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By Robert Linton

AT its meeting on May 21, the Board of Directors approved the recommendations submitted by President Lovejoy and named a nominating committee for the year that is especially well distributed as to major technical interest and as to geography. As may be seen from the list published herewith, two come from the East, two from the Midwest, and three from the far West. Mr. Linton, the chairman, has been a valued member of the Institute for 36 years. A memorandum drawn up by Paul D. Merica, representing the A.I.M.E., and W. R. Barclay, representing The Institute of Metals of Great Britain, setting forth a reciprocal plan for special memberships in each Institute for members of the other, was discussed. Both societies already have formally approved the plan in principle and Dr. Merica was clothed with full authority to agree on a few small details that have not been settled, and conclude the arrangement in behalf of the Institute.

Jan 1, 1936

By R. D. Pehlke, W. M. Small

The effect of added alloying elements on the solubility of nitrogen in a liquid alloy of 74 wt pct Fe. 18 wt pct Cr, and 8 wt pct Ni has been studied. At 1600°C and 1 atm nitrogen pressure, aluminum, chromium, columbium, manganese, molybdenum, tantalum, and vanadium increase the solubility of nitrogen in the liquid 18 Cr-8 Ni alloy, whereas additions of cobalt, copper, nickel, silicon, tin, and tungsten decrease the solubility. Interaction parameters based on a constant ratio of iron, chromium , and nickel have been used to quantitatively define the changes in nitrogen solubility. INVESTIGATIONS on the solubility of nitrogen in liquid iron have usually treated the pure metal, and often the influence of a single added element. Pehlke and Elliott1 carried out a comprehensive study on the influence of alloying elements on nitrogen solubility in liquid iron, and compared their results with previous studies. Turnock and pehlke2 have recently reported a study on the solubility of nitrogen in multicomponent liquid iron alloys. The original interaction parameter proposed by wagner3 was successfully employed in these studies. The need for obtaining information on the gas solubility in more complex alloys is indicated by the increasingly significant role played by dissolved gases in the processing of iron-based materials. The commercial stainless steels, for example, represent an area where data on nitrogen solubilities are required as a function of composition, temperature, and gas pressure. The nitrogen solubility in liquid Fe-Cr-Ni alloys has been studied by Humbert and Elliott.4 Their investigation covered the entire ternary system, and attempted to use data from binary alloys to predict nitrogen contents in liquid ternary alloys. Langenberg,5 using a method developed by Morris and buehl,6 calculated the nitrogen solubilities in multicomponent iron alloys from data available on binary iron alloys at 1600°C. Nelson,7 using a regular-solution model, extended this prediction to include the effect of temperature. Chipman and corrigan8 and Turnock and pehlke2 also employed empirical methods to calculate nitrogen solubilities in liquid multicomponent iron alloys. In order to provide a basis for predicting nitrogen solubility in liquid stainless steels, and to extend the data available for thermochemically modeling systems of this type, the following investigation was carried out. The study involved measupements of nitrogen solubility in a nondilute alloy and examination of the influence of a small addition of an alloying element on the nitrogen solubility. The study was focused on a very narrow range Of composition in the vicinity of 18 wt pct Cr-8 wt pct Ni (herein referred to as 18-8) with special emphasis on the change of.nitrogen solubility with the addition of a fourth metallic element. EXPERIMENTAL PROCEDURE A Sieverts' apparatus, described previously,1,2 was used in this investigation. The charge materials were contained in a recrystallized alumina crucible and inductively melted. Additions were dropped from a glass side-arm. All materials were hydrogen-refined and vacuum-treated to eliminate or minimize the oxygen content. The source and purity of materials is given in Table I. The initial metallic charge was heated to approximately 1000°C and held under vacuum for 30 to 45 min as a means for removing hydrogen and adsorbed gases. Argon was used as the calibrating gas to determine the "hot volume". The average hot volume of the system was approximately 50 cu cm STP and a few experimental runs were made using a reaction bulb with a hot volume of 80 cu cm. The temperature range studied was from 1550° to 1700°C and the temperature was measured optically. THEORETICAL PRINCIPLES General Discussion. Nitrogen dissolves in liquid iron alloys according to the reaction: 1/2N2(g) = N (wt pct in solution) [l] The equilibrium constant, K1, for this reaction is written as:

Jan 1, 1969

By AIME AIME

Gentlemen, we have ten questions on our list and a number of us here probably have other questions to ask. Therefore, we arc going to allocate the time to these different questions so that we will try to be finished by 4:30; from then until 5 o'clock, if any of you have any questions on subjects not related, you will be free to ask them. Mr. Morton and I have changed the order of the questions. Instead of what is now first, we have put No. 3 first, progress on basic-roof construction, inasmuch as we feel that it is quite important. At this time, H. S. Robertson, of Harbison-Walker, is going to give you a short talk, with slides, on what has been done to date on this subject.

