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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 Fred K. Stieler

The largest updraft sinter machine built has a roasting area of 132 sq.m. Sinter machines with an area of over 200 sq.m. are under consideration. These large machines require new techniques with regard to drive thrust and sealing mechanism. To optimize the results of sinter plants, the charge preparation systems and the return fines circuits were improved. Special crushers were developed for crushing return sinter of up to 800 ºC to a size of minus 6 mm. Strand cooling of Pb-sinter was investigated. Different methods for plant ventilation were compared. New processes, which might replace sintering in the future, are in the stage of large - scale testing.

Jan 1, 1977

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 AIME AIME

THAT the copper industry is in serious straits is admitted. So are the lead and zinc industries, and both lead and zinc are tariff protected. Conditions in the Western lead, zinc and silver mining districts are desperate. These are illustrated by the present disastrously low prices of these metals compared with the levels of two and three years ago and by the comparative rates of metal production and the present activity of the western mining camps. The January production of copper was 32.2 per cent of the 1929 average rate, of steel was 32.3 per cent of the 1929 average, while January production of pig iron was only 27.6 per cent of the monthly average of 1929. The copper industry is only one of many industries that are suffering in this tragic depression. It has been exceedingly prosperous and will be prosperous again when business recovers. Its difficulties are not unique and it does not deserve the tariff equivalent of a dole. The burden of subsidizing through an effective tax or tariff the minority group of high-cost mines would be a heavy one on the consuming public of the United States. On the basis of 1929 consumption figures, an effective 5c. tariff would result in an increased copper bill to the country of over $109,000,000, of which the electric manufacturers would contribute $26,000,000, the telephone and telegraph companies $16,000,000, and the automobile manufacturers $14,000,000. And of this total bill, only $33,000,000 would go to those who are urging a tariff and the balance of $76,000,000 would be paid to those who do not admit either its necessity or usefulness under present conditions.

Jan 1, 1932

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 E. J. Roberts

Prior to 1916, thickening was an art, and any accurate decision as to what size of machine to install to handle a given tonnage of a specific ore must have been one of those intuitive conclusions, based on both intimate and extensive acquaintance with thickners and ore pulps. Then in 1916 "knowledge of acquaintance," became "knowledge about" with the publication of the Coe and Clevenger paper.' The unit operation of thickening had graduated to the status of an engineering science. The fundamental similitude relationships for the two major phases of the operation were defined so clearly that batch tests on models as small as liter cylinders could serve to specify protypes as large as 325 ft in diameter.

Jan 1, 1949

By J. B. STEWART

T HE position of each hole of any series of holes was carefully located by the surveyor, plotted in plan and elevation, and numbers assigned to them. The second series was staggered halfway between the first and so on for each succeeding series. Two and a, half to three months was frequently required to grout a series of holes calculated to insure safety for sinking a distance of 10 meters or say 33 ft. Naturally the initial operation called for drilling and grouting two series of holes but thereafter each sinking campaign required only one series of holes as shown in Fig. 8. The pump handling about 75 gal. per minute corresponds to 46 tons of grout or 6 tons cement delivered as a fair over-all average for the 24 hours. Thus 600 to 800 tons of cement was used for a single section of 33 ft. depth of shaft. One campaign used nearly 1100 tons of cement. Some holes were effectually sealed with as low as 30 tons, while many required 60, and some 80 to 90 tons.

Jan 1, 1926

By D. M. Clippinger, J. J. Eidel, J. E. Frost

At Ithaca Peak, one of three peaks situated on Duval Corporation's Mineral Park property, a 'Single pulse of quartz monzonite magma intruded the isoclinally folded Precambrian Cerbat complex consisting of schist, quartzfeldspar gneiss, and amphibolite. The magma reacted with the Precambrian rocks, yielding a quartz-diorite rim. As the magma crystallized within the rim, it produced, progressing in'ward, annular bodies of quartz monzonite porphyry, quartz porphyry, and a distinctive quartz porphyry containing abundant podiform, tabular, and irregular quartz masses. These quartz masses are spatially associated with discontinuous crenulate and tabular quartz veinlets. Following crystallization, the Ithaca Peak stock and part of the surrounding Cerbat complex were fractured, producing a stockwork of predominantly northwest and northeast-trending fractures. The stockwork was then mineralized during a period of decreasing temperature by: ( 1) quartz-pyritechalcopyrite veinlets without distinct selvages and with distinct alteration envelopes in which orthoclase was altered to quartz and sericite; (2) quartz-pyrite-chalcopyrite veins and veinlets with distinct selvages and without distinct alteration envelopes; and (3) pyrite-chalcopyrite- galena-sphalerite veins. The Ithaca Peak stock was pervasively sericitized or argillized. The Mineral Park property is surrounded by a zone of propylitization Alteration and mineralization are coextensive in space and were, at least in part, coincident in time. The Mineral Park property possibly was elevated and exposed twice prior to its present exposure. The protore has been enriched by a factor of three. The Ithaca Peak supergene ore body is conformable with present topography and lies above the present water table. The shape of the ore body was controlled by erosion and oxidation of the previously enriched mineralization. Hematite, goethite, and jarosite present at the surface indicate a high primary pyrite environment, but it is possible to recognize that a small proportion of veins and veinlets contained greater amounts of chalcocite and chalcopyrite than pyrite. The igneous facies on Ithaca Peak are the products of differentiation within a solidifying rim according to Bowen's Reaction series. The innermost facies contain more quartz and orthoclase than the outer facies. Late magmatic quartz in the case of Ithaca Peak contains small amounts of pyrite and molybdenite. It is suggested that this quartz-sulfide mineralization, may also have been the product of differentiation the product of differentiation, may indicate that the later hydrothermal mineralization may also have been the product of differentiation.

Jan 1, 1968

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. 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 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 J. Campbell

The hydrostatic tensions in pure metals and long freezing range alloys are evaluated theoretically considering the viscous flow of residual liquid and the general plastic collapse of the casting. The former mechanism operates until the radii of the remaining liquid channels are of the order of 10-3 cm radius, after which the second process limits the attainable negative pressures which range from —1500 to -4200 atm for nickel, aluminum, copper, and iron. In previous reports1'2 the maximum possible hydrostatic tensions developed during solidification have been evaluated employing spherical symmetry. This inevitably involves the concept of totally confined liquid regions, the reality of which has recently been brought into serious doubt3 following observations of the freezing of various transparent materials. Regions of interdendritic liquid which are apparently about to be isolated by the growth of dendrite arms are observed not to be cut off because the advance of the dendrite tips is arrested until the region has completely solidified. Although this may be the result of the very low melting point liquid available at this late stage in freezing, or the result of a higher rate of heat transfer to the dendrite tips because of the acceleration in liquid flow rate near the completion of solidification, this point need not concern us. In this paper an attempt is made to approximate more closely to real conditions; thus the assumption of confined liquid regions is abandoned and cylindrical symmetry adopted. It has also been necessary, therefore, to consider the viscous flow of residual liquid through the remaining channels in the pasty zone. Furthermore, the rather special conditions of spherically symmetric solidification led to extremely high solidification rates and hence unusually high tensile stresses as solidification neared completion; the cylindrical Symmetry appears to yield solidification rates and stresses which are more typical of real castings. The rate of solidification is first considered since this critically affects both the flow of the liquid and the creep of the solid.

Jan 1, 1969

By R. F. Favreau, G. E. Larocque

The objective of the study is development of an analytic process based on rock and explosive properties which allows the prediction of the stress distributions and fracture zone around explosive charges. A computer program has been written for this purpose, using a finite difference approximation to the momentum equation, to describe the motion of the surrounding rock by step-wise integration. Its application provides ground motion data such as stress components, displacement and particle velocity as a function of position and time with respect to a spherical charge. The effect of tensile and shear failure on these ground motion parameters can be included. The analytic procedure contained within this program has been used to predict peak stress as a function of distance in rock for rock-explosive combinations for which field data exists to determine peak stress levels directly. Two methods have been used for the spherical charge which in the analysis replaces the actual cylindrical charges: instantaneous application of peak pressure to a charge having a radius equivalent in contained volume to a charge having a length to diameter ratio of 1, and gradual pressure build up to peak pressure with a charge radius equivalent in contained volume to the total charge used in the field test. The agreement with field data would appear best with the latter treatment. The peak stresses obtained are in closer agreement with field measured peak-stress and a better approximation to field stress wave duration exists in the case of gradual pressure build up. One of the assumptions made in the analysis is that, in the case of cylindrical charges, the explosion pressure rather than the detonation pressure of an explosive establishes the peak stresses in the

Jan 1, 1971

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

By G. J. Saulnier, K. E. Goddard

The US Geological Survey has developed mathematical modeling techniques to assess potential impacts on the hydrologic system resulting from mining of coal and oil shale in northwestern Colorado. Potential impacts of oil shale mining on groundwater hydrology have been simulated using a quasi-three-dimensional, finite-difference model. The model has been used to simulate hypothetical mine dewatering operations at two proposed oil shale mines. Dissolved-solids distribution and solute transport were simulated with a three-dimensional, finite-difference model that was used to evaluate the effects of flooding in hypothetical abandoned mines. The potential impacts of coal and oil shale mining on water-sheds were investigated using a digital watershed model. This deterministic physical-process model predicts responses of the hydrologic system to modifications resulting from mining in basins having snow as the major source of runoff and recharge. Potential impacts of coal development on surface water quality were investigated using a river quality assessment model. The models developed by the US Geological Survey have been useful in defining the potential effects of mining in northwestern Colorado. They are an important first step in analyzing the existing data base and in predicting impacts and new data needs. The models may not be totally applicable on the local scale but are very meaningful on the regional level.

Jan 1, 1983

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 R. L. Bowron

GROWING use of basic slag in the agricultural industry is of special interest and importance to the iron and steel industry of the Birmingham district, providing an increasing outlet for this by- product. Such use of basic slag was first developed in England about fifty years ago where it soon became wide- spread. About twenty years ago the Tennessee Coal, Iron & R. R. Co. began its production in the United States after exhaustive tests had demonstrated its high value here as in Eng- land. It is particularly valuable on heavy clay and acid soils, improving their physical condition, increasing their alkalinity where that is desirable, and providing soluble phosphate and small amounts of manganese to soils deficient in these respects. Its use in agriculture is growing as its advantages are becoming more widely known.

Jan 1, 1937

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 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