Search Documents

By George C. Lipsey

FOR many years the mining of coal and other minerals occurring in sedimentary formation has been planned according to information obtained by diamond drilling. Placer mines have used churn drills for sampling and estimating deposits, and some metal mines have used the method either wholly or in part. The diamond drill method, therefore, although not entirely new, has come into more prominence in the last few years. Several large-scale developments now being planned or in the production stage have used this method in varying degree. This paper will describe the operations of the Howe Sound Exploration Co., Ltd. at Snow Lake, Manitoba, where a gold mine was developed and equipped for mining and milling wholly on the basis of diamond drilling results. The property has been in operation nearly 3 years, so that results can be compared with the estimates and assumptions made from the diamond drilling. In developing the subject it is necessary to outline as briefly as possible the normal method of mining, from discovery to production stage. Table I sets forth these various stages, the methods of financing, the percentage of total cost, and the relative time interval required for each section of the work. The plan presented in this table cannot be applied to all mining developments, but it gives a basis on which to work. In Stages 2 and 3 the investment is moderate, and if the property does not respond favorably the options or interests can be terminated. Stage 4 represents a sizeable investment, and it is more difficult to drop the property after this part of the operation is completed. Stages 2, 3, and 4 may be repeated several times by different groups or companies before Stage 5 is reached, but when Stage 5 has been started the total capital investment must be provided and there can be no turning back. It will be noted that there is a total of 116 units of investment in the overall cost if one company proceeds with Stages 2, 3, 4, and 5 in order. Stage 4, however, consumes time which is important when markets are being considered. Bringing a property into production when metal prices are low is much more difficult than when the prices are high, and if this stage can be set aside production is attained more quickly. When Stage 4 is eliminated the total becomes 86 units of cost for Stages 2, 3, and 5, but an adjustment of 14 units must be added to Stage 5. These 14 units represent the useable value of Stage 4. Thus Stage 4 represents an investment of 30 units if carried on before Stage 5, but 14 units if included with Stage 5. This is accounted for by the fact that development work can be accomplished more cheaply under Stage 5, since everything that is done is planned on a production basis with permanent services. Stage 4 must of necessity be governed by expenditures, and a great many temporary services, including housing and power plants, must be provided. The prospect shaft is a small one, and drifts,

Jan 1, 1952

By R. L. Robinson, P. DeFrank, R. F. Hatfield

Production equal to or greater than open-hole completions is possible through perforated completions if the flow paths throughout the perforations are free of obstructions.' Previous investigations have indicated that, in some cases, maximum productivity is not obtained due to plugging of the perforations with debris from the wellbore fluid and/or debris from the perforating process. Control of wellbore fluid and proper well completion techniques are very important in the elimination of perforation plugging.' Selection of properly designed perforating equipment also is quite important for maximum well productivity. Studies made with the use of the perforation flow laboratory, in which perforating tests are run under simulated well conditions, have shown that perforation plugging occasionally occurs. Plugs in perforations result from various causes. One of the contributors to plugging that can be controlled by the perforating device is the slug or carrot resulting from jet perforating. This condition has undergone extensive study, and previous developments have resulted in a decrease in carrot size—but not in complete elimination. It was felt that, if a method of completely eliminating the carrot could be found (with charge performance equal or better than previously obtained), a significant contribution would have been made. Such a charge has been developed and is reviewed in this paper. THEORETICAL CONSIDERATIONS OF CARROT ELIMINATION There are several popular theories of how a shaped charge jet is formed."," One of these is the hydrodynamic theory, wherein the liner is treated as a fluid under the impact of the explosive forces. Another is the spalling theory, where the interaction of shock waves in the liner causes small particles to be torn from the inner surface of the liner and projected at high velocity toward a common focal line to form the jet. It is not unreasonable to assume that both of these theories may be partially correct. For purposes of our discussion, it is simpler to use the second theory. Generally speaking, when a plate of metal is attacked by an explosive shock pulse, the pulse is transmitted through the metal at high velocity, as shown in Fig. 1(A). Pressures at the shock front may be 6 X 10' psi or higher. This shock wave (a compression wave) upon striking a free surface, Fig. 1(B), is partially reflected back as a tension wave. The interaction of the tension wave with the decaying compression wave causes a stress concentration in the metal; when this stress becomes great enough, a failure occurs and a small particle flies off and is projected away from the metal surface, as shown in Fig. 1(C). A new free surface is then presented to the shocks; the process repeats itself until there is no longer enough energy to cause failure, leaving a portion of the metal relativcly untouched. See Fig. 1(D). To apply this theory to a perforating shaped charge, consider the flat plate to be folded into a somewhat conical shape, with a column of explosive placed in intimate contact with the cone (Fig. 2). An explosive initiator detonates the charge, creating a shock wave

By Henry Faul

Measurement of radioactivity of rocks and ores has developed into a complete method of geophysical exploration. The problem falls into three natural categories: (1) surface radiation measurement in the field and underground; (2) radioactivity logging of drillholes; and (3) laboratory analysis of samples. Portable counters are used in the field for tracing radioactive formations and for quanti tative estimates of radioactive content of surface rock. Wartime advances in battery design make it possible to construct small precision counters that can be carried easily. Even the smallest conventional diamond drillholes can be logged, and actual content of the radioactive element can be measured. The exact grade of very thin active layers or veins can be estimated only approximately. Samples may be analyzed for their alpha, beta or gamma activity, and beta assaying is the best suited to rapid routine operation. Simple calibration with standard samples is sufficient for rough work, but a density correction is necessary when greater precision is required. Samples with high emanating power should be fixed and stored for three weeks to reestablish the radium-radon equilibrium prior to precision assaying. Carno-tite can be fixed by sintering or molding with plastics. Among the important new developments, a low-voltage Geiger tube has been 'designed recently, and crystal scintillation counters hold great promise for high-efficiency gamma-ray measurement. Introduction During the last few years we have witnessed an unprecedented rise in the demand for radioactive minerals, coming as a direct consequence of the discovery of nuclear fission and subsequent developments, the full impact of which is by no means in sight even now. It remains certain, however, that radioactive minerals, and especially those which can be used as ores of uranium and thorium, will continue to be of great interest not only to the prospector and mine operator but also to the various inspection parties envisioned by most of the schemes advanced for the control of nuclear energy on a world-wide basis. It is clear, therefore, that the measurement of radioactivity in rocks in the field as well as in samples in the laboratory has become a method of geophysical exploration of some importance. If we consider the recent technological advances in the general fields of nuclear physics and electronics it seems proper to review the present status of the methods and procedures of radiological exploration. General Considerations Equilibrium If we take a fixed amount of uranium that has just been purified, some of it will immediately start decaying into its first daughter element which in turn decays into its own daughter and so on down the line until the last daughter is reached—a stable isotope of lead. The rate at which a radioactive element decays is frequently expressed as the half-life, which is the length of time required for the decay of half the atoms in a given amount. The half-life is inversely proportional to the decay constanl X which is defined as the number of atoms which decay in one unit of time divided by