Jan 1, 1940

By S. D. Baumer, P. M. Hulme

B. M. Larsen (U. S. Steel Co., Kearny, N. J.)—Could we have some sort of an estimate on the cost of calcium carbide for this treatment? Also, if you used calcium carbide on carbon-saturated iron, would some effect hold back the reaction? P. M. Hulme (authors' reply)—As far as I know we have not worked on carbon-saturated iron. About 3½ pct C is the highest we have worked on. As far as cost goes, it depends on how much sulphur You have to take out. It will run anywhere from $0.75 to $2.00 a ton overall cost including the cost of carbide per ton of metal treated. That is, using the efficiencies that we show in this paper. We are getting better efficiencies now. T. L. Joseph (University of Minnesota, Minneapolis, Minn.)—Did you say 2 lb of calcium compound? Mr. Hulme—That is based on 10 lb of calcium carbide per lb of sulphur removed. D. E. Babcock (Republic Steel Corp., Youngstown, Ohio)—We have run this type of desulphurization process at the blast furnace. We have desulphurized between 100 and 200 at the blast furnace and taken the iron right out of the furnace with an injector system. The process gave us quite varied results; stratified iron, high sulphur on the top and low sulphur on the bottom, and we decided it was impractical. Mr. Hulme Just what was the type of injection equipment? Dr. Babcock—Nitrogen gas for one injection method and we used fairly large gun equipment. We followed the same general type of pattern you show only on a larger scale. Mr. Hulme—It sounds very much like what we call the gale of wind method. We have very erratic results with that. Dr. Babcock—We have tried several types, and particle sizes. The operation was not satisfactory. Mr. Hulme—Of course, we have never tried it on that scale except on a continuous basis. That is the reason we abandoned the idea of working on 35 ton ladles on a batch principle and took it as it came out of the cupola and treated it on the fly on a continuous basis. Dr. Babcock—When you compare the relative effectiveness of the process as carried out, it does not give a very favorable comparison of that process which is already developed. We did not feel it was worth the

Jan 1, 1952

By L. G. Austin, P. T. Luckie, R. R. Klimpel

Application of the Charles and Bond Laws to batch grinding can be deduced as a special case of solution of the first-order batch grinding equation. With additional limitations on the feed size distribution, this special case also gives Rosin-Rammler distributions of constant slope, with the 36.79% retained point a simple function of time. In general, the assumptions concerning the breakage parameters S and B which are necessary to get Charles' and Bond's laws, are not valid except as a crude approximation. With less restrictive assumptions, it is still possible to get approximate Rosin-Rammler distributions, but the slopes change with time of grinding, and the 36.79% retained point is a more complicated function of time. An alternative form to Charles' law is suggested, which accomplishes the same result and has the same errors, but avoids the use of energy of grinding and deals with specific rate of grinding.

Jan 1, 1973

By Alan Kissock

MOLYBDENUM is thought of as one of the rarer elements, for though it occurs in almost every country of the world it is seldom found in commercial quantities. In this country, however, there is one deposit which makes the term "rare" a misnomer. There are two commercially mined molybdenum minerals: molybdenite, the sulfide, and wulfenite, a lead molybdate. Of these two only the former is of any con- sequence but wulfenite is of mineralogical interest. It is frequently associated with oxidized lead ores and particularly in arid sections. Deposits, however, are usually quite superficial. It is highly oxidized and yet it does not appear to be an alteration product for it apparently disappears below water level and I know of no case where primary molybdenite, or in fact any molybdenum whatever has been identified below these surface wulfenite occurrences. The genesis of wulfenite would make an !- interesting study.

