 to 270 cal per mole. At constant temperature the cellular growth rate is proportional to the cube of the dissipated free energy. At 600°C newly 100 pct of the equilibrium segregation is achieved during cellulm precipitation whereas at 400°C only 85 pct of the equilibrium segregation is attained. During cellular growth, mass transport of zinc occurs by grain boundary diffusion; excess zinc remaining in the a! phase after the completion of growth is removed slowly by volume diffusion. A modified Cahn theory of cellular precipitation predicts the observed interlamellar spacing within a factor of two. In cellular precipitation reactions such as pearlite formation or discontinuous precipitation, the basic problem is to predict the variation of growth rate G, interlamellar spacing S, and degree of segregation P with composition and temperature. To accomplish this we need three independent equations relating these quantities. One of these equations comes from the diffusion solution. To obtain two additional independent equations, some assumptions must be made. cahnl has suggested recently that two plausible assumptions are 1) that growth rate is proportional to the dissipated free energy and 2) that the spacing which occurs is that which maximizes the dissipated free energy. According to the first assumption, this spacing also maximizes the growth rate and the rate of decrease of free energy per unit area of cell boundary. The present work was undertaken to test these assumptions. To test the first assumption it is necessary to study a cellular reaction over a wide range of supersatura-tions to establish a relationship between G and the dissipated free energy at constant temperature. This is possible only in discontinuous precipitation reactions since in pearlite reactions constituents other than pearlite form if the composition of the parent phase deviates even slightly from the eutectoid composition. The Fe-Zn system was chosen for study because 1) discontinuous precipitation proceeds to completion over a wide temperature and concentration range, 2) the degree of segregation within the cell can be measured by lattice parameter measurements,2 and 3) the thermodynamics of this system have recently been determined by Wriedt.3 In this system the cells consisting d alternate lamellae of a and r phases form from supercooled iron-rich a phase. The a phase within the cells is bcc as is the original a phase, cia, but has a different orientation and a slightly lower zinc content than the original a phase. The r phase has a zinc content of about 70 at. pct and a crystal structure isomor-phous with T brass. EXPERIMENTAL PROCEDURE Four Fe-Zn alloys with 9.7, 15.2, 23.5, and 30.5 at. pct Zn were prepared from carbonyl-iron powder (400 mesh, 99.8 wt pct Fe) and zinc powder (200 mesh, 99.99 wt pct Zn). The powders were ball-milled together and cold-pressed under 60,000 psi to discs $ in. thick by $ in. diam. The cold-pressed discs of the alloys with 9.7 and 15.2 at. pct Zn were sealed in evacuated silica capsules and heated slowly to 1100°C over a period of 1 week (3 days at 600°C, then 3 days at 80O°C, then 1 day at 1100°C). The alloys with 23.5 and 30.5 at. pct Zn were treated similarly except that the final homogenization temperatures were 1000" and 85O°C, respectively, to prevent melting. The alloys were quenched in iced brine from the final homogenization temperature. Specimens of each alloy were subsequently aged in salt pots at temperatures of 400°, 450°, 500°, 550°, 600°, and 650°C for times that varied from a few minutes to several hundred hours. At a late stage of this work, an alloy containing 11.2 at. pct Zn was prepared by vapor-impregnation of iron foil with zinc vapor at 890°C. This alloy proved useful for electron microscope studies because it was free of porosity. The homogenization and aging conditions were based on the recent Fe-Zn phase diagram of Stadelmaier and Bridgers4 rather than the earlier diagram of ansen.5 They consist of a homogenization heat treatment in the homogeneous a field followed by an aging treatment in the two-phase a + r field. The aged specimens were metallographically polished and etched in 2 pct nital and the radius of the largest cell in the microstructure determined. This radius plotted vs time gave a straight line whose slope is the boundary migration rate or growth rate G of the cell. To determine the interlamellar spacing, specimens were examined by surface-replica and thin-section electron microscope techniques. Because of the irregular nature of the lamellae within the cell, the average interlamellar spacing S .of the cell was measured by the method of Cahn and Hagel,6 where S is defined by:

Jan 1, 1969

By Alice S. Allen, James Paone

SURFACE SUBSIDENCE IN MINING The consequences of subsidence are becoming progressively more serious as the consumption of an ever- increasing quantity of minerals conflicts with the needs of an expanding population for surface land area. The inevitable result will be an increase in the amount of land area on which undermining is restricted because surface subsidence must be prevented. Designation of land areas that may be undermined, provided that subsidence is controlled and kept within specified limits, also will increase. The decisions as to which land is to be used for urban, agricultural, or other purposes clearly involve considerations that go far beyond the scope of mining technology. Under the authority assigned to the US Bureau of Mines (USBM) by the Organic Act (May 16, 1910) and its succeeding amendments and pursuant regulations (30 U.S.C. 1-1 1 ), the Bureau conducts scientific and technologic investigations concerning mining and its related problems. Subsidence-control demonstrations that have been conducted under this authority include projects in Wyoming, West Virginia, Illinois, and the Pennsylvania anthracite region. The problenls for technology arc to devise methods for extracting minerals with controlled subsidence in varying geologic settings and for calculating or predicting the amount, extent, and characteristics of the subsidence that accompanies the extraction of our principal minerals mined by high-tonnage methods so that logical choices can be made from the possible alternative approaches. Where underground minerals and fuels are mined and removed, the voids that are created underground generate strong imbalanced stresses in the surrounding and overlying rock strata. The resulting readjustments in the rock masses may cause subsidence of the ground surface. Subsidence implies vertical collapse, because the most conspicuous component of movement is downward. However, the downward component is accompanied by differential horizontal strains that may be more damaging to man-made surface structure.; than the more apparent vertical displacements. SUBSIDENCE CONTROL The most widespread method of alleviating potential subsidence problems in undermined areas has been to backfill mine voids with mine refuse or some other in- expensive material that provides lateral support to the remaining mine pillars and vertical support to the mine roof and overburden. Most USBM backfilling work has been conducted jointly with the Pennsylvania Department of Environmental Resources in the anthracite region of northeastern Pennsylvania. These joint projects were conducted under the authority of the Anthracite Mine Drainage Control Act of July 15, 1955 (Public Law 84-162, as amended), and the Appalachian Regional Development Act of 1965 (Public Law 89-4. as amended). In addition to the Appalachian and mine-drainage projects, the Bureau has either conducted or participated in demonstration projects to develop the "pumped- slurry" method of backfilling mine voids. Three of these projects were conducted in Rock Springs, WY. The first was a field test of the pumped-slurry technique, conducted in 1970 under the combined participation of the City of Rock Springs, the US Department of Housing and Urban Development, the Dowell Div. of the Dow Chemical Co., and USBM (Candeub, Fleissig, 1971). The objective was to demonstrate that a large quantity of sand could be hydraulically injected under pressure through a single borehole, and that filling of the mine voids would be essentially complete. In the earlier "blind-flushing" methods, involving sluicing material through boreholes by gravity, quantities per injection hole ranged from 15 to 765 m3 (20 to 1000 cu yd), and the mine voids were only partially filled. In the first test of the pumped-slurry method, approximately 14 900 m (19,500 cu yd) of sand were injected success- fully through a single borehole. Subsequent information obtained through 43 monitoring boreholes indicated that mine voids below 11 330 m2 (2.8 acres) had been filled. As a result of the successful initial test of the pumped-slurry process, USBM conducted additional backfilling demonstration projects using the new technique. The first full-scale demonstration was carried out in the Green Ridge section of Scranton, PA, between 1972 and 1974. This project proved the feasibility of using the new hydraulic-injection technique to back- fill dry mine voids, as well as flooded voids, and it demonstrated that crushed anthracite refuse could be used as easily as sand in the process. Stabilization was pro- vided for about 202 300 m2 (50 acres) of Scranton, having a population of approximately 1000. Additional demonstration projects at Rock Springs, WY, between 1973 and 1975, resulted in the stabilization of about 364 200 m2 (90 acres). The work was necessary to preserve the physical and economic well- being of the city, because there have been numerous occurrences of subsidence during recent years. A demonstration project completed in 1976 in Rock Springs brought the total stabilized area in that city to 647 500 m2 (160 acres). At a cost of about $3,000,000, a population estimated at 7000 persons and property values exceeding $18,000,000 were protected. Presently. USBM is participating in subsidence- control projects involving backfilling in Pennsylvania, Illinois, and West Virginia. Eleven demonstration projects were in progress by USBM in 1977-eight in the Pennsylvania anthracite region and three in bituminous areas of West Virginia (one) and Illinois (two). The estimated total property value is over $100,500,000, and the total cost of the projects is expected to be nearly $20,000,000. The demonstration projects are designed to adapt the pumped-slurry technique to a variety of subsurface conditions, to increase efficiency, and to reduce costs.

Jan 1, 1982

By Leonard Obert

Part I of this report3/ proposed a sonic method for determining pressure, in situ, on rock pillar, arches, and other structures. The method required (1) a laboratory determination of the variation of the velocity of sound with pressure in samples of rock taken from mine pillars to establish a velocity pressure relationship for the type of rock, and (2) a measurement of the velocity of sound in the mine pillars from which the laboratory samples were taken. The pressure on the pillars could be determined by correlating the velocity in the pillar with the velocity-pressure relationship established in the laboratory. The success of this method depends on the extent to which velocity of sound varies with pressure. Furthermore, Young's modulus of elasticity, E, can be calculated from longitudinal velocity, V, by the relationship, E = V2d ...(1) where d = density. Hence, the variation of the elastic modulus with pressure also can be determined. It pointed out that this method might be extended to assist in the prediction of rock bursts, inasmuch as the rock is subjected to extreme pressure in these instances.