Jan 1, 1949

By Henry Faul

Measurement of radioactivity of rocks and ores has developed into a complete method of geophysical exploration. The problem falls into three natural categories: (1) surface radiation measurement in the field and underground; (2) radioactivity logging of drillholes; and (3) laboratory analysis of samples. Portable counters are used in the field for tracing radioactive formations and for quanti tative estimates of radioactive content of surface rock. Wartime advances in battery design make it possible to construct small precision counters that can be carried easily. Even the smallest conventional diamond drillholes can be logged, and actual content of the radioactive element can be measured. The exact grade of very thin active layers or veins can be estimated only approximately. Samples may be analyzed for their alpha, beta or gamma activity, and beta assaying is the best suited to rapid routine operation. Simple calibration with standard samples is sufficient for rough work, but a density correction is necessary when greater precision is required. Samples with high emanating power should be fixed and stored for three weeks to reestablish the radium-radon equilibrium prior to precision assaying. Carno-tite can be fixed by sintering or molding with plastics. Among the important new developments, a low-voltage Geiger tube has been 'designed recently, and crystal scintillation counters hold great promise for high-efficiency gamma-ray measurement. Introduction During the last few years we have witnessed an unprecedented rise in the demand for radioactive minerals, coming as a direct consequence of the discovery of nuclear fission and subsequent developments, the full impact of which is by no means in sight even now. It remains certain, however, that radioactive minerals, and especially those which can be used as ores of uranium and thorium, will continue to be of great interest not only to the prospector and mine operator but also to the various inspection parties envisioned by most of the schemes advanced for the control of nuclear energy on a world-wide basis. It is clear, therefore, that the measurement of radioactivity in rocks in the field as well as in samples in the laboratory has become a method of geophysical exploration of some importance. If we consider the recent technological advances in the general fields of nuclear physics and electronics it seems proper to review the present status of the methods and procedures of radiological exploration. General Considerations Equilibrium If we take a fixed amount of uranium that has just been purified, some of it will immediately start decaying into its first daughter element which in turn decays into its own daughter and so on down the line until the last daughter is reached—a stable isotope of lead. The rate at which a radioactive element decays is frequently expressed as the half-life, which is the length of time required for the decay of half the atoms in a given amount. The half-life is inversely proportional to the decay constanl X which is defined as the number of atoms which decay in one unit of time divided by

Jan 1, 1949

By D. A. Dahlstrom

WITHIN the past decade, certain economic, legal, and technological changes have intensified the necessity for an efficient classifier or deslimer of particles at and below the 200-mesh point. First, mechanization of mining has increased the percentage of clays, silts, refuse, and high-ash material in the run-of-mine coal. Second, many washing plants must rely on smaller percentages of fresh make-up water because of shortages, economic factors, or legal restrictions on waste water disposal. Finally, the crushing of middlings and coarse refuse plus the trends toward the finer sizes in sales has further increased the amount of fines produced per ton of mine-run coal. A joint consequence of these circumstances has been the augmented percentage and build-up of —200-mesh material in the finer coal fractions and tipple circulating water. This results in the following disadvantages to the operator: (1) Higher ash percentages in the fine coal caused in large measure by the greater ash content of the —200-mesh material, (2) more difficult or expensive recovery of the fine coal from plant slurries, (3) lower filtration, drainage, and drying rates caused chiefly by lower permeability and the greater surface exposed per unit mass of the finer particles, (4) greater difficulty of beneficiatiing the fine coal because of lowered equipment efficiency and the partial "coating-out" of the —200-mesh fraction on the coarser coal particles. Obviously some cheap but effective fine-size classifier must be developed to aid in solving the above problems. To eliminate the generally unwanted —200-mesh fraction from coal slurries, several methods are in use. Probably the most popular with coal operators is the settling cone wherein the larger particles settle downward to the apex of an inverted cone while the slimes and fines are carried upward by vertical fluid currents caused by the overflow of the bulk of the water at the outer rim. The main advantage of the unit is the low initial and operating cost since no moving parts are involved. Disadvantages are primarily low capacity per square foot or floor space and the loss of good coal in the overflow above the 100-mesh size due to localized, high-velocity gradients and nonuniform, fluid-flow patterns, especially at high loadings. Also, when high slime concentrations are present, desliming efficiency is severely injured.', ' Other methods employed are the drag and spiral classifiers and the hydroseparator.' These involve mechanical separations achieved by moving parts and naturally result in higher initial costs. As they are all based on the settling velocity of the particles at the mesh of separation under operating conditions, area requirements are again large per gallon of throughput if they are to classify at the 200-mesh point. Usually the selection is made to operate at a higher throughput and thus sacrifice some coarser coal in the slime overflow. Cyclone Possibilities in Desliming To overcome the disadvantages of present deslirning methods, it is necessary to use a device that exhibits a very stable and uniform flow pattern while greatly increasing the settling velocity of