Jan 1, 1933

By J. Gurland

The wmk is concerned with the estimate of the degee of continuity of a particulate phase dispersed in a matrix. The first section is a verieu: of the parameters nzeasurable by quantitative rnetallogvaphy, namely the contact area and, under certain conditions, the number of contacts between particles of one phase. In the second section, the frequency of particle contacts is related to the degee of continuity of a randomly embedded phase orz the basis of the lattice-percolation model of Hamrnersley and others. THE extent of continuity of the constituent phases may greatly influence the structure-sensitive properties of phase mixtures, particularly if the phase properties differ significantly. An example from the literature1 is shown in Fig. 1. The marked decrease of the electrical resistivity of sintered silver-alumina bodies near 25 pct by volume of silver presumably occurs because the more conductive phase changes from a discontinuous to a continuous structure. Other example of electrical- and thermal-conductivity transitions are cited by kingery In view of these observations, a knowledge and understanding of the connectivity of the microstructure may be of considerable practical importance. This study is concerned with one aspect of the genera1 problem, namely the estimate of the degree of continuity of a particulate phase dispersed in an opaque matrix. The first section reviews the parameters measurable on a polished cross section by auantitative metallography, such as the contact area and under certain conditions, the number of contacts between particles of one phase. In the second section, an attempt is made to relate the frequency of particle contacts to the degree of continuity of a randomly embedded phase. For the purpose of this discussion, the aggregate has two constituents: the particulate phase and the matrix. The latter is considered to be homogeneous. The term "particle" is used with the meaning of "grain", i.e., a minute unit of matter which is usually a single crystal of a readily distinguishable substance within the confines of a simply connected boundary. While many of the concepts which are used here to describe the structural variables characterizing the microstructure of phase mixtures are qualitative and ill-defined, they are presented with the intent of stimulating further work of a more general and rigorous nature. I) PARTICLE CONTACT, CONTIGUITY, AND MEAN FREE PATH 1) Measurement of Contiguity. The fraction of the total internal surface of a phase a shared with particles of that same phase in a two-phase mixture of and $ is indicated by the contiguity ratio:3 whereis the area of the interface between o particles, and Sap is the area of the interface between a particles and matrix P, per unit volume of mixture. The contiguity ratio ranges from 0 to 1 as the distribution of the a phase changes from a completely dispersed to a completely agglomerated structure. The interfacial areas may be obtained from simple intercept measurements on a random plane, after Smith and Guttman where and N are the numbers of intercepts between a random line of unit length on a plane of polish and the interparticle and interphase interfaces, respectively. The equations are valid for any distribution of particle sizes and shapes.

Jan 1, 1967

By John Fritz

MEMBERS of the American engineering societies who were in London and Paris during the last days of. June and early July were present at many interesting gatherings. The official delegates of the Founder societies were the specially honored guests of the foreign engineering societies, but other American engineers took part in the meetings. The members of the delegation were: Arthur S. Dwight, Charles F. Rand, William Kelly, representing A. I. M. E.; Ambrose Swasey (general chairman of deputation), Ira N. Hollis, Jesse M. Smith, representing A. S. M. E.; F. B. Jewett, A. E. Kennelly, G. O. Squier, representing A. I. E. E.; Charles T. Main, Robert A. Cummings, John R. Freeman, D. C. Henny, representing A. S. C. E.

