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By R. W. Parsons

This discussion concerns fitting an ellipse to angle, where the permeability k is measured for a series of plugs cut from varying angles in the horizontal plane. The conclusions and procedures given by Greenkorn, et al. are in no way affected by these comments, since the degree of anisotropy (ratio of maximum to minimum permeabilities) they found is generally quite close to one (see their Table A-1). However, since large anisotropies may arise, I feel that the comments should be made. Both Marcusl and Scheidegger2 have shown that there are two limiting expressions for the permeability of an unbounded anisotropic media. If the permeability is measured in the direction of the pressure gradient, the result is equivalent to the author's Eq. 11. This is the equation used to find kl1 and k22 from the experimental data. If, however, the permeability is measured in the direction of flow, then the resulting equation for the measured permeability is From these equations it follows that plots on angle will be ellipses. The principal axes will correspond with the major and minor permeability axes. The question arises as to what permeability is actually being measured on an anisotropic sample in the laboratory. An overall pressure gradient and a gross flow rate are determined and a permeability calculated by Darcy's law. The details of the macroscopic flow within the sample will be determined by the boundary conditions of the sample and the orientation and magnitude of the major and minor permeability axes. The situation is shown in Fig. 1 for the two-dimensional case. Using the method of conformal mapping, Marcusl has solved the problem of determining the overall apparent (or measured) permeability for this experiment for any condition of W-L ratios and orientation and magnitude of the principle permeability axes. For W-L ratios approaching zero, the streamlines are all forced by the boundary conditions to be in the direction of the measured permeability which is equal to kg. For W-L ratios approaching infinity, the effects of the no-flow bounds are negligible, and the pressure gradient and measured permeatility direction are parall;l. The streamlines are

Jan 1, 1965

By W. A. Thomas

STEAM is, and doubtless always will be, the basic power in the anthracite industry, either directly applied through engines and pumps or electrically. The rapidity with which electric power is being applied to the mining and preparation of anthracite leaves little doubt as to its utility and economy. There is practically no operation, except possibly the drilling of hard rock, that has not been successfully electrified and the matter of drilling is handled economically with electrically driven compressors and the standard compressed-air drills. Conservation and the seemingly ever-increasing cost of production, justify careful consideration of the possible results of the more extensive use of electric power in the industry. It is not planned to discuss, herein, the relative economies of individual applications but to consider the relative economy and flexibility of steam and electric power. The reports available for the year 1920 indicate a production of slightly over 80,000,000 tons of anthracite, of which 8,843,500 tons were used for power and heating purposes; in other words, over 11 per cent. of the production was used at the mines. Some of this was used to generate electric power at the mines, about 160,000,000 kw-hr., while 150,000,000 kw.-hr. was purchased from the public utility companies. One large producer, who is buying some electric power and generating some, used over 13 per cent. of the production for power and heating at the mines. Another producer, where considerable electric power is generated on the properties, is using less than 9 per cent. of the output for power and heating. One of the largest producers has a fuel consumption of over 15 per cent. of the output, besides which considerable power is purchased. A careful study of the field indicates that steam-operated plants use upwards of 13 per cent. of the production for power and heating, while electrically operated plants use about 1 per cent. for heating; so that, for steam-power purposes at least 12 per cent., or 268.8 lb., are used for power out of each ton produced.

Jan 9, 1921

By Charles Schwab

WHEN I leave this life, as an employer of labor there is no one thing that I want so much to be engraven upon my monument as the fact that I have been one of the men who have worked, whether with my brains or my hands, for the advance-ment of industry in the United States, and to be regarded, as I should like to be regarded by labor with love and kindly remembrance for our years of association. In my relationship with my employees, I have always contended that every man should be paid for what he does. I will not employ or be associated with any kind of labor of which the organization insists that a man shall receive a fixed sum of money for his day's wages, however poor or good his day's labor may be. What applies to the common laborer applies equally well to the management, the brains of an industrial establishment. I have always thought of industry as a three-legged stool, with capital, management and labor as the legs. All three legs of that stool must be there before that stool can stand up. If you take any one of those legs away, it is going to topple over. The. three must stand together for successful business accomplishment. I believe that a man who has his labor to sell, to put it plainly, has just as good a right to bargain for the best price for that labor as those who employ him. It is a perfect right for him to do so. If the government of our United States is right, our method of handling labor in our Bethlehem plants is right because it is representative and democratic. We have our labor elect delegates; elect representa-tives from their own ranks, who shall meet frequently. That constitutes, as. it were, the House of our great Congress in Washington, the industrial house. We have our management, who represent the Senate, who meet independently. Working men have learned to express their views freely as to whether they are treated fairly and whether conditions are right or wrong.