Jan 1, 1952

G. O. SMITH,* Washington, D. C.—This program is itself a demonstration of the widespread interest in the subject of more efficient development and operation of oil fields. The statements of fact already presented and the views expressed are sufficient evidence that the engineering and economic principles fundamental to the idea of unit operation are much more generally accepted than they were even one year ago. And real progress has been made in putting these principles to the test. No one realizes better than I that the fight against waste, whether physical or economic, must be waged in the front-line trenches: Washington at best is only an observation post. Yet viewed from this distance the advance made in the last few months has been notable and to a high degree encouraging. Five years ago the Federal Oil Conservation Board, in its first contacts with the industry, raised the question, "How can new pools be explored without exploitation until new production is warranted by the country's needs?" Three and one-half years ago the Board, in its first report to the President, paid largest attention, among what it listed as fundamental conservation measures, to "cooperative methods in sane development of new fields." A year ago, in its third report, the Board again endorsed as the essential factor in conserving the nation's reserves of oil, "full cooperation in unit control, whether by voluntary agreement or by state enforcement." The function of this federal agency has been purely fact-finding and educational. The Oil Conservation Board early declared that the major part of its task was to help the industry "formulate the broader by-laws in the sense of conservation." And now, after these five years of eventful history in the oil industry, may I size up the situation as I see it from the Washington point of view? First of all, and fundamental to a successful issue, I have observed in the industry a gradual acceptance of technical opinion as the guide to executive action; and this has at last led to some agreement as to what procedure offers most promise. Analyzed in more detail, this highly desirable progress has consisted in "self-restraint actuated by enlightened self-interest." We see emerging the general desire of oil producers for efficiency in operation, which means low costs and high recovery. As shown by facts presented at this meeting, competition in drilling is in direct opposition to efficiency in oil production, in that such competition increases costs and lowers recovery. The only route to conservation of this natural resource and also of the capital and labor involved in making it available is through unit control, which substitutes cooperative, well-planned action for competitive, unregulated activity, forced by the individualism of a small minority of landowners or perhaps by a single operator. Or, in terms of actual experience, development of an oil field and its subsequent operation under unit control means the substitution of majority rule for minority rule, yet the minority enjoys to the full the resulting economic benefits. So far as it affects the public interest, unit operation of oil properties deserves unqualified approval. It means more oil from the same area over a longer period

Jan 1, 1930

By P. L. Allsman

A method of analyzing blasting action indicates that major cost savings are possible by revising practice and bringing the classical blasting formulas up to date; difficult problems such as taconite and throw-breakup can be attacked by engineering Denonation pressure — the peak pressure developed by the explosive reaction; "shatter blasting." Average detonation pressure — the average pressure over the time the reaction continues; "heave blasting." Radial presswe — the stress in a radially expanding sphere about the detonation. Corresponds to the diminishing peak pressure. Lateral stress - the stress in a circumferential direction, induced by outward movement of the rock away from the detonation. Limiting stress or Slim— the rock property at which failure will occur in the case at hand; may be tensile or compressive strength, or combinations. Recent developments in explosives technology, following the advent of nuclear explosions, have led to a rather complete understanding of their action. A survey was made for the purpose of uncovering any knowledge which might be applicable in the mining and quarrying industries, the principal users of explosives. Although no revolutionary techniques have become apparent, much basic data pertaining to blasting — fortunately not classified — have been developed by recent research, from which a good general understanding of the process can be derived. It is hoped this explanation of the scientific principles governing explosive action, together with a proposed analysis of blasting practice, will further the development of mining engineering. NATURE OF THE REACTION: The phenomenon of explosion, termed detonation, is in reality a rather complex chemical reaction, and as such is completely explainable by physical and chemical laws. Detonation is characterized by high temperature and pressure, extreme rapidity — often being complete within one microsecond — and most importantly by formation of a shock pulse which accompanies the reaction zone. As with any chemical reaction, there is a critical value of temperature and pressure below which detonation can not occur; this is termed the "critical point". Whereas an explosive substance will decompose slowly at ordinary temperature, with the formation of gases which readily dissipate; at elevated temperature the reaction is considerably speeded. If pressure is suddenly applied at some point in an explosive medium, adiabatic compression results, and the temperature is locally raised. When the temperature becomes high enough the decomposition will be so rapid that the gaseous products do not have time to dissipate, but contribute to a further build-up of pressure. This, in turn, furthers adiabatic temperature rise over a widening area, and with any heat generated by an exothermic reaction, causes a rapid increase in temperature. Both temperature and pressure are spontaneously built up in this manner until the critical point is reached, resulting in detonation. The values of pressure and temperature necessary to create detonation are established by the unusual nature of the process. If temperature alone is raised above the critical point, without sufficient increase in pressure, deflagration or "low-order detonation" occurs. In the true, or "high-order detonation", the shock pulse resulting from the violent decomposition becomes an integral part of the reaction, and itself aids in increasing the detonation

Jan 1, 1961

By R. Gibson, F. S. Turneaure

The tin deposit of Monserrat, Bolivia, consists of one major vein 1600 m in length. The ore is unusual because of the notable quantity of teallite, even though cassiterite is the principal tin mineral. The deposit, formed at shallow depth under a wide range in temperature, may be classed as xenothermal. Polished sections reveal a complex history of replacement, with low-temperature minerals deposited before high-temperature minerals. THE Monserrat mine1'2 of the Compania Minera Monserrat is located 11 km east of Callipampa, a station on the Antofagasta and Bolivia Railway some 55 km south of Oruro. The mine lies in a group of low hills on the east slope of a broad valley of northerly trend. The valley is separated from the altiplano to the west by a prominent ridge that rises several hundred meters above the surrounding country (fig. 1). Callipampa is situated a short distance west of the ridge, near the east margin of the altiplano. The elevation at Callipampa is 3700 m and at the mine 4100 m. Monserrat is near the northeast limit of what may be called the Poopo-Pazna district, an area of northwesterly trend about 25 km in length and 15 km in width. The district includes the tin-silver prospects of Poopo to the north, the zinc-tin deposits of Salvador to the southwest, and the tin-tourmaline veins of Avicaya to the south. Along the prominent ridge already mentioned, which forms the western limit of the district, there are several tin and antimony prospects including the Trinacria mine, which is famous for its fine specimens of cylindrite. Although the principal ore mineral is cassiterite, the tin deposit of Monserrat contains an unusual amount of teallite. Cassiterite occurs in part as a finely granular mixture with sulphides of iron and zinc and in part as needle tin of late hypogene age. The regional structure is dominated by broad folds typical of the Andes of Central Bolivia and similar to those of the Llallagua district. The area adjacent to the Monserrat mine is underlain by thin-bedded shale somewhat variable in color but predominantly dark grey to black. The beds strike northwest and dip at low angles southwest, forming part of the northeast limb of a major syncline. A resistant sandstone formation, which underlies the black shale, crops out along a ridge east of the mine and also accounts for the high ridge to the west that was previously mentioned. No igneous rock of any kind is found at Monserrat. The nearest intrusive bodies known are the irregular dikes and masses of quartz porphyry at Avicaya about 10 km to the south. Mine Workings The Monserrat property includes one main vein, which can be followed on the surface for 1600 m, and two or three branch veins of distinctly minor importance. The principal haulage tunnel, referred to as the San Carlos adit or level III, extends northeasterly for 1700 m, following the main vein throughout most of this distance. Above the adit, there are