Jan 1, 1921

By Stanley D. Michaelson

The selection of the proper size of grinding mill for a given installation has long been a subject of discussion by mill manufacturers, consulting engineers, and their clients. It would be presumptuous for the author to propose a simple cure-all. Instead this paper describes a method that has been used over a period of years to determine mill size with a reasonable degree of accuracy. Furthermore, this method claims no theoretical background, for ball milling, as practiced today, is still an art. The selection of the proper size and type of mill must, therefore, be made on an empirical basis. A ball mill, in order to grind, must deliver power through its grinding media to the material to be ground. Assuming that a mill is properly supplied with grinding media, ore pulp, and power, the amount of grinding done is closely proportional to the power expended. With this as a basic premise we can, for the moment, forget about type of mill (i.e., overflow, low-level mill, or other design) and base our calculation of mill size on the total power required to grind a ton of a particular ore from a given size of feed to a given size of product. When this is determined the type and size of mill, its speed, and the weight of the load of grinding media can be selected to absorb this amount of power. Development of Technique Some years ago, W. L. Maxson, F. Cadena, and F. C. Bond, working in the Minerals Research Laboratory of Allis-Chalmers Co., Milwaukee, developed and described a laboratory technique for comparing grindability (or relative ease of grinding) of various ores at different mesh sizes of grind.' Additional papers2 have described the work done since that time, and the most recent paper3 summarizes results for hundreds of tests on many ores and materials. The essentials of the testing method are: a 12 by 12-in. laboratory mill operating at a constant speed of rotation, using a constant volume of ore, and a constant ball load having a relatively constant grinding surface. Tests are run at 250 per cent circulating load with the feed, circulating load and products screened on calibrated Tyler standard screens. Certain ores from operating mills, representing various mineralogical types, were selected for reference and designated "standard ores." In most cases, detailed studies of the grinding circuits were made when samples of mill feed were taken for grindability tests. Each analysis formed a basis for comparison of the grindability in the laboratory with actual power requirements for a particular grinding circuit. In addition, a critical study was made to evaluate the differences in the various grinding circuits so that two or more field installations having different layouts (single-stage grinding, two-stage grinding, or others), and using mills of various diameters operating at various speeds with differing circulating loads, etc., could be compared. Large-scale tests by mill operators complemented by laboratory investiga-

Jan 1, 1947

By Stanley D. Michaelson

The selection of the proper size of grinding mill for a given installation has long been a subject of discussion by mill manufacturers, consulting engineers, and their clients. It would be presumptuous for the author to propose a simple cure-all. Instead this paper describes a method that has been used over a period of years to determine mill size with a reasonable degree of accuracy. Furthermore, this method claims no theoretical background, for ball milling, as practiced today, is still an art. The selection of the proper size and type of mill must, therefore, be made on an empirical basis. A ball mill, in order to grind, must deliver power through its grinding media to the material to be ground. Assuming that a mill is properly supplied with grinding media, ore pulp, and power, the amount of grinding done is closely proportional to the power expended. With this as a basic premise we can, for the moment, forget about type of mill (i.e., overflow, low-level mill, or other design) and base our calculation of mill size on the total power required to grind a ton of a particular ore from a given size of feed to a given size of product. When this is determined the type and size of mill, its speed, and the weight of the load of grinding media can be selected to absorb this amount of power. Development of Technique Some years ago, W. L. Maxson, F. Cadena, and F. C. Bond, working in the Minerals Research Laboratory of Allis-Chalmers Co., Milwaukee, developed and described a laboratory technique for comparing grindability (or relative ease of grinding) of various ores at different mesh sizes of grind.' Additional papers2 have described the work done since that time, and the most recent paper3 summarizes results for hundreds of tests on many ores and materials. The essentials of the testing method are: a 12 by 12-in. laboratory mill operating at a constant speed of rotation, using a constant volume of ore, and a constant ball load having a relatively constant grinding surface. Tests are run at 250 per cent circulating load with the feed, circulating load and products screened on calibrated Tyler standard screens. Certain ores from operating mills, representing various mineralogical types, were selected for reference and designated "standard ores." In most cases, detailed studies of the grinding circuits were made when samples of mill feed were taken for grindability tests. Each analysis formed a basis for comparison of the grindability in the laboratory with actual power requirements for a particular grinding circuit. In addition, a critical study was made to evaluate the differences in the various grinding circuits so that two or more field installations having different layouts (single-stage grinding, two-stage grinding, or others), and using mills of various diameters operating at various speeds with differing circulating loads, etc., could be compared. Large-scale tests by mill operators complemented by laboratory investiga-

Jan 1, 1947

By J. S. Blakemore, E. O. Hall

SERRATED yielding, commonly called the Portevin-Le Chatelier effect, is known to occur in a wide range of alloys where the moving dislocations are considered to be locked by the diffusion of solute atoms to them. From the work of Cottrell,1 an approximate relationship between strain rate, E, dislocation density, p , and solute diffusion coefficient, D, giving rise to serrated yielding is: e =Dp

