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By Clarence Zener

The heat-treatment of steels will not pass from the stage of an art into that of a science until the mechanism of the phase transformations associated therewith is thoroughly understood. Such an understanding will involve two distinct types of inquiry: (I) the direction of the transformation, (2) the speed of the transformation. The first type of inquiry, which involves only the equilibrium relations between the various phases in steel, is the subject of the present and of a succeeding article; the second type, which involves the kinetics of transformations, will be the subject of a later article. In order that a theory of equilibrium relations may give useful information, certain experimental data must be put into the theory. It is customary for these data to be of a thermal nature. In the system discussed herein—steel—such thermal data are very scarce. In medium-alloy steels another type of data might possibly be used. In these steels the concentration of solute is so low that it may be possible to use the approximation of dilute solutions, an approximation that leads to linear relationships. The unknown functions, or constants, may then be determined in some cases from the binary iron-alloy system and in other cases from the ternary iron-carbon-alloy system. From the functions, or constants, so determined may then be computed the equilibrium relations in more complex medium-alloy steels. In the present article this second approach is used, the necessary experimental data 'being obtained from the simple binary and ternary equilibrium diagrams. Fundamental Equations In this article are derived the equilibrium relations between the gamma, alpha and cementite phases. In these three phases all important alloying elements except carbon form substitutional solutions, each alloying atom occupying a potential site for an iron atom. Carbon, on the other hand, forms an interstitial solution. In the face-centered cubic gamma phase, the interstitial positions are at the center of the unit cube and at the centers of the cube edges. There are therefore four interstitial positions per unit cell, or one per iron atom. In the body-centered cubic alpha phase, the interstitial positions are probably at the face centers, and also at the centers of the edges, which are crystallographically equivalent to the face centers. There are therefore six interstitial positions per unit cube, or three per iron atom. In cementite, (Fe, Alloy)3C, the carbon atoms may. be regarded as filling all the lattice positions of a given type. The introduction of the fundamental the of equilibrium has presented the author with a quandary. These equations are formally so similar to other equations that have been derived1,2 for related systems that they may appear to some readers

Jan 1, 1947

By William H. Hamilton

Tunnel-driving capabilities in terms of feet per hour have advanced several hundred percent in the last century. Indications are that this capacity will double each decade for the next three decades. These future improvements will be made possible by a greatly increased need for underground excavation and the awareness of industry of its own latent capabilities. The factors of most interest to any potential user of a tunneling machine are those making possible an increase in production. The first and most obvious factor, and probably the one that has created the excitement and interest of most people, is the increase in boring rate through faster rock fracture and improved support systems. Boring rate can be increased by more horsepower in the form of increased thrust, torque, and cutterhead speed, and by better cutters and cutting methods, better material-handling systems, and better ground-control systems. All of these items are extremely important and are the basis of much discussion and effort by everyone concerned with the excavation of tunnels by machine. The second factor, just as obvious and equally important as an increase in boring rates, is "increase in production through better utilization of the tunneling machine and related equipment," which means a reduction in downtime through greater reliability of equipment. Mechanical rock attack is presently capable of cutting hard rock at 5 fph and soft rock at more than 20 fph. In current practice, however, this advance rate is maintained for less than 50% of the shift time. I predict that tunnel-driving contractors and their supporting manufacturers of tunnel-boring equipment will provide machinery that will maintain these advance rates for more than 90% of the shift time. This increase in effective production will be due to the increase in reliability of the tunneling machine and tunneling system components. Such a percentage increase in utilization of the tunneling equipment is just

Jan 1, 1970

By Zavis M. Zavodni, Carl D. Broadbent

During the past 15 years, rock mass structural properties have become acknowledged as the focal point for rock slope design. Re- search into the methods for measuring and incorporating structure into the design effort has taken many directions and has produced a variety of useful techniques. Among the findings at Kennecott has been the realization that structure-controlled instability can be categorized as regressive or progressive depending on the relative geometries of slope and structure and on the strength characteristics of the structures. It was further identified that a failure condition can initiate as regressive and in time become progressive should external stimuli increase or structure characteristics be altered due to extensive displacements. The transition point has been shown to be a useful indicator of the expectable collapse time. This paper defines three fundamental failure types as dictated by structure and offers discussion on the practical value of such a classification to slope design and failure control. Several prominent examples of failures typifying the categories of structure- controlled instability are discussed. An application of f allure prediction based on the regressive-to-progressive transition point is described.