Jan 1, 1951

By S. S. Cole, D. L. Armant

The smelting of titaniferous ores for the past hundred years has not been successful because of thickening of the slags in the furnaces. The interest in the utilization of these iron bearing materials has been due to the extensive deposits which have been found in various parts of the world.1 The literature contains references2 to the use of titaniferous ores in blast furnaces. None of these attempts has met with success due to the lack of sufficient information as to the cause of failures and to economic considerations. Evaluation of the published information, as to the probable cause of failure in various attempts, has indicated that the presence of lower valence titanium oxide and formation of titanium carbide contributed to the failure. It was apparent that severe reducing conditions which are present in a blast furnace had been one of the factors for such failures. The previous smelting operations had as a primary objective the recovery of iron from these ores. Under these conditions the highest recovery of iron could only occur with severe reducing conditions. Application of Melting Point Data The application of the melting point data, from the system CaO-MgO-Al2O3-TiO2 reported in another paper,3 to the smelting of titaniferous ores required an extrapolation of these melting points due to the presence of other impurities from the ores. Likewise, a study of the effect of varying FeO and reduced titanium content of the slag was necessary before the data could be fully applied. The melting point data under oxidizing conditions had clearly indicated that a fluid slag with a melting point in the temperature range of 1370°C could be obtained. The smelting of a titaniferous ore had as a primary objective the production of a high titanium slag which was fluid and substantially free of iron and low in trivalent titanium. Since production of a high titanium slag was of primary importance, rather than the recovery of iron from these ores, the smelting concept could be altered to a marked degree. The ferrous oxide could be maintained at higher levels than is the normal practice in iron ores slags without being detrimental to the slags, although a loss in iron recovery resulted. During the reduction of the molten mixtures of ferrous iron and titanium, carbon acts as the reductant for changing the valence of the titanium and iron. The reaction can go further than in the case of the reduction which occurs in iron-titanium solutions. The formation of metallic iron removes iron from the system, so that the titanium oxide can be reduced to trivalent and divalent titanium oxides and also form titanium carbide which removes the titanium from the slags as a soluble constituent. The problem, therefore, resolves itself into maintaining the proper amount of ferrous iron in the slag thereby avoiding the formation of appreciable amounts of trivalent titanium and no divalent titanium oxide or titanium carbide. The latter two have been found in the viscous

Jan 1, 1950

By F. S. Turneaure, R. Gibson

The tin deposit of Monserrat, Bolivia, consists of one major vein 1600 m in length. The ore is unusual because of the notable quantity of teallite, even though cassiterite is the principal tin mineral. The deposit, formed at shallow depth under a wide range in temperature, may be classed as xenothermal. Polished sections reveal a complex history of replacement, with low-temperature minerals deposited before high-temperature minerals. THE Monserrat mine1'2 of the Compania Minera Monserrat is located 11 km east of Callipampa, a station on the Antofagasta and Bolivia Railway some 55 km south of Oruro. The mine lies in a group of low hills on the east slope of a broad valley of northerly trend. The valley is separated from the altiplano to the west by a prominent ridge that rises several hundred meters above the surrounding country (fig. 1). Callipampa is situated a short distance west of the ridge, near the east margin of the altiplano. The elevation at Callipampa is 3700 m and at the mine 4100 m. Monserrat is near the northeast limit of what may be called the Poopo-Pazna district, an area of northwesterly trend about 25 km in length and 15 km in width. The district includes the tin-silver prospects of Poopo to the north, the zinc-tin deposits of Salvador to the southwest, and the tin-tourmaline veins of Avicaya to the south. Along the prominent ridge already mentioned, which forms the western limit of the district, there are several tin and antimony prospects including the Trinacria mine, which is famous for its fine specimens of cylindrite. Although the principal ore mineral is cassiterite, the tin deposit of Monserrat contains an unusual amount of teallite. Cassiterite occurs in part as a finely granular mixture with sulphides of iron and zinc and in part as needle tin of late hypogene age. The regional structure is dominated by broad folds typical of the Andes of Central Bolivia and similar to those of the Llallagua district. The area adjacent to the Monserrat mine is underlain by thin-bedded shale somewhat variable in color but predominantly dark grey to black. The beds strike northwest and dip at low angles southwest, forming part of the northeast limb of a major syncline. A resistant sandstone formation, which underlies the black shale, crops out along a ridge east of the mine and also accounts for the high ridge to the west that was previously mentioned. No igneous rock of any kind is found at Monserrat. The nearest intrusive bodies known are the irregular dikes and masses of quartz porphyry at Avicaya about 10 km to the south. Mine Workings The Monserrat property includes one main vein, which can be followed on the surface for 1600 m, and two or three branch veins of distinctly minor importance. The principal haulage tunnel, referred to as the San Carlos adit or level III, extends northeasterly for 1700 m, following the main vein throughout most of this distance. Above the adit, there are

