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By Canada Ernst & Whinney

INTRODUCTION The principal objective of this chapter is to provide guidance in understanding and making effective use of published financial statements of mining companies. Although financial statements only represent a gathering together of information relevant to an enterprise's financial position and results of operations, they increasingly reflect the complexities of the modern day business environment. In the mining industry, a wide variety of accounting practices have evolved to deal with the unique aspects of the business in the absence of authoritative accounting pronouncements from standard-setting bodies. Given such complexities and divergent practices, financial statements of mining companies are not easily understood and cannot be used effectively as a tool in decision-making without a basic knowledge of fundamental accounting concepts, generally accepted accounting principles and their application in the mining industry. As such, this chapter is not restricted to a discussion of the accounting for the unique aspects of the mining industry. These aspects are dealt with at length but the text also includes a discussion of general accounting concepts and practices which have wide application in the mining industry. Frequent references are made in the text to international practices and to specific rules and requirements of individual countries. In this respect, the scope is restricted to four major mining environments, Canada, the UK, the US and Australia, all of which have made significant contributions to the development of generally accepted accounting principles. Acknowledgements This chapter was prepared under the direction of Ian Murray of Ernst & Whinney in Toronto, Canada. Contributions were received from Graham Butt, Mark O'Sullivan and Dominic Robinson, all from Ernst & Whinney in Toronto and Bill Duvall and Tom Schoenbaechler from Ernst & Whinney in Louisville, Kentucky, US. The authors would also like to acknowledge the assistance received from David Ferguson, Don MacLean, David Taylor and Susan Zoutman from Ernst & Whinney, Toronto, from the UK and Australian practices of Ernst & Whinney and from Ernst & Whinney International. FINANCIAL STATEMENTS - ELEMENTS AND CONCEPTS Financial statements were originally a stewardship report to owners to satisfy them that management had given a proper account of the resources with which it had been entrusted. Today, the number of users of financial statements has expanded to include the following groups: - shareholders, present and prospective; - lenders, present and prospective; - financial analysts; - regulatory authorities; - taxing authorities; - labour unions and employees; and - creditors and customers. The expectations of these user groups differ. For example, investors and lenders require information that will indicate the adequacy of their security and the profitability of the enterprise; taxing authorities will be concerned with the concept of income; labour unions and employees will focus on the profitability of the enterprise. Ground Rules for Preparing Financial Statements Responsibility for Financial Statements Financial statements are the representations of management. This may strike the reader as a rather obvious statement. It is important to remember, however, the role of judgment in accounting. Notwithstanding the promulgation of accounting standards in the developed countries, considerable flexibility remains in the selection of accounting practices and methods and in their presentation. Arguably, there is more flexibility in the mining industry as few countries have developed standards to deal with its unique aspects. Accordingly, financial statements are as much a product of management's perspective and philosophy as they are a compliance document.

Jan 1, 1985

By J. D. Price, W. M. Bertholf

Making maximum possible use of available equipment and material, CF&I placed a high-efficiency, high-capacity washery unit in the existing buildings to gain simplified operation, reduced manpower requirements, raised efficiency, and boosted production. THE central washing plant of the Colorado Fuel & Iron Corp. was first operated in 1918 to furnish coal for two 60-oven batteries of Koppers design. Prior to that time the coke for the blast furnaces had been secured from a number of beehive operations, some with their own washeries. The original washery had four Elmore jigs, which did not do as well as had been expected. In 1923 the jigs were taken out and replaced by Deister Plat-O tables, with various changes in auxiliary equipment. In 1936 the throughput of the plant was considerably increased by the installation of vibrating filters for recovery and drying of fine coal, thereby reducing a large recirculating load of fine coal which passed through the screens of the Carpenter driers. In the late 1940's it became evident that something would have to be done to keep the washery abreast of progress at Pueblo and other washeries. Coal requirements were nearly two and one-half times the original requirements of the 120 small ovens, as additional ovens had been built in 1930, 1938, and 1945, whereas no additional coal washing capacity had been added since the 1936 modifications. Much of the equipment, moreover, had seen better days. One answer to the problem was to build a new and completely up-to-the-minute plant, with sep- arate sections for coarse, medium fine, fine, and very fine coal. On the other hand, it might be financially more profitable to place a high-efficiency, high-capacity unit in the existing buildings. The washing problem at Pueblo is not unique, but it is approached differently. There are no established specifications for coke ash and sulphur. The plant is operated on the principle that if it is more economical to wash it out than to slag it out, wash it out; otherwise, send it on to the blast furnaces. With coals at the Pueblo plant costing what they do, the benefits of washing at a low gravity are questionable. Hence the primary problem is to eliminate the high-ash components and retain all the low and medium-ash components. Sulphur is no problem at all, averaging 0.6 pct in the coke. Since a good proportion of the free impurity in the coals is as large as the coal or bone or larger, it appeared that it would be safe to go to jig washing and discontinue crushing the coal before washing, as had been the practice with the tables. This would boost the throughput. It might also decrease manpower requirements and washing costs considerably. It was assumed at the start and has been demonstrated in the first year of operation that a jig will do as well on the fines as the tables had been doing. It would not be necessary to discount any gains due to increased efficiency in washing the coarse material because of incapacity to wash the fines. By washing at —2+0 mesh it would be possible to dewater, by screening, a considerable tonnage satisfactorily, for example, at —2 + % or —2 + l/4, and then crush it in a hammer mill. The —?+20 could be dewatered in centrifugal driers and the remain-

Jan 1, 1955

The following discussion was held on August 2.2, 1927, during the meeting of the Institute at Salt Lake City. It. IIIatch, Garfield, Utah.—The object of the meeting, I belicve, is to determine the proper chemical composition to insure long wear of a ball in a mill. This might also apply to rods. I feel that there is only one way of determining accurately the wearing qualitites of balls in mills, and that is by actually wearing out 400 or 500 tolls of different kinds of balls. Two things add to the wearing qualities of a ball: (1) the tenacity, with which the particular particles of iron or steel are cemented together; (2) the individual hardness of each partiele. My cxperience is that a ball that has a Brincll of 550 to 650 will give comparatively long life. There is considerable difference of opinion as to the best steel and iron alloys neressary to produce a satisfactory grinding ball. Some favor chrome, somr manganese and some nickel. Personally, I believe cast iron alloyed with chrome makes the best ball for small sizes; for large sizes (31/2 in. arid up) steel alloyed with mangancse is best. L. O. Howard, Pullman, Wash.—How would this method apply to hning? A. Kissock, New York, N. Y.—I do not subcrribe to that exactly. As I am connected with the Climax Molybdenum Co. you would hardly expeet me to. We have been much interested in trying to work out this grinding-ball problem and have spent a, great deal of thought and effort on it. As we are associated with the American Metal Co. and have a number of Proprrties of our own, we are in poaition to try our ideas on ourselves first. We have tried different compositions, but, have not done anything very remarkable with them yet,, except that we have been able to say that the combination of chrome and molybdenum has given the best, wearing ball. Mr. Hatch feels that iron is the best. medium for grinding, hut since toughtiess as well as hardness is a factor in wearability we feel that a well made steel is better for this purpose. Another difficulty with cast, or chilled, iron is lack of uniformity front day to day or from heat to heat, and the impossibility of protlucing balls free from blow or shrinkage holes which matcrially affect their life. If cast iron is made hard it becomes brittle and lacks strctc11 or toughness. This prevents its use in the larger sizes, aS break:lgc, is a sorious hantlic:~p to efficiency. We have trirtl hard and brittle balls and softer but very tough balls and have proved that maximum hardness, with sufficient toughuess to prevent hrcakagv, is the most important factor. in resistarnce of a steel to abrasion. It is well known that. chromium is one of the best hardening agencies for cither iron or. steel; and molybdenum is recognized as an clerment capable of introducing tonghness. In other words, the proper combination of chromium and molybdenum permits unusual hadness with still sufficient touglmess to stand up under severest impact. It is yet too early to give any actual mill results of this new Combination and though we are proving our ideas at our own properties we have to depend on others to help us make the data more general and more worth while and would like to get some testa 206