Jan 1, 1940

By Ivan Kovac, Pero Marjanovic, Branko Bozic

"The paper deals with results of geophysical survey of the rock masses, and the blasting parameterswhich have effect on the fragmentation in quarries . A relationship between the burden, which is the distance between the charged borehole and the free rock surface, and the velocity of longitudinal waves in the rock mass is established based on measured values for different quarries. It is shown that the drilling geometry can also be detemrined from this relationship."

Jan 1, 1999

By Fletcher G

This paper will outline the design process, and the constmction and rehabilitation techniques proposed for the external waste dumps associated with the mining of the Wolfang Basin coal measures.The basin is located to the north of Clermont and immediately west of the Bowen Basin in Central Queensland, Australia. Development of the Clermont Coal Mine is planned for the mid-1990s after adequate sales commitments are obtained to allow the project to proceed.

Jan 1, 1992

President Zakotnik opened the final session of the Institute at 9:30 a.m. Kenneth R. Herrmann of the Resolutions Committee presented the following resolutions to the group for their approval: RESOLVED: That the members of the Rocky Mountain Coal Mining Institute, in convention assembled, hereby ratify and confirm all official actions of the officers, the executive committee and other committees of the Institute between July 4, 1962, and the present time.

Jan 1, 1963

By A. M. Milev

A stand-alone instrument especially designed for recording strong ground motions was used in this study to create a large data base of peak particle velocities measured on stope hangingwalls. A total number of 58 sites located in stopes mining the Carbon Leader Reef, Ventersdorp Contact Reef, Vaal Reef and Basal Reef were monitored. The peak particle velocities were measured at the surface of the excavations to identify the effect of the free surface and the fractures surrounding the underground mining. From the monitoring results obtained, the velocity criterion of 3 m/s will bean adequate value that support systems have to sustain during a rockburst in most cases. However, PPVs in excess of this criterion were recorded, and some PVDs were damaged or irretrieveable, a consequence of which was that possible extreme events were not recorded. The data recorded on the skin of the excavations were compared to the data recorded by the mine seismic networks to determine the site response, defined as the ratio of the measured peak ground velocity to the peak ground velocity inferred from the mine seismic data. The site response measured at all mines studied was found to be 9 ±3 times on average. Values of the site response did not vary significantly with seismological parameters such as peak ground velocity, hypocentral distance and source radius.

Jan 1, 2005

By J. H. Jackson, J. G. Kura, M. C. Udy, L. W. Eastwood

The melting and casting of any commercial metal depends upon the success with which the problems attendant to the handling of the specific metal are overcome. Common difficulties encountered in the handling of commercial metals are tendencies to burn or oxidize excessively, low fluidity, entrapment of dross, hot cracking, cold embrittlement, cold shuts, and un-soundness caused by gas evolution. Beryllium is not only subject to these same difficulties but is generally more sensitive to them than are the more common metals, thus necessitating more exact founding precautions. Characteristics of Beryllium Certain characteristics of beryllium which make it particularly difficult to handle in the plant or laboratory are as follows: 1. The melting point of beryllium is about 2400°F and pouring temperatures vary from 2600 to 2900°F, depending upon the degree of fluidity required. The extreme chemical activity, combined with the high temperatures, makes necessary the use of inert atmospheres, slags, or vacuum for protection during melting and pouring. Furthermore, beryllium tends to react with the melting crucibles, tools, and molds, thus requiring the selection of suitable materials and proper maintenance of these items. 2. The very marked absorption of gas by molten beryllium and the subsequent evolution of gas during solidification causes a great deal of difficulty with unsoundness. Beryllium castings may also be subject to unsoundness as a result of gas formation by chemical reaction with the mold surface during solidification. 3. The very marked chemical affinity between molten beryllium and the normal atmosphere causes the formation of dross. On relatively quiescent melts, this dross forms a very tenacious film which, if carried over to the mold, can cause defects such as the formation of skins or dross in the interior or on the surface of the casting; folds or defects similar to cold shuts may also be found. 4. During the pour and attendant turbulence, the dross may be mixed into the metal and then carried to the casting. There it may be entrapped during freezing and cause internal defects, or it may float to the surface and cause severe dross defects on the cope side of the castings. 5. Solid beryllium is very weak at a temperature near the solidus line; brittleness is also a problem at lower temperatures. Thus, hot and cold cracking must be guarded against. Scope of the Experimental Work All of these problems have been given consideration in the work at Battelle. Before the work was started at Battelle, it was customary to melt beryllium in a vacuum furnace or under flux in graphite-lined induction furnaces. Because of the difficulty of preventing gas unsoundness in beryllium castings, an investigation was undertaken primarily to study this particular problem and, secondarily, to devise methods of overcoming the other difficulties encountered in the founding of beryllium. The objectives of the laboratory work were multifold: 1. To devise a practical method of melting, other than vacuum melting, by means of which consistently sound beryllium castings could be produced. 2. To obtain fundamental data which would promote a better understanding of the causes of gas unsoundness in beryllium. 3. To develop a casting technique which would permit low melt temperatures, minimize dross defects and hot cracking, and eliminate reaction with the mold surface. Up to the present time, a considerable amount of progress has been made on the first and third objectives, with the result that it can now be safely stated that consistently sound castings can be made by the open-pot melting method, using argon as a protective atmosphere and a proper casting technique. There has been some progress on the second objective of the project, but it cannot yet be safely stated that the problem is completely understood. Most of the experimental work has been conducted in an enclosed melting and pouring apparatus in which very-close control of melting variables has been effected. Findings based on the results of the work conducted in this small experimental apparatus have been applied to large-scale melts which have been made by techniques suitable for duplication on a commercial scale.