Jan 1, 1951

By L. D. Jaffe

IN 1945-1946, the author measured temper brittle-ness in ingots of varying composition prepared by remelting SAE 3135 bar stock under vacuum. Since other investigators1-" have been referring to this hitherto unpublished work, belated publication seems worthwhile. Steel in 7 lb batches&apos; was induction melted with nickel additions in a magnesia crucible under a pressure of 10 microns of Hg or less and cast into a 2 in. diam chill mold. The compositions of ingots A, B, and C prepared in this way are shown in Table I; the vacuum melting lowered carbon, manganese, sulphur, and nitrogen content. Ingot B was homogenized 24 hr at 1095°C; ingots A and C were not homogenized. Blanks, M in. square, were cut from each ingot, austenitized 1 hr at 870°C, water quenched, tempered 1 hr at 595"C, and water quenched. Half the blanks from each ingot were given an embrittling treatment of 50 hr at 455°C and water quenched. A few specimens, embrittled and unembrittled, were finally heated 1 hr at 580°C and water quenched. Charpy specimens machined from the blanks were V-notched on the side closest to ingot mid-radius and paired as to radial position in the ingot, one of each pair having received the embrittling treatment and the other not. Specimens were broken in a standard impact machine at various temperatures, the same temperature, in general, being used for both bars of a pair. Fig. 1 shows the results. It is evident that temper brittleness, as measured by the increase in temperature of transition from brittle to tough failure introduced by the 455°C treatment, was decreased but not eliminated by reducing carbon to 0.006 pct, manganese to < 0.004 pct, and nitrogen to 0.001 pct. In another study," temper brittleness was apparently eliminated by -reducing carbon to 0.003 pct, with 0.80 pct Mn and 0.0015 pct N (ingot D)." • Analyses given for carbon below 0.02 pct were made by Jensen method; for nitrogen, by Kjehldahl method. Values quoted else-where&apos;-&apos;." for carbon content of ingot C and for nitrogen content of ingot D were obtained by conventional combustlon and by vacuum fusion methods, respectively, and are considered much less accurate. The 580°C final treatment removed previous em-brittlement (ingot B particularly), indicating that this was "reversible embrittlement."" Metallographic examination of ingot C with 1 pct nital etch revealed only ferrite; ingot A had carbide spheroids uniformly distributed in a ferrite matrix. With ethereal picric acid plus zephiran chloride etchant.&apos; no differences between embrittled and un- embrittled specimens of ingots A, B, or C were noted. Fracture of C and D below the transition temperature was predominantly transcrystalline, with traces of intercrystalline, in both embrittled and unembrittled conditions. The corresponding fracture of unembrittled B was mixed transcrystalline and intercrystalline; of embrittled B, predominantly intercrystalline with some transcrystalline areas. . . . Acknowledgment The efforts of J. C. Leschen, formerly of National Research CO~P., in preparing the ingots are acknowledged with thanks.

Jan 1, 1956

EDUCATION is a business; and, if the consumer is satisfied with the product, no one has any grounds for complaint. "Pretty well satisfied" about slims up the attitude of the mining industry toward the young graduates presently entering the field. Yet, with amazing regularity the controversy over the five-year curriculum recurs in educational circles. The educational mills grind out their products, some higher grade than others, depending on individual proclivity. The cycle continues almost without interruption. Arthur Taggart, however, takes his bowie knife to some of the sturdiest limbs of the tree of knowledge in his article on engineering education on page 770 of this issue. As all who know him might suspect, the pruning is not restrained. With one swing of the knife, such practically-labeled courses as Mine Design are stricken from the tree and transplanted to the field of Graduate Education. Emphasis on the basic sciences and the study of man through courses in history, economics, and literature are recommended to constitute an undergraduate curriculum. Such courses show the least depreciation and therefore give the student the most for his money. Partial justification for this dramatic departure from tradition is explained by the increasing number of companies giving training courses. Why make the student pay for what he can get paid for learning after graduation, the author argues. Other leading educators espouse this doctrine. Dean S. C. Hollister of Cornell, for the Committee on Adequacy and Standards of Engineering Education, recommended to ECPD in 1951 that emphasis be placed on the basic sciences. "Sifting back through the curriculum, it seems clear that instruction in the basic sciences, if taught in a manner such that knowledge of them makes available working tools, contributes the most sustaining part of the curriculum", reported the committee. Note that he would utilize practical problems from the fields of engineering in teaching the sciences. In our opinion the important thing to learn in college is how to think the engineering approach. Recitation on traditional solutions of classic problems sharpens the memory, not the ingenuity of the student. The engineering schools of the nation recognize the advantage of great emphasis on fundamentals, and the inclusion of cultural courses in the curriculum. However, industry presently requires graduates to have a substantial knowledge of the tools of the trade upon graduation. This means that the colleges are forced to pare to the bone to encompass the required studies in the time allotted. In addition, there is the limitation of student money. College educators also claim that entering freshmen are often improperly prepared in such subjects as writing and mathematics. These considerations make it difficult for the critic of the engineering graduate to make any headway with the educators. Actually with the mounting problems imposed by lower grade and more complex ores, a new breed of engineers is needed. The impetus for this new product must come from industry. If there was a concerted effort from industry for an improved type of engineer, the colleges would modify their aims to meet the demand for the kind of individual desired.