Jan 12, 1922

By John Drew Ridge

For the early conservationists before the first decade of the twentieth century, conservation meant largely the planned preservation of water, forest, soil, and wildlife resources-renewable natural resources-through efficient use. Water was primarily thought of as a resource that must be kept available for humans to enjoy in a recreational sense and as a habitat for wild animals, birds, and fish. Little thought was given to conserving water for actual human or industrial needs. If industry entered the picture it was as an actual or potential polluter of water, which was the basic necessity for drinking and bathing and for outdoor recreation. Even the forests were only held worthy of conservation largely because they were areas in which wildlife could, or should, abound and in which human beings might find relief from the cares and problems of everyday life. Finally, the soil was to be preserved in its natural fertility through proper fertilization, cultivation, and crop rotation; in this the conservation program did enter the realm of the practical. For through soil conservation, the world was to be able to raise increasingly large crops of constantly improved organic foods which the steadily growing population of the earth was to need to survive and improve. Relatively little time, however, was spent on considering what profit the farmers of the world were to make from the scientific farming on which they were to embark. In short, the appeal of the early conservationist program was much the same as that of the entire era of Theodore Roosevelt of which it was a part; it was to the heart rather than to the mind; it was to emotion rather than to reason. Because of the way in which the problem was approached, conservation took on the character of a crusade rather than that of a practical plan for the betterment of man through economically sound methods of improving his environment and the uses he makes of it. Conservation had, then, all the advantages of a stirring appeal to the emotions, but it did not have the lasting results that would have been achieved by a program with a sound economic basis.

Jan 1, 1959

The great additional clue given to professional books by adequate alphabetical and analytical indexes has been recognized from the beginning in the publication of the Transactions of the American Institute of Mining Engineers The bust ten volumes, edited by Dr Thomas M Drown, the secretary, were provided with indexes, as well as tables of contents Moreover, he included in Vol V (published in 1877), a consolidated index of Vols I to V , inclusive, and, at the triune of his resignation in 1884, he had prepared a similar index of Vols I to X , inclusive, which was published in the following year The indexes of the annual volumes from Vol XI on were made much more extensive than their predecessors Casual mentions, as well as important discussions, were included, the geographical localities of mines of works alluded to in the text were added in the index, and numerous cross-references were introduced, with the purpose of making it easy for the student to discover at once what the Transactions contained, either of thorough treatment or of hints and clues to further inquiry, concerning any topic, locality, or person within their scope In order to extend the advantages of this larger scheme to the earlier volumes a consolidated index of Vols I to XV , inclusive (published in 1888), was prepared by compiling the several annual indexes and introducing additional items for those preceding Vol XI Vol XX , issued in 1892, contained a similar consolidated index for Vols XVI to XX , inclusive, in 1897, a sepal ate index-volume, covering Vols XXI to XXV, inclusive, was published, and on 1902 appeared another for Vols XXVI to XXX These four indexes, for Vols I to XV, XVI to XX, XXI to XXV , and XXVI to XXX, respectively, bound together into a book of about 950 octavo pages have constituted for the last five year s a consolidated index to the Transactions which they cover-with the single difference that the book had to be consulted four times, where a single comprehensive index would require to be consulted but once This disadvantage, however, was relatively small, compared with the great convenience of finding by four trials m one book what would otherwise call for 30 separate searches in 30 books Moreover, the possession of the one might be most useful to a student who did not possess the 30, by indicating to him what they con-