Jan 1, 1969

By Stanley D. Michaelson

THE selection of the proper size of grinding mill for a given installation has long been a subject of discussion by mill manufacturers, consulting engineers, and their clients. It would be presumptuous for the author to propose a simple cure-all. Instead this paper describes a method that has been used over a period of years to determine mill size with a reasonable degree of accuracy. Furthermore, this method claims no theoretical background, for ball milling, as practiced today, is still an art. The selection of the proper size and type of mill must, therefore, be made on an empirical basis. A ball mill, in order to grind, must deliver power through its grinding media to the material to be ground. Assuming that a mill is properly supplied with grinding media, ore pulp, and power, the amount of grinding done is closely proportional to the power expended. With this as a basic premise we can, for the moment, forget about type of mill (i.e., overflow, low-level mill, or other design) and base our calculation of mill size on the total power required to grind a ton of a particular ore from a given size of feed to a given size of product. When this is determined the type and size of mill, its speed, and the weight of the load of grinding media can be selected to absorb this amount of power. DEVELOPMENT OF TECHNIQUE Some years ago, W. L. Maxson, F. Cadena, and F. C. Bond, working in the Minerals Research Laboratory of Allis-Chalmers Co., Milwaukee, developed and described a laboratory technique for comparing grindability (or relative ease of grinding) of various ores at different mesh sizes of grind.1 Additional papers2 have described the work done since that time, and the most recent paper3 summarizes results for hundreds of tests on many ores and materials. The essentials of the testing method are: a I a by 12-in. laboratory mill operating at a constant speed of rotation, using a constant volume of ore, and a constant ball load having a relatively constant grinding surface. Tests are run at 250 per cent circulating load with the feed, circulating load and products screened on calibrated Tyler standard screens. Certain ores from operating mills, representing various mineralogical types, were selected for reference and designated "standard ores." In most cases, detailed studies of the grinding circuits were made when samples of mill feed were taken for grindability tests. Each analysis formed a basis for comparison of the grindability in the laboratory with actual power requirements for a particular grinding circuit. In addition, a critical study was made, to evaluate the differences in the various grinding circuits so that two or more field installations having different layouts (single-stage grinding, two-stage grinding, or others), and using mills of various diameters operating at various speeds with differing circulating loads, etc., could be compared. Large-scale tests by mill operators complemented by laboratory investiga-

Jan 1, 1945

By R. E. Head

UNDER existing economic conditions, the treatment of gold ores occupies an outstanding position in metallurgical activity. The increased price of gold has automatically brought about a reclassification of old dumps and low-grade orebodies, and many that formerly were considered valueless are now being vigorously investigated and worked. This situation has developed problems that have taxed the ingenuity and ability of metallurgists to the utmost. Some gold ores give a high extraction of gold by standardized methods of treatment; other ores of apparently similar character yield but a relatively small percentage of their gold when treated by the same processes.

Jan 1, 1936

By R. L. Sandefur

Professor Kim's excellent review article1 on ultimate pit planning contains a statement of the apparently widely held but incorrect belief that "kriging provides information on the confidence limits around the estimate which no other extension technique is capable of supplying." If a geostatistical approach is valid, there are several good reasons for using kriging but uniquely provided confidence limits is not among them. If kriging is possible, then confidence limits can be calculated for any other linear estimate. Kriging is the procedure which finds the weighing coefficients which minimize the estimation variance [ ] is the variance of blocks like the one being estimated. wi is the covariance of the estimated block and the i-th sample. oij is the covariance of the i-th sample and the j-th sample. Ai is the weighing coefficient for the i-th sample. N is the number of samples. The formula for estimation variance can also be used to calculate confidence limits on blocks estimated by any linear combination of sample grades including: 1) Polygons a. Hole nearest block centroid is assigned weight 1. All other holes are assigned weight 0. b. Polygonal area weighing. The samples are assigned weights directly proportional to their polygonal areas of influence within the block. The weights sum to 1. 2) Inverse distance weighing. The samples are assigned weights directly proportional to the reciprocal of their distance to some power from the block centriod. The weights sum to 1. Often only holes nearer than a certain distance and/or the nearest hole in each octant is considered. If there is an antisotropy to the ore body, an elliptical transformation often gives more accurate results. The most popular inverse distance weighing function is inverse squared distances (IDS). 3) Simple average. All samples within a certain distance are assigned the same weights which sum to 1. As an example of the applicability of the above formula to any linear estimate an internal verification on copper grade of 665 blast holes on the 1040 bench at Utah's Island Copper mine is presented. Internal verification is simple and applicable to non-linear estimates as well. Any sample (drill hole, bulk, trench, or chip) is dropped out, the value of the dropped sample estimated and compared with the actual value. If this is done for all samples, a measure of confidence limits is obtained. Internal verification is used in Cominco's MEPS system4 and in the BLUEPACK system.5 Comparison of block values against estimates697 is not dissimilar to internal verification. Results of the internal verification are presented below:

Jan 1, 1980

By Arthur Knapp

THE economics of fuel oil must be considered in two aspects; viz., fuel oil as one of a number of competitive fuels and fuel oil as a refined product of petroleum. "Fuel Oil" is usually defined as that product of petroleum which is used in furnaces for the generation of heat by simple combustion. It is true that any product of petroleum used in an internal combustion engine is essentially a fuel oil, but such oils are not included in the class known as "Fuel Oil." FUEL OIL AS ONE OF SEVERAL FUELS For a number of years, the petroleum industry has believed that the petroleum reserves were very limited and that petroleum products would begin to become scarce within a few years. The industry has believed FIG. 1.-DAILY AVERAGE CONSUMPTION OF FUELS BY PUBLIC UTILITY POWER PLANTS. that there would be a gradual but continual rise in prices as a result of this impending scarcity.

Jan 1, 1928

Executive Committee. ROBERT H. RICHARDS, Chairman. ALBERT SAUVEUR, Vice-Chairman. TIMOTHY W. SPRAGUE. HENRY A. WENTWORTH. AUGUSTUS H. EUSTIS, Secretary, 131 State St.., Boston, Mass. The twelfth meeting of the Boston Section was held at the Engineers' Club, 2 Commonwealth Ave., Boston, on Monday evening, May 5, 1913, fifteen members and three guests being present.. The meeting Was called to order by Vice-Chairman Albert Sauveur, Who introduced the speaker of the evening, Utley Wedge, of Ardmore. Pa. Mr. Wedge read a, paper on the chloridizing roasting of copper ores, Which lie illustrated With a stereopticon, and Which will be published in full by the Institute. After Mr. Wedge had closed his paper, Which was received with great interest by all the members. the following discussion took place: Professor Hofman said that iron was formerly thought to bo essential to the Henderson process, but that it is not absolutely essential. Mr. Wedge said that., as far as he knew, the presence of some sulphates Was an absolute necessity, and that lime ores can be worked satisfactorily, although lime has an effect on the subsequent leaching. He added that the low fuel consumption of the latest Wedge furnace is a result of direct application of the combustion gases to the ore. Ten per cent. of fuel Was needed with the muffle furnace. By direct firing lie has been able to reduce the fuel needed to from 2 to 4 per cent. In the latest types of furnaces the uses from the upper hearth are not used. There is doubtless some tem-perature at Which the loss from copper chloride volatilization becomes serious, but in practice on Rio Tinto ores troublesome losses are not experienced.. The copper apparently does not become volatile at the temperatures where we would expect that it would. There are doubtless ores in which there would be serious danger from losses from this source. The latest furnaces do not use the high temperatures that Were needed in the former single-hearth Wedge furnaces.

Jan 7, 1913

By George S. Rice

MECHANIZATION in coal mining is a phrase which has attracted world-wide attention, and those persons not engaged on the practical side of coal-mine operations seem to regard mechanization as a panacea for many difficulties of the industry. The term "mechanization" as now employed by experienced mining men is restricted to mean mechanical loading of coal at the face. In a broader sense mechanization in coal mining was begun nearly half a century ago with the introduction of compressed-air undercutting machines and locomotives and, later, of electrically operated locomotives, undercutting and shearing machines, and drills. The use of coal-cutting machines, which has increased until now about three- fourths of the bituminous coal produced is being mechanically cut, first displaced hand-picking except in a few long-wall mines, and later gradually displaced the dangerous and coal-smashing method of blasting-off- the-solid.