Jan 1, 1950

MEMBERSHIP In terms of By-laws7.7.1 members are hereby notified that the membership of the following has been cancelled and their names removed from the Membership Register: H. Digby Roberts, R. J. P. van Tonder, S. D. Williams (Members), A. I. de Klerk, M. F. Ford, A. W. Hine, H. A. Johnston, R. T. Joyner, H. D. Small, P. White (Associate Members), J. de F. Agrella, A. Leibowitz (Affiliates), J. P. M. Maidment, C. P. J. M. Steyn (Graduates), H. C. M. Fourie, B. C. P. Job, J. P. Kruger, E. C. Nicolas, A. P. Prinsloo, B. Schuitema (Students). INSTITUTE PUBLICATION Transactions of the Seventh Commonwealth Mining and Metallurgical Congress Requests are still being received by the Institute for sets of the above transactions. Members who wish to dispose of sets in good condition are requested to contact the Secretary, P.O. Box 1183, Johannesburg or 304 Kelvin House, Johannesburg. As these are for resale a price of not more than R10.00 can be offered per set of three volumes. INTERNATIONAL SOLVENT EXTRACTION CONFERENCE, 1971 The Society of Chemical Industry has agreed to sponsor an International Conference devoted to the science and technology of solvent extraction. This will be held in The Hague in the Netherlands during the week commencing 19th April, 1971. The organisation of the Conference will be in the hands of a Conference Committee under the chairmanship of Mr B. F. Warner. The emphasis of the Conference will be on technical solvent extraction of both inorganic and organic substances, including basic principles and practical applications. Sessions are envisaged covering the chemistry of solvent extraction, industrial processes and economics, the performance of contacting equipment and the underlying phenomena involved, and the response characteristics and control of such units. The Conference Committee will welcome offers of papers for presentation at the Conference. In the firts instance, intending authors should submit a preliminary title and an indication of the scope of their contribution to the Conference Secretary: Dr. C. Hanson, University of Bradford 7, U.K. It is hoped to have available in September a first circular giving further details of the Conference. Copies may be obtained from: International Solvent Extraction Conference, 1971, 14, Belgrave Square, London, S.W.1. SOUTH AFRICAN TUNNELLING CONFERENCE TO BE HELD IN 1970 A South African Tunnelling Conference, to be known as TUNCON 70, is to be held in Johannesburg during the latter half of July, 1970. South Africa, in common with other countries, has a vital and growing interest in tunnelling. Apart from the 800 kilometres of tunnel excavated annually in the gold mines, the 82 kilometre-long Orange-Fish Tunnel is being driven to carry water from the Orange River to the drier areas of the Eastern Cape Province. This and the realization that services such as water and transport are increasingly being carried underground in various parts of the world have emphasized the need for a forum at which the technology and potential of tunnelling can be discussed. TUNCON 70 is being supported by three professional engineering societies representing some 6,000 engineers, namely, the South African Institute of Mining and Metallurgy, the South African Institute of Civil Engineers and the South African Institution of Mechanical Engineers, as well as by the Geological Society of South Africa. The conference has three main purposes. Firstly, to pool and exchange expertise by authorities on tunnelling; secondly, to discuss future developments in tunnelling technology with the aim of making those concerned with long-term planning aware of the potentialities of tunnelling; thirdly, to propose standards for tunnel shapes and sizes. Adoption of standards for tunnelling equipment and accessories would be a major step forward in improving tunnelling efficiency. The five themes of the conference will be: tunnel utilisation, exploration, design, construction and standards. The conference will be held at the University of the Witwatersrand, Johannesburg, from 21st to 24th July, 1970. Arrangements will be made for visits to be paid to mines and tunnelling projects before the start of the conference. Further information is obtainable from the Organising Committee, TUNCON 70, P.O. Box 1183, Johannesburg, South Africa.