Jan 1, 1928

By R. B. Basham, C. W. Macune

Very small anomalies in oil well temperatures are detected and measured by recording the difference in temperature existing between two thermally sensitive elements which are Spaced several feet apart on a small diameter carrier and lowered into the bore hole. The elements operate in a balanced electrical circuit, sending only a reference signal to the surface as long as normal gradient temperatures are encountered. When either element enters a temperature disturbance, the circuit is unbalanced and produces large recorder deflections for minute anomalies. The system is highly sensitive to small changes in thermal gradients caused by gas or fluid movements and to the boundaries between beds of different thermal conductivities. Typical logs show successful applications of the process to locating gas and water leaks in casing, gas entry. and gas-oil contact. INTRODUCTION Temperature measurements in oil wells have been made for many years in the search for additional sub-surface data. It has been generally agreed that many conditions in the bore hole cause temperature anomalies which, if located and measured, could yield useful information. But it has been recognized also that many of the irregularities, such as those produced by small water and gas leaks in casing, might be very small and therefore difficult to detect with certainty.' Although tem-perature measurements have been made which show anomalies of less than one degree deviation from normal gradient temperatures, the results are usually uncertain and unconvincing. The general practice, therefore. has been to apply temperature surveys only to well conditions known to produce appreciable disturbances, such as large gas movements and cement hydration. The recent demand for additional data has focused attention on the problem of making dependable temperature measurements in the many instances when anomalies are small but none the less significant. It was for the purpose of detecting these minute irregularities in bore hole temperature that the program described in the following text was undertaken. It is emphasized that this presentation is not offered as a complete study, for the process is still partially experimental. But it is believed that the results obtained thus far will be of interest to the industry and will indicate some of the system's potential usefulness. DIFFERENTIAL MEASUREMENT The chief value of differential measurement lies in the fact that a difference between two factors can be measured with a scale of values best suited to that difference, a scale that might be cumbersome or impractical if applied to the factors themselves. When it is known that all factors will be nearly equal in value, and therefore the differences consistently small, it is possible to select a scale which will permit very accurate measurement of the differences. This advantage is in contrast to the difficulty of measuring small increments of a single variable, where the increment must be measured by a large scale which is suitable for measuring the entire factor. Gradient temperatures are usually measured under this latter condition where temperature measuring apparatus must have a scale extending over the range of expected temperatures, say 50' to 25O°F, and it becomes almost impossible to detect small irregularities of less than one degree. The use of large strip chart recorders and divided multiple ranges tend to amplify the small anomaly and make it more pronounced. But the instability of instruments, both

Jan 1, 1952

By R. E. Smallman

R. E. Smallman (University of Birmingham, England)—Recently, Tedmon and Ferriss11 have determined the yield stress parameters oi and ky for tantalum by measuring the lower yield stress as a function of grain size 2d and fitting the results to a relationship of the form They report that although ky , which is taken to be a measure of the dislocation locking strength, is small (- 2 to 4 x 106 cgs units) a substantial yield drop is nevertheless observed in a normal tensile test. Niobium gives a similar result,12-14 as pointed out in the original work by Adams et a1.,12 and in order to check this apparent anomaly the yield-stress parameters of electron beam-melted niobium have recently been reanalyzed15 by the Luders strain technique. In this method the strain hardening part of the stress-strain curve is extrapolated to zero plastic strain; the intercept on the preyield portion of the curve is taken to give oi, whilst the difference between oi and the lower yield stress gives kyd-1/2. The results indicate that ky increases with increasing grain size and hence, a plot of vs d-112 yields an apparent ky, which is lower than the true value. A similar effect could account for the small ky found in the relatively pure tantalum used by Tedmon and Ferriss. The variation of ky with grain size shows that dislocations are more strongly locked in coarse-grained specimens than in fine-grained samples. In niobium, this may be attributed to the fact that the dislocation density in the fine-grained material is higher than that found in the coarse-grained samples which are given a sufficiently prolonged anneal to remove any residual substructure and, since the metal contains only a small amount of interstitual impurity, a variation in locking occurs. By contrast, application of both the grain size analysis and the Luders strain method to yield-stress data from commercially pure vanadium containing a large amount of interstitial impurity gives consistent values of oi and ky, with ky independent of grain size and temperature. Electron microscope observations show minor variations in dislocation density from grain size to grain size, but in any case in this material the dislocations are heavily locked with precipitate. On yielding new dislocations are generated and, as a consequence, the importance of any differences in dislocation density between the various specimens of different grain size is considerably reduced. It is perhaps significant that Adams and lannucci,16 working with a grade of tantalum containing a higher interstitial content than that used by Tedmon and Ferriss, prepared the specimens of different grain size by annealing in the temperature range 1500" to 2000° C to minimize any differences in dislocation structure, and found that ky had a value of 1.04 x 107 cgs units, independent of testing temperature. Such behavior is consistent with the dislocations being locked by carbide precipitates so that the generation of free dislocations is an athermal process. The recent work of Gilbert et al.17 also shows that in tantalum there is no significant variation of ky with grain size provided it contains 150 ppm of oxygen. In this case, however, the dislocations are not locked by precipitate and ky is temperature dependent. C. S. Tedmon and D. P. Ferriss (authors' reply)— We would like to thank Dr. Smallman for his interesting comments and discussion to our paper, "The Dependence of Yield Stress on Grain Size for Tantalum and a 10 pct W-90 pct Ta Alloy".18 It was suggested that perhaps the relatively small values obtained by us for ky of tantalum could be attributed to the same cause that accounts for the apparently small values of ky that result when it is determined by the Luders Strain technique. Since our values were obtained by plotting the lower yield stress vs the reciprocal of the square root of the grain size, it is not clear how this could be the case. The values of ky in this experiment have been calculated, using the Luders strain technique. With this method, values for ky on the order of 2 x 105 to 5 x lo6 cgs units were obtained. In spite of this rather large variation, the magnitudes are still small, and there appeared to be no good correlation between ky and the grain size or the yield stress, probably because of the difficulty in accurately extrapolating the work-hardening portion of the curve back to zero plastic strain. As was shown in the original data,18 there was little work hardening in any of the curves, at any temperature. In his discussion, Dr. Smallman also points out how ky has been observed to increase with increasing grain size, when determined by the Luders strain technique. There are at least two possible explanations for this. In the first case, if it is assumed that the bulk of the interstitial impurities are concentrated at the grain boundaries, then, of course, the available grain boundary area would decrease with increasing grain size, thus presenting less area for the interstitials, which would then presumably increase the concentration within the grains, thereby increasing the locking of the dislocations. In the second case, the increase in ky with increasing grain size would be attributed to the nature of the grain boundary itself. One of the several ways of deriving the Hall-Petch equation19 is based on the stress concentration arising from a pile-up of dislocations at the boundary. The ability of the stress concentration to unlock a source in a neighboring grain would depend on the strength of the grain boundary. As is well-known, the nature and struc-