Jan 1, 1929

By J. G. EVANS

It is with a certain amount of trepidation that a man considers gathering his family of six, traveling across a continent, two oceans and a sea, and going to live in a foreign land. But "pioneering" is still possible, and a nation is "foreign" only until one has lived in it. We've had eight years of pleasant experiences in South Africa on which to base that contention. It was in January 1938 that my brother first extolled the virtues of South Africa to us, and just three months and a thousand small problems later my wife, two daughters, three sons and I were boarding a ship in New York, bound for Cape- town. We arrived in May, and within a week had found a furnished home with beautiful lawns, tennis, court, a garden, and everything needed for family comfort. The children were soon located in school, and I settled down to the task of entering the importing business, as had been my original intention. The importing business started out slowly but was coming along nicely when on Sept. 3, 1939 the war started in Europe and a declaration of war by the Union of South Africa was made on the following day. With the declaration came import and export controls. In fact we were controlled completely out of business in one stroke. Liquidating the business, I went to Johannesburg intending to find work in one of the mines. Government health and safety regulations require that all underground mine employees have a red ticket from the Miners Phthisis Bureau certifying that they are not susceptible to silicosis, the scourge of the Witwatersrand gold mining industry. I made application to the Bureau, completed a form covering my whole past life, and was given a complete physical examination. The following day I was startled to find that I had not only been refused a red ticket but had been classified with a green ticket which for- ever bars one from working underground in any scheduled mine in the country, and all this in spite

Jan 1, 1949

By H. S. Cannon, F. N. Rhines

IT. H. Hausner( Sylvunia Electric Products Inc., Bay-side, N. Y.)—The results reported by the authors are interesting because they contribute some information on the principles of sintering and also because of a certain practical value for the manufacture of powder . metallurgical parts in production. Pressed powder compacts are frequently sintered in a boat in several layers so that the lower layers are actually sintered under the weight of the upper layers and, therefore, are frequently deformed. The authors gave information on this deformation process by determining the effect of the load on the density. The experimental setup of the authors, however, has a disadvantage—that the sintering occurred in a cylindrical container so that the deformation was stopped in the direction of the diameter. The authors should have rather called this a study of hot-pressing than of sintering. I want to call the attention of the authors and others interested in sintering under load to the experimental work described by Dawihl and Rix.' These authors sintered cylindrical compacts under a lengthwise ten-sional load of 0 to 25 g and found that the shrinkage in length and diameter changes considerably with the load, but that the volume shrinkage is affected very little by the load. H. S. Cannon (authors' reply)—We wish to thank Dr. Hausner for calling attention to the experimental work of Dawihl and Rix. The fact that they found volume shrinkage little affected by varying tensional loads during sintering seems to support the contention that the forces due to loading are equivalent to and additive to the forces which cause densification. This contention is the basis of our comparison of sintering and creep.

Jan 1, 1952

By Colin M. Sargent

THE electron diffraction pattern as seen in the electron microscope represents an approximately plane section of the reciprocal lattice. Identifying the zone axis of a diffraction pattern is often laborious for a crystal of low symmetry, especially if several crystal orientations are present. A program has been written in FORTRAN II (Version 2) for the IBM 1620 which will assist in plotting out reciprocal lattice sections. The calculation is performed by choosing a zone axis [uvw] and determining the d spacings and angle between the planes (Ow - v) and (wO - u). (This procedure is varied slightly for zone axes with Miller indices equal to zero.) Cartesian coordinates for the reciprocal lattice points are then calculated for these planes and for (ww— [u + v]). Any reciprocal lattice points within the parallelogram formed by the origin and these three points are then computed. The input and output are best described by use of an example (Monoclinic ZrO2) used in the author's work. Input consists of two cards: 1) The crystal parameters: a, b, c, a, 8, y, e.g., 5.1477, 5.2030, 5.3156, 90.000, 99.380, 90.000. 2) Three numbers, D, E, F, e.g., 5, 3, 3.0710-the first is the limit of the sum (u + v + w) considered; the second limits the absolute value of u, v, and w to (E- 1), e.g., 2, so that zones (221) would be the highest order considered in this example; the third, F, is the required constant, yL. Output is in the format shown in Table I. The reciprocal lattice sections may then be plotted on standard graph paper and compared with the unknown diffraction pattern. (Note that no attempt has been made to deal with systematically absent reflections.) A card deck for the program is available from Colin M. Sargent, ARZ, W-PAFB, Dayton, Ohio, 45433.

Jan 1, 1969