By Charles Barron McIntosh

BECAUSE of the well known relationship between a low atmospheric pressure and increased amounts of methane in coal mines, attempts have been made to find associations between low pressures and coal mine explosions. An early study by F. Able disclaimed this relationship. He found that half the explosions took place during a rising barometer when methane content in a ventilated mine would be decreasing. A more recent study by C. L. Hosler revealed a close relationship between anthracite mine explosions and low atmospheric pressure. Bituminous coal mine explosions, where coal dust as well as methane can be exploded, were found to be more closely associated with a rising pressure following a low barometer than with the falling or low pressure period. Emphasis on rising pressure periods led the present writer to investigate the effect of the post-frontal or post-cyclonic weather on coal dust. Dust samples of uniform mesh were exposed in an Illinois mine. Analyses of the dust taken before and after the passage of cold fronts and their accompanying low pressure troughs indicated an increase in coal dust moisture prior to a frontal passage and loss of moisture from coal dust following the time of lowest pressure. Thus the warm, moist air (Tropical Maritime)found over the mine during the falling pressure period was cooled in the mine and moisture added to the dust. Cold, dry air (Polar Continental) accompanies' the rising barometer following a frontal passage and the air, being heated in the mine, subtracted moisture from the coal dust. A similar but seasonal phenomenon occurs from summer to winter, the dry dust period in winter being recognized as the more dangerous period. The question arose as to how the short period or daily changes in moisture content of coal dust compared with the seasonal changes. Temperature and humidity statistics were obtained from the surface and mine entry for the months of July andJanuary. The gain and Ioss of mdisture. to the hine were calculated for these two morths as were the gain and loss associated with mT (Tropical Maritime) and cP (Polar Continental),air masses as observed in the investigatiynsl. Temperature and humidity statistics tigatiynsl.taken from [the two air masses represent atimp span of only 34 hr. The redultant changes as presented in the accompanying table verify the supposition that short period changes in coal mine moisture may attain the magnitude of recognized dangerous seasonal changes. Statistics for the mT air and the July air have been placed next each other and those for cP and January air together for easier comparison of the two warm and two cold air units. The last column readily reveals that there is little difference in the amount of moisture being deposited in the mine from the mT air and the July average air. A similar conclusion is reached regarding the loss of moisture from the mine when it is affected by the cP air mass or the average January air. This comparison would be particularly true from late fall through early spring when strongly contrasting air mass changes are a common occurrence. The invasion of cold, dry air following a cold front thus produces a drying period that could well receive more attention as a danger period for coal dust explosions. This drying period would compare with Able's explosion period during a rising barometer. The drying period is also perltinent in Hosler's findings, where it was noted that the average explosion day (for bituminous coal mines) followed the day of lowest atmospheric pressure. Moreover, the author's investigation of some 365 coal mine explosions also indicates that this post-frontal period should receive special attention in connection with mine safety measures. Two comparatively recent major explosions in Illinois (Centralia and West Frapkfort) occurred after the mines had been under the influence of cold-dry air for 25 to 30 hr-following a deep low pressure passage. Dust was the important explosive element in both these explosions.

Jan 1, 1958

By H. P. Leighly, F. C. Perkins

THERE have been several statements in the literature about the difficulty of producing single crystals of high-purity (99.99pct) by the strain-anneal method. Consequently, investigators tend to employ low-purity aluminum for their single-crystal experiments, or else resort to the Bridgman or other techniques which depend on solidification for the production of single crystals. The following paragraphs describe a solid-state method for the manufacture of single crystals of high-purity aluminum which should provide crystals with greater perfection than those formed by solidification. The starting material for this method of producing single crystals was secured from the Aluminum Corp. of America and has a purity of 99.99 pct, the balance being trace amounts of impurities. Specimens approximately 1 by 4 in. are sheared from sheet having a nominal 0.050 in. thickness. The specimens are given a preliminary anneal at 640°C for approximately 3 hr in order to remove fabrication strains and to produce an average grain size of about 1/4-in. diam. The specimens are then etched in Tucker's etchant (45 pct HC1, 15 pct HNO3, 15 pct HF and 25 pct H2O) to remove the oxide film. Critical strain is applied by wrapping each specimen about a 1 3/4 in. round and subsequently straightening it against a flat surface. The high-purity aluminum sheet is sufficiently soft that this operation can be accomplished by ordinary finger pressure. The specimens are immediately annealed again at 640°C for 3 hr and reetched for examination. Ordinarily, considerable growth of certain of the grains will have occurred, and occasionally a single crystal will be produced on the first attempt. The procedure of alternately straining, annealing and etching is repeated until the majority of a batch of specimens contains usable crystal sizes. Typical examples are illustrated in Fig. 1. The greatest changes in crystal sizes are produced in the initial treatments. As the average crystal sizes get coarser in the later treatments, the sever- ity of the strain must be increased in order to produce grain boundary- migration. This increase in severity is effected by decreasing the diameter of the round used for straining (to 1 1/2 in., for example) and/or wrapping the specimens about the round twice, with opposite faces in contact with the round, before flattening. Usually the strain treatments described are not severe enough to produce nucleation in coarse grain high-purity aluminum. The growth of grains occurs by strain-induced grain-boundary migration. It has been observed that the grain boundaries move most readily during the first hour or so of each annealing treatment and that the rate of movement decreases with extended holding times at temperature. Prolonged annealing treatments are therefore not usually beneficial. Similarly, the rate of growth of each crystal appears to depend upon the orientation of the crystal with relation to those of its neighbors. Frequently island grains are formed after the initial heat treatment as the result of slow grain-boundary migration. These sometime become stationary during later heat treatments. Twin orientation interfaces are frequently developed during annealing. These imperfections can usually be removed by increasing the severity of strain to produce actual nucleation of new grains of more favorable orientation at the imperfection interfaces. The largest single crystals produced in our laboratory by the above method measured 4 by 1 by 0.050 in. Examination of Laue back-reflection patterns from a limited area of the specimens, gave no evidence of polygonization. Probably there is some indication of polygonization in the original grain area provided a more sensitive technique is used for detection. Experiments to produce wider specimens were less successful, possibly because wider sheets increase the complexity of the strains induced by deformation and promote widespread nucleation. Grain boundary migration occurs preferentially in a direction parallel to the longitudinal axis of the specimen. The choice of specimen geometry with respect to the rolling direction of the sheet appears to be immaterial in regard to the production of single crystals.