Jan 1, 1952

By A.A. Gustafson

In diamond or churn drilling for the prospecting of ore bodies, two products can be used for quantitative analyses; i.e., the core and the cuttings, or sludge. Some operators prefer an assay of core; others insist that the sludge be sent for analysis. Both must recover the sludge for check samples or for assays if there is no core recovery for any particular section of the diamond-drill hole. The Freeport Sulphur Co. found that numerous cumbersome sludge boxes or barrels were inconvenient and delayed drilling. The sludge settled slowly in the ordinary shallow, rectangular diamond-drill boxes and its total removal from the container was difficult, time-consuming and unsatisfactory; and sampling of deposits of which the metal contents are easily concentrated, as in ores containing cinnabar, gold and silver, is subject to error if the boxes are not thoroughly cleaned. Mr. Charles Ballard, dredge superintendent at Freeport&apos;s operation at Grande Ecaille, Louisiana, developed a portable cone while he was in charge of prospecting a mercury property near Tonopah, Nevada, in 1940. Ballard later improved the design while the settler was being used at a manganese property in New Mexico. The operation is based on the "free settling" of comminuted particles in a quiescent liquid. The ordinary sludge box discharges over a lip 8 to 12 in. wide and consequently the water has sufficient velocity so that the discharge water is never clear. The cone settler used by Free-port Sulphur Co. discharges over the entire circumference, which gives a discharge ltp of 9 ft. in the size of cone used at the cinnabar and manganese prospects. Consequently, the velocity of the discharge water is slow enough so that all fine material settles and the discharge water is clear. When the drilling of a sample interval is complete, the sludge can be drawn off by the time the driller is ready to cut another sample. Thus, only one cone is needed per drill. The cone has been used by Freeport successfully in churn-drilling a manganese deposit, in diamond-drilling a mercury deposit, and in diamond-drilling a deposit of ore containing tin, copper, gold and silver. The cone is greatly superior to the old methods of collecting sludge and is now used on all Freeport drilling campaigns in the western states, both for surface and underground drilling. Fig. I shows the important details of a cone designed for average operating conditions, 30 in. in diameter and 36 in. high. Abnormal conditions may require some revisions in design. Fig. 2 shows a working drawing. A hose or pipe larger than the one used to convey water from the pump to the swivel head of the string of rods is connected to the stuffing box at the collar of a diamond-drill hole, The usual T-connection is used. It is not necessary to use a stuffing box if the top of the cone can be placed lower than the collar of the hole. Then the return water and cuttings can be made to flow by gravity. This hose or pipe discharges the cuttings of the drill bit and the return

Jan 1, 1943

By J. H. Slater

There is nothing particularly new about the use of sinter in a blast-furnace burden. For many years flue dust has been sintered at the various blast-furnace plants to put it in a form that could be recharged to the furnace with some assurance that it would stay in the second time. This paper is not a compilation of scientific data but only an attempt to give our experience with sinter. There are some unexplained discrepancies in the tests given here, such as the flue dust produced and iron tonnages, but the results given are the results that were obtained. Sinter is not the cure-all for blast-furnace problems. A good coke is still the most important factor in good furnace operation. However, good sinter is a big help. As flue dust is produced under varying operating conditions and fluctuates greatly in carbon content, it is difficult to produce from flue dust alone a uniform sinter, which will give satisfactory blast-furnace results. Sinter made from flue dust alone will change in character with every carbon variation. Almost invariably flue dust contains carbon in excess of the amount necessary to make good sinter, and this excess must be burned out. This, of course, necessitates reduction of the speed of the sintering machine and thus a decrease in tonnage. An alternative is to dilute with a carbon-free material to use up the excess carbon. High-carbon material when sintered results in a very dense, undesirable product, since the excess heat melts the charge into a hard, nonporous mass, and tends to form refractory silicates. At the Republic Steel plant in Cleveland, there had existed since 1924 a continuous-type sintering machine having 224 sq. in. of grate-bar area, capable of producing 9 to 11 tons of sinter per hour from flue dust. Under a former management, it was the practice to sinter only the flue dust from the blast furnaces, no attempt being made to recover the gas-washer sludge. As the dust was limited, the plant operated only part time. In other words, there was a sintering plant representing considerable investment, with good facilities for screening and preparation of raw materials, and flue dust available for only part-time operation. The sinter produced was costly and not a very desirable product. The first step was to install equipment for the recovery of the gas-washer sludge. This was done in the form of two large thickeners and a filter, which have proved very satisfactory. It was then decided to screen out the minus 3/8-in. fines from the Mesabi ores, mix them with flue dust and coke breeze when necessary, and operate the plant at full capacity. The tonnage increased to 18 tons per hour and the product was much improved. Sintering costs fell and the product was a preferred blast-furnace material, which on an iron unit cost compared very favorably with good Mesabi ores. On the average, 40 per cent flue dust and 60 per cent fine ore was being used. The oversize from the ore screening also became a preferred product for the

Jan 1, 1941

By H. J. Hager, P. H. Haskell

For the past 20 years Alabama has probably led all other coal-producing districts in the proportion of coal prepared by wet washing. All rail mines, with one exception, and a high percentage of truck mines now wash all coal under 3 or 4 in. Most of this coal has been cleaned by jigs, though within the past few years one or more wet-table plants have been installed for cleaning some of the coal from practically all the major seams. These tables supplement either single or multicell jigs usually for one of the following purposes: (I) to wash the finer sizes of raw coal, (2) to rewash the finer sizes of washed coal after they leave the jigs, (3) to re-treat jig middling products. The main purpose of this paper is to present cleaning data and operating factors involved in the preparation by tables of the finer sizes of raw coal. The specific information and data pertain to the Mary Lee seam of coal as extracted at three of the mines of the Alabama By-Products Corporation, which produces coal for both byproduct coke-oven consumption and the general steam market. The Mary Lee seam has long been recognized by coal-preparation authorities as presenting a very difficult ash-reduction problem. Washing characteristics and physical structure of the seam vary widely. Its thickness ranges from 32 in. with no rock parting to 100 in. with from one to three partings. Over the entire field the bed contains bands of bone and rash. These are distributed throughout the seam cross section. A portion of this bone or bony coal occurs as definite bands I in. or more thick; however, some bone is interlocked as layers within the coal and consequently breaks into flakes when the coal is broken. This material is relatively high in ash and its specific gravity varies widely. It has been observed that in most cases in the washing of Mary Lee coal from 10 to 18 per cent of the raw coal is within +0.10 specific gravity of the attempted point of separation. Therefore with the average wet jig, only with extreme difficulty can a low-ash product be obtained without loss of much marketable coal in the refuse. Without entering into the controversial question as to whether modern washing equipment can efficiently handle simultaneously all sizes up to a given maximum, suffice it to state that the comparatively old-type jigs in service at the Alabama By-Products Corporation&apos;s mines could not clean the he sizes satisfactorily. Therefore, in 1931 that company installed six diagonal-deck tables at the Barney mine, which were installed as an addition to a complete plant cleaning all coal below minus 3 in. The primary washed product is shipped as steam coal and another portion may be prepared as washed nut for domestic trade. These tables are washing 9 tons per hour of raw coal sized to ¼ in. to o. The washed table coal is mixed with the 3 to ¼-in. jig product to constitute a 3 in. to o steam coal.

Jan 1, 1942

By John T. Galey

A combination of two factors in addition to the current Federal drilling and price regulations reduced by 120 from I942 the number of new wells completed in southwestern Pennsylvania during the year. No new prospective area was found comparable to the Armbrust gas pool, Westmoreland Co., which accounted for a seventh of the last year&apos;s drilling, and it is considered doubtful whether, even if such a field had been discovered, the drilling contractors, who have been seriously affected by the manpower shortage, could have drilled many more wells than they did during the year. New wells totaling 445 were completed for a combined footage of over 1,250,000 Dry holes numbered 131 and had a footage of nearly 400,000. Old wells drilled deeper total 55, of which 17 were found nonproductive. Shallow Development Oil Oil completions yielded an initial of only slightly over one hundred barrels from 21 wells for about the same per well average as that of a year ago. However, 65 bbl. of this total came from two wells in two new strikes. The footage drilled of nearly 33,000 gives an average of nearly I00 ft. less per well than a year ago. Two old wells drilled deeper had a combined initial of only 4 bbl. The largest oil well of the year, 40 bbl., was found in the Speechley (U. Dev.) sand at a depth of 3492 ft. in German township, Fayette County. The location is several miles east of the nearest oil production hut as the area that could be occupied by a pool is fairly well delimited by drilling, the extent of this production would not appear to be large. The other notable oil production was found in the Dunkard (Penna.) sand, Cumberland township, Greene County, at a depth of 1065 ft. Here the first well had an initial of g bbl. and the second 25 bbl. However, a third proved dry, and as this area is also delimited by drilling, the pool doubtless will be of small area. Gas Gas completions yielded an initial of slightly over one hundred million cubic feet per day and new wells, totaling 303, account for all but four million of this volume. Old wells drilled deeper, numbering 35, found gas. The per well average in new drilling of approximately 320 M cu. ft. is nearly 165 M cu. ft. less than a year ago and the total footage is slightly over 830,000, which is about a quarter of a million feet less than in 1942. Average rock pressure of new completions is approximately 365 lb. per sq. in. The entire volume of new production is from the Upper Devonian or higher stratigraphic horizons. In Allegheny County, the grestest activity was in Forward and Plum townships. Wells there account for more than one half the completions. One large well was drilled in Forward township and an open flow of 3.623 million cubic feet was found in an old Gantz (U. Dev.) sand area. The top of the sand was encountered

Jan 1, 1944

By W. J. Tait, A. E. Craig, E. P. McCurdy

The Circuitron herein described applies current from the classifier motor circuit and energy from the sound of grinding media to move an oscillating disc. The disc through a photoelectric cell controls the speed of the ore feeder and consequently the amount of ore delivered to the mill. THE Telluride Mines, Inc. at Telluride, Colo., is operating a flotation concentrator recovering lead, zinc, and copper with some gold and silver. The nominal capacity is 500 tons per 24 hr. The ore, which comes from several different mines, is characteristic of the district, changing from extremely tough rhodonite to soft gouge. The ore is crushed and washed, recrushed, and delivered to the mill bin at about % in. to serve as ball mill feed. Details of Grinding Circuit The material from the fine-ore bin is carried over a belt feeder driven by a Reeves vari-speed device. The belt discharges to a No. 86 Marcy ball mill, which operates in closed circuit, with jigs and a 6-ft Dorr classifier. A Massco density controller holds the density of the classifier overflow at a predetermined point. The overflow, which contains about 5 pct + 48 mesh, goes to the flotation section. Manual Operation Before the adoption of automatic control for the grinding circuit, this work was all done manually by the operator. The bin gate was set to deliver 40 to 50 lb of ore per lineal foot of belt, and the speed of the belt was adjusted by the operator to deliver at the rate of about 500 tons per 24 hr. As the ore changed in hardness, specific gravity, or in the amount of moisture, or, as the size of the feed changed due to bin segregation, crusher setting, or other causes, the operator made corresponding changes in the rate of feed by manipulation of the Reeves control. Sometimes these changes occurred so rapidly that they caused sharp, pronounced surges, which created very considerable unbalance from normal conditions in the grinding circuit and, more critically, in the flotation section. Since it is not possible for the operator to anticipate or follow such changes closely, a very considerable unbalance from normal conditions may occur, which will require a drastic change in the rate of feed. These surges upset the fineness of grind and interfere radically with the balance between tonnage and reagent control. All of this tends to reduce the grade of concentrates, increase the loss in the tailing, and thus decrease the recovery of values. The Circuitron Some years ago, The Mine and Smelter Supply Co., in conjunction with the operators at the Telluride Mill, undertook to build a device that would give close control of the rate of feed from the ore bin and to coordinate with this a related device that would regulate the amount of sand load on the classifier rakes. Emphasis was strongly laid on the fact that such a control must apply to the entire grinding circuit and must include the feed to the ball mill, the classifier sand return, and the density of the classifier overflow. If this could be done successfully, we felt that the circuit could be operated under best possible conditions, with compensations

Jan 1, 1951

By Edward H. Robie

SUBJECT to formal Board approval in September, AIME dues will continue indefinitely at the present scale of $20 for Members and Associate Members; and $12 for Junior Members for the first six years of such membership, thereafter $17. That is the wish of the Board of Directors and the will of the membership, as expressed in the recent referendum. Though the official returns, as reported by the tellers, are not available as this is written, the data are sufficiently complete to give the results to the nearest percent. The total vote was approximately 93 pct for continuance of the present scale. Metals Branch members voted about 94 pet for continuance; Mining Branch members slightly less than that; and Petroleum Branch members about 91 pct. As in the referendum of 1949, there was no particular difference of opinion on the question of dues as between the three Branches. The percentage of ballots returned was about 61 pct, only slightly more than in the 1949 referendum. But the affirmative vote was much greater, 93 pct against 78 pet in the earlier referendum. Looked at another way, a majority of those to whom ballots were sent, 56 pct, voted to continue the present scale of dues, whereas only 46 1/2 pct of those canvassed voted to raise the dues in the previous referendum. But approximately 600 , members expressed their opinion that the dues should go back to $15 for Members and Associate Members and to $10 and $15 for Junior Members. This is a substantial group of members who either do not think they are getting their money&apos;s worth from AIME membership, or believe that a more businesslike conduct of Institute affairs should provide current services at less cost, or believe that more service is being provided for their dues than they need. As to the first point, it is probable that some members are not getting their money&apos;s worth. They perhaps have only a fringe interest in. AIME activities, or they do not or cannot participate in such activities at the Local Section, divisional, or national level, so that all they get out of the Institute is the monthly journal, which they do not have time to read. If they did read it intelligently, and acted upon the professional information thus gained, it is difficult to understand how one or more ideas cannot be gained that would be worth $20 to them professionally. But the 600 members who want lower dues offer a challenge to the Directors and to the staff to do a better job; to make Institute membership so rewarding that every. member will be thoroughly satisfied. Twenty 1935 to 1939 dollars are worth only $11.60 now, according to the latest government data, but that makes it all the more necessary for us all to appraise the value of what we buy. To the 93 pct of Institute members who expressed a vote of confidence, so to speak, the staff, and we feel sure the Directors as well, are most appreciative and thankful.

Jan 1, 1952

By C. E. Birchenall, F. N. Rhines, L. A. Hughes

A. J. Shaler—I should like to congratulate the authors on the presentation of this paper, which we have been awaiting a long time. The view they have taken of the sintering process, namely that voids are comparable to particles of a second phase capable of dissolving to a slight extent in the metal phase, is an interesting but dangerous one. Pines has given a first-order solution of the problem and Has predicted a decrease in the concentration of pores near the surface. The authors of the present paper have experimentally demonstrated the second-order effects, that is, the diffusion of the space phase between voids of different sizes, and also the movement of atoms from highly curved parts of a void to less curved regions of the same void. I have recently calculated the rate of vapor-phase transport within odd-shaped pores in copper at 850 °C and found that probably at that temperature volume flow is the more rapid shape-changing mechanism. The authors of the present paper have found that at 1000°C, vapor-phase transport is dominant. There is no disagreement here, I think. The dominant phenomenon can easily be different&apos; in different temperature ranges and with different pore shapes: I used a cylinder going to a sphere, not a tricuspid going to a "triple cuspidor." But there is some danger in the other art of the explanation, that is, in the study of the diffusion of vacancies from a pore to its neighbors and to the outside. A vacancy is very different in dimension from an atom. When it moves, it carries a field of elastic strain. This has to be taken into account. A concentration gradient of vacancies can cause them to diffuse. But in the absence of such a gradient, they can be made to diffuse by a field of stress. What happens in some of our experiments, and it is most noticeable in the shrinkage of synthetic cylindrical pores, like the tubes we have been working with, may be broken down into the following steps. (1) The presence of a surface establishes a vacancy concentration gradient, as Dr. Kuczynski and the present authors have shown. (2) Vacancies then diffuse in order to decrease this gradient. (3) In so doing, they cause an elastic field to be built up in such a direction as to oppose the diffusion, so that it slows down until it reaches a steady-state rate. (4) The steady-state flow is due to the fact that the surface tension acting on curved surfaces applies a stress to the system which itself mechanically causes vacancy diffusion. What I have said is that the diffusive types of flow of the authors and of Dr. Kuczynski are transient phenomena which soon stop if the pores are large. We are then left with a mechanical problem only, and our researches have not yet disclosed just what the mechanism of flow is; it may be a slip process, or it may take place by a vacancy diffusion of the type Nabarro has de-

Jan 1, 1951

By H. J. Hager, P. H. Haskell

For the past 20 years Alabama has probably led all other coal-producing districts in the proportion of coal prepared by wet washing. All rail mines, with one exception, and a high percentage of truck mines now wash all coal under 3 or 4 in. Most of this coal has been cleaned by jigs, though within the past few years one or more wet-table plants have been installed for cleaning some of the coal from practically all the major seams. These tables supplement either single or multicell jigs usually for one of the following purposes: (I) to wash the finer sizes of raw coal, (2) to rewash the finer sizes of washed coal after they leave the jigs, (3) to re-treat jig middling products. The main purpose of this paper is to present cleaning data and operating factors involved in the preparation by tables of the finer sizes of raw coal. The specific information and data pertain to the Mary Lee seam of coal as extracted at three of the mines of the Alabama By-Products Corporation, which produces coal for both byproduct coke-oven consumption and the general steam market. The Mary Lee seam has long been recognized by coal-preparation authorities as presenting a very difficult ash-reduction problem. Washing characteristics and physical structure of the seam vary widely. Its thickness ranges from 32 in. with no rock parting to 100 in. with from one to three partings. Over the entire field the bed contains bands of bone and rash. These are distributed throughout the seam cross section. A portion of this bone or bony coal occurs as definite bands I in. or more thick; however, some bone is interlocked as layers within the coal and consequently breaks into flakes when the coal is broken. This material is relatively high in ash and its specific gravity varies widely. It has been observed that in most cases in the washing of Mary Lee coal from 10 to 18 per cent of the raw coal is within +0.10 specific gravity of the attempted point of separation. Therefore with the average wet jig, only with extreme difficulty can a low-ash product be obtained without loss of much marketable coal in the refuse. Without entering into the controversial question as to whether modern washing equipment can efficiently handle simultaneously all sizes up to a given maximum, suffice it to state that the comparatively old-type jigs in service at the Alabama By-Products Corporation&apos;s mines could not clean the he sizes satisfactorily. Therefore, in 1931 that company installed six diagonal-deck tables at the Barney mine, which were installed as an addition to a complete plant cleaning all coal below minus 3 in. The primary washed product is shipped as steam coal and another portion may be prepared as washed nut for domestic trade. These tables are washing 9 tons per hour of raw coal sized to ¼ in. to o. The washed table coal is mixed with the 3 to ¼-in. jig product to constitute a 3 in. to o steam coal.

Jan 1, 1942