Jan 1, 1983

By Raymond Mancha

TO determine the optimum inside dimension of an airshaft, it is necessary to strike the proper balance between the cost of power for air friction and turbulence losses within the airshaft, on the one hand, and the cost of sinking the airshaft, on the other. An airshaft larger than necessary is extravagant because the increased amortization rate exceeds the reduction in power cost. Conversely, an airshaft smaller than is needed wastes more power than can be justified by the corresponding reduction in amortization rate. The airshaft power is a function of the air volume Q cu. ft. per min. and the air- pressure drop p in. water gauge across the airshaft The airshaft pressure drop is calculated by the formula: [ ] wherein R is the air-friction coefficient for the portion of the airshaft 1 ft. deep, 0 ft. in perimeter and A sq. ft. in area, when passing an air volume of Q cu. ft. per minute. The airpower w kw. dissipated within the airshaft is computed by the formula: [ ] With a combined over-all unit efficiency of the fan and drive of E per cent, the air- shaft power requirements become [ ] There are 8760 hours per year, so with energy at I cent per kilowatt-hour the annual cost per kilowatt for continuous operation is $87.60; therefore, with energy at C$ per kw-hr. the annual airshaft power cost MI dollars per year becomes: [ ] which, expressed by the component parts, becomes : [] The amortization rate a dollars per year per dollar invested includes both depreciation and interest on the investment. These in turn involve the expected useful life of the airshaft and available monetary interest rate. For the purpose of this discussion the cost of airshaft sinking is assumed to comprise a fixed move-in charge Y dollars, which covers the cost of transporting, erecting and removing the shaft-sinking equipment and is independent of the air- shaft size. Also, it is assumed that the cost of shaft lining, if any, is fixed at $X per cu. yd. of lining material and that excava-

Jan 1, 1946

By Maureen G. Johnson

At least three million ounces of placer gold has been mined from hundreds of deposits in the Southwest since the 1600's. Nevada, the most productive state, produced about 1,700,000 oz. Most of the placers are in accumulations of unsorted, coarse, angular gravel in gullies, dry washes, and on hill- sides. The gold is mixed with debris from the adjacent bedrock and is typically concentrated at or near bedrock or on clay layers or caliche-cemented gravels above bedrock. The most extensive and productive placers are thick accumulations of debris in well-defined stream or gully channels, in alluvial fans at the edges of the range fronts, or as thick residuum on hillsides. With very few exceptions, the gold has been transported only a very short distance from the source. Numerous small high-grade gold-quartz veinlets, rather than single large lode deposits, are by far the most common source of placer gold. Most placer mining consisted of small-scale one-or two-man operations using rockers, pans, and sluices when water was available and dry-washing machines when it was not. Where gravels were deep the miner would dig shafts to bedrock and drift along the paystreak. Large-scale operations in the major districts utilized floating or stationary dredges or, more rarely, hydraulicking to mine extensive deposits. An adequate supply of water for efficient concentration of gold has always been a major problem in the Southwest. To overcome this problem, difficult and sometimes ingenious techniques were used-pipelines to carry water for miles to the placer area were constructed, or gravel was hauled to central washing plants where water was available; numerous designs of dry-washing machines were invented. In some districts, attempts to mine the gravels by large- scale operations failed because the gravels were too wet for dry concentration methods, and no water was available for wet methods; or, when water was available, the gravels were too porous to retain sufficient water in the dredging pond.