Jan 1, 1961

By G. M. Schwartz, G. A. Thiel

THE first record of the quarrying of dimension stone in Minnesota dates back to 1820 when limestone was quarried locally for part of old Fort Snelling. Limestone quarries were operated at Stillwater, Mankato, and Winona as early as 1854. Granite was quarried first at St. Cloud in 1868, and within a few years thousands of tons were shipped to widespread points. Rough dimension stone for large buildings furnished the first important market, but beginning in 1886 paving blocks were in demand. The largest shipment was in 1888, when 1925 cars were shipped from the St. Cloud area. Quartzite was quarried first at New Ulm in 1859 and somewhat later at Pipestone and elsewhere in southwestern Minnesota: The productive dolomite quarries at Kasota were opened first in 1868 and have continued as large producers of a variety of stone to the present time. At present, the industry is controlled by relatively few operators, and for that reason detailed figures on dimension stone are not released for publication. A general idea may be obtained from the data in the Minerals Yearbook for 1948. The figures for total stone produced in Minnesota are 1,804,000 tons valued at $5,090,652. Probably the largest item in the latter figure is received from dimension stone. A better idea of the situation, in relation to the country as a whole may be gained by using the data for 1930 when more companies were operating in Minnesota, and complete figures were published. In that year Minnesota produced granite valued at $2,668,119 and ranked third among the states in value. Minnesota's production of granite was almost exclusively for dimension stone. In the same year Minnesota produced 300,000 tons of limestone (dolomite) valued at $840,860, and this likewise was mainly dimension stone. In finished limestone Minnesota ranked second among the states in 1930. Sandstone and minor amounts of quartzite are the only other dimension stones that have been produced in Minnesota, but the quarries are now inactive. The commercial stones of Minnesota have been described in two reports by Bowled and by Thiel and Dutton.2 The early history of quarrying in Minnesota and extensive notes on the various rocks are given by N. H. Winchell.3 Small limestone and dolomite quarries were numerous throughout the area of Paleozoic rocks in southeastern Minnesota. Early production was largely dimension stone. With the increased use of Portland cement, most of these ceased production, and today only those at Kasota and Winona remain in operation. In recent years many quarries have reopened and new ones started, but these are devoted to the production of crushed rock and agricultural lime. As the application of modern quarrying and finishing methods increased, small companies in the granite business have dropped out, and the remaining companies have modernized their plants, purchased old quarries, and opened up new ones, thus furnishing a wide variety of granites suitable for most of the customary uses. It is the purpose of this review to present notes on' the geology and operations of each of the quarries now operating within the state. Granites and Related Igneous Rocks The term granite as used in this report includes granites, gneisses, diorites, gabbros, and other igneous rocks. The granites of greatest economic importance are found in three widely separated regions, see Fig. 1. 1-Central Minnesota in the region of the city of St. Cloud, 2-the upper Minnesota River valley region, 3-the northeastern portion of the state, commonly referred to as the Arrowhead region. The St. Cloud Region: The rocks of the St. Cloud region are mainly granites and related rock types such as monzonites and quartz diorites. The stones may be grouped into three major types, namely, pink granite, red granite and gray granite. Most of the pink granite occurs in the area to the southwest of St. Cloud. The rock is best described as stone with large pink crystals set in a finer grained black and white background. The minerals of the matrix occur in remarkably uniform sizes, and the pink crystals are sufficiently uniform in. their dis-