Jan 1, 1963

In the special number of the magazine Ocl und Kohle (Oct. 22, 1940) dedicated to the petroleum industry of southeastern Europe, there appeared several articles relating to Yugos1avia.t In the first article the author calls attention to the fact that there are two fields in Yugoslavia that have been explored and exploited by more than 100 wells at variable depths during the last 50 years. These are on the Island of Mura, at the confluence of the Mura and the Drave. Additional exploration drilling in various other parts of the country has been without result, but certain ones of these explorations have revealed the existence of gas horizons that have some value, which are being used for industrial purposes. Table I shows the production of oil and gas between the years 1900 and 1939. TaBle —Production of Oil ad Gas in Yugoslavia . n,i Natural Natural Years oil Years Gas, Cu. Years Gas, Cu. Meters Meters T900-1929 10,000 1923 507,000 1934 1,388,000 1930 277 1924 567,000 1935 1,401,000 1931 202 I92S 691,000 I936 1,483,000 t932 289 19261 1.472,000 1937 1,841,000 1933 629 1927! 1,732,000 193s 2,431,000 1934 324 1928 2,515,000 1939 2,628,000 1935 260 1929 4,700,000 1936 137 I93O 5.345,200 1937 463 1931 6,376,153 1938 1,091 1932 1,661,737 1939 1,114 1933 959,574 The number of wells drilled from 1884 to the present time was in the neighborhood of zoo and the total footage drilled was 65,000 m. up to July I940 They may be divided roughly as follows: 12 wells in the region produce natural gas; 134 wells in the two old fields of the Island of Mura and about zo wells for the exploration of the zone of the Flysch, in Bosnia. The depths of 50 of these wells was roo to 200 m.; the depths of 50 wells oil the Island of Mura ran from zoo to 500 m.; of 2 ; wells, 500 to I000 m.; of ro wells from rooo to 2000 meters. The deepest well drilling in Yugoslavia has reached 2000 in. In the Valley of Morave, in the region of Selnica, I00 km. south of Nish, there have been discovered asphalt-bearing limestones, which contain about 15 per cent asphalt and are associated with bituminous clays. Later, at Paclenica, important deposits of natural gas were found. A zone rich in gas and oil seepages runs through Pracec, Miklenska, Bujavica, etc. The author studied these regions from the point of view of stratigraphy and tectonics. The anticlines and gas-bearing domes are concentrated in two zones: the region of the Drave (Island of Mura) and the region of the Save, southeast of zagreb, which contains the gas field in exploitation in the region of Bujavica. A full description of the productive region of the Island of Mura has been given by Egon Bohm, a geologist, in the second article mentioned. This includes a history of the exploration and exploitation that took place under Hungarian rule in this region. The most important exploration was effected between 1884 and 1885 by Stavenov and Stinger, of Vienna. In 1927 work was done in the same region by Anton Raky, a well-known explorer who makes work of this kind his specialty.

Jan 1, 1941

In the special number of the magazine Ocl und Kohle (Oct. 22, 1940) dedicated to the petroleum industry of southeastern Europe, there appeared several articles relating to Yugos1avia.t In the first article the author calls attention to the fact that there are two fields in Yugoslavia that have been explored and exploited by more than 100 wells at variable depths during the last 50 years. These are on the Island of Mura, at the confluence of the Mura and the Drave. Additional exploration drilling in various other parts of the country has been without result, but certain ones of these explorations have revealed the existence of gas horizons that have some value, which are being used for industrial purposes. Table I shows the production of oil and gas between the years 1900 and 1939. TaBle —Production of Oil ad Gas in Yugoslavia . n,i Natural Natural Years oil Years Gas, Cu. Years Gas, Cu. Meters Meters T900-1929 10,000 1923 507,000 1934 1,388,000 1930 277 1924 567,000 1935 1,401,000 1931 202 I92S 691,000 I936 1,483,000 t932 289 19261 1.472,000 1937 1,841,000 1933 629 1927! 1,732,000 193s 2,431,000 1934 324 1928 2,515,000 1939 2,628,000 1935 260 1929 4,700,000 1936 137 I93O 5.345,200 1937 463 1931 6,376,153 1938 1,091 1932 1,661,737 1939 1,114 1933 959,574 The number of wells drilled from 1884 to the present time was in the neighborhood of zoo and the total footage drilled was 65,000 m. up to July I940 They may be divided roughly as follows: 12 wells in the region produce natural gas; 134 wells in the two old fields of the Island of Mura and about zo wells for the exploration of the zone of the Flysch, in Bosnia. The depths of 50 of these wells was roo to 200 m.; the depths of 50 wells oil the Island of Mura ran from zoo to 500 m.; of 2 ; wells, 500 to I000 m.; of ro wells from rooo to 2000 meters. The deepest well drilling in Yugoslavia has reached 2000 in. In the Valley of Morave, in the region of Selnica, I00 km. south of Nish, there have been discovered asphalt-bearing limestones, which contain about 15 per cent asphalt and are associated with bituminous clays. Later, at Paclenica, important deposits of natural gas were found. A zone rich in gas and oil seepages runs through Pracec, Miklenska, Bujavica, etc. The author studied these regions from the point of view of stratigraphy and tectonics. The anticlines and gas-bearing domes are concentrated in two zones: the region of the Drave (Island of Mura) and the region of the Save, southeast of zagreb, which contains the gas field in exploitation in the region of Bujavica. A full description of the productive region of the Island of Mura has been given by Egon Bohm, a geologist, in the second article mentioned. This includes a history of the exploration and exploitation that took place under Hungarian rule in this region. The most important exploration was effected between 1884 and 1885 by Stavenov and Stinger, of Vienna. In 1927 work was done in the same region by Anton Raky, a well-known explorer who makes work of this kind his specialty.

Jan 1, 1941

By J. D. Miller, J. E. Sepulveda

Tar sand deposits in the state of Utah contain more than 25 billion bbl of in-place bitumen. Although 30 times smaller than the well-known Athabasca tar sands, Utah tar sands do represent a significant domestic energy resource comparable to the national crude oil reserves (31.3 billion bbl). Based upon a detailed analysis of the physical and chemical properties of both the bitumen and the sand, a hot-water separation process for Utah tar sands is currently being developed in our laboratories at the University of Utah. This process involves intense agitation of the tar sand in a hot caustic solution and subsequent separation of the bitumen by a modified froth flotation technique. Experimental results with an Asphalt Ridge, Utah, tar sand sample indicated that percent solids and caustic concentration were the two most important variables controlling the performance of the digestion stage. These variables were identified by means of an experimental factorial design, in which coefficients of separation greater than 0.90 were realized. Although preliminary in nature, the experimental evidence' gathered in this investigation seems to indicate that a hot-water separation process for Utah tar sands would allow for the efficient utilization of this important energy resource. The projected increase in the ever-widening gap between the domestic energy demand and the domestic energy supply for the next few years has motivated renewed interest in energy sources other than petroleum, such as tar sands, oil shale and coal. Although a number of research programs on the exploitation of national coal and oil shale resources have already been completed, very few programs have been initiated on the processing of tar sand resources in the United States. In recognition of their significance as a domestic energy resource, investigators at the University of Utah have designed an extensive research program on Utah tar sands. An important phase of this program, and the main subject of this publication, is the development of a hot-water process for the recovery of bitumen from Utah tar sands, as a preliminary step toward the production of synthetic fuels and petrochemicals. The term "tar sand" refers to a consolidated mixture of bitumen (tar) and sand. The sand in tar sand is mostly a-quartz as determined from X-ray diffraction patterns. Alternate names for "tar sands" are "oil sands" and "bituminous sands." The latter is technically correct and in that sense provides an adequate description. Tar sand deposits occur throughout the world, often in the same geographical areas as petroleum deposits. Significantly large tar sand deposits have been identified and mapped in Canada, Venezuela and, the United States. By far, the largest deposit is the Athabasca tar sands in the Province of Alberta, Canada. According to the Alberta Energy Resources Conservation Board (AERCB),2,3 proved reserves of crude in-place bitumen in the Athabasca region amount to almost 900 billion bbl. To date, this is the only tar sand deposit in the world being mined and processed for the recovery of petroleum products. Great Canadian Oil Sands, Ltd. (GCOS) produces 20 million bbl of synthetic crude oil per year. Another plant being constructed by Syncrude Canada, Ltd. is expected to produce in excess of 40 million bbl of synthetic crude oil per year. According to the Utah Geological and Mineral Survey (UGMS), tar sand deposits in the state of Utah contain more than 25 billion bbl of bitumen in place, which represent almost 95% of the total mapped resources in the United States.4 The extent of Utah tar sand reserves seems small compared to the enormous potential of Canadian tar sands. Nevertheless, Utah tar sand reserves do represent a significant energy resource comparable to the United States crude oil proved reserves of 31.3 billion bbl in 1976.5 Tar sands in Utah occur in 51 deposits along the eastern side of the state.4 However, only six out of these 51 deposits are worthy of any practical consideration (Fig. 1). As indicated in Table 1, Tar Sand Triangle is the largest deposit in the state and contains about half of the total mapped resources. Information regarding the grade or bitumen content of Utah deposits is still very limited. The bitumen content varies significantly from deposit to deposit, as well as within a given deposit. In any event, the information available6-8 seems to indicate that Utah deposits are not as rich in bitumen as the vast Canadian deposits which average 12 to 13% by weight.9 Although many occurrences of bitumen saturation up to 17% by weight have been detected in the northeastern part of the state (Asphalt Ridge and P. R. Spring), the average for reserves in Utah may well be less than 10% by weight. Separation Technology As in any other mining problem, there are two basic approaches to the recovery of bitumen from tar sands. In one

Jan 1, 1979

By D. J. Carney, R. L. Stephenson

A sampling technique has been developed for procuring a sample of sinter representative of the entire depth of the sintering bed. The sampling method involves the use of an open-bottom metal basket that rides on the grate of the sintering machine and when removed contains a sample of the sintered product. Additional data have been obtained to indicate that the tumbler test is a suitable means of measuring sinter strength. IN the last few years additional sintering facilities have been installed in both the Pittsburgh and the Chicago district of the United States Steel Co. Since the construction of these sintering plants made possible the use of higher percentages of flue-dust sinter in our blast-furnace burdens, it became important to study means of controlling the quality of sinter to obtain optimum results in the blast furnace. For controlling an operating process, it is necessary first to establish standards by which the quality of the product can be judged. For sinter, it appeared that an important property was its strength or its resistance to degradation during transportation and charging into the furnace. Consequently work was undertaken to establish a standard for sinter strength that could be used both for controlling sintering-plant operations and for correlating sinter quality with blast-furnace performance. The first problem in setting up a standard was that of procuring a sample that would be representative of the sinter made under any particular set of conditions at the sintering plant. Since the United States Steel Co. sintering plants discharge the finished sinter either into a large pit or onto a rotary cooler, the sinter becomes inseparably mixed with material sintered 2 hr before or 2 hr afterwards. For this reason the exact identity of the sinter is lost. A sample selected as the cooler is discharged, or as the sinter is removed from the pit, cannot be said to be truly representative of the sinter made at any specific time. Sampling The first attempt to procure a sample that would be representative of a specific sinter mix and of specific operating conditions was made by stopping the Dwight Lloyd sintering machine and removing an entire pallet full of sinter. This method, however, proved very difficult to perform and interfered considerably with the operation of the plant. To overcome this difficulty, a sampling method was devised by technologists at South Works enabling them to secure, without interrupting the sintering operation, a sample of about 1 cu ft of sinter, representative of sinter for the full depth of the sintering bed. The South Works method involves the use of a steel-frame-work basket. A typical basket is shown in Fig. 1. The basket has been used both with and without crossbars along the bottom. As long as the crossbars are in the same direction as the grate bars on the sintering machine they do not interfere with the sintering process. The basket is set on an empty grate of the Dwight Lloyd sintering machine before it passes under the swinging feed spout, see Fig. 2. When the basket is removed after it has travelled the length of the sintering machine, it contains the sample. Just before the basket is removed, the sinter is scored and chipped to facilitate removal of the sample from the sinter bed. A view of the basket after its removal is shown in Fig. 3. Although the sampling method was originally designed for use on a Dwight Lloyd sintering machine, it can also be used on the Greenawalt type of machine. When used on the Greenawalt-type machine, the basket is placed on the sintering grate before the charging car passes over it, and finally it is removed just before the pan is dumped. Testing After a method of obtaining a representative sample of sinter had been developed, the next step was to select a method of measuring its strength. The irregular shape and size of the sinter pieces precluded the use of a simple compression test for determining strength; consequently, the shatter test and tumbler test were investigated. To perform the shatter test, a sample of sinter, approximately 5 lb, is dropped from a hinged-bottom box at a height of 3 ft onto a steel plate. The broken sinter is sieve-analyzed after a specified number of drops. The tumbler test is performed with the use of a standard ASTM coke-tumbling drum. The drum is 3 ft in diam and is equipped with two lifter bars diametrically opposite one another on the inner periphery of the drum. The drum is rotated at a speed of 24 rpm for 200 revolutions, and after tumbling the sample is sieve-analyzed. To express as single numbers the results of sieve analyses after shattering or tumbling, the method suggested by R. E. Powers1 was employed. This method involved plotting the size of the sieve openings on a logarithmic scale and the cumulative per cent larger than each sieve on a probability scale as described by J. B. Austin.' By interpolating from the plotted data, which in most cases approximated

Jan 1, 1954

By C. D. Stahl, S. M. Farouq Ali, W. E. Culham

One- and two-dimensional mathematical models have been developed that simulate transient, two-phase flow of hydrocarbon mixtures in porous media in a manner that accounts for interphase mass transfer. Numerical simulations of one-dimensional depletion-drive experiments using a two-component hydrucarbon fluid were used to establish the validity of the mathematical models. In addition, the experimental and numerical data were used to demonstrate that production rate had a relatively insignificant effect on the recovery of individual hydrocarbon components from the experimental system, and that attainment of equilibrium between phases is possible for a wide range of liquid and vapor velocities in reservoirs containing light hydrocarbon fluids. Results of some two-dimensional numerical simulations are also presented. INTRODUCTION This study was undertaken to develop a mathematical model that would simulate transient, two-phase flow of hydrocarbon mixtures in porous media under conditions that result in interphase mass transfer and to test the validity of the assumptions used to set up the model. In addition, the study was designed to determine if production rate influences the recovery of individual hydrocarbon components from reservoirs producing by depletion drive. Two-phase flow in porous media, with interchange of components between the two phases, is important in many petroleum recovery processes. Studiesl-5 conducted within the last 3 years have outlined methods of solving multiphase flow problems incorporating mass transfer. Some of these studies have also indicated the importance of accounting for mass transfer under various producing conditions. An earlier work6 first demonstrated the importance of combining relative permeability data with equilibrium ratios in compositional balance methods. The mathematical model presented in this paper is formulated so that a phase behavior package, as described in previous papers,5,17 is not required as an integral part of the routine employed to solve for the primary dependent variables. The finite difference formulation is designed so that all the primary variables can be solved for simultaneously. This is accomplished by utilizing one basic set of equations. These innovations, which are in contrast to other models2,3,5 but are similar in some respects to the approach used by Taylor,l render the total problem computationally simpler than any of the previously referenced formulations. The mathematical model was developed by combining Darcy's law with a continuity equation for each hydrocarbon component. The principal assumptions invoked in the formulation were (1) that capillary forces and diffusional effects are negligible, and (2) that thermodynamic equilibrium exists in the reservoir at all times. No assumption as to the type of vaporization process was made in formulating this model. Experimental data were required to complete this study. These were generated by conducting several depletion drive experiments. The experimental apparatus consisted of a sandstone core enclosed in a pressurized casing. The apparatus was designed in such a manner that the core could be charged with a liquid hydrocarbon mixture and depleted at different production rates. The experimental tests were designed to determine the effect of production rate on component recovery. In addition, direct comparison of experimental and mathematically predicted

Jan 1, 1970

By J. G. Eenink, R. A. Cunningham

One phase has been completed of a laboratory invesrigation of formations with relatively high permeability under conditions of overburden, formation and mud coltrmn pressures. The following statements are based on these tests within the limitations described. Drilling rate decreased when mud column pressure was greater than formation pressure. The decrease was primarily due to a layer of cuttings and mud particles held to the hole bottom by the difference in pressure. Much of the force of the bit was wasted in this layer and was unavailable for penetrating virgin formation. Adequate jet velocities helped clean the filter cake and chips away and resulted in increased drilling rate — the higher the jet velocity the faster the drilling rate. Drilling rate increased slightly when formation pres sure was greater than mud column pressure. The i11(reuse resulted from cleaning the hole bottom as formarion fluid flowed into the borehole. Overburden pressure had practically no effect on drilling rate. INTRODUCTION Some formations are difficult to drill in a wellbore hole but are easily drilled under atmospheric conditions in the laboratory. The reasons for this apparent difference in drillability are not fully understood. An increase in formation compressive strength causes a reduction in drilling rate. It has been shown by Griggs1, Handin2 and others3-1 that formation strength increases under conditions of high hydrostatic stress. Based on these results, a decrease in drilling rate would be expected with increased drilling fluid pressure in formations of low permeability having low formation fluid pressure. Such a decrease has been shown by Murray. et a1 and Eckel" This effect holds even when using nitrogcn under high pressure as the drilling fluid (unrc-ported tests from the Hughes laboratory). Poor bottom-hole cleaning may account for decreased drilling rates. Even though it may not be easily explained, evidence of poor cleaning can easily bc seen. One evidence is the increased drilling rate that accompanies decreased balling obtained with addition of oil in some muds used to drill shales.4,7,3 Many variables can affect drilling rates. These may include mud composition and properties, formation properties, bit types and operating conditions, formation fluid, mud column and overburden pressures. These must be studied a few at a time under carefully controlled conditions which simulate actual conditions as closely as possible. The purpose of this work was to investigate effect of overburden, formation and mud column pressures on drilling rates in permeable formations. These tests were suggested by conditions shown in Fig. I. Represented is a borehole drilled into a formation with relatively high permeability. In the vicinity of the hole bottom there exists a mud column pressure, P a formation fluid pressure, P and an overburden pressure, The mud has low water-loss properties and sufficient weight to more than overbalance the formation pressure. Laboratory investigation included a range with formation pressure greater than mud column pressure. NOMENCLATURE Pm mud column pressure = pressure of drilling fluid or mud at hole bottom Pf, formation pressure = pressure of fluid in intersticies of formation

By S. R. Mitchell, M. Weissman, J. H. Brown

It has been demonstrated that for fine grinding operations, the energy input (E) and the product size modulus (k) are related by an equation of the form E = Ak1-n (1) where "A " and "n" are constants.(1) Also, it was suggested that "A" is determined by the comminuting device and that "n" depends primarily on the rock. It was the purpose of this investigation to study these constants by determining the form of an energy-size reduction relationship over an extreme range of energy inputs, and at the same time, to evaluate the minimum particle size produced during grinding. The comminution system consisted of a stainless steel laboratory ball mill in which quartz was ground at 60% solids for times ranging from 1 minute to 700 hours. The mill which measured 10" diameter by 12" long was loaded with about 60 pounds of balls ranging in diameter from 3/8" to 1-1/8" and was driven at 55 r.p.m. Quartz was selected for the tests because it is a brittle material for which size distribution data were expected to show a constant distribution coefficient regardless of product size.(1) This permitted specification of product size by a size modulus alone. The quantities measured were the time of grind (measure of energy input) and the specific surfaces of the products. Since no accurate method is available yet for determining particle size directly in the very fine size ranges specific surface data were used for this purpose. The relationship between these two quantities may be expressed as where Cis a constant which includes a shape factor, p is the specific gravity of the solid,_S is the specific surface and D is an equivalent particle diameter. For a material in which the distribution coefficient does not change with size the term (l/S) may be substituted for k in equation L if the proportionality constant "A" is changed correspondingly. This substitution was made in this work but in addition, screen analyses were made on several products to verify the equivalence of the two methods of size specification. EXPERIMENTAL DETAIL QUARTZ Quartz from two sources was tested. One series of tests was made on a clear, high purity crystalline quartz from Brazil and Arkansas, and another series was run on massive quartz fromNewHampshire. Grinding charges of the high purity material consisted of 2000 gram samples of -3-8 mesh particles, while the charges of the New Hampshire quartz were 500 gram samples of a -35+48 mesh screen fraction. The difference in the sizes of these two series of charges was not significant in view of the extreme size reduction accomplished, and the New Hampshire samples were smaller to permit ginding to finer sizes in shorter times than required for 2000 gram lots. SAMPLING AND ANALYSIS Separate charges were ground for all tests in which the grinding times were less than one hour and for all tests on New Hampshire quartz since these products were difficult to sample accurately without affecting the entire charge. During the long time grinding tests the mill was

Jan 1, 1961

By Raymond D. Sloan

During 1944, substantial gains were recorded in practically every phase of the petroleum industry in Oklahoma. With the spotlight of activity focused on other states during the more recent years, the production momentum that was achieved in Oklahoma through the discovery and development of such outstanding pools as Oklahoma City, Fitts. and major pools in the Seminole area during the late 1920's and early 1930's was practically spent by the end of 1943. However, with much of the industry's exploratory efforts greatly reduced in Illinois, Indiana and Kentucky, the strong demand for Mid-Continent crude and the optimism created by some discoveries in the state toward the end of 1943, Oklahoma has again become the recipient of much activity and interest, with its 1944 drilling activity, production and refinery runs to stills registering substantial gains over the preceding year. Development and Exploration Since the discovery of the West Edmond pool in April of 1943, with the substquent drilling and development boom, the scene of exploratory activity has shifted from the central. south and east portions of the state to the general trend of the . Granite Ridge, where leasillg and wild-catting were greatly activated within a wide radius of this high subsurface feature. Leasing activity, however, jumped the confines of the Granite Ridge trend and extended into all of the northwestern counties in the state. Seismograph activity at once began following and keeping pace with this shift in the leasing and exploration locale, by the concentration of a large number of seismograph crews in the area. At one time during the year,35 crews were concentrated in Oklahoma, the majority being west of the Granite Ridge and extending into the northwestern counties. To illustrate the increase in geophysical work done in the state during the year, records indicate that 377 crew months were worked in 1944, which is a 21 per cent increase over the 311.5 crew months activated in 1943. With the discovery in 1943 of several encouraging prospects and the finding of West Edmond as an incentive, drilling operations increased some 54.3 per cent, for a total of 1986 completions during the year. This is compared with a figure of 1287 for wells drilled in 1943, an increase of 699 wells. The industry, with a total of 1064 oil producers, enjoyed a success percentage of 53.6 per cent. Gas wells completed were 223, while 699 dry holes were drilled. In 1943, completions included 587 oil wells, 127 gas wells, and 573 dry holes. The number of oil producers completed during the Year was almost double the numher of wells in 1943. Of the total drilling wells completed in the state, 350, or 17.6 per cent, were wildcat operations—an increase of 5.1 per cent over the 194.3 total of 333. The wildcat success percentage was 13.4 per cent, 47 successful wells being completed for an average initial potential of 160 bbl. In the state, 23 gas wells were completed with an average initial production of

Jan 1, 1945

By J. F. Gallie

From the standpoint of exploration and new discoveries, South Arkansas had a discouraging year in 1944. In January there were 73 untested drilling blocks, and 60 additions were made during the year. Of the total, 65 were condemned by dry holes on or adjacent to them, while 5 proved productive. The latter resulted in the discovery of the Salem Church, Calhoun Wilks, and Strong fields, and the extension of the Village field to the southeast. The year ended with 63 untested blocks, 41 of which were scheduled for tests to the Smackover lime, the state's deepest oil-producing formation. Geophysical activity in 1944 totaled 164 crew weeks, of which 110 were seismic, 43 gravity, and 11 radioactive (Fig. I), in contrast with 674 crew weeks reported for 1943, of, which 563 were seismic and in gravity, Five seismograph and two gravity crews were engaged in exploration in South Arkansas during January 1944. but only two seismograph crews were working there at the end of the year. In addition, exploratory core drills were used for a total of 38 weeks during the year. The year's few discoveries, none of which is considered as having added appreciably to state reserves, were the meager results obtained in the drilling of 62 wildcat wells 2000 ft. or deeper. Of this total (a detail of which appears in Table 2), 42 were drilled to the Smack-over lime or below, The average depth of the 62 wells was 5384 ft., as compared with 4992 ft. in the preceding year, while the average depth of the 42 wildcats in the lime was 5913 ft. in 1944. Fig. 2 has been prepared to show all , wells penetrating to the Reynolds oölitic lime. The 1944 dry holes are circled lightly, while discoveries are heavily circled and shown as producers. Smackover lime fields, producing from the Reynolds oolitic lime member, are indicated in black, and the locations of dry wildcat wells drilled prior to 1944 are also shown. Table 3 provides a listing of all Smackover lime wells drilled through the end of 1944. Production for the year was slightly more than in 1943, totaling 29,282,747 bbl., as compared with 27,521,943 bbl. for the preceding year. The increase was 1,760,804 bbl., or 6.4 per cent, bringing production to a peak exceeded only in the 1923 to 1928 period of flush production from the older shallow fields, principally Smackover. The increase was largely attributable to additional development in the Dorcheat-Macedonia, Atlanta, Fouke, Stephens-Smart, Haynesville, and Village fields, little production deriving from 1944 discoveries. These six fields accounted for loo of the 122 field wells drilled during the year. Annual and cumulative production for the producing life of the state are presented graphically in Fig. 3. Fig. 6 has been prepared by the author to show the producing ages and present known productive limits of all fields in South Arkansas. Production data, as usual, are assembled in Table 2, and provide a further breakdown of producing formations in the individual fields.

Jan 1, 1945

By W. J. Millard

It is well known that the excellent mica from Brazil played a most vital role in World War 11. Increased production from Brazil was necessary and with the assent of the Brazilian Government engineers from the united' States went down to study the problem. One of the means advocated by a former Chief of Metals and Minerals Branch of the Board of Economic Warfare, Mr. Carl Schwegler, was the use of stripping machinery. On a pegmatite dike favorable for stripping, a small number of miners could produce as much mica per day as an underground mine employing many times the same number of men. Type of Deposit Favorable The most favorable type of pegmatite for stripping is that which lies more or less parallel to the surface of the slope of a hill. Pegmatite dikes that cut directly into a hill are not suitable for stripping except where a long outcrop extends along the slope. Then just the outcrop could be stripped to some advantage. The ratio at no time should be greater than 1: 3; i.e., I cu. m. of pegmatite to 3 cu. m. of waste, unless the pegmatite was extremely rich throughout the dike. The Economics of Stripping When a mineral is badly needed to win a war, not too much heed can be paid to costs. The object is to produce rapidly and in quantity. Again, a deposit containing a superior grade of mica may be worked although the pegmatite is more costly to work than another deposit containing a considerably greater amount of poor-grade mica. For example: one deposit may have a little mica that will run 11 per cent good stain and better when trimmed and contain 2 kg, of crude mica per cubic meter of pegmatite, while another may run 0.2 per cent of good stain and better, and have a mica content of 10 kg. per cubic meter. A number of stripping projects were found in the vicinity of Raul Soares, Minas Gerais. These were sampled in the various old workings by cutting out a cubic meter of pegmatite and saving the usable mica. Because of the proximity to the surface of most stripping projects, about 34 per cent only of the mica is good for trimming. The other 66 per cent is too badly weathered to be of use. In Table I are given data collected on five properties near Raul Soares. In Table I, "value per cubic meter of pegmatite" means the value of the qualified mica obtainable. This does not include the half trimmed and splittings. By "value per cubic meter of matter" is meant the value of the qualified mica per cubic meter of material excavated; i.e. both waste and pegmatite. This is the figure that should be used to determine whether the operation will be profitable or not, as the cost of stripping is about $0.14 per cubic meter. With new machinery not subject to breakdowns, the cost per cubic meter should be around $0.10 to $0.15. To this should be

Jan 1, 1948

By W. J. Millard

It is well known that the excellent mica from Brazil played a most vital role in World War 11. Increased production from Brazil was necessary and with the assent of the Brazilian Government engineers from the united' States went down to study the problem. One of the means advocated by a former Chief of Metals and Minerals Branch of the Board of Economic Warfare, Mr. Carl Schwegler, was the use of stripping machinery. On a pegmatite dike favorable for stripping, a small number of miners could produce as much mica per day as an underground mine employing many times the same number of men. Type of Deposit Favorable The most favorable type of pegmatite for stripping is that which lies more or less parallel to the surface of the slope of a hill. Pegmatite dikes that cut directly into a hill are not suitable for stripping except where a long outcrop extends along the slope. Then just the outcrop could be stripped to some advantage. The ratio at no time should be greater than 1: 3; i.e., I cu. m. of pegmatite to 3 cu. m. of waste, unless the pegmatite was extremely rich throughout the dike. The Economics of Stripping When a mineral is badly needed to win a war, not too much heed can be paid to costs. The object is to produce rapidly and in quantity. Again, a deposit containing a superior grade of mica may be worked although the pegmatite is more costly to work than another deposit containing a considerably greater amount of poor-grade mica. For example: one deposit may have a little mica that will run 11 per cent good stain and better when trimmed and contain 2 kg, of crude mica per cubic meter of pegmatite, while another may run 0.2 per cent of good stain and better, and have a mica content of 10 kg. per cubic meter. A number of stripping projects were found in the vicinity of Raul Soares, Minas Gerais. These were sampled in the various old workings by cutting out a cubic meter of pegmatite and saving the usable mica. Because of the proximity to the surface of most stripping projects, about 34 per cent only of the mica is good for trimming. The other 66 per cent is too badly weathered to be of use. In Table I are given data collected on five properties near Raul Soares. In Table I, "value per cubic meter of pegmatite" means the value of the qualified mica obtainable. This does not include the half trimmed and splittings. By "value per cubic meter of matter" is meant the value of the qualified mica per cubic meter of material excavated; i.e. both waste and pegmatite. This is the figure that should be used to determine whether the operation will be profitable or not, as the cost of stripping is about $0.14 per cubic meter. With new machinery not subject to breakdowns, the cost per cubic meter should be around $0.10 to $0.15. To this should be

Jan 1, 1948

By R. Taggart, R. H. Ericksen, D. H. Polonis

Transmission electron microscopy techniques hare been used to study the spontaneous marfensite phase that forms during the thinning of Ti-Cr alloys. The structure of this phase has been found to differ. from the hcp martensite and has been identified as a clisforted fcc lattice. The spordaneous marlensile in this system is interpreted as a metastable isomorph of the equilibriunz TiCr, phase. WHEN thin-foil specimens for transmission electron microscopy were prepared from metastable alloys of the Ti-Cr system, a spontaneous transformation occurred' when the foils contained the untrans-formed 0 phase or two-phase mixtures of the 0 and w phases. The purpose of this paper is to examine the relationship that exists between the crystal structure of the spontaneous product and that of the matrix. The differences between the martensite transformation that occurs in the bulk material and the spontaneous reaction that occurs during the preparation of thin foils are compared with the observations that have been reported for similar phenomena in a commercial titanium alloy.' A spontaneous reaction that exhibits characteristics similar to a martensite transformation has been reported to occur in thin foils prepared from the retained metastable structures of several alloy systems.1'3"7 ~ull~ observed such a spontaneous reaction in 0 brass and attributed the apparent increase in the martensite transformation temperature to a decrease in the non-chemical free energy associated with bulk strain energy. He also showed that the spontaneous product did not have a specific habit plane, but that its orientation relationships to the parent matrix were dependen on the thickness of the transforming region. It has also been pointed out by Breedis6 that in thin films the absence of the constraints imposed by a bulk matrix can cause different requirements for the martensite plate interface and lattice distortion which is reflected bv a dependence of the transformation mechanism on the specimen geometry. In the case of the spontaneous reaction reported for Au-Zn alloys7 a definite habit plane was determined and shown to correspond with the habit plane calculated on the basis of the Bowles and Mackenzie crystallographic theory of the martensite transformation.'- lo EXPERIMENTAL PROCEDURE A series of titanium alloys containing 5.5, 5.9, 6.4: 7.6, 8.0, 11.3, and 18.7 at. pct Cr were prepared from high-purity (iochrome) chromium and iodide-grade titanium by levitation melting. The experimental procedure used for sectioning the alloys and for thin-foil preparation has been reported in an earlier paper.' X-ray diffraction patterns were obtained using a 114-mm Debye-Scherrer camera and Cu-Ka radiation. In order to minimize fluorescent radiation from the specimen, a strip of aluminum foil 0.002 in. thick was placed in contact with the X-ray film. After a specimen was removed from the camera: an additional X-ray experiment was performed, to insure that no spurious lines were obtained due to the presence of the aluminum foil. The interstitial impurity analysis for the alloys used in this study was as follows: oxygen, 100 ppm: hydrogen, 30 ppm: nitrogen, 28 ppm. EXPERIMENTAL RESULTS Electron diffraction patterns were obtained from several individual crystals of the spontaneous marten-

Jan 1, 1970

By J. R. Johanson

With the continuing development of automatic operations, it is important that reliable discharge rates through hoppers and bins be obtained for bulk solids. In most applications a feeder is used to control the feed rate. The selection of this feeder size is usually based on the rated ability of the feeder to move material without regard to the capability of the material to flow through the opening in the hopper onto the feeder. The dis charge rate through a given hopper opening may vary from zero to a theoretical maximum depending on the flow properties of the material. This paper presents an equation for the maximum discharge rate for a bulk material as a function of the hopper shape, the opening size and the flow properties of the material. Discharge-rate tests on various materials indicate good agreement with calculated values for opening sizes from a fraction of an inch to several feet. As part of a program to develop improved design criteria for bulk storage and handling facilities, a specific study of the methods involved in calculating the rate of discharge of bulk solids from hoppers and bins has been made. This study is closely related to earlier work done on the prevention of flow problems through the choice of proper design geometry of hoppers and bins1. The present paper is an extension of this earlier work, and therefore many of the concepts previously derived and explained will be used without explanation. Two types of flow will be considered: plane-strain flow in a wedge-shaped hopper and axial-symmetry flow in a conical hopper. The literature contains many studies on discharge rate for regular-shaped dry particles2*3 such as glass beads, lead shot and coarse dry sand. However, the work is empirical in nature, and it is difficult to extrapolate the given information to apply to moist, sticky materials typical of many industries. Some studies4 have shown a decrease in discharge rate with an increased moisture content of the material. However, no attempt was made to correlate the discharge rate with moisture. The method of calculating discharge rate presented in this paper was derived analytically from the basic knowledge of the flow of frictional, cohesive, granular solids, The equation includes the effect of material cohesion (stickiness), hopper configuration, opening size and bulk density of the material. THEORETICAL WORK The basic assumptions made in deriving the discharge rate equation are as follows: 1) The material is a frictional, cohesive, granular solid. 2) The flow is considered continuous and without significant density change. 3) The material at the hopper opening forms a continuously failing arch that is in dynamic equilibrium. 4) The strength of the material in the arch, which must be overcome for the arch to fail, is caused by the pressure in the flowing mass near the hopper opening. 5) This pressure at the hopper opening can be calculated5. If we consider the dynamic equilibrium of an arch of uniform thickness such as that shown in Fig. 1,

Jan 1, 1965

By W. J. Millard

IT is well known that the excellent mica from Brazil played a most vital role in World War II. Increased production from Brazil was necessary and with the assent of the Brazilian Government engineers from the United States went down to study the problem. One of the means advocated by a former Chief of Metals and Minerals Branch of the Board of Economic Warfare, Mr. Carl Schwegler, was the use of stripping machinery. On a pegmatite dike favorable for stripping, a small number of miners could produce as much mica per day as an underground mine employing many times the same number of men. TYPE OF DEPOSIT FAVORABLE The most favorable type of pegmatite for stripping is that which lies more or less parallel to the surface of the slope of a hill. Pegmatite dikes that cut directly into a hill are not suitable for stripping except where a long outcrop extends along the slope. Then just the outcrop could be stripped to some advantage. The ratio at no time should be greater than I:3; i.e., I cu. m. of pegmatite to 3 cu. m. of waste, unless the pegmatite was extremely rich throughout the dike. THE ECONOMICS OF STRIPPING When a mineral is badly needed to win a war, not too much heed can be paid to costs. The object is to produce rapidly and in quantity. Again, a deposit containing a superior grade of mica may be worked although the pegmatite is more costly to work than another deposit containing a considerably greater amount of poor-grade mica. For example: one deposit may have a little mica that will run II per cent good stain and better when trimmed and contain 2 kg. of crude mica per cubic meter of pegmatite, while another may run 0.2 per cent of good stain and better, and have a mica content of 10 kg. per cubic meter. A number of stripping projects were found in the vicinity of Raul Soares, Minas Gerais. These were sampled in the various old workings by cutting out a cubic meter of pegmatite and saving the usable mica. Because of the proximity to the surface of most stripping projects, about 34 per cent only of the mica is good for trimming. The other 66 per cent is too badly weathered to be of use. In Table I are given data collected on five properties near Raul Soares. In Table I, "value per cubic meter of pegmatite" means the value of the qualified mica obtainable. This does not include the half trimmed and splittings. By "value per cubic meter of matter" is meant the value of the qualified mica per cubic meter of material excavated; i.e. both waste and pegmatite. This is the figure that should he used to determine whether the operation will be profitable or not, as the cost of stripping is about $0.I4 per cubic meter. With new machinery not subject to breakdowns, the cost per cubic meter should be around $0.10 to $0.I5. To this should be

Jan 1, 1946

SEC. 1. The membership of the Institute shall comprise seven classes, namely: 1. Members; 2. Honorary Members; 3. Senior Members; 4. Associates; 5. Junior Members; 6. Rocky Mountain Members; 7. Junior Foreign Affiliates. All shall be equally entitled to the privileges of membership, excepting that no member of any class whose residence shall be outside of the United States of America, except those residing in Mexico and Canada, shall be entitled to vote. SEC. 2. Members: A person to be eligible for election or transfer into the. class of Member must be at least 27 years of age and must have had at least six years' employment in the practice of engineering, mining, geology, metallurgy, or chemistry, during at least three years of which he must have held positions of responsibility in one or more of these fields. This shall not apply to Rocky Mountain members. SEC. 3. Honorary Members: Each candidate for Honorary Membership must be recommended by at least ten Members and must be elected by ballot at a meeting of the Board of Directors by the unanimous vote of the Directors present; provided, however, that the number of Honorary Members shall at no time exceed twenty. Such election shall automatically act as a transfer of such member from any other class of membership to which he may belong. SEC. 4. Senior Members: The Directors may, at the special request of any Member or Associate who has reached the age of seventy years and who has paid dues for thirty years, transfer such Member or Associate to the class of Senior Member; and in special cases may waive either the requirement as to age or as to the number of years for which dues have been paid. SEC. 5. Junior Members: A person to be eligible for election to the class of Junior Member must be qualified -through education or experience to hold a subordinate position in engineering, mining, metallurgical, geological or chemical work. He shall be not more than 33 years of age at the time of election. .SEC. 8. Associates: A person eligible for election or transfer into the class of Associate shall be one, who, in the opinion of the Committee on Admissions and the Board of Directors, is suitable for such election or transfer by reason of his interest in or connection with mining, geology, metallurgy, or chemistry. SEC. 7. Rocky Mountain Members: Membership in a special class of life memberships without annual or other dues and without right to receive the printed literature distributed to other members or the right and privilege of wearing the official badge of technical members, is given to all members of the Rocky Mountain Club at the time of consolidation. These special Rocky Mountain memberships shall be claimable by any member of the Rocky Mountain Club, not already a member of the Institute, availing himself of the privilege within ninety days after mailing of notice to him, by the Secretary of the Institute. SEC. 8. Junior Foreign Affiliate: Upon appropriate action by the Board of Directors and with its approval, members in good standing of any technical society having its headquarters outside of the United States of America, Canada, or Mexico,

Jan 1, 1940