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By A. C. Losekamp, R. M. Fincel, J. B. Conway

TANTALUM like several other metals exhibits a great affinity for or reactivity with certain gases. Tested in atmospheres which are not completely pure this metal becomes contaminated by certain impurities (notably nitrogen, oxygen, and hydrogen) with a resultant change in various properties. Well-known are the drastic changes in hardness, ductility, and electrical resistivity. Also previously recorded are the distortions of the lattice1,2 accompanying the increased content of oxygen and nitrogen in tantalum. This pronounced affinity for certain gases along with the significant property variations thereby produced makes it necessary to observe additional experimental precautions in making various property determinations with tantalum. Some recent measurements of the linear thermal-expansion characteristics of tantalum have made this point quite emphatically. Failure to observe the greatest precautions in reducing the contaminating influence has led to some extremely distorted results. Not only does such operation lead to different heating and cooling curves during the thermal-expansion measurements, but upon cooling to room temperature the sample is found to have suffered a permanent (actually semipermanent since the specimen can be returned to its original length by a 3-hr degassing treatment in vacuum at 2300°C) elongation. In addition, subsequent heatings will not follow the initial heating curve unless the expansion measurements are based on the actual specimen length at the beginning of the heating cycle. These thermal-expansion measurements were made by positioning the specimen horizontally at the longitudinal midpoint of a 20-in.-long, 4-in.-diam resistively heated tungsten tube furnace. The specimens were in the form of 20-mil sheet formed into a sharp U-shape about 0.3 in. in width and 0.3 in. in depth. Fiducial marks in the form of 0.010-in. holes drilled at either end of the 2.5-in.-long sheet specimens were viewed by a paired filar micrometer telescope system mounted on an Invar bar on the top of the furnace. Tests were made in helium with temperatures measured by a Pt/Pt-lORh thermocouple to 1000°C; above 1000°C temperature was measured by means of an optical pyrometer sighted into a black-body hole drilled into the specimen-holding fixture in the vicinity of the specimen. Previous tests with black-body holes drilled in a 0.25-in.-diam rod specimen had confirmed the temperature uniformity between specimens and specimen-holding fixture in this furnace. A typical thermal-expansion curve for arc-cast tantalum in helium is shown in Fig. 1 for the case where no attempt is made to remove all the impurities from the cover-gas atmosphere. The initial heating curve appears smooth and has a shape characteristic of all metals. However, it is known that the sample begins absorbing impurities at temperatures above 1000°C which are sufficient to distort the lattice and lead to specimen lengths which are somewhat larger than would be obtained in the absence of contamination. During the cooling cycle the expansion curve is seen to be significantly different from the heating curve. This is due to the impurity content causing a semipermanent (i.e., permanent until the impurities are removed by vacuum degassing) deformation which will not allow the speci-

Jan 1, 1965

By J. L. Wyatt

J. W. Tomlison—It seems probable that the author's conclusion, that the conductivity of the slags decreases with increasing content of FeO, is erroneous due to the method of plotting the data. Thus, the results of other workers2,3,13,14 show that the conductivity of molten oxide slags increase with increasing temperature, indicating that the trend of the curve shown by the authors is due to the fact that the data for the slags containing relatively small amounts of FeO are at higher temperatures than those for the slags containing relatively large amounts of this oxide. In this connection the work of Wejnarth2 may be mentioned in which in an investigation of a series of iron silicate slags, the conductivity was found to increase with increasing content of FeO; the fact that the author's slags contained TiO2 probably does not affect the issue since it has been shown15,16 that, at least as far as conductivity is concerned, this oxide behaves essentially the same as silica and that mixtures of the oxide with silica have a very low conductivity of the order of 10-4 mhos at 1750°C. In connection with the experimental technique it may be noted that bubbles of carbon monoxide would probably be produced at the surface of the graphite electrodes in contact with the melts thus rendering the accuracy of the measurements somewhat open to question. J. L. Wyatt (author's reply)—Dr. Tomlinson makes reference to works by Martin and Derge, which indicate that slag systems containing no TiO2 are charac-terized by an increase in conductivity with an increase in temperature. Further, I have cited his examples, which indicate a drastic increase with small temperature change. The results of Dr. Derge's work have no bearing whatsoever on the results of this paper, since the slags under study here contained up to 70 pct TiO2. One cannot generalize on slag systems as a whole on the basis of any one family, for there are definite compounds of ionic nature formed in each system, and the ionic mobilization of these is something that cannot be determined without actual experimentation. Dr. Tomlinson makes reference to an unpublished work which, he claims, shows that titanium dioxide behaves the same as silica in regard to conductivity. Since I have not had 'an opportunity to see the paper I cannot make any comments, but unless the same slag system was studied there can be no specific comparison of results. If the decrease in conductivity of the slag was due entirely to temperature phenomena, as Dr. Tomlinson indicates, then one is without an explanation of the values shown in the paper for slags containing from 10 to 20 pct FeO. I cannot agree that the deviation in findings was appreciably related to temperature change; further, I believe this point was completely covered in the paper. References 13 Endell and Hellbrügge: Glasstech. Ber. (1942) 20, 277. 14 Bockris, Kitchener, Ignatowicz and Tomlinson: Faraday Society. Discussion (1948) 4, 265. 15 Tomlinson: Thesis, London 1949. 10 Bockris, Kitchener, Ignatowicz, and Tomlinson: Unpublished work.

Jan 1, 1951

By A. A. Burr, D. Caplan

IN carrying out experiments in metal-gas systems, it is often difficult to estimate how long a reaction period must be allowed for equilibration. This is especially true in systems for which the diffusion coefficient is not known. In an investigation of the Cr-O system, the necessary information was obtained in the way described subsequently. Two pieces of chromium sheet of different thicknesses were packed in mixed powders of chromium and Cr2O3 and sealed into an evacuated metal capsule by welding. Austenitic stainless steel proved suitable as a capsule material up to 1350°C and molybdenum at higher temperatures. The capsule then was heated at a specific temperature for a time sufficient to saturate with oxygen the thinner but not the thicker specimen. The time is readily estimated from a preliminary experiment if a sufficient thickness differential is chosen. Under these experimental conditions, the metal sheet is exposed to a partial pressure of oxygen equal to that of the dissociation pressure of the oxide and hence can become saturated with oxygen without developing any second phase. After cooling, both samples were analyzed for oxygen by the vacuum fusion technique. Assuming that the surface of the specimens is kept saturated with oxygen throughout the experiment, the diffusivity can be calculated by application of the appropriate solution to Fick's second law. This is most conveniently accomplished by use of a plot of mean fractional saturation (Cm-Co) / (Cs-Co) vs (Dt) ½/L as given by Darken,' where t is time, L the half-thickness of the sheet, and D the diffusion coefficient. Cm, the mean concentration at time t, is given by the oxygen percentage in the thicker specimen; Cs, the constant surface concentration, is the oxygen percentage in the thinner specimen; Co is the oxygen content of the original chromium sheet. (Cm-Co) / (Cs-Co) is calculated and the corresponding (Dt) ½/L is read from the graph and D thereby calculated. By performing the experiment at two temperatures and plotting log D vs 1/T, the D for any temperature is established. (The activation energy for diffusion may be obtained from the slope of the line.) Once D is known as a function of temperature, it is a simple matter to estimate the correct reaction time for experiments with single specimens at other temperatures by equating (Dt) ½/L to its value for effective saturation. Darken' points out that this is approximately unity. The solubility of oxygen in chromium, as determined from the thin samples at each temperature, proved to be approximately 0.03 pct at 1350°C and less at lower temperatures. Due to low solubility and to the known difficulties of the vacuum fusion method in this low range, the values obtained for D (5x10-0 sq cm per sec at 1350°C) are estimated to be correct only within an order of magnitude. Nevertheless, the method has proved useful in estimating reaction times. Its usefulness will be greater in metal-gas systems where the solubility is appreciable. A second function of the method is to permit the determination of diffusion coefficients in systems for which the conventional techniques are experimentally difficult or impossible.

Jan 1, 1956

Open pits and quarries are the major sources of all hard rock tonnages mined today. Normally, ore is fractured from the pit bench face by blasting and then truck-hauled to a primary crusher on the pit rim for conveying to the concentrator; overburden (or waste) is generally truck- hauled from the pit to dump areas. This system of truck haulage of blasted ore and waste accounts for some 25 to 50% of total mining costs in a typical open pit operation and commonly is the largest single cost factor. In an effort to reduce these costs mining companies are continually replacing their older and smaller trucks with larger ones and at the same time are studying conversions to conveyor haulage, since this method of transport seems to offer possibilities of lower costs per ton of material handled. In these studies it seems logical to consider installing these conveyors as close as possible to the mining face for maximum use of conveyors and minimum use of trucks, since a conveyor is much less labor intensive than a trucking system. However, direct loading onto a conveyor near the retreating bench face is impractical because of excessive lump size. Consequently, it is conventional practice to load the blasted rock into trucks at the face for haulage to the crusher. After crushing, the rock is fed onto a conveyor. The closer the crusher is to the mine, or if it is in the mine, the shorter the truck haul and the greater the potential for cost saving. At first appearance, in-pit conveying might seem to be an obvious solution, yet many mining companies have considered such a move while few have implemented it. To determine why, it is necessary to consider the factors influencing such a decision. A conveyor system tends to be capital intensive but can have a lower operating cost per ton than trucks. Trucks, on the other hand, tend to cost less to buy but more to operate; thus, economics tend to favor the conveyor if it can be operated close to capacity and if a sufficiently large tonnage can be passed over the system to mitigate the effect of its high capital cost. In a typical hard-rock, open-pit mine it is necessary for a variety of reasons to mine several faces of both ore and waste at the same time; and any mining system must have sufficient flexibility to permit loading from one face or another at short notice. This makes it impractical to place small mobile crushers and their consequent network of conveyors at each working face. To date no alternative has been found to using trucks for bringing all rock of the same class to some convenient marshalling point from which it may be crushed and conveyed from the pit. Another advantage of trucking when compared to conveying is flexibility. Many mines wish to deliver their shot-rock to different locations, depending on the mineral content. For instance, a typical copper mine may have to simultaneously mine milling ore, leach ore,

Jan 1, 1978

By C. F. Seaman, R. G. Clay

The mines of The Tennessee Copper Co. are in the Ducktown Basin, in southeastern Tennessee. The ore is a heavy sulphide consisting principally of chalcopyrite, pyrite and pyrrhotite and in places running as high as 95 per cent sulphur and iron. Operating mines are the Burra Burra and Eureka. Development ' of the Boyd ore body was started in 1940. Late in 1939 the decision was made to change from fuse and cap to all electric blasting and at the same time electric cap lamps were installed to replace carbide. These changes were made to improve the safety and working conditions in the mines. Both the Bureau of Mines and explosives manufacturers recommended electric blasting as a safety measure. Until that rime, electric blasting had been used only for shaft sinking and in places where it was absolutely essential for safety. Very few men had had experience in this type of blasting, as the places where it was required were those that also required direct supervision. It was necessary for the mining staff to work out methods, procedure, and equipment to be used. In this they were ably assisted by representatives of the explosives manufacturers. Many mistakes were made and it is the purpose of this paper to describe the original plan and installation, show the major difficulties encountered. the remedies applied, and present the practices now being followed. Comparison of costs of the two methods will also be discussed. About six months were required to make the change and eliminate fuse and cap blasting entirely. The first electric blasting stations were installed in November 1939 and the last fuse was taken underground in May 1940. A very unusual condition forced the use of fuse for a short time in one heading after May, but that will be discussed later. Original Equipment and Installation A power line is generally recommended as a source of current where permanent blasting stations are used. The plan worked out for electric blasting was to tap 220-volt alternating-current lines when they were available and use 250-volt direct current from the trolley wire at all other places. A blasting machine would be available at each shaft for places where there were no regular blasting stations and could be used if a regular station should fail. The equipment decided on was No. 16 duplex reinforced lamp cord for lines to within about 50 ft. of the working place and No. 20 lead wire from that point to the shots. An open double-pole, double-throw switch with slate base was put in the circuit, so connected that the down position would close the line to the shots 2nd the up position would connect it to the Iine from the power contactor. This switch was located in the circuit as close to the blast as was safe (should the shots explode when it was placed in the shooting

Jan 1, 1943

By Brooks J. Klimley, Daniel M. Higgins, Grover R. Castle

Measuring The Economic Viability Of Resource Projects (A Western Surface Coal Project) The bankers approach to the analysis of a new "green fields" mining project is similar but different from the approach that the pining company might take. The purpose of this paper shall be to discuss the techniques that are used by mining companies and bankers as they try to measure the economic viability of mining projects. The analysis of these new projects (by both the mining companies and the banker) is based on a projected revenue stream and not on historical earnings or past perfomance because these projects are start up situations with no history to look to. The starting point in the analysis of these projections is the choice of the assumptions to be used in preparing these projections. After the assumptions have been determined and the projections are prepared then the projections are submitted to various tests in order to determine whether the project measures up to previously agreed upon minimum standards. In order to illustrate the analytical procedures used in evaluating a mineral project the writer has constructed a hypothetical case involving the development of a western surface coal mine (Property). The basic data in connection with this hypothetical new project is described in the next section of this paper. After the Description of the Project is a section which shows the type of analysis that a mining company might prepare in order to approach a bank for a loan and the final section illustrates the type of analysis that a bank might prepare in evaluating this loan request. Description Of The Project or Basic Data In Connection With Undeveloped Western Surface Coal Property An independent mining/geological Consultant's report has been completed on the Property which confirms that the Property contains 175 million tons of economically recoverable coal subdivided into the following reserve categories in accordance with U.S. Bureau of Mines and U.S. Geological Survey definitions: Measured or Proven 90,000,000 Indicated or Probable 50,000,000 Inferred or Possible 35,000,000 The owners of the reserves have chosen to drill up those reserves that are nearest to the subcrop and have not spent as much time on those portions of the Property which are on the fringes of the deposit where the overburden is thicker because the stripping ratio and hence the operating costs and capital requirements are higher down dip. The proven reserves are based on an average thickness of 80 feet, of which a mining recovery of 90% is projected. The overburden thickness varies from 0' to 280' with an average thickness of 127'. The overburden consists mainly of scoria, sandstone, gravel, shale and silty clay. 134 holes were cored (and analysis done on five foot increments) in the measured reserve on the following spacing: 100 foot centers to define the burn line; 1,320 foot centers to define the early mining areas; 2,640 foot centers to identify the later mining areas.

Jan 1, 1985

By Ralph M. Hunter

Until recent years, maintenance of surface coal handling facilities was a relatively simple task. Equipment consisted principally of conveyors, screens and crushers of comparatively simple construction. Electrical equipment consisted of motors and starters, with little, if any, interlocking. In most cases failure of any component affected only the grades of coal loaded so that even a prolonged breakdown did not affect mine operation materially. However, with the introduction of modern cleaning plants, the picture has changed. The equipment is much more complex and therefore more difficult to maintain. In many cases failure of any component of the plant prevents loading of coal. Even if provision is made to bypass, the quality of coal loaded is lowered, affecting markets. In the case of plants designed to permit full seam mining, marketable coal cannot be loaded without operation of the cleaning plant. Under these conditions it is apparent that considerable thought and effort toward better maintenance are justified. In order to assure continuous, efficient and economical operation, maintenance should receive consideration, not only after the plant is in operation, but while it is being designed. It is hardly necessary to say that the designer should strive for sturdi-ness and wearing qualities. It is sometimes not realized, however, that even the best parts will sooner or later require replacement or repair. The designer should, there- fore, give consideration to accessibility. Anyone who has had to do with cleaning plant maintenance is familiar with installations where several hours are required to get at a part that can be repaired or replaced in a few minutes, once it is accessible. As in most cases, experience is the best guide. The men who operate and maintain this equipment know the problems encountered. Whenever available, the opinions of these men should be sought and heeded. As an example, we might consider the refuse elevator on a Chance Cone. Our company is now operating three of these cones, the first of which was installed in 1941, and the latest in 1946. Exact records are not available, but well over 50 pct of delays charged to the Chance Cones have been caused by the refuse elevator. This elevator receives refuse from the cone and carries it to a point above the water level of the cone, discharging onto a dewatering and desanding screen. Since the elevator operates under water and necessarily has close clearances in the lower boot, it tends to jam on large pieces of refuse or tramp iron. The chain, buckets and boot are also subjected to the action of water, sand and refuse, which accelerate normal wear. In case of jamming or chain-breakage it is necessary to drain the elevator and the cone so that even a minor fault in the elevator causes a considerable loss of time for the cone. In our new plant, which will contain two Chance Cones, the refuse elevator has been eliminated. This was accomplished by raising the cone so that the refuse gates could discharge directly onto the dewater-

Jan 1, 1949

By R. G. Clay, C. F. Seaman

The mines of The Tennessee Copper Co. are in the Ducktown Basin, in southeastern Tennessee. The ore is a heavy sulphide consisting principally of chalcopyrite, pyrite and pyrrhotite and in places running as high as 95 per cent sulphur and iron. Operating mines are the Burra Burra and Eureka. Development ' of the Boyd ore body was started in 1940. Late in 1939 the decision was made to change from fuse and cap to all electric blasting and at the same time electric cap lamps were installed to replace carbide. These changes were made to improve the safety and working conditions in the mines. Both the Bureau of Mines and explosives manufacturers recommended electric blasting as a safety measure. Until that rime, electric blasting had been used only for shaft sinking and in places where it was absolutely essential for safety. Very few men had had experience in this type of blasting, as the places where it was required were those that also required direct supervision. It was necessary for the mining staff to work out methods, procedure, and equipment to be used. In this they were ably assisted by representatives of the explosives manufacturers. Many mistakes were made and it is the purpose of this paper to describe the original plan and installation, show the major difficulties encountered. the remedies applied, and present the practices now being followed. Comparison of costs of the two methods will also be discussed. About six months were required to make the change and eliminate fuse and cap blasting entirely. The first electric blasting stations were installed in November 1939 and the last fuse was taken underground in May 1940. A very unusual condition forced the use of fuse for a short time in one heading after May, but that will be discussed later. Original Equipment and Installation A power line is generally recommended as a source of current where permanent blasting stations are used. The plan worked out for electric blasting was to tap 220-volt alternating-current lines when they were available and use 250-volt direct current from the trolley wire at all other places. A blasting machine would be available at each shaft for places where there were no regular blasting stations and could be used if a regular station should fail. The equipment decided on was No. 16 duplex reinforced lamp cord for lines to within about 50 ft. of the working place and No. 20 lead wire from that point to the shots. An open double-pole, double-throw switch with slate base was put in the circuit, so connected that the down position would close the line to the shots 2nd the up position would connect it to the Iine from the power contactor. This switch was located in the circuit as close to the blast as was safe (should the shots explode when it was placed in the shooting

Jan 1, 1943

By I. R. Kramer

[The final paper in the diffusion symposium was given by I. R. Kramer. The author called attention to the diffusion characteristics of chromium deposited on the surface of steel by treatment with chromium chloride and alloyed with the surface layer of the steel by thermal treatment. As the data cited were previously published in the paper " Chromizing of Steel," by I. R. Kramer and Robert H. Hafner (Metals Technology, October 1942, and in the Transactions, volume 154, 1943, pages 415-420), it is unnecessary to reproduce these data here. The following paragraphs are taken from the author's presentation.— Editor.] Aside from the well-known practical application of the diffusion phenomena to carburizing and calorizing, -diffusion also finds application in adding alloying elements to the surfaces of steel parts. The chromizing of steels is a process by which the surface of a low-alloy steel may be enriched with chromium to form a highly corrosion-resistant case; This chromized case differs from an electroplated coat in that it is a diffused layer integral with the base metal and has all of the properties of a high-chromium steel. Recently a method of chromizing has been developed in Germany by Daeves, Becker and Steinberg in which gaseous chromium chloride and hydrogen are passed over the steel at temperatures of 1650° to 1800°F. The chromium chloride reacts with the iron to give chromium and iron chloride. The chromium atoms are deposited on the surface and diffuse in as the iron atoms diffuse out, forming a layer rich in chromium. The principal role of hydrogen is to remove the carbon from. the surface of the steel. Of the two mechanisms, chemical reaction and diffusion, the one operating at the slower rate will determine the rate of formation of the layer. In this case, diffusion is slower and therefore any factors influencing its rate will affect the process. The rate of diffusion is influenced by the temperature and crystal structure of the parent metal. Depth-concentration studies show that the concentration falls off gradually for a considerable distance and then drops sharply at about 13 per cent chromium, the point coincident with the boundary of the gamma loop. The influence of the composition of the base metal upon diffusion is quite important. The depth of penetration of the chromium decreases as the carbon content increases and, of more importance, the chromized layer becomes brittle and is likely to spa11 if the carbon content is too high. It appears that in order to chromize successfully the carbon must be fixed by strong carbide-forming elements such as chromium or molybdenum, which decrease the diffusion of carbon. If this is not done, the carbon will diffuse from the interior faster than the chromium can diffuse inward, thus forming a carbide barrier. The corrosion resistance of the chromized layer is high, and these steels offer some unique advantages over other corrosion-resisting coating or alloys. While chromizing is not meant to replace stainless steels, it may serve in the saving of alloying elements. The chromizing process uses soft or low alloyed steels that can be easily hot or cold formed and since the chromized layer is soft and firmly adherent, chromized parts can be subjected to cold deformation without spalling or chipping and without loss of corrosion resistance.

Jan 1, 1944

By I. R. Kramer

[The final paper in the diffusion symposium was given by I. R. Kramer. The author called attention to the diffusion characteristics of chromium deposited on the surface of steel by treatment with chromium chloride and alloyed with the surface layer of the steel by thermal treatment. As the data cited were previously published in the paper " Chromizing of Steel," by I. R. Kramer and Robert H. Hafner (Metals Technology, October 1942, and in the Transactions, volume 154, 1943, pages 415-420), it is unnecessary to reproduce these data here. The following paragraphs are taken from the author's presentation.— Editor.] Aside from the well-known practical application of the diffusion phenomena to carburizing and calorizing, -diffusion also finds application in adding alloying elements to the surfaces of steel parts. The chromizing of steels is a process by which the surface of a low-alloy steel may be enriched with chromium to form a highly corrosion-resistant case; This chromized case differs from an electroplated coat in that it is a diffused layer integral with the base metal and has all of the properties of a high-chromium steel. Recently a method of chromizing has been developed in Germany by Daeves, Becker and Steinberg in which gaseous chromium chloride and hydrogen are passed over the steel at temperatures of 1650° to 1800°F. The chromium chloride reacts with the iron to give chromium and iron chloride. The chromium atoms are deposited on the surface and diffuse in as the iron atoms diffuse out, forming a layer rich in chromium. The principal role of hydrogen is to remove the carbon from. the surface of the steel. Of the two mechanisms, chemical reaction and diffusion, the one operating at the slower rate will determine the rate of formation of the layer. In this case, diffusion is slower and therefore any factors influencing its rate will affect the process. The rate of diffusion is influenced by the temperature and crystal structure of the parent metal. Depth-concentration studies show that the concentration falls off gradually for a considerable distance and then drops sharply at about 13 per cent chromium, the point coincident with the boundary of the gamma loop. The influence of the composition of the base metal upon diffusion is quite important. The depth of penetration of the chromium decreases as the carbon content increases and, of more importance, the chromized layer becomes brittle and is likely to spa11 if the carbon content is too high. It appears that in order to chromize successfully the carbon must be fixed by strong carbide-forming elements such as chromium or molybdenum, which decrease the diffusion of carbon. If this is not done, the carbon will diffuse from the interior faster than the chromium can diffuse inward, thus forming a carbide barrier. The corrosion resistance of the chromized layer is high, and these steels offer some unique advantages over other corrosion-resisting coating or alloys. While chromizing is not meant to replace stainless steels, it may serve in the saving of alloying elements. The chromizing process uses soft or low alloyed steels that can be easily hot or cold formed and since the chromized layer is soft and firmly adherent, chromized parts can be subjected to cold deformation without spalling or chipping and without loss of corrosion resistance.

Jan 1, 1944

By M. H. Caron

D. C. Ralston—The fact that none of the organizations that have worked on these ammoniacal leaching processes have contributed discussion of Mr. Caron's papers today is a matter of some disappointment. Adaptations, modifications and improvements have been made by several organizations or individuals and to complete the record I want to document a number of them. The Nicaro Nickel Co. operated a Government owned plant at Nicaro, Oriente, Cuba, during the past war. A report on the project was written for Reconstruction Finance Corp. and is in the document file collected by the Office of Technical Services, Commerce Department and filed with the Library of Congress, where it can be consulted or bibliofilm or photostat copies obtained by asking for copies of PB 97271. Nicaro Nickel Co. has obtained a series of U.S. Patents, my list (probably incomplete) of which bears the following numbers: 2,400,098; 2,400,114; 2,400,115; 2,400,461; 2,400,612; 2,458,902 and 2,473,795. International Nickel Co. has also studied the problem and U.S. Patent 2,478,942 shows that some of the complications involving occasional low extractions have been unraveled. Forward, Samis, and Kudryk, of the University of British Columbia, have also published a study of copper-nickel sulphide concentrate.' Others are interested and it is evident that this type of hydrometallurgy is likely to find a happy application in the not-too-far distant future. Caron's two papers practically amount to a summation of his life's work on the important subject of nickeliferous iron ores and the American Institute of Mining and Metallurgical Engineers is fortunate to have been chosen the medium of publication. M. Rey—The ammonia leaching process has been the subject of certain studies in New Caledonia, but at the present time, the conditions are not very favorable to its adoption. The process is a combination of a thermal and a hydrometallurgical process. As a result, the first costs and the operative costs are rather high. To counteract this unfavorable situation, it would be necessary to make a very high extraction. But unfortunately, the process is delicate and in spite of all of Caron's valuable work, there are still many unknown factors. It is difficult to obtain regularly a high extraction, exceeding 75 to 80 pct. It would be interesting to have the opinion of Inter-national Nickel on the process and to know what the Dutch are doing in Celebes. M. H. Caron (author's reply)—I am quite familiar with the different types of nickel and cobalt ores as found on New Caledonia, and it may be pointed out that the Celebes ores are quite the same in their properties. Their behavior when treating the ores by the ammonia-leaching process is also alike and in my opinion, there is no reason why a high extraction of the New Caledonian ores could not be obtained if the process is properly applied. For the sake of completeness I may add that one of my patents has not been discussed in the papers. This patent deals particularly with the treatment of serpentine nickel ores. (Netherlands No. 62693 40a, French No. 965.786 April 21. 1948. and Cuban No. 13.625 Jan. 1950.)

Jan 1, 1951

By R. C. Gifkins

R. C. Gifkins (CSIRO)—In this paper evidence is put forward to support the idea of grain boundary shearing in aluminum at 4.2°K and the phenomenon is explained in terms of a low-temperature "equicohesive temperature". It is stated that these findings show that grain boundary shear cannot be ruled out in any situation by temperature arguments alone. We believe, from our own recent study of "apparent grain boundary sliding'' at temperatures below 0.45 Tm, that it can be ruled out on temperature arguments alone.45 Our experiments were made principally with magnesium and a Ms-A1 alloy ("Magnox A12"). Careful examination of offsets of marker lines or steps at grain boundaries produced by deformation at room temperature (-0.3 Tm) showed many of these to result from curvature over narrow zones along the grain boundaries. A significant residuum did, however, appear as a sharp offset or step. It was found that the mean value of this step rose to a small value during the first 0.5 to 1 pct strain and then remained at this level with further strain. Moreover, exactly the same values were obtained (at the same stress) at 75°C instead of 21°C. That is, in this range of temperature the "apparent sliding" was not temperature-sensitive. At 125°C and above there was sliding which increased with strain and which was temperature-dependent. When this temperature-dependent sliding was used to calculate the sliding to be expected at room temperature, this underestimated the "apparent sliding" by a factor of 104. The conclusion drawn is that the "apparent sliding" consists of shear localized in a zone of a width measured perhaps in microns, whereas it seems that true sliding occurs on the boundary surface.46 A model for such localized shear has been developed in general terms, following a similar model of Urie and Wain.47 Other strong evidence for low-temper- ature sliding has been based on the offsets of marker lines consisting of a substructure seen using the electron microscope to resolve these.48 We have repeated this work and found that the effects can be accounted for in terms of cracks in an oxide film which is produced when the substructure is etched up; these cracks form over zones of shear rather than at actual interfaces. However, in aluminum at room temperature (-0.3 Tm) there is no "apparent sliding" which is optically detectable. We suggest that, when the temperature is much lower (and possibly the rate of strain higher, too), as in the experiments of Chin et al., then aluminum behaves like magnesium at room temperature, and a narrow zone of shear develops to accommodate differences in strain across the grain boundary. Such a zone could well give rise to triple-point cracks. It is hoped to investigate these suggestions further. B. Y. Chin, W. F. Hosford, Jr., and W. A. Backofen (authors' reply)—It is interesting to have Dr. Gif-kin's comments, although it is somewhat difficult to reply in detail without the reference containing the work that he describes. Apparently though, he does find a low-temperature temperature-independent sliding that would seem to be related to the observations reported here. Perhaps some of the difficulty has semantic origins, growing out of the distinction between ('true sliding" and "apparent sliding".

Jan 1, 1965

By H. H. Kellogg

0. C. Ralston—The fact that none of the organizations that have worked on these ammoniacal leaching processes have contributed discussion of Mr. Caron's papers today is a matter of some disappointment. Adaptations, modifications and improvements have been made by several organizations or individuals and to complete the record I want to document a number of them. The Nicaro Nickel Co. operated a Government owned plant at Nicaro, Oriente, Cuba, during the past war. A report on the project was written for Reconstruction Finance Corp. and is in the document file collected by the Office of Technical Services, Commerce Department and filed with the Library of Congress, where it can be consulted or bibliofilm or photostat copies obtained by asking for copies of PB 97271. Nicaro Nickel Co. has obtained a series of U.S. Patents, my list (probably incomplete) of which bears the following numbers: 2,400,098; 2,400,114; 2,400,115; 2,400,461; 2,400,612; 2,458,902 and 2,473,795. International Nickel Co. has also studied the problem and U.S. Patent 2,478,942 shows that some of the complications involving occasional low extractions have been unraveled. Forward, Samis, and Kudryk, of the University of British Columbia, have also published a study of copper-nickel sulphide concentrate.' Others are interested and it is evident that this type of hydrometallurgy is likely to find a happy application in the not-too-far distant future. Caron's two papers practically amount to a summation of his life's work on the important subject of nickeliferous iron ores and the American Institute of Mining and Metallurgical Engineers is fortunate to have been chosen the medium of publication. M. Rey—The ammonia leaching process has been the subject of certain studies in New Caledonia, but at the present time, the conditions are not very favorable to its adoption. The process is a combination of a thermal and a hydrometallurgical process. As a result, the first costs and the operative costs are rather high. To counteract this unfavorable situation, it would be necessary to make a very high extraction. But unfortunately, the process is delicate and in spite of all of Caron's valuable work, there are still many unknown factors. It is difficult to obtain regularly a high extraction, exceeding 75 to 80 pct. It would be interesting to have the opinion of Inter-national Nickel on the process and to know what the Dutch are doing in Celebes. M. H. Caron (author's reply)—I am quite familiar with the different types of nickel and cobalt ores as found on New Caledonia, and it may be pointed out that the Celebes ores are quite the same in their properties. Their behavior when treating the ores by the ammonia-leaching process is also alike and in my opinion, there is no reason why a high extraction of the New Caledonian ores could not be obtained if the process is properly applied. For the sake of completeness I may add that one of my patents has not been discussed in the papers. This patent deals particularly with the treatment of serpentine nickel ores. (Netherlands No. 62693 40a, French No. 965.786 April 21. 1948. and Cuban No. 13.625 Jan. 1950.)

Jan 1, 1951

By Bunting S. Crocker

Bunting S. Crocker (Executive Vice President, Kilborn Engineering Ltd., Toronto, Ont., Canada)-This is an interesting report on laboratory-scale testing with pebble charges in 12-in.-diam mills, and the author correctly concludes that useful laboratory testing can be done. In the opening paragraphs he states, "laboratory pebble-milling tests have been considered too difficult to perform and not relevant enough to the actual grinding situation"!! This is definitely not so! The writer published detailed tests done in a 12-in.-diam mill comparing pebble and steel balls in 1952 6 and again in 1954 7. We note that our review paper in 1963 was the only one mentioned in the References. The author states that "a smooth lining is more frequently used in commercial-scale pebble grinding." The writer has designed some 22 plants treating 134,000 tpd in North America and Australia. He has also visited pebble installations in Norway, Sweden, Finland, and South Africa. All these plants correctly use heavier lifters than would be considered standard for ball mills of a similar size. Smooth liners are not recommended in wet ore treating plants. In the 1952 paper6 it was clearly demonstrated in the test work that the individual ore pebbles had to be the same weight-not the same diameter or size-as the steel balls if comparable grinding results were to be obtained. In the tests performed in this study most of the pebbles "passed 50-mm round hole screen openings and remained on a 40-mm round hole sieve. Correspondingly most of the ball mill experiments were performed with 50 or 40-mm steel balls". Thus, the balls and pebbles were essentially the same size or diameter. This means we are comparing apples and oranges. When the feed being ground is fine enough the pebbles will be superior. When the feed is coarse the steel balls will be the better. This is similar to grinding tests in which two sizes of steel balls are compared. In the middle of the article, "Crocker5 has reported that equal weights of pebbles and steel balls achieve the same grinding capacities under similar conditions …", then he proceeds to question this postulate. In the eight reports that I have published, I have never made this statement. What I have said is "for comparable grinding, the individual pebbles in the mill charge should be the same weight as the individual steel balls." This is not the first time that I have been misquoted on this point! It is most unfortunate that this test work has been done with a ball and pebble of the same diameter and not the same weight. It is impossible to properly evaluate the results. I also regret the fact that the screen analyses of the tests were not shown. Many of us who have been 40 years or less in the grinding game can easily read screen analyses at a glance. We can tell whether the top meshes are being ground by the ball charge, etc., etc. The graphs are fine for the summary or to illustrate conclusions but are often misleading in the report itself. However, we do agree that useful pebble grinding tests can be done in a 12-in. mill.

Jan 1, 1973

R. E. Powers (Mellon Institute of lndustrial Research, University of Pittsburgh, Pittsburgh)—Several of the processing methods at the Josephtown smelter might well be considered by plants producing iron blast furnace sinter. In making blast furnace sinter the objectives are much the same as at Josephtown: To produce a strong yet porous agglomerate which has a size consist in a preferred range. The emphasis however is on production rate rather than on the size or other quality characteristics of the sinter. Therefore the technique is entirely different. Blast furnace sinter plants attempt to make their product in one pass with a minimum recirculating load and without sizing treatment after the sinter leaves the machines. Many plants are beginning to feel that more attention paid to quality characteristics would make the product a more desirable blast furnace material. The pelletizing of very fine material before placing it on the bed is not deliberately done at iron sinter plants, although the installed equipment sometimes gives a simila; action. Such pelletizing seems to be a promising way of accommodating the down-draft sintering Drocess to very fine materials. he recirculating load at Josephtown is extremely high. Two plants' producing blast furnace sinter from raw material of comparable size have a recirculating load of about 100 pct of production. At Josephtown it is notable that the size and the rate of feeding of return fines are under control. This is not the case in most blast furnace sinter plants. The sinter slicer appears to be a worthwhile aid in effecting such control. It is similar in action to rakes and plows used at a number of plants, but it has the advantage that the sinter dislodged is positively removed and not left to the inefficient action of a stationary grizzly screen. Blast furnace size specifications for sinter are not at present so rigid as the Josephtown specifications. If the sizing of blast furnace sinter does become important, then the Josephtown crushing practice, in which a large recirculating load is maintained on the crusher in order that minimum fines be produced, may be of interest. H. K. Najarian (authors' reply)—We note that Mr. Powers appreciates some points that we tried to emphasize in our paper as being of importance in making zinc sinter of good quality and which may well be applied to iron blast furnace sinter. Fig. 2 shows the pelletized nature of the feed going on to a Dwight-Lloyd type sintering machine in our plant. Practically, the entire feed consists of more ,or less rounded particles which permit more uniform air passage through the charge, while the larger particles in the charge roll on to the surface of the grate, which helps to keep the grate openings from being clogged by excessive fines. An agglomerated charge of this character may preferably be made by rolling action while being mixed in a rotating drum and wetted by water sprays to bind the fine particles together and give the charge the desired moisture content. From our observations of several iron ore and blast furnace flue dust sintering operations, we believe that if the fines from the sinter discharged from the machines were to be recirculated and agqlomerated with the new feed, not only sinter of better quality but larger capacity from the machines may be realized.

Jan 1, 1952

By Luis Toron, Salvador Cortes-Obregon

John D. Price (Colorado Fuel and Iron Corp., Pueblo, Colo)—The paper on the coal fields of the Oaxaca district as prepared by engineers Toron and Cortes-Obregon of the staff of the Bank of Mexico bears witness to the thorough and careful way in which the men associated with this organization perform their work. There is little to be added to their paper in way of discussion other than to confirm and amplify some of their statements. Since the only extensive and well-developed field of coking coal lies in the northeastern section of the country adjacent to Sabinas in the state of Coahuila, it follows that blast furnace plants would be located in that same region. Two such plants are now operating at Monterrey and Monclova, using coke produced at the Sabinas district mines. But the nearer of these two plants is 600 miles from Mexico City and even farther from the center of population. Transportation of products from these mills to the market area is therefore expensive, both because of the distance and the difficulty of the terrain over which it must be carried. The development of an integrated steel industry closer to the center of population has therefore long been a goal toward which the Mexican technicians have been striving. While the presence of coal of some grade has been reported in many of the states, and many ideas have been advanced regarding its possible uses in iron and steel production, deposits of anthracite in Sonora and the various coals of the Oaxaca district as reported on in this paper are the only ones that have been explored in a serious manner. The coking coal from the Mix-tepec zone appears to offer promise of producing a coke which could be used in a standard blast furnace. Several problems are indicated, however: 1—The ash in the coal is high as mined, but indications are that it can be washed to an ash content of 15 pct with a recovery of 70 pct of washed coal. 2—Such washing would increase the volatile content from 17.4 pct to about 20 pct, and in a byproduct oven this should give a coke yield of close to 80 pct with an ash content of coke under 20 pct. 3—A free swelling index of 5 appears low for a good coking coal, and below that of the coals from the Sabinas district, which show between 6 and 9. But washing of the coal should result in an improvement in this regard; in the United States coals from Utah with an index even lower than 5 have made a usable coke. 4—A coal with volatile as low as 17.4 in raw coal and 20 in washed coal would come close to being classed as a low-volatile rather than medium-volatile coal, and low-volatile coals are notorious for their high expansion properties. Several plants in the United States are making coke from straight medium-volatile coal of 26 to 28 volatile content, and one at Rosita,, Mexico, from coal of 25 volatile. But no plants to my knowledge are using coal as low as 20 volatile. Since the Rosita coal appears to be a borderline coal from the angle of its expansion properties the coking of one of the straight lower volatile must be approached with caution. 5—There are few coals possessing any degree of coking properties which cannot be used in coke production by careful attention to its preparation and blending. The fact that coals of other types are available in this same region make improvement through blending very possible. 6—There are other workable methods of reducing iron ore other than the conventional coke-blast-furnace method. These will not be discussed here but it is known that their use has been considered. The technicians of not only Mexico but also of the other Latin American countries are keenly aware of their natural resources and their national needs. This paper emphasizes the fact that the Mexican technicians are working on their problem and attempting to speed the day of self-sufficiency for their country. Salvatore Cortes-Obregon (author's reply)—I wish to thank Mr. Price for his kind remarks. The Mixtepec coal as shown in Table II has 30 pct ash and a free swelling index of 5, but when the same coal is washed to 15 pct ash it has a free swelling index of 8 to 9 and the volatiles increased from 17.4 to 20.7 pct. A satisfactory coke has been produced from blends made in the Mexican laboratory using at least 40 pct of the Mixtepec coking coals with the other Oaxaca non-coking coals. Koppers in Germany report good coke obtained from the Oaxaca coal with a blend of 80 pct Mixtepec coal. Consideration is being given the possibility of using methods other than the conventional blast furnace for the reduction of iron ore near the Oaxaca area; electric furnaces appear promising. The non-coking coals could be used to produce cheap electric energy and the coking coals to make metallurgical coke.

Jan 1, 1955

By Carl Wagner

This paper contains solutions of the differential equation of diffusion in binary alloys if the diffusion coefficient is an exponential function of the concentration of one of the components. THE general differential equation for one-dimensional diffusion in a binary system can be written as: where c is the concentration of one of the two components in any suitable units; x, the distance from a reference plane; t, time; and D, the diffusion coefficient. Hartley and Crank&apos; have pointed out that eq 1 with conventional length units for measuring the distance x holds only if the alloy can be regarded as a rigid framework. In general, this is not the case, especially if the lattice parameter varies appreciably with composition. If a modified coordinate system is used instead of an orthodox system based on the centimeter a:: length unit, eq 52 is obtained, which has the same general form as eq 1 and is generally valid. The necessary transformations are presented in the appendix. If&apos; D is independent of concentration, analytical solutions of eq 1 are known for numerous boundary conditions. If D depends on concentration, solutions based on known analytical functions are not available. Among others, Boltzmann2 and Matano3 have shown that the concentration c depends only on (x/t1/2) in the case of diffusion between two semi- infinite spaces with uniform initial concentrations: c = c, for x < 0, t = 0, i.e., x/tl/2 = — x [21] c = c2 for x > 0, t = 0, i.e., x/tl/2 = + a Moreover, Boltzmann and Matano have shown how D may be calculated as a function of concentration over the whole range of composition from c -c, to c = c2 if c as a function of (x/t1/2) for the initial conditions stated in eq 2 has been determined experimentally. Thus, for instance, in principle, the diffusion coefficient in silver-gold alloys may be calculated as a function of concentration for the whole range of composition from only one diffusion experiment with a couple of pure silver and pure gold welded together. If the diffusion coefficient D as a function of concentration c is known, a solution of eq 1 for boundary conditions as stated in eq 2 with arbitrarily chosen values of c, and C2 may be wanted. For instance, if the diffusion coefficient in an alloy A-B as a function of concentration has been derived from an experiment with a diffusion couple involving 0 and 30 pct B as initial. concentrations, the calculation of the concentration distribution in a diffusion couple involving other initial concentrations, e.g., 5 and 20 pct B, may be desired. This problem may be solved by numerical or graphical integration of eq l. The solution of an ordinary differential equation of second order, which results from using (x/t1/2) as an independent variable, can be computed quite rapidly if two conditions for the same value of the independent variable are given. In eq 2, however, two conditions are given for two

Jan 1, 1953

By T. W. Healy, P. Somasundaran, D. W. Fuerstenau

The flotation behavior of quartz in the presence of alkyl ammonium acetates as a function of alkyl chain length has been interpreted in terms of hemi-micelle formation at the solid-liquid interface. The van der Waals cohesive free energy that is responsible for hemi-micelle formation has been found from these results to be 1.0 kT per CH2 group. This is in excellent agreement with literature values obtained from research on the properties of micelles in bulk systems. Collectors of the type CH3(CH2)n-2CH2-P, where P is a polar group and n is the number of carbon atoms in the straight alkyl chain, will induce flotation of an appropriate mineral at lower concentrations as n increases. This reduction in collector concentration with increasing n has been reported for such systems as xanthates on sulfides1,2 and amines on quartz.&apos; The effect of chain length of alkyl surfactants on such solution properties as conductivity,&apos; surface tension,&apos; and solubility6 and such interfacial properties of solids as contact angle7 and electrokinetic potential8 have received considerable attention, but only limited quantitative descriptions of the effects have been given. The electrokinetic studies of alkyl amine salts on quartz indicated that the solution concentration of collector at which the marked change in electrokinetic potential of quartz occurs decreases systematically as the length of the alkyl chain increases. The aim of the present work was to obtain a quantitative description in terms of current theories of flotation of the effect of the hydrocarbon chain of alkyl ammonium acetates on the flotation behavior of quartz. BASIC PRINCIPLES The marked changes in electrokinetic potentials of quartz in dodecylammonium acetate (DAA) solutions were postulated to result from the association of adsorbed collector ions into two-dimensional aggregates, called hemi-micelles.9 Subsequently, the rapid increase in flotation rate of quartz was shown to coincide with hemi-micelle formation." We would like to discuss briefly the concept of hemi-micelles in terms of recent interpretations of micelles in bulk solution. Micelle hypothesis: In dilute solutions, alkyl ammonium and related salts behave as ordinary strong electrolytes, but at a certain concentration there is a marked change in physical properties of the solutions, e.g., in equivalent conductance, transport number, freezing point lowering and viscosity. To account for these phenomena, McBainl1 introduced the concept of micelles, which are aggregates of long-chain ions. It is believed that the ionic heads of the constituent ions of the micelle are in contact with water, whereas the nonpolar groups turn away from it and are in contact with each other. Work must be done to bring these charged polar groups together, but energy is gained (i.e., so called van der Waals cohesive energy) when hydrocarbon chains are expelled from the water and allowed to associate. For a given polar group, the tendency to form micelles will therefore strongly depend on the hydrocarbon chain length. Theoretical treatments of micelle formation by Shinoda12 and Phillips13 and others lead to the following expression for the concentration at which micelles form (the critical micelle concentration or cmc) in aqueous solutions of long- chain electrolytes

Jan 1, 1964

By W. Seymour

Uranium was cold rolled to a reduction in thickness of 90 pct and the preferred orientation of the grains was determined from X-ray intensity data. Complete pole figures for a large number of atom planes were plotted, and the preferred orientation was found to be duplex (102) [0101 + (012) [021]. These preferred orientations are compared with those reported for many of the hexagonal-close-packed metals. THE preferred orientation of a highly reduced cold-rolled uranium strip was determined in order to obtain additional experimental data which would be useful in further understanding the deformation mechanisms of uranium. The preferred orientations reported herein are not in agreement with those theoretically predicted by Calnan and Clews&apos; for cold-rolled uranium sheets, and differ somewhat from those reported from experiments by Fisher, and Burke and Turkalo. Such variations might be accounted for by differences in the fabrication temperature and the reductions. Experimental Procedure The uranium used in this investigation had been hot rolled to 0.060 in. at 575°C from 0.393 in., and from a 0.060 to 0.020 in. thick sheet at 275°C. From this point the sheet was cold reduced 90 pct to 0.002 in. at room temperature. Three specimens were used to obtain complete pole-figure data. The first was a 1/2 in. square section illuminated by the X-ray beam on the rolling surface. The other two were l/2 in. square longitudinal and transverse sections each made as follows: Approximately 200 1/2 in. square sections of the foil were piled one on top of the other, so that the rolling direction for each piece was kept coincident. A 1/2 in. square by 1/8 in. thick steel section was placed on the top and bottom of a pile and the entire assembly was then bolted tightly together through a 1/8 in. diam hole drilled down through the center. Each pile was then surface-ground from the two proper sides to sections 1/2 in. square by 3/16 in. thick. These specimens were then polished flat and electro-polished to remove 0.005 in. of the surfaces to be examined. It can be shown that uranium is so opaque to a Cu Ka X-ray beam that only a small percentage would be transmitted by one micron of uranium. Therefore, a method of measuring the X-ray intensities diffracted from specimen surfaces was used. The specimens were mounted in a modified Schulz reflection holder&apos; which automatically changed the position of the specimen at a constant rate. The holder, in turn, was mounted on a General Electric XRD-3 Spectro-Goniometer and an automatic record made of the diffracted X-ray intensity vs specimen position. While a specimen moved automatically through 10" inclination, at a rate of l° per min, it simultaneously rotated 360" about a normal to the illuminated surface. (The inclination and rotation angles are referred to as F and a, respectively.) The data were then plotted along the resulting spiral path on a stereographic projection, Fig. 1. Because of defocusing effects on the diffracted beam which occur when diffracting surfaces are in-

Jan 1, 1955

By J. C. Williams

THE operations described in this paper are those of the flotation plant of the Valley Forge Cement Co. at Conshohocken, Pa., and concern the use of a particular cationic reagent-namely, DP-243. The role of this reagent in this plant may be better understood by first considering the purpose of beneficiating cement rock in general and the specific problems at the Valley Forge plant. The ultimate purpose of beneficiating cement rock is to obtain a kiln feed of the desired chemical composition in a form suitable for the most efficient kiln operation. The chemical composition is governed by the type of cement to be made. Before government restrictions (the purpose of which is to simplify operations during wartime) limited the types of portland cement a manufacturer is permitted to produce, five principal types of cement were generally recognized by the industry. They are based upon chemical and physical specifications proposed by the American Society for Testing Materials, federal and other governmental agencies, state highway departments, water boards and other users of large quantities of cement. The physical as well as chemical specifications for any particular type of cement necessitate proper proportioning of the major components-that is, the silica, alumina, iron and lime-and it is necessary to establish and maintain definite limits for each in the raw mix. The manufacturer must then prepare a kiln feed of a composition that will yield a final product that will fulfill the desired specification. This step is further complicated for kilns using coal as fuel, by the amount and analysis of the ash introduced. The amount and nature of the dust lost during burning will also affect the composition of the kiln feed. The problem of preparing a kiln feed of the desired chemical composition is not a simple one, and it is the final product that governs the methods used to obtain such a feed. For example, in starting with a material deficient in lime for the mix sought, it may not be sufficient to raise it to the required lime content. The other constituents must also be present in the proper proportions. In some cases, it may be necessary to raise the lime content to a point greatly exceeding that of the final mix, then, by addition of one or more other materials lower in lime, obtain the mix. The rock used at the Valley Forge plant is a metamorphosed argillaceous limestone of the Jacksonburg formation, the same formation as that of the Lehigh Valley cement rock (which occurs approximately 50 miles north), and its oxide composition is similar to that of the latter. The only practical differences between the two are that, as a result of metamorphosis, crystallization of the Valley Forge rock is considerably coarser and it contains less carbon than the Lehigh Valley rock as well as slightly more magnesia in the form of magnesian mica. The calcium carbonate content of the various strata in the Valley Forge quarry ranges from about 65 per cent to over 85 per cent, the average for the entire quarry being somewhat under the requirement for manufacture of cement, which usually is from 75 to 76 per

Jan 1, 1946