Jan 3, 1974

E. T. Turkdogan (British Iron & Steel Research Assoc., London, England)—I would like to suggest a different method for the treatment of the authors' data on the Si-0 equilibrium in liquid iron. Because the activity coefficient of silicon dissolved in iron is influenced by the concentration of silicon as well as that of oxygen, the equilibrium constant K2 of reaction 2 becomes a function of metal composition, and, as shown in Fig. 1, a plot of log KP1 at 1600°C against the weight percent of silicon gives a smooth curve. A similar curve is obtained when log Kzl is plotted against the oxygen content of iron. Before

Jan 1, 1955

By Joseph R. Crump, Francois Ville, Frank M. Tiller

Several simplifications have been accepted in filtration theory during the past fifty years in an effort to avoid apparently intractable mathematical problems. In effect, these simplifications ignored the variation in average cake porosity and average filtration resistance with variation in pressure drop or with time. This study examines the traditional approach and shows that, in many cases, it cannot give correct design guidance. Using a set of dimensionless parameters because of the large number of significant variables, a revised theory is developed which does not require that porosity and resistance values be set constant. Graphs developed in the traditional approach (e.g., dt/dv versus v or t/v versus V) have substantial curvature near the ordinate; but this curvature is frequently ignored in theory and is easily missed in experiments. The revised theory allows more accurate design calculations, especially for compressible cakes.

Jan 1, 1980

By Harlan W. Nelson, James B. Purdy

A study of data obtained during laboratory tests to determine the suitability of bituminous coals for use in residential underfeed stokers of the clinkering type has led to the following general conclusions: 1. Performance data obtained during the test furnish useful and practical information upon which to judge the performance of stoker coals. 2. Conclusions derived from the results of evaluation tests are in general agreement with use experience of the coals in the field. The method is of practical use in this regard. 3. Few direct relationships were found between data furnished by the evaluation test and determinations by standard laboratory tests. This is particularly true with regard to coke formation and clinker formation, two of the most important factors in determining the degree of satisfaction to be obtained in applications of the small underfeed stoker. It has not been the intention in this paper to prescribe this evaluation test as the best method for the evaluation of coals for use in residential stokers. That there are other test methods having similar objectives serves to point out the necessity for full-scale laboratory tests. At the present time, a standard test procedure for the evaluation of coals for use in residential stokers is not available. For the past several years a joint committee, appointed from the membership of the Residential Stoker Committee of Bitulninous Coal Research, Inc., and from the Engineer- ing and Research Committee of The Stoker Manufacturers Association, has been working to develop a standard procedure for testing and evaluating bituminous stoker coals. This procedure has been completed to the extent that it has been drawn up in tentative form. It is now planned to install the equipment in several laboratories and to run check tests, using portions of identical lots of coal, to determine the degree of reproducibility offered by the method. Completion of this standard test procedure should greatly facilitate the evaluation of coals for use in residential stokers. Acknowledgments The cooperation of coal operators sponsoring the coal evaluation tests, in granting permission to use data and results from their tests, is gratefully acknowledged. Acknowledgment is also made to Carroll F. Hardy, Chief Engineer, Appalachian Coals, Inc., for his interest and timely suggestions on the preparation of this paper. Special thanks are due to Ralph Sherman, Assistant Director, Battelle Memorial Institute, who was largely responsible for development of the evaluation procedure, for helpful suggestions and advice. References I. H. A. Sherman: The Evaluation of Coal for Use in Domestic Stokers. Univ. of Ill., Eng. Experiment Station, Circular No. 39 (1939). 2. . Q. Shotts: The Relation of Free- swelling Indexes to Other Characteristics of Some Alabama Domestic Stoker Coals. TP 2314, Coal Tech., (Feb. 1948); Trans. AIME (1948) 177, 502. 3. . J. Helfinstine and C. C. Boley: Correlation of Domestic Stoker Combustion with Laboratory Tests and Types of Fuels 11. Combustion Tests and Preparation Studies of Representative Illinois Coals. Ill. State Geol. Survey, Re1. of Zntlestigations No. 120 (19467 ) p. 35. DISCUSSION E. R. Kaiser*—AS a former fuel engineer assigned to residential stokers at Battelle, the writer has read with considerable interest the paper by Messrs. I'urdy and Nelson. It is gratifying to note that earlier test techniques have been developed and that criteria found important in the years 1935 to 1938 are still important in judging the performance of stoker coals for residential heating. The authors have plotted a number of the primary coal data singly against performance data of the stoker in an effort to establish relationships. Unfortunately, wide scattering of points usually resulted, which did not illustrate what experience generally indicates to be the case. For example, Fig 3 indicates almost no trend in the percentage of released ash converted to clinker with change in ash-softening temperature. Other factors, such as fuel bed temperatures and zones of heat release must have influenced the results. In the present state of our knowledge, we cannot explain all of the reasons why suitable stoker coals perform satisfactorily, nor why one satisfactory coal may be better than another. It would therefore