Jan 1, 1907

By A. W. Schillinger

Drilling was not entirely satisfactory in the search for native copper in the Osceola amygdaloid footwall, for oreshoots mined were more continuous than drilling indicated. The Geophysics Dept of Michigan Technological University developed a portable induced polarization drill hole probe with a 3-electrode configuration and a sampling radius of 5 ft ±. A 20 ma current of alternately reversed polarity is pulsed at 3.5 sec intervals and readings are taken every 2 ft. Potential is read during the "on" cycle for resistivity determinations. During the "off" cycle, after 10 ms delay, the 1P potential is sampled for 10 ms and recorded. Resistivity (ohm-ft) and "S" values (mv/v) are plotted on semi-log paper. A lithologic log is plotted based on resistivity and anomalous S values noted. The low resistivity amygdaloid contains sills of high resistivity basalt. Verified by core logs, contacts between amygdaloid and basalt could be picked to ± 0.5 ft from resistivity alone. Lithologic logs of long-steel holes, based on resistivity, then replaced sludge logging. An S value of 30 mv/v was established as the lower limit of diagnostic 1P effects in mineralized amygdaloid and 50 mv/v as a threshold value for ore. Of 914 footwall zones drilled, 315 were probed yielding 203 anomalies. Success increased from 34% for drilling to 72% for drilling and probing. As checked by mining, 46 probe results were 78% correct. Failures were due to the instrument not detecting copper or to the misinterpretation of an anomaly. No false anomalies were found. During the past ten years, the Calumet Div. of Calumet &Hecla, Inc has conducted extensive exploration and mining in the footwall of the Osceola amygdaloid lode in the native copper district in Michigan's Upper Peninsula. In an attempt to delineate more clearly the extent and nature of high-grade footwall copper mineralization, 876 diamond drill and long steel holes aggregating slightly more than 47,000 ft have thus far been drilled from the accessible hanging-wall workings in the Osceola No. 6, 13, 14 and 15 shafts. Early in this program it became obvious, through inspection of footwall workings in progress, that due to the erratic and bunchy occurrence of the native copper in the footwall zone, the exploration drilling was failing to outline completely the extent of the oreshoots. Early drilling re-

Jan 1, 1964

By C. D. Woodward

THE subject of electric signals for the despatching off mining cages through shafts has received cousiderable attention recently from various mining companies. The Anaconda Copper Mining Co. has found it necessary to develop a system that is reliable, safe and simple in operation, and easy to repair. The movement of all mine cages or skips is under the direction of a station tender, who rides on the cages when men or supplies are being lowered or raised. When ore is hoisted, he is stationed at a skip pocket for loading and despatching the skips. For the station tender's use, two signal systems are installed in each hoisting compartment, and he is the only employee authorized to use them. One system is known as the "AC" or Station bells, as it is operated from the 110-volt, 60-cycle, power service; the other system, called the "DC" or shafts bells, takes its energy from closed-circuit primary batteries. As a supplement to the two main signals, a buzzer system is installed. The general arrangement is similar to "the main signal, except that there is a mine buzzer at each station. When a signal switch connected into the buzzer system is pulled down, all buzzers are operated, thus notifying the station tender that he is wanted at some particular level. Placed on the station posts, and within easy reach of the men on the cages, are the signal pull switches for the main signals. The station, or "AC," pull switches are furnished with a short rope as shown in Fig. 1, while the shaft, or "DC," pull switches are equipped with ropes that extend down the shaft past the next station. To avoid all confusion in the use of the three signal systems, a standard arrangement has been adopted in all mines for the location of the pull switches. Fig. 2 illustrates the position of the pull switch at the various mine stations for the main signal systems, but it does not show the buzzer pull switch; this is fixed to the roof timbers about 25 ft. back from the shaft. If the skip pockets are a considerable distance below the station, a pull switch for each main shaft compartment is connected to the "AC" system and placed at this point. However, if the distance is short, a light cord is attached to the station pull switches and dropped down to the skip pocket.

Jan 2, 1922

By Walter Showak

The published phase diagrams for the copper-zirconium system1,2 indicate that the solid solubility of zirconium in copper is not accurately known. Studies made on relatively impure alloys by Pogodin, Shumova, and Kugucheva3 yielded a maximum solubility value of 1 wt pct Zr at the eutectic temperature. Allibone and sykes5 reported that an as-cast 0.2 wt pct Zr alloy is single phase at room temperature. Limited results obtained by Raub and Enge14 indicated that zirconium solubility in copper is less than 0.13 wt pct Zr at 940°C. Tn 1960, on the basis of a microstructural examination of several copper-rich Cu-Zr alloys, saarivirta6 placed the maximum solubility of zirconium in copper at 0.15 wt pct at the eutectic temperature, 980°C. The present note describes results obtained during a re-investigation of the Cu-Zr system for the purpose of accurately establishing the solidus and solvus curves and the eutectic temperature in copper-rich alloys. High-purity copper (99.99 pet) and zirconium (99.64 pet) consolidated reactor grade sponge were used to prepare several alloys, containing up to 2 wt pct Zr, by induction melting in graphite crucibles in vacuum and chill casting into a split steel mold. The melting procedure consisted of first melting the copper charge and then adding the bulk zirconium wrapped in pure copper foil. Chemical analyses were made of both the cold-rolled alloy ingots after homogenization at 925°C for 450 hr in evacuated quartz capsules and of small specimens which had been used in the phase diagram studies. 'The carbon content of each alloy was also determined and was found to be less than 0.001 wt pct in every case. Metallographic and electrical resistivity measurements were used as the primary methods for this investigation, although some thermal analysis studies were also performed. Heat treatments of all the alloys were carried out with the specimens wrapped in molybdenum or alloy foil (of the same composition as the specimen) and sealed in quartz tubes either in vacuum or under a purified helium atmosphere. Following high-temperature anneals for long periods of time, slight depletion of zirconium from the specimen surface was observed. Therefore, careful preparation of samples for microscopic examination and chemical analysis of the samples themselves was necessary to check the composition of samples used in determining phase boundaries. Points on the solvus curve were determined on quenched specimens by both isothermal reaction after solution heat treatment at 950°C for 8 hr and by isothermal annealing of alloy specimens at successively higher temperatures for times ranging from 24 to 500 hr. Liquation studies for positioning of the solidus and eutectic temperatures were performed by holding a pre-

Jan 1, 1962

The object of crushing and grinding for leaching is obviously to permit contact of the solvent and mineral and subsequent removal of pregnant solution in a reasonable time. Heap leaching is the leaching of ores that have been mined and transported to prepared, relatively impervious pads for leaching by sprinkling or ponding, and percolation of the lixiviate through the ore. Dump leaching is the leaching of dumps or low-grade ore or overburden without the use of prepared pads under them. Run-of-mine material, e. g., low -grade ore from open pit stripping operations, may be leached without crushing, and a fairly high recovery made by heap leaching. An alternate method for better grade gold/silver ores is to crush to 9 to 13 mm (318 to112 in. ) size for pad heap leaching, as used by Smoky Valley Mining Company at Round Mountain, Nevada, on gold ore, and to 38 rnrn (1 -1 12 in.), which will be done by Candelaria Partners, at Candelaria, Nevada, on silver ore. Heap leaching of copper probably began at Rio Tinto, Spain, about the year 1752. This was done by the leaching of weathering piles of ore in the open air followed by iron cementation of the liquor that drained off the heaps. Its application to the low -grade southwestern United States copper ores was overlooked for many years, and low- grade ores from Sacramento Hill at Bisbee, Arizona, were the first to be heap leached in 1923. Site preparation used a half century ago is the same today. "The need for some form of waterproofing of the pad was evident. Before leaching ore was piled, the site was cleared of cactus and brush. After clearing a gentle sloping area, it was then covered with a blanket of rolled and tamped clay or slime tailing. When heaps are built, a layer of coarse fragments may be put down first to assist drainage. A second method of leaching without grinding, usually for higher grade ores, is by vat or tank leaching of material crushed to 9 to 13 mm (318 to 112 in.) size. Such a plant is considerably more expensive than a heap leaching installation but much less costly than grinding for agitation leaching, and a recovery higher than that obtained by heap leaching is to be expected. Vat leaching has been widely used by the copper industry on higher grade copper ores but has been used in only a few small operations on that size gold or silver ores. However, such a plant should be considered for a modest tonnage of a few million tons, 3 to 7 g per mt (0.10 to 0.20 oz per st), of Carlin-type leachable ores crushed to 9 mm (318 in.).

Jan 1, 1981

By L. Kaufman

An equation is derived relating AF a", the difference in free energy between austenite and martensite, to temperature and composition in the iron-chrmnium and iron-chromium -nickel systems. This equation is used to calculate at M,for low-carbon stainless steels. A method for utilizing the energetic equations to calculate M, in stainless steels is suggested. In addition, an explanation is given for the anomalous effect of chromium in forming a y loop concurrent with the lowering of M, in iron-base alloys. It has been demonstrated that the effect of alloying elements on the Ms temperature is strongly dependent upon the influence of the alloying element on the thermodynamic properties of the y and a phases. Notwithstanding the fact that alloying elements affect the Ms temperature by changing the lattice parameters, elastic constants, and inter-facial energies of the a and y phases, their primary role is a thermodynamic one. Thermodynamic analyses have been presented for the iron-carbon'-4 and iron-nickele systems in which equations are derived for Fa and FY, the free energies of the b.c.c. (or b.c.t.) and f.c.c. phases. These equations, which can be obtained with the aid of the equilibrium diagrams, can be used to help predict the Ms vs composition curves. This paper deals with a calculation of the thermodynamic properties necessary to compute AFffLY for iron-chromium alloys. In view of the lack of direct information concerning the activity of chromium in the a and y phases of iron, the regular solution approximationa,g was utilