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By M. E. Lombardi

ThE Coalinga oil field is located on the west side of the San Joaquin Valley, California. The structure is in general a monocline, the edges of the oil horizon resting on the foot hills and dipping gently toward the east. One prominent anticline occurs plunging southeast. The earlier drilling was done in the foot hills comparatively near the outcrop and the wells were shallow. The sands were followed eastward and, in the case of the anticline, along the plunge, the wells becoming deeper and deeper until the depth of 4,000 ft. was reached and passed. There is nothing to show that the oil will not be found in quantity at still greater depth. In fact, some of the best producers have tapped the sand at close to 4,000 ft. The recovery of oil still farther to the east, and therefore at greater depth, seems to be mainly a question of drilling. In this territory the formations drilled through are chiefly sands and shales; they will not "stand up" in an open drilling hole; the casing has to be carried close to the bit, and it is always difficult to keep the casing free for any considerable distance. Ability to carry casing of comparatively large diameter without conductor pipes for distances of 2,000 or 3,000 ft. or over is desirable in such territory chiefly for two reasons. It makes it possible to enter the oil sand with a pipe of ample diameter; it eliminates one or more expensive strings of casing which act only as conductors for the water string, and furthermore, in territory where waters are encountered which corrode steel rapidly, it makes possible the construction of a rust- and alkali-resisting water string. It is always desirable to shut off top waters, which may lie within 100 ft. or less of the oil sand, with 10-in. pipe. Where the depth is so great that a practical weight of 10-in. pipe will not withstand the probable collapsing pressures, 81/4 in. at least is desirable. About the limit of rotary drilling to date in California seems to be the setting of the 10-in. string at 3,200 ft., although the rapid advance in rotary work during the past year seems to indicate that this depth may soon be increased. It is my purpose now, however, to treat only of cable-tool drilling. The problem is to reach a depth of 4,000 ft. or more with a string of

Jan 1, 1915

Discussion, at the Virginia Beach Meeting, February, 1894, of the Paper of Prof. Posepny. (Trans., xxiii., 197, 587.) Including communications subsequently received. a T. A. Rickard, Denver, Colorado (communication to the Secretary) : The paper of Professor Posepny was printed so short a time

Jan 1, 1895

By William A. Johnson

The fundamental role of diffusion in many reactions occurring in solid metals has long been recognized, and there have been careful measurements of rates of diffusion in numerous alloy systems; but our understanding of the atomic mechanism by which diffusion occurs is scarcely in an advanced state.l-3 It is doubtful that this mechanism can ever be directly ascertained; rather, all imaginable mechanisms must be examined and the one that is in best accord with such data as can be obtained from suitably chosen experiments should be selected. Because of its relative simplicity, diffusion in a system containing only one kind of atomic species—commody called self-diffusion—is of considerable theoretical interest. It may be expected that diffusion theory will not be advanced appreciably until a wide variety of diffusion experiments—of which diffusion in a chemical concentration gradient and self-diffusion are two important types—are carefully performed and inter-compared. Aside from the purely theoretical interest in self-diffusion measurements, such studies will be of considerable engineering importance. The rate of any process in a pure metal requiring the movement of atoms, of which recrystallization and creep are two familiar examples, will be understood more readily with the aid of self-diffusion data.† Self-diffusion implies the diffusion of something into its exact counterpart; the detection of such a process is patently impossible, and it is necessary to substitute for it the diffusion of one isotope (in this case radioactive) into another isotope (in this case stable) of the same chemical element. There is, however, no reason to suppose that a radioactive atom will not diffuse at the same rate as its stable isotope, for all of the isotopes of the same element have exactly the same chemical properties; neither the difference in nuclear energy nor the difference in mass would be expected to exercise an important influence in this phenomenon. In this sense, self-diffusion may be thought of as the diffusion of radioactive atoms down a radioactivity concentration gradient, just as usual diffusion is the diffusion of one kind of atom down its chemical concentration gradient. The first measurements of self-difiusion were made by von Hevesy4-6 using the radioactive lead isotope thorium B in ordinary lead. The recent development of means for producing artificial radioactive elements has made possible similar measurements on gold,7, 8 copper,2, 9 and zinc.10, 11 Many of the data that have been obtained are rather poor, and no adequate reason has been given for their poorness; it appears, however, that the unsatisfactory results are characteristic of the particular experimental procedure employed, and not of self-diffusion measurements in themselves. The experimental technique that has yielded the poorer results is the deposition of a thin layer of radioactive atoms on the surface of a specimen and the determination of the rate of inward diffusion

Jan 1, 1941

By John C. McDonald

This report is a continuation of an earlier one with a similar titlell to which the reader is referred for such details of procedure as do not appear here. A brief summary will be given of the objects and methods of the research at this time. It was pointed out in the previous paper that properties of magnesium alloys, as reported by different workers, often differ. This was thought to be due to differences in the preparation of the specimens tested. The ideal way to compare alloys would be to prepare each of them in so many ways that reasonable assurance would exist that the best properties of each alloy had been developed. The alloys would then be compared on the basis of these properties. The treatment actually given the alloys of these studies is necessarily a compromise with completeness. Nevertheless, enough treatments have been given so that a considerable degree of assurance exists that the properties obtained are representative of the alloy as such, and not merely of the manner in which it was worked and heat-treated. Procedure The alloys were cast in the form of cylindrical ingots, which were extruded into bar before being rolled. Some of the ingots were made in 2-in. diameter molds whereas all of the alloys of the previous paper were cast in 4.in. molds. For the larger ingots, the extruded bar was 2 by 0.140 in.; lor the 2-in. ingots, it was I by 0.100 in. Preliminary work indicated that strip rolled from both sizes of the bars would have substantially the same properties. The bar was rolled at some temperature in the range 400' to 800°F., for a total reduction of 50 per cent. The temperature and reduction per pass were so selected that no cold cracking occurred, but were such as to leave the metal in a semihard condition. The hot-rolled metal was annealed at several temperatures in the range 400' to 800°F.; it was also cold-rolled, and given solution and aging treatments. Two standard test bars corresponding to each treatment were pulled in tension, and the average was obtained. Twelve treatments, in all, were given each alloy. The properties of the annealed specimen having the highest percentage of elongation were selected to represent the alloy in the annealed condition. The properties of the cold-rolled specimen having the highest yield strength were selected to represent the alloy in the cold-rolled condition. These properties were then plotted against percentage composition by analysis. Smooth curves were drawn among the points, thus graphically averaging the results over the entire range studied. The percentage of elongation in the cold-rolled state was not plotted, since it was always low, and varied in an irregular manner with composition. Alloys Studied Magnesium-calcium Alloys.—Fig. I presents the results on the alloys of the rnag-

Jan 1, 1941

By H. Brevers, Maurice Rey

Through circumstances connected with the war, the senior author lost his records, therefore it has been impossible to include numerical data in this paper. The arsenical quicksilver ore investigated had been roasted, with poor metallurgical results. The vapor-tension temperature curves of mercury and arsenious trioxide are similar, and the arsenical compound hampers the coalescence of mercury drops in the condenser and is the cause of large stack losses. The ore was a gray limestone carrying small amounts of carbon and mineralized with cinnabar and two arsenic minerals, realgar (As2S2) and arsenopyrite (FeAsS). The proportion of realgar often exceeded that of cinnabar; the proportion of arsenopyrite was smaller. The grade of the ore varied between 0.8 and 4 per cent Hg. The problem was to make a high-grade mercury concentrate, as free from arsenic as possible. The water used throughout the testing was hard, having a pH of about 8. Flotation of Realgar Realgar showed a very strong tendency to float. It formed a film at the surface of the pulp and floated in abundance with only pine oil. Therefore an attempt was made to float the realgar before the cinnabar, but this was not very successful; the percentage of mercury in the realgar concentrate could never be brought below 2 or 3 per cent. Cinnabar slimed easily and floated with pine oil, so that some of it always accompanied the realgar. There was also the question of the cinnabar-arseno-pyrite separation. On further investigation it was found that realgar can be floated not only with pine oil but even better with a combination of hydrocarbons and pine oil. Washing the ore with ether revealed that it contained hydrocarbons, which helped to float the realgar. In this connection it belongs to the same class as carbon, sulphur and talc. It can be floated also with a xanthate, when activated by a heavy-metal salt. The conditions favoring the two types of flotation are opposed, as activation by a copper salt facilitates xanthate flotation (in fact, is even necessary to it), whereas it hinders pine oil and petroleum flotation. One type of flotation can be termed non-polar and the other polar, a distinction that has been stressed in a previous paper.' Depression of Realgar and Flotation of Cinnabar Many chemicals were tested in an attempt to float the cinnabar and depress the realgar. Finally starch rendered soluble with caustic soda was tried, also several grades of commercial soluble starch, and dextrin was found to be the best and the most convenient reagent. Several of these starch preparations have a strong flocculating effect on the pulp, which is a distinct disadvantage. Dextrin, with its low flocculating power, is excellent. It is noteworthy that starch preparations are also depressants for carbon and talc, so that the fact apparently has a general

Jan 1, 1943

By H. C. Ray

ABOUT the beginning of 1930, coal preparation had reached the stage where the cleaning of coal by liquid or air had become the order of the day. Since that time many new preparation plants have been installed in spite of the years of depression in the coal industry, and this is indicative of a trend wherein the coal producer found it expedient to manufacture a cleaner product in order to compete with other fuels. It was also about 1930 that, because of the difficulties encountered with fine coal in cleaning plants, dust collectors came into use. Both cleaning plants and dust collectors, therefore, were much discussed at the engineering meetings of the early 1930's. During these meetings the many advantages to be gained through the use of dust-collecting systems were pointed out by Hebley,1 Appleyard2 and others. The great possibilities of new markets for this product were emphasized and discussed at the same time; however, the disposal of the collected dust remained a problem. Since the purpose of dust collecting is to improve the preparation and, consequently, the salability of the coal, it seems illogical to return the collected dust to any of the coal being prepared for market use, as is done in preparation plants that include dust-collecting systems. The logical thing is to dispose of the dust in some useful and practical manner outside of the tipple. As a practical matter, this can be done in one or both of two ways; either by consuming the dust at the mine's own power plant or by marketing the dust direct to the consumer. The latter method not only offers the greatest opportunities, but is the most important, because of inequalities that may exist between the quantity of dust collected and the consuming capacity of the mine's power plant. In considering the more important marketing possibilities, we are confronted immediately with the problems connected with the handling and transportation of so fine a material. It is this problem, rather than the problem of finding new markets, that has been responsible for the lack of progress in the matter of dust disposal. It is the purpose of this paper to point out again some of the potential market outlets, and also to describe an existing system of handling and transporting coal dust. To the best of the writer's knowledge, this is the first and only operation in this country that has handled fine-mesh coal in bulk 50-ton lots direct to the consumer. POtential Markets for Fine-mesh Coal Dust In general, the potential market outlets would include any operation of which the pulverization of coal is a necessary part. To be more specific, a few of the better known operations of this type must include: colloidal-fuel plants, briquetting plants, "sea coal" distributing plants, water-purification plants and, obviously, pow-

Jan 1, 1944

By G. G. Gunther, C. L. Boeke

(Any discussion resulting from presentation at the Los Angeles Extractive Metallurgy Division Meeting will appear in the November 1953 issue.) I. M. Symonds (Cia Minera de Penoles, Monterrey, N. L., Mexico)—This is a most interesting paper describing zinc deleading operations. Mr. Gunther was kind enough to show me his operation last fall. Outside of the excellent results, I was most impressed by the ease of control. I had expected a difficult operation requiring careful control and supervision, but found one requiring little attention. The only disappointing fact of the separation is the poor copper recovery in the concentrate of the deleading float. The final zinc concentrate still has about 3 pct chalcopyrite (assuming the copper is chalcopy-rite) which adversely affects the zinc grade. If the copper mineral is free, the poor recovery is probably due to the high cyanide concentration. Fortunately, about 75 pct of the copper is recovered before the pulp goes to the zinc circuit. Mr. Gunther's remarks about proper cyanide to zinc sulphate ratios cannot be overemphasized. Although our experience has been mostly confined to the conventional Pb-Zn separation, we have found these ratios most important, and more important in the mill than in the laboratory. The laboratory appears to give optimum results over a wider range of reagent control than the mill, especially in regard to the effect of cyanide. In the mill in handling relatively high grade ores, we usually find that ratios of cyanide to zinc sulphate of 1 to 10 or 20 or higher work best. Where there is a substantial amount of chalcopyrite present, it has been noted that cyanide in the range of 0.02 to 0.10 lb per ton of ore is often beneficial in zinc rejection and possibly in copper recovery in the lead concentrate, while larger amounts, say 0.20 to 0.25 lb per ton, will float more zinc and cause a serious rejection of chalcopyrite. On an ore with 15 to 20 pct Zn content, we have had to use 0.8 to 1.2 lb of zinc sulphate per ton of ore with the above cyanide to obtain optimum results. Sometime ago we made a laboratory study following Parral and Santa Barbara practice on an ore which never had a very satisfactory rejection of zinc in the lead concentrate. The results were not encouraging and a mill test was not run. Since then the study has been reopened using fresh zinc concentrate from the mill in the laboratory instead of one made in the laboratory. A reasonably clean lead float from the zinc concentrate was obtained. It has been checked several times and each time the zinc concentrate made in the mill responds and the one made in the laboratory does not. Although undoubtedly some reasonable explanations can be given, it is presented here only

Jan 1, 1953

Using 5-p Ni powder and 0.018 ,u A1203, oxide dispersion strengthened nickel alloys were prepared by mechanical mixing of powders, followed by compaction, sintering and extrusion. Processing variables such as cooling of the powder during mixing, electrostatic discharging, mixing time, amount of oxide, and so forth, were studied to determine the effects on room-temperature tensile and high-temperature creep-rupture properties and on the reproducibility of alloys. MECHANICAL mixing of metal and metal oxide powders provides one of the simplest and most inexpensive methods for preparing dispersion strengthened metal-metal oxide alloys;" however, because of large differences in specific gravity, powder particle diameters and other variables, difficulties have been encountered in reproducing alloys and alloy properties. Using 5-p Ni powder and 0.018 p A1203, the following variables were investigated: a) Blending techniques b) Blending time c) Volume fraction of oxide for the selected pow- ders d) Sintering of compacts before extrusion Cooling of Powders During Blending— Cremens and Grant had indicated that dry ball milling of the powder mixture gave superior results over other procedures. When ultra-fine metal and oxide powders became available and comminution was unnecessary or offered little gain, the use of dry mixing in a Waring Blendor gave optimum results.' In the current study, observations of the powder mixtures indicated that considerable heat was generated due to the high speed of the Waring Blendor (15,000 rpm). This heating appeared to produce some degree of clustering and pelletizing. To minimize such agglomeration, a Pyrex cooling coil was installed in the mixing vessel. A double coil, water cooled, proved to be quite satisfactory, providing a fair degree of cooling along with good erosion resistance. Discharge of Electrostatic Charge—Electron-microscopy showed that agglomeration was minimized by the cooling procedure but was not entirely eliminated. It appeared that some of the ag- glomeration had to be explained on the basis of an electrostatic attraction due to the high-speed blending operation. These charges had the effect of promoting the growth of alumina clusters. At the end of the blending procedure such clusters had an effective size which was as much as 50 times that of the individual powder particles and were very tightly bunched. Several methods were utilized to try to discharge the electrostatic forces to minimize the clustering. One successful method, utilized in alloy 17, involved continuous exposure during blending to a radium source of 50 millicuries at a distance of 1.5 ft from the container. A number of other methods were noted to work equally well. Fig. 1 shows the more open oxide structure which results from such a treatment, permitting the matrix metal to penetrate into the cluster more effectively. The combination of powder cooling and discharging led to noticeably improved structures. Fig. 2 shows a cross and a longitudinal section in alloy M-17, one of the best alloys, illustrating the marked decrease

Jan 1, 1961

By F. H. Wilson, E. W. Palmer

G. L. Bailey (British Non-Ferrous Metals Research Association, London, England)—I was glad to see this further work on the structure and properties of the Cu-Ni-Fe alloys of low nickel content in which we have been so interested in the United Kingdom since before the war." The authors' data on the variation in solubility of iron in both 5 and 10 pct Ni alloys with temperature is based on more detailed observations than we have made but is generally in line with our own work. The difficulty in these matters is in the identification of the structures obtained. The precipitation of the second phase which we presume to be the a' solid solution, richer in nickel and iron, from the a solution with which it is in equilibrium occurs in such a fine state of division that its resolution by ordinary micro-graphic methods is impossible in the early stages. We have published some electron micrographs which illustrate the effects observed more clearly but we agree with the authors when they say that dark staining characteristics when the alloys are subjected to sea-water corrosion are more sensitive in detecting the presence of the a' phase than is the microscope. After taking this point into consideration, however, I confess to some difficulty in understanding Fig. 4. I can only assume that the precipitates visible at such a low magnification as 150 diam in these relatively coarse forms are produced from iron-rich areas of the initial solid solution due to inadequate homogenization. If this is so I find it difficult to accept the authors' contention that such local variations in iron concentration as are capable of giving the heavy precipitates shown in Fig. 4 do not in fact materially affect the solvus surfaces of their equilibrium diagrams. I share the authors' enthusiasm about the resistance to corrosion of these alloys by seawater but I am not fully in agreement with them on the vexing question of the best iron content for these alloys. In Fig. 3 of ref. 8, we give many results showing that the resistance of these alloys to impingement attack in the Jet Test increases as the iron content rises to 2 pct. With higher iron contents pitting at shielded areas is likely to occur. Admittedly the addition of 0.75 pct Fe produces a large improvement in corrosion resistance compared with a pure 90/10 Cu-Ni alloy, but the further improvement on raising the iron to 2 pct is substantial. We recommend from 1.5 to 2 pct Fe as the desirable range from the point of view of corrosion resistance. While it is true that British practice generally uses a hard tube, the corrosion resistance of the 10 pct Ni alloys has in our work been little affected by the structural condition. Although the annealed alloys blacken in seawater, this effect seems to be superficial and not to be accompanied by any real decrease in

Jan 1, 1953

By James Terry Duce

In order to facilitate interpretation of the data in this chapter, we print the following excerpts from circulars to authors, compiled by Mr. Frank A. Herald when he was Vice-chairman for Production of the Petroleum Division: Generally in Table 1 the unit for presentation of data is a field. For our purposes a field is defined as the whole of a surface area wherein productive locations are continuous. Such unit commonly includes and surrounds nonproductive areas. Such unit commonly includes a great variety of geologic conditions-—several units of continuous productive reservoirs of distinctly different structure and of distinctly dzerent stratigraphy. Therefore it is hoped that our authors will subdivide "field" so as to enable students to make analyses that may have scientific and/or commercial value. As to each space in the tabulation, it is either (1) not applicable, (2) the proper entry is not determinable, (3) the proper entry is determinable, but not determinable from data available to the author, (4) the proper entry is determinable by the author. In spaces not applicable, the author will please draw horizontal lines; in spaces where the proper entries are not determinable, the author will please insert x; in spaces where the proper entries are determinable but not determinable from data available to the author, the author will please insert y; in spaces where the proper entries are determinable by the author he will, of course, make such entries. Generally, y implies a hope that in some future year a definite figure will be available. Inability to determine precisely the correct entry for a particular space should not lead the author to insert merely y. Contributions of great value may be made by the author in many cases where entries are not subject to precise determination. In such cases the author should use his good judgment and make the best entry possible under the circumstances. For many spaces, the correct entries represent the opinion of the author (for example, "Area Proved") and in such cases the entries need not be hedged to such extent as in cases where the quantities are definite yet can be ascertained only approximately by the author. In cases under definite headings but where figures are only approximate, the author may use 2. For example, if the total production of a field is known to be between 1,800,000 and 1,850,000, the author may report l,8x,xxx; or if the production is between 1,850,000 and 1,900,000, the author may report 1,9xx,xxx. Where a numeral is immediately to the left of x or y, such numeral represents the nearest known number in that position. As to quantity of gas produced from many fields the question will arise as to whether the figures should include merely the gas marketed or should include also estimates of gas used in operations and gas wasted. Although rough approximations

Jan 1, 1937

By J. J. Lillie

Sublevel stoping was first developed in the Michigan iron mines many years ago. Since that time this method, and modifications with long hole drilling, have been used in a number of non-ferrous mines and have been described in various papers and articles. With a few exceptions, the operations where sublevel methods have been applied are producers of large daily tonnages from ore bodies of considerable size. In selecting mining methods for small properties or relatively small detached ore bodies, sublevel methods often have been overlooked. The method has been used successfully in relatively small operations at Copper Canyon, Nevada, and, more recently, in mining small ore bodies at Darwin, Calif. As a result of this experience, we believe that a review of the general procedure, the advantages noted, and the results obtained will be of interest to many operators of relatively small mines. General Description Briefly, sublevel stoping consists of, first, driving sublevel drifts at predetermined vertical intervals through the length of the stope block; second, breaking an initial mining slot across the stope; third, driving crosscut benches from each sublevel by slabbing into the slot; and, fourth, breaking the ground between benches. Applicability Since this is an open stope method, it requires wall rock that will stand over reasonable stope lengths without caving or excessive slabbing. Ore must be moderately firm but need not be hard. The method is best adapted to steep dipping, tabular or lenticular ore bodies but may be applied to ore bodies of almost any shape if their vertical extent is appreciable. It can be used with dips that are too flat for successful shrinkage stopes and can be modified for use with slushers on dips below about 40". Initial mining at the Copper Canyon operation consisted of sublevel stopes for the wider ore sections and shrinkage stopes for the narrow sections where ore widths averaged about 15 ft. After some experience and a study of comparative results on completed stope blocks, sublevels were substituted for shrinkage stopes even in the narrow ore widths. Practical stope lengths depend upon the character of the walls and the continuity of values within the ore body. Where blocks of low grade or waste are included in the stoping limits, these areas can readily be left as pillars for intermediate wall support and longer stope lengths may be planned. Raises can be spotted in narrow or low grade areas and stope lengths adjusted accordingly. Stoping Details To illustrate stoping details, assume that we have an ore shoot 15 to 30 ft wide and about 200 ft long, with walls that would stand over a stope length of 80 to 90 ft (Fig1). PRELIMINARY DEVELOPMENT Preliminary development for sublevel stoping would be simple. A service raise is driven at about the center of the block and stope limits planned to leave a pillar around this raise. Sublevel drifts driven each way from this central raise will outline the

Jan 1, 1950

By John C. McDonald

This report is a continuation of an earlier one with a similar titlell to which the reader is referred for such details of procedure as do not appear here. A brief summary will be given of the objects and methods of the research at this time. It was pointed out in the previous paper that properties of magnesium alloys, as reported by different workers, often differ. This was thought to be due to differences in the preparation of the specimens tested. The ideal way to compare alloys would be to prepare each of them in so many ways that reasonable assurance would exist that the best properties of each alloy had been developed. The alloys would then be compared on the basis of these properties. The treatment actually given the alloys of these studies is necessarily a compromise with completeness. Nevertheless, enough treatments have been given so that a considerable degree of assurance exists that the properties obtained are representative of the alloy as such, and not merely of the manner in which it was worked and heat-treated. Procedure The alloys were cast in the form of cylindrical ingots, which were extruded into bar before being rolled. Some of the ingots were made in 2-in. diameter molds whereas all of the alloys of the previous paper were cast in 4.in. molds. For the larger ingots, the extruded bar was 2 by 0.140 in.; lor the 2-in. ingots, it was I by 0.100 in. Preliminary work indicated that strip rolled from both sizes of the bars would have substantially the same properties. The bar was rolled at some temperature in the range 400' to 800°F., for a total reduction of 50 per cent. The temperature and reduction per pass were so selected that no cold cracking occurred, but were such as to leave the metal in a semihard condition. The hot-rolled metal was annealed at several temperatures in the range 400' to 800°F.; it was also cold-rolled, and given solution and aging treatments. Two standard test bars corresponding to each treatment were pulled in tension, and the average was obtained. Twelve treatments, in all, were given each alloy. The properties of the annealed specimen having the highest percentage of elongation were selected to represent the alloy in the annealed condition. The properties of the cold-rolled specimen having the highest yield strength were selected to represent the alloy in the cold-rolled condition. These properties were then plotted against percentage composition by analysis. Smooth curves were drawn among the points, thus graphically averaging the results over the entire range studied. The percentage of elongation in the cold-rolled state was not plotted, since it was always low, and varied in an irregular manner with composition. Alloys Studied Magnesium-calcium Alloys.—Fig. I presents the results on the alloys of the rnag-

Jan 1, 1941

By H. R. Hendricks

The ore of the Iron King mine, being very hard and having a very fine crystalline structure, presents many problems in milling that are not present in ordinary lead-zinc ores. This very fine crystallization causes much interlocking of minerals and is most prevalent between the gold and pyrite and the sphalerite and pyrite. The recovery of gold is extremely important in the economic value of the ore and therefore constitutes a large part of the metallurgical problem to be solved. All of the gold is free but it has not been observed in any of the mill products except under very high magnification. Some of the gold is tarnished, and this condition also makes recovery difficult. The gold values have been determined to be highest in association with chalcopyrite and arsenopyrite. The silver occurs mostly in association with galena and tennantite. A small amount of native silver has also been observed occasionally. The galena, being softer than the other minerals, frees much better in grinding. Some oxidation of galena is present in the ore, as the non-sulphide lead content amounts to approximately 10 pct of the total lead. Some oxidation of sphalerite is also present in the ore. Microscopic examination of the zinc concentrate shows many grains in the minus 325-mesh to be half pyrite, together with the fact that some of the zinc is marmatite, makes production of a high-grade zinc concentrate very difficult unless recovery is sacrificed. In the grinding, it has been determined that the flotation feed must be maintained at 85 pet minus 200-mesh to free the minerals sufficiently to make the best economic recoveries. Table I illustrates the effect of fine grinding. It shows that the bulk of the recovered minerals is in the minus 325-mesh size and is the part of the minerals that is more readily freed. The pyrite concentrate is not given in the table because it contains practically no minus 325-mesh. Development of Present Flowsheet and Milling Practice The first milling operation started in 1938, with a mill capacity of 140 tons daily, producing a bulk gold-silver-lead concentrate. After several stages of expansion and changes of the mill flowsheet, the mill is currently treating 380 tons of ore daily and three separate concentrates are produced. In order of their importance and value, these are: (I) lead concentrate, (2) zinc concentrate and (3) pyrite concentrate. The present average grade of mill heads is: gold 0.11 oz, silver 4.00 oz, lead 2.50 pet, zinc 7.60 pct, copper 0.22 pct, iron 22 pct and 30 pct insolubles. Specific gravity of the ore is 3.8. During the period when bulk flotation mas practiced, the flowsheet was very simple. The ore was ground fine enough to liberate as much gold, silver and lead as possible. Necessary collectors and frothers were added in sufficient quantity

Jan 1, 1949

By William A. Johnson

The fundamental role of diffusion in many reactions occurring in solid metals has long been recognized, and there have been careful measurements of rates of diffusion in numerous alloy systems; but our understanding of the atomic mechanism by which diffusion occurs is scarcely in an advanced state.l-3 It is doubtful that this mechanism can ever be directly ascertained; rather, all imaginable mechanisms must be examined and the one that is in best accord with such data as can be obtained from suitably chosen experiments should be selected. Because of its relative simplicity, diffusion in a system containing only one kind of atomic species—commody called self-diffusion—is of considerable theoretical interest. It may be expected that diffusion theory will not be advanced appreciably until a wide variety of diffusion experiments—of which diffusion in a chemical concentration gradient and self-diffusion are two important types—are carefully performed and inter-compared. Aside from the purely theoretical interest in self-diffusion measurements, such studies will be of considerable engineering importance. The rate of any process in a pure metal requiring the movement of atoms, of which recrystallization and creep are two familiar examples, will be understood more readily with the aid of self-diffusion data.† Self-diffusion implies the diffusion of something into its exact counterpart; the detection of such a process is patently impossible, and it is necessary to substitute for it the diffusion of one isotope (in this case radioactive) into another isotope (in this case stable) of the same chemical element. There is, however, no reason to suppose that a radioactive atom will not diffuse at the same rate as its stable isotope, for all of the isotopes of the same element have exactly the same chemical properties; neither the difference in nuclear energy nor the difference in mass would be expected to exercise an important influence in this phenomenon. In this sense, self-diffusion may be thought of as the diffusion of radioactive atoms down a radioactivity concentration gradient, just as usual diffusion is the diffusion of one kind of atom down its chemical concentration gradient. The first measurements of self-difiusion were made by von Hevesy4-6 using the radioactive lead isotope thorium B in ordinary lead. The recent development of means for producing artificial radioactive elements has made possible similar measurements on gold,7, 8 copper,2, 9 and zinc.10, 11 Many of the data that have been obtained are rather poor, and no adequate reason has been given for their poorness; it appears, however, that the unsatisfactory results are characteristic of the particular experimental procedure employed, and not of self-diffusion measurements in themselves. The experimental technique that has yielded the poorer results is the deposition of a thin layer of radioactive atoms on the surface of a specimen and the determination of the rate of inward diffusion

Jan 1, 1941

Discussion of the paper of GEORGE K. BURGESS and SIR ROBERT A. HADFIELD, presented at the New York meeting, February, 1915, and printed in Bulletin No. 98, February, 1915, pp. 455 to 468. ALBERT SAUVEUR, Cambridge, Mass.-There is a definiteness and finality in the investigations that have been conducted by the Bureau of Standards and in the method of presenting the results that is most satisfactory and refreshing.. A year ago Dr. Burgess and his co-workers investigated the question of critical points in carbonless iron, and they settled the question once for all. No one can now claim that the thermal point A2 in carbonless iron does not exist. At the past meeting of this Institute in Pittsburgh, Dr. Burgess showed the worthlessness of the shrinkage clause. To my mind after his demonstration no one can again claim that the shrinkage clause renders any service. He also showed us that it was a practicable proposition to use pyrometers in steel work in order to ascertain the finishing temperatures of rails. No one can again come before this body and make the claim that it is not a practical proposition. And now Dr. Burgess shows us that steel ingots can be made in a commercial way, practically free from pipe, from segregation, aid from blow holes. It places again before us in a very clear manner the question of piped ingots and of segregation, a question which has been debated for 20 years or more. I know very well the attitude of the standpatters in the matter. They claim that most of the pipe is removed in the discard, and that if a little is left it is so thoroughly welded that it does no harm. I for one am not convinced of the soundness of the argument. I do not believe there is any one present who would claim that, given an ingot free from pipe, free from segregation, and free from blow holes, better rails will not be made out of that ingot than out of the ordinary piped or segregated metal which is used by almost all of our steel mills. In my opinion, if we give our sanction to the casting of steel in the shape of piped or segregated steel ingots, we are guilty. If we give our sanction to the rolling of rails from such metal, we are guilty. It is our duty as experts, as steel makers, and as steel consumers to make and to use the best steel obtainable on a commercial basis, for the manufacture of rails. Now, we have been shown that sound ingots can be produced from which superior rails can be made. If we do not insist upon these sound ingots, then we fail in our duty. I think that the art of casting steel ingots has advanced to such a point that the casting of steel ingots piped and segregated ought no longer to be tolerated.

Jan 5, 1915

By J. W. Downey, M. V. Nevitt

MUELLER and Knott' have reported that Ti2Cu has the Si2Mo-type (Cllb) structure. The lattice parameters of Ti2Cu and the positions of the titanium and copper atoms are given by them as follows: ao = 2.944A; co = 10.786A 6 atoms per unit cell: 2 Cu (a) 000; 1/2: 1/2, 1/2 4 Ti (e) 002; 1/2, 1/2, 1/2 + z 00 -2 ; 1/2. 1/2. 1/2- z where z = 0.341 We have found eight other binary compounds which appear to be isostructural with Ti2Cu. All of them involve the combination of a titanium group element with a copper group element or a titanium group element with palladium. The compounds and their lattice parameters are given in Table I. The alloy samples were prepared by arc melting metals having a purity of 99.9 pct or better. They were homogenized for 72 hr at 900°C in evacuated capsules and quenched in tap water. Filings were prepared, and these were reannealed for 4 hr at 700°C, also in evacuated tubes. The lattice parameters were determined by a least squares Bnalysis of the back reflection lines of patterns made with cobalt radiation. It was not possible to make a precise determination of the lattice parameters of Zr2Au because the filings underwent rapid oxidation at room temperature. The evidence supporting the large tetragonal cell containing 6 atoms involves the presence of the reflections 002, 101, 004, 112, 114, 202, and other reflections with higher indices for which I = 3n. Three or more reflections corresponding to these indices, which cannot be indexed on the basis of a small body-centered tetragonal cell containing 2 atoms, are observed in the patterns of all of the compounds listed in Table I. They are strongest and most numerous when the components have large differences in their atomic scattering factors, as is the case in Ti2Pd, Zr2Cu, and Hf2Cu. The values of the z parameter should be regarded as approximate They were selected on the basis of the best agreement between the observed intensities of two pairs of closely spaced powder pattern lines (004): (112) and (114): (202) and the corresponding intensities calculated for several values of z. An ordering scheme was assumed in which titanium, zirconium, or hafnium atoms occupy the titanium positions of Ti2Cu, and copper, silver, palladium, or gold atoms occupy the copper positions. Calculated and observed intensities of the aforementioned lines and the other lines of the patterns are in satisfactory agreement, but a refinement of the z parameters and a final confirmation of the ordering arrangement must await a more detailed study. It is interesting to note that in transition metal systems at A2B stoichiometry the Si2Mo-type structure competes for stability with two other fre-

Jan 1, 1962

By D. H. Baldwin, R. E. Reed-Hill

DURING the course of experiments involving oxygen equilibrations with a high-purity Pd-5 at. pct Rh alloy, the appearance of a subscale was noted. Most of the heat treatments in a pure oxygen atmosphere were performed at 900°C. The presence of the subscale was noted when the oxygen partial pressure was in excess of approximately 95 mm, up to 1 atm, the highest pressure used. According to available thermodynamic information' on the free energy of formation of Rh2O3 (the only reliably reported solid oxide of rhodium), this oxide should not be stable under these conditions, nor, of course, any oxide of palladium. Extraction of the subscale from the matrix was affected by treatment with aqua regia, and the resulting X-ray diffraction powder pattern is shown in Table I. This pattern is not that of Rh2O3, nor that of any known rhodium or palladium oxide. It can be indexed on a hexagonal lattice with a. = 5.22Å and co = 6.0Å. In addition to this unknown material, there was a weak Rh2O3 pattern. An X-ray fluorescence analysis of the powder gave 3214 palladium counts per sec and 5381 rhodium counts per sec, thus indicating the likelihood of a double oxide of palladium and rhodium. In view of the fact that the above result was obtained on a single sample, it is perhaps best not to insist upon the above ratio of about 0.6, but possibly to tentatively assume 0.5 or 1/1. The small amount of Rh2O3 in the X-ray pattern might have occurred on cooling, and it seems possible that the double oxide, below some temperature level, is not as stable as Rh2O3. To examine this point, some of the extracted double oxide was heated on a sheet of pure palladium in air at 900°C, and allowed to furnace-cool over a period of several hours. The X-ray diffraction pattern of the powder showed approximately the same intensity of Rh2O3 and the "hexagonal" oxide as before, but, in addition, a rather faint pattern of another crystalline species made its appearance. This finding again suggests partial transformation of the major (double) oxide to other oxides on cooling from 900°C. Application of the phase rule shows that only one equilibrium oxide (subscale) could have been present at temperature. Vacuum-fusion, oxygen analysis of the extracted subscale showed 7.7 ± 0.7 pct O. Combining this information with that of the X-ray fluorescence analysis suggests the formula PdRhO. This compound should have an oxygen content of 7.07 pct, which is within the range of error of the oxygen analysis. However, until a more careful stoichiometric study is made, it is suggested that the formula be written

Jan 1, 1965

By R. Schuhmann

As a response to rapidly growing specialization in various branches of metallurgy, it is proposed that undergraduate education in metallurgical engineering should be built up around four principal scientific themes: 1—thermodynamics, 2—structure, 3—mechanics, and 4-—rate processes. THE following quotation from a recent article by von Bertalanffyl describes very well the situation which today challenges curriculum designers in all branches of science and engineering: "Modern scientific education is presented with a well known dilemma. The amount of facts in modern science, and in any of its smallest branches, is enormous. Life in general, and the academic curriculum in particular, is short. The abundance of factual data, as well as the intricacy of modern scientific techniques, experimental and theoretical, necessitates utmost specialization. This specialization, unavoidable though it is, involves serious danger for both the education of the scientist and the social function of science. "In so far as the advancement of science is concerned, history shows that interconnection of different fields and problems is a most important basis of progress. Many of the paramount achievements of science arose on borderlines, and from the synthesis of formerly separate fields . . ." Specialization in subfields of metallurgy has become a serious educational problem only in about the last decade. The growth in specialization, almost as a necessity, has paralleled the rapid growth in metallurgical knowledge during this period. Moreover, in planning future educational policies, the educator must be concerned not only with the rate of increase in metallurgical knowledge and with the concomitant rate of increase in specialization, but also with the second derivatives, that is, with the rates of increase of the rates of increase. These second derivatives are positive and large; the growths being dealt with are self-catalyzing, or like a chain reaction. As a measure of specialization, the length of the list of different kinds of metallurgies can be considered, the kinds of metallurgy which represent one person's particular interests and those of other practicing metallurgists. To start with, there are physical metallurgy, process metallurgy, and mineral dressing, then mechanical metallurgy, metal processing, foundry metallurgy, corrosion, solid state physics, high temperature metallurgy, welding, powder metallurgy, nuclear metallurgy, etc. Each of these fields is growing so rapidly that any one metallurgist can hope at best to keep up to date in only one or two. What is the impact of all these "metallurgies" on metallurgical education and metallurgical curricula? In answering this question, it can be said, first, that educators have resisted fairly well the temptations to split up metallurgy, with different curricula, different options, degrees with different labels, and the like—degrees are still given in plain Metallurgy and Metallurgical Engineering. In other ways, though, the pressures of specialization have had noticeable effects. There is a tendency to think of these "metallurgies" as the important subdivisions of Metallurgical Engineering and to work up

Jan 1, 1956

• In 1949 the United States imported 7,400,000 tons of iron ore; Chile, Sweden and Canada, in that order of importance, supplied over 80 pct of this amount. U. S. imports have increased from 3 pct of the total supply in 1940 to 8 pct in 1949. • Mica has been successfully synthesized by scientists of the National Bureau of Standards. quartz, magnesite and bauxite, plus a fluorosilicate compound to act as a crystallizing agent, are heated to 2500°F in a platinum-lined crucible and as the furnace cools mica crystals grow. The key to parallel growth to make sheets is in precision control of temperature. • Inland Steel Co. of Chicago has taken an option to lease from Steep Rock Iron Mines, Ltd., of Canada -- a large undeveloped acreage in the vicinity of Steed Rock Lake, Ontario -- for the purpose of exploring and developing a new iron ore mine. The exploration program is expected to begin at an early date. •A 20-yr test conducted by the U. S. Bureau of Mines with the cooperation of Cleveland-Cliffs indicated that chemically treated timbers last from 2 to 3-1/2 times as long as untreated ones set in the same passageway. Of the three reagents used, zinc chloride gave the best result -- 13.3 yr as compared to 3.8 yr for untreated timber. •Senator Pat McCarran charged that President Truman used a "shotgun approach" to the taxation problem when he accused the mining industry of enjoying excessive depletion exemptions when actually mining taxes should be lowered. He introduced to the Senate a report of the National Advisory Council -recommending a decrease * The country's first graduate program of research and instruction in conservation of natural resources will begin at Yale in September. The Conservation Foundation is supplying $20,000 a year to finance the program. •Felix E. Wormser urged that adequate tariff protection should be provided to offset effects of devaluation on the domestic mining industry in an address before the Colorado Mining Association. He advocates tariffs rather than subsidy legislation, such as S. 2105, for the protection of the lead-zinc miners because the latter would promote inefficiency and "impair our priceless free market." •"It may be startling to realize that for the world's white population the per capita requirements for minerals have multiplied 30 times in 30 yr, or that more minerals have been mined in the last 40 yr than in all prior history." -- Evan Just, in "Mineral Depletion and Metal Supply." •The National City Bank in its monthly letter states that both management and the unions concede that there are too many miners trying to make a living out of the industry for its present market. •On Feb. 3 it was reported that 6000 Anaconda employes in Chile quit jobs in sympathy with striking telephone and electric workers. The government has ordered them back to work on penalty of discharge. It appears that the walkout will not be prolonged in view of the government stand.

Jan 3, 1950

By H. L. Talbot, A. J. Hall, R. S. Young

A method for the chemical differentiation of oxidized cobalt from sulphide cobalt in complex mine and concentrator products is of considerable interest to a number of mining companies throughout the world. At Rhokana Corporation, in Northern Rhodesia, for instance, cobalt is found chiefly in the sulphide form, as carrollite, but a cobalt oxide mineral also occurs irregularly underground, and relatively small quantities in the mill feed from time to time are difficult to detect visually at the concentrator. The oxide cobalt is not amenable to the sulphide flotation process used at Rhokana and nearly all of it passes into the tailing. The occasional period when high tailing losses occur with no appreciable change in the cobalt content of the mill feed has naturally been disturbing to the concentrator staff, who had no means of determining whether an influx of cobalt oxide ore or laxity on the part of the mill flotation operators as responsible for these results. A procedure for determining oxidized copper minerals, based on their solubility in dilute sulphuric acid saturated with SO2, has been used for many years in the copper industry. In the presence of cuprite the procedure fails, but otherwise it gives results quite satisfactory for control pur- poses. Empirical methods for determining oxidized forms of zinc and lead have been proposed, but they are generally unsatisfactory except in isolated cases of a very simple ore. No published procedure for determining oxidized forms of cobalt or siinilar elements has come to the authors'notice. Experimental Procedure There are obviously two alternatives in attacking a problem of this nature. One is to employ a solvent for oxide cobalt that will not dissolve cobalt sulphide. The other is to find a substance capable of oxidizing the cobalt sulphide to sulphate, which of course is water-soluble and can then be leached from the water-insoluble oxide cobalt. A representative sample of Rhokana cobalt oxide ore was secured, ground to minus 200-mesh, and carefully rolled to furnish a standard sample for this work. The analysis of this material gave the following percentages: Co, 0.55; Cu, 1.64; Fe, 2.98; S, 0.08; SiO2, 41.42; Al2o3, 11.20; CaO, 13.82; MgO, 3.80. As a source of cobalt sulphide a similar large sample of carrollite was prepared, analyzing Co, 23.36 per cent; Cu, 27.32; and S, 31.13. This sample of course was not pure carrollite, but was the best obtainable in quantities sufficient for a standard intended for considerable work. It was a pure sulphide sample, microscopic examination showing only carrollite, a little bornite, and gangue, with no visible sign of oxidized ore present. A wide variety of reagents as used in an effort to selectively dissolve oxide cobalt or to oxidize sulphide cobalt to

Jan 1, 1949