Jan 1, 1950

By A. M. Gaudin

A GREAT number of hypotheses has been advanced to explain the complex phenomena that are encountered in flotation. In the days of bulk-oil flotation, when a large quantity of oil was employed, it was observed that the oil-mineral aggregate floated above the ore-pulp because of its lesser density. With the advent of froth-flotation this phase of the theory had to be modified to take into account the presence of air bubbles. The reason why the mineral-bearing froth of today rises to the top of flotation cells is, of course, because the air-mineral aggregate, with whatever reagent may be associated with it, is lighter than the pulp. In an attempt to explain froth-flotation the fact has been emphasized that certain minerals are wetted and therefore remain in water; while others are not wetted and therefore become associated in preference with air. Such an "explanation" does not explain facts, it merely states them; for instance, it does not explain why minerals are wetted under some circumstances and not wetted under others. Going somewhat further, the wetting or nonwetting of minerals by water has been ascribed' to differences in the angles of contact between oil droplets and mineral surface at the surface of the minerals. This explanation, however, was abandoned when soluble flotation reagents came to replace the so-called "insoluble" reagents. Another theory, sponsored about 10 years ago by Ralston,2 involves the effects of electric charges at the surface of the minerals particles and air bubbles. This theory, unfortunately, was not developed as fully as it seems desirable.

Jan 1, 1927

By J. H. Carpenter

LMENITE, rutile, zircon, and other heavy minerals are found in small amounts in the sand that covers most of Florida. Small enriched deposits occur on the beaches and coastal dunes. One beach deposit east of Jacksonville was worked during World War I and another near Vero Beach during World War 11. Both ceased operation when the resumption of imports reduced prices. Early in World War II reduction of rutile imports made a domestic source of the mineral necessary to maintain the expansion of armaments. Explorations made near Jacksonville discovered a relatively large body of low-grade ore about 10 miles inland from the beaches. A plant installed by Rutile Mining Co. of Florida using both flotation and tables encountered difficulties because of low plant capacity and organic slimes in the ore. In 1944 Humphreys Gold Corp. solved the problem of concentrating the low-grade ore by using the newly developed Humphreys spiral concentrator and by mining with a hydraulic dredge. This combination of low cost mining and concentration permitted the plant to continue operation in competition with foreign ores when imports were resumed following the war. The success of the Jacksonville operation renewed interest in Florida as a source of titanium. An extensive survey of the state was made by E. I. du Pont de Nemours and Co., Inc., under the direction of J. L. Gillson in 1945 to 1946. Sampling was carried on by drive pipe drilling to depths of 20 to 25 ft. Encouraging values were discovered near the location of the present Trail Ridge mine, 50 miles southwest of Jacksonville, near the town of Starke. The United States Bureau of Mines was interested in exploring for heavy minerals in Florida, and on the recommendation of both the state geologist of Florida and of Du Pont geologists, drilled deeper where marginal indications were found on Trail Ridge. In 1947 it was discovered that sufficiently higher values occurred at depth to make the deposit commercial. The orebody at Trail Ridge is a long sand deposit averaging about 1 mile in width and 35 ft in depth. It is approximately 1 1/2 miles wide at the north end of the portion now being worked, tapering to a blunt point 3 miles to the south. The ore reaches a maximum depth of 65 ft near the center and feathers out into barren sand at the edges. The ground surface is relatively flat and is covered with vegetation including pine, scrub oak, palmetto, and occasionally in swampy areas, cypress. The remain of a dense woody growth underlies and marks the bottom of the Ore 'One. It is difficult to drill through this layer, but in some cases where it is less than 5 ft thick it has been possible to drive through the wood. Sand lying below is coarser than that in the ore zone and contains no values. The wood layer consists of unconsolidated carbonized fragments of branches and other forest trash associated with a considerable amount of ar-

Jan 1, 1954

By Owen S. Roberts

During the design phases of a conveyor system, the question arises regarding the type of drive required for each application. The paper shows that the correct drive for any particular application is the one that satisfies the criteria established for that application in the most economical manner consistent with such considerations as site location, availability of spare parts, area environmental conditions, power supplies available, etc. Following discussions of various mechanical arrangements for the drives, motor types and their characteristics are considered along with the control means of modifying these characteristics to satisfy individual requirements. These characteristics are all illustrated graphically. After selecting the motor and its starter, control accessories and interlocking arrangements as applied to material handling systems are discussed.

Jan 1, 1973

By N. Jaffe, J. E. Dorn

The effect of stress on the creep rate of high-purity poly crystalline aluminum in the intermediate temperature range was investigated by conducting creep tests at a series of different stresses and by decreasing the stress abruptly during the course of primary and secondary creep. Three series of decreases in stress tests were conducted: the stress was decreased in primary and secondary regions of creep at 347.5°K and the stress was decrensed in primary creep at 327°K. The results demonstrate the effects of structure and temperature on the stress dependence of the creep rate. The trends in the creep rate after a drop in stress are complicated and show that recovery takes place. The experimental results were compared with Friedel's theory for cross-slip, the experimentally determined energy for a constriction being in good agreement with the theoretical estimate proposed by Stroh. PREVIOUS investigations have shown that the ap-- parent activation energy for creep of polycrystalline aluminum, over the range from about 250° to 380°K, is approximately 27,400 cal per mole,' a value that is to a first approximation insensitive to the applied stress, the creep rate, and the creep strain. Several facts strongly support the thesis that creep, in this temperature range, is controlled by the rate-determining mechanism of cross-slip: 1) Recovery of the cold-worked state can take place by a number of mechanisms prominent among which is climb and cross-slip. Astrom2 reported rapid recovery of cold-worked polycrystalline aluminum at 340° to 370°K giving an activation energy of 28,000 cal per mole and additional recovery at the higher temperatures of 455° to 478°K giving an activation energy of 36,000 cal per mole. Whereas the latter process was clearly identified with the climb mechanism, the former process. which has an activation energy in good agreement for that of creep in the same temperature range, is most likely due to recovery by thermally-activated cross-slip. 2) Single crystals of aluminum exhibit sharp slip-

Jan 1, 1962

By Auguste J. Rossi

As is well known, practically all the phosphorus of the ironores smelted in the blast-furnace passes into the pig-metal, increasing its fluidity, but diminishing its strength to such an extent that, if phosphorus exceeds a certain amount, the metal is only fit for fine castings, not required to resist strains, and it cannot be used for the ordinary purposes of good foundry-iron. No phosphorus being eliminated in the acid Bessemer process, the quantity admissible in pig-metal destined for that process must be very small, and hence the ores smelted must be strictly Bessemer ores, of which, in many districts, the supply is taxed to the utmost. On the other hand, very highly phosphoric pig, a metal of little or no value by itself, when refined by the Thomas-Gilchrist process in the basic converter, furnishes an excellent steel; but such pig-iron must be made from ores particularly high in phosphorus; and these ores are not everywhere abundant, and are much sought after for this special purpose. But there are many iron-ores containing a moderate, though still important, amount of phosphorus, which yield a pig-metal

Jan 1, 1897

By Edwin R. Phelps

In the year 1900 about 212 million tons of bituminous coal were produced in the United States. Almost none of this was produced by surface mines (Fig. 1), because there was no large equipment suitable for stripping overburden at that time. By 1915 the surface mines had built up their capability largely through use of the steam shovel to the point that they produced 2.8 million tons of the nation's total production of 443 million tons. Over the intervening years between 1915 and 1970 surface mining provided a continually increasing percentage of the nation's bituminous coal, stimulated by the growing availability of giant surface mining machines, some of which are presently capable of removing 220 cu yd of earth with one bite of their mighty buckets. In 1970 the nation's surface mines produced 244 million tons of bituminous coal, over 40% of the total national production. On the other hand, the nation's underground mines hit a peak of 560 million tons in 1920, dropped to a low of 273 million tons in 1961, and are presently on a very slow rise. Environmental pressures generally oppose surface mining, and have forced coal companies to consider more underground mining in future planning. Also, since strippable reserves are decreasing in the eastern states, an increase in underground mining is expected over the coming years. However, in the western states there are enormous reserves to be extracted by surface methods. Why has surface mining increased, while underground mining has decreased? It's largely a matter of economics. The cost per ton is greatly reduced by surface mining because of the relative ease of obtaining the coal. Also, there is a much greater percentage of coal recovery-nearly 85% in surface mines within the minable area, as compared to an average of about 50% in underground mines. Much surface mining recovers coal that could not be extracted by underground methods because of coal bed thinness, or multiple beds close together, split seams, the nature of the roof, and other geologic conditions. What are some of the limitations of surface mining? First of all, with presently available machinery it is difficult to surface mine at depths

Jan 1, 1973

By D. K. Crampton, H. l. Burghoff, J. T. Stacy

The study of copper alloys of the age-hardening type has received considerable attention, and, among the alloys which the authors have, considered, those containing small amounts of nickel and phosphorus as essential constituents seem especially interesting. Such alloys have attractively great capacity for being hardened and also have relatively high electrical and thermal conductivity in the hardened state. In addition, it has been found that the increase in hardness and strength imparted by aging can be greatly augmented by cold-working, so that excellent combinations of conductivity and strength are obtainnble. In view of the possible importance of the alloys, it was considered desirable to investigate the solubility relationships in the copper-rich portion of the copper-nickel-phosphorus system. This paper present's some of the results obtained in this investigation to date and also gives propertics of typical alloys that appear to be of commercial interest,. Review of literature Copper-nickel System,.—The binary coppcr-nickcl system has long been known to consist of a series of simple solicl solutions, copper and nickel being miscible at all tempcrat'ures and concentrations. This was establisl~cd by Guertler arid Tammannl in 1907 and confirmed by Tafel2 in 1908. Copper-phosphorus System.—The copper phosporus system has been studied by numerous investigators, so that the copper-rich portion is now definitely known. As most recently determined by Lindlief," his system has a cutectic at 8.38 per ccnt P, cuteetic temperature being 714° C. Components of the cutectic are Cu3P and alpha solid solution. Solid solubility of phosphorus in copper, according to Mertz arid Mathewson decreases from a maximum of about 1.75 per cent at 714' C. to about 0.6 per cent at 300' C.

Jan 1, 1940

By D. K. Crampton, H. l. Burghoff, J. T. Stacy

The study of copper alloys of the age-hardening type has received considerable attention, and, among the alloys which the authors have, considered, those containing small amounts of nickel and phosphorus as essential constituents seem especially interesting. Such alloys have attractively great capacity for being hardened and also have relatively high electrical and thermal conductivity in the hardened state. In addition, it has been found that the increase in hardness and strength imparted by aging can be greatly augmented by cold-working, so that excellent combinations of conductivity and strength are obtainnble. In view of the possible importance of the alloys, it was considered desirable to investigate the solubility relationships in the copper-rich portion of the copper-nickel-phosphorus system. This paper present's some of the results obtained in this investigation to date and also gives propertics of typical alloys that appear to be of commercial interest,. Review of literature Copper-nickel System,.—The binary coppcr-nickcl system has long been known to consist of a series of simple solicl solutions, copper and nickel being miscible at all tempcrat'ures and concentrations. This was establisl~cd by Guertler arid Tammannl in 1907 and confirmed by Tafel2 in 1908. Copper-phosphorus System.—The copper phosporus system has been studied by numerous investigators, so that the copper-rich portion is now definitely known. As most recently determined by Lindlief," his system has a cutectic at 8.38 per ccnt P, cuteetic temperature being 714° C. Components of the cutectic are Cu3P and alpha solid solution. Solid solubility of phosphorus in copper, according to Mertz arid Mathewson decreases from a maximum of about 1.75 per cent at 714' C. to about 0.6 per cent at 300' C.

Jan 1, 1940

By C. W. Goodale

THE motives behind the organization of the Bureau of Safety were twofold. First, there was the policy of the company toward its employees, which was one of fairness and consideration for their welfare and which entails the prevention of injuries; second, in order to prevent injuries, a study of their causes was essential, and this study could be made only by a thorough systematic classification of accidents, which required a special department to form and carry out the plans. The Bureau of Safety was organized to begin operation Jan. 1, 1914, but practically all of the first year was spent in collecting data, making plans, etc. This department is semi-independent of others and reports directly to the president and general manager. It consists of the president of the company and the managers of the following departments: Mines, Washoe Reduction Works, Great Falls Reduction Works, Coal Department, Lumber Department, B. A. & P. railway, and the chairman of the Bureau of Safety. Working under this group is the General Safety Committee consisting of the manager of mines, as chairman, the general superintendent, two assistant general superintendents, and the heads of the following departments: mechanical, geological, purchasing, sampling works, and framing plant, as well as the safety engineer. Operating under the General Safety Committee are the group safety committees, consisting of the Anaconda group of mines, Boston & Montana group of mines, zinc group of mines,.Washoe reduction works at Anaconda, B. & M. reduction works at Great Falls, coal department, lumber department and the B. A. & P. railway. Each of these group safety committees was in charge of the superintendent or manager of the group represented and the committee consisted of the foremen of the mines or works that the committee served, as well as the safety engineer.

Jan 2, 1922

By G. P. Chatterjee

HARDNESS is one of the most common properties of metals and alloys but is one of the most difficult to define or express in terms of the fundamental units. In the Brinell or Vicker's hardness test, the hardness number N is defined as N = L/A where L is the load in kilograms and A is the area in mm'. This does not necessarily imply that a pressure of N kilograms per mm2 actually acts on the specimen.' It is essential that for a true comparison of hardness values the angle of indentation be kept constant,' a which is not done in the Brinell test. For this and other reasons, the Brinell hardness value is not based on a satisfactory physical concept.' The concept of hardness based on the work done per unit volume of indentation3-6 offers a simple and scientific definition. An attempt has been made here to develop this concept (at constant angle) and to compare the values so obtained with the Bhn for different metals and alloys. In Fig. 1 a vertical load L is applied to a compara- tively hard ball of diameter D creating an indentation of diameter 6 and depth h on the specimen S. The work done, W, is then given by W = So Ldh. [I] Since L is not independent of h, Honda and Takahasip and Matsumura" assumed different functions leading to equations which do not apply equallye to a wide variety of metals and alloys. A better relation is obtained from Meyer's Law L = ad" [2] where A and n are constants depending on the nature of the material but independent of the ball diameter D. From Fig. 1 it is obvious that h = R(l — cosq) and d = 2R sin 8.

Jan 1, 1957

The American Institute of Mining Engineers, the second of the four great national engineering societies established in the United States, was organized in 15,71. Its membership is composed of men engaged in the different branches of mining, metallurgy, geology, and chemistry. It includes both those who have worked their way up from the ranks and graduates from mining schools and universities, and at all times has counted in its membership leading technical educators. By a letter ballot of the members of the Institute, it was voted, at the annual business meeting of the Institute on February 18, 1919, to change the name of the American Institute of Mining Engineers to the American Institute of Mining and Metallurgical Engineers. The vote for the amendment to the Constitution making this change was a heavy one for such matters. It is the consensus of opinion that this marks an epoch in the history and usefulness of this Institute, and it is expected that the change will presage a large and healthy growth in metallurgical fields. Many metallurgists who knew nothing about min¬ing were loath to join a society of mining engineers, despite the fact that over 35 per cent. of the papers published in our Transactions during the past five years have dealt with metallurgical subjects.

Jan 1, 1923

By H. Reginald Hardy, Y. S. Kim

Recent studies of inelastic behavior and acoustic emission in geologic materials further validate the concept that such materials may be considered to consist of an elastic media containing an array of pores and cracks. In the present paper this concept is investigated further by experiments in which the mechanical behavior of a number of geologic materials were studied in detail at low stress levels using ultrasonic pulse methods. Experiments were carried out on cylindrical test specimens of Indiana Limestone, Tennessee Sand- stone and Barre Granite under relatively low confining pressures (0-5000 psi). Analysis of data from these experiments indicated that there existed a low-level critical stress at which there appeared to be a transition in the mechanical behavior of the material. Under triaxial stress, results indicated that the differential critical stress was relatively independent of confining pressure, i.e., the mechanism responsible for the transition in mechanical behavior was governed by a maximum shear stress criteria.

Jan 1, 1971