Jan 1, 1952

By P. J. Spencer, J. N. Pratt

The vapor pressure of manganese over liquid Mn-Sn alloys has been determined by a high-temperature torsion-effusion technique. Alloys containing from 8 to 100 at. pct Mn were investigated in the temperature range jrom 1280" to 1580" and the measured pressure values were used to calculate the partial and integral thermodynamic properties of the liquid alloys. The activities show small negative departures from ideality while the integral heats and excess entropies of mixing are asymmetric inform, changing from positive to negative with increasing manganese content. The possible contribution of various factors to the observed thermodynamic properties is discussed. COMPARATIVELY few thermodynamic data are available for manganese alloy systems.' Therefore, as part of a continuing program of studies of the thermodynamic properties of transition metal alloys, measurements have been made on various binary alloys involving this element. In recent publications,2~3 investigations of liquid Mn-Cu alloys and of the Mn-Au system in both solid and liquid states have been reported. For the first-mentioned system the work suggested that magnetic interactions may be responsible for the observed form of the thermodynamic properties, while in the second the influence of the electrochemical factor appears to be dominant. The present paper describes a similar study of liquid Mn-Sn alloys. Again the thermodynamic properties have been obtained from vapor pressure measurements made by use of a high-temperature torsion-effusion technique.4 A detailed description of the apparatus and of the experimental procedures used in alloy preparation and pressure measurement may be found elsewhere.2'4 EXPERIMENTAL RESULTS Sixteen alloys, ranging in composition from 8 to 100 at. pct Mn, were prepared from spectroscopically standardized manganese of 99.99 pct purity and tin of 99.999 pct purity, both supplied by Johnson-Matthey and Co., Ltd. One-gram samples of the alloys were obtained by carefully weighing appropriate amounts of the pure components into an effusion cell; this was then suspended in the apparatus and the metals melted together in situ by heating under vacuum to approximately 1550°K. After allowing sufficient time for a homogeneous liquid alloy to be formed, vapor pressure measurements were commenced. These were determined as rapidly as possible at a series of steady temperatures within the range of interest. The duration of experimental runs on individual samples was kept sufficiently short to ensure insignificant varia- tion of alloy composition during investigation. After completing pressure measurements, the alloys were rapidly cooled and their compositions checked by weighing or chemical analysis. All experiments were conducted using effusion cells machined entirely from boron nitride. Measurements were made using a variety of cells with orifice areas ranging from 0.0032 to 0.0075 sq cm and lengths of the order of 0.04 cm; the usual effusion correction factors for orifice geometry and molecular distribution were calculated using Freeman and Searcy's equation5 and had values between 0.6 and 0.75 for the orifices employed here. The vapor pressures of manganese over the alloys were measured at approximately 20°K intervals in the temperature range 1280" to 1580°K. In view of the close approximation of the measured pressures to Clausius-Clapeyron behavior in the experimental temperature range, the data for each alloy have been expressed by equations of the form: logp(atm) =-A/T + B A least-squares computer treatment was applied to the vapor pressure values in order to obtain the coefficients A and B with their associated error. The resulting equations are listed in Table I, together with the equation for pure solid manganese obtained from a previous study.4 To minimize the effect of possible apparatus calibration errors, the activities and partial free energies of manganese in the alloys were calculated by initial reference to the latter equation, obtained from identical torsion-effusion measurements. The immediately resulting thermodynamic quantities, based on a solid manganese reference state, were then converted to refer to the more appropriate supercooled pure liquid manganese standard; tabulated values of the free energies of solid and liquid manganese from Hultgren et al.' were used for this purpose. Partial entropies of solution of manganese were calculated from the temperature coefficients of the free energies and partial heats from the Gibbs-Helmholtz relationship.

Jan 1, 1969

By J. M. Karpinski, R. O. Tervo

A method and equipment have been developed for measuring the impact strength of grains of brittle materials. It is shown that brittle materials develop a characteristic particle size distribution when fractured by impact. A simple mathematical model has been found to describe this distribution, and one of the parameters of the model has been designated the r value. The r value is directly proportional to impact energy and, for fixed impact energy, it becomes a useful criterion of grain strength. Data are presented on the variation of the r value as a function of sample history, initial size, energy input and number of impacts. The results are interpreted in terms of the flaw theory of brittle fracture. The view is now widely held that brittle materials comprise a strong matrix, throughout which flaws or points of weakness are scattered at random. Failure always starts at one or more of these imperfections, and planes of fracture are established by the geometry of the flaw distribution. This view has been developed extensively by Weibull 1,2 and experiments of a confirmatory nature have been reported.3,4 Thus the process of fracture or size reduction of brittle materials can be regarded as a process of elimination or creation of flaws, and the concept of strength is intimately connected with the size and history of the specimen tested. One consequence of the point of view described above is that brittle materials will fracture to yield similar size distributions, although the impact energy to produce a given amount of fracture may vary greatly from one material to another. Generally speaking, experimental investigations of brittle fracture in solids have been confined to two extreme experimental situations. On one hand, we have tests on single specimens of well-defined, but artificial, geometry and, on the other, empirical data have been obtained on the feed-product relation in a variety of mills. The first situation necessitates specimen preparation, which in the case of brittle materials is equivalent to a priori size reduction, while the second one suffers from the disadvantage that little knowledge is gained about individual fracture events. Also, the statistical significance of single specimen tests may be questioned. Our approach lies somewhere between these two extremes. We test a large population of sized specimens and yet subject each specimen to a single fracture event. In 1956 G.H. Fetterley of Norton Co., Chippawa, Ont., made the empirical observation that product size distributions obtained from impacting sized grits can be described by the following equation where R is the proportion by weight of the grain that remains on a screen having an opening x on a side, x, is a parameter having the dimensions of length, and r is a number. This equation has since been derived on a theoretical basis by Gaudin and Meloy.5 They identify x, with the initial size of the test piece and r with the number of flaws per unit length. As will be shown below, our experiments indicate that both x, and r are functions of energy and hence they should more properly be called the effective initial size and the effective number of flaws respectively. Also, our empirical data give us no basis for assuming that r represents the effective number of flaws per unit length, rather than per unit volume. METHOD AND MATERIALS The impact test machine has been designed to feed specimens, essentially one at a time, into an evacuated chamber, where they fall freely and are struck with random orientation by one or the other of a pair of flat steel vanes. The vanes are mounted at opposite ends of a steel bar, which rotates at a closely controlled angular velocity about a vertical axis. Fig. 1 shows the flat circular vacuum chamber with the cover removed, and the horizontal steel bar with the steel vanes at the ends. The cylindrical pot in which the grains are collected after impact appears at the left. Fig. 2 shows the equipment assembled for a test. The vacuum chamber occupies the central position with the variable speed drive below the table surface, the vacuum pump at the extreme left

Jan 1, 1964

By James L. Poole

INTRODUCTION Companies in the mining industry are subjected at times to currency exchange rate risks and interest rate risks. The former occurs any time a firm deals in more than one currency. The latter occurs anytime a firm is financing itself with borrowings that have a floating interest rate (although a fixed rate can be a problem at times). FOREIGN EXCHANGE RISKS There are two general reasons why exchange rate risks (FX risk) can occur: First, a firm may be building a mine or plant in another country where financing for the plan is denominated in funds other than the country hosting the plant. For example, if a Canadian firm is building a smelter in another country which is being financed with Canadian dollars, the C$100 million for labor, which must be paid in the local currency, has subjected the firm to a C$100 million FX exposure risk. If after the cost estimates were made the local currency appreciates 30% against the Canadian dollar, the cost of labor has increased to C$130 million. (Of course, the local currency could also devalue during this period making the labor cost component less than estimated.) The other reason exchange rate risk arises is due to the fact that the costs of production in one country producing minerals may be incurred in one currency while the production is sold in another currency, such as coal being exported from Canada to Japan under contracts denominated in Yen. In this example, the Canadian firm is exporting coal to Japan and being paid in yen in 30 days at a contract price of 100 Y per ton. Furthermore, assume that the mining firm anticipates a depreciation of the yen against the Canadian dollar of 5% in the next 30 days. Therefore, the mining firm stands to lose 5% of the value of its shipment in a month's time not including carrying costs. In both cases, the mining company risks an adverse change in the exchange rate, that is, the currency the company holds may decrease in value relative to the other currency. There are a number of ways to mitigate the exposure to foreign exchange which can be utilized by the mining company. Optimally, a company should try to match all their inflows and all their outflows in the same currency. For example, if a Canadian firm has mining costs and debt service costs both denominated in Canadian dollars then 100% of their receipts would need to be in Canadian dollars. If a Canadian firm has 70% of its cash outflows in Canadian dollars and 30% of its cash outflow in American dollars in the form of a loan amortization then their receipts would need to be 70% in Canadian dollars and 30% in U.S. dollars. This general principle of matching cash receipts and cash costs in the same currency works fine in a perfect world but what of the real world when these cannot be matched? In actual practice, however, the financial markets have developed a number of ways in which foreign exchange risks can he managed or reduced. FORWARD MARKETS One method is to use the forward markets which are made on a world-wide basis by the commerical banks for foreign exchange. If a firm wanted to lock in the current exchange rate for the receipt of a foreign currency at some point in the future, they could do so by contracting on the forward market with a bank selling that foreign currency at the current rate at the future date. For example, assume that a Canadian firm is exporting coal to Japan which would pay yen in 90 days. If a firm wanted to lock in today's exchange rate they could contract with a bank to sell on the forward market yen in 90 days. When the yen was received the contract would be executed by selling the yen to the bank and receiving the previously agreed upon number of Canadian dollars. Generally speaking, the forward market can be used to sell forward about 12 months. The costs of selling (or buying) forward

Jan 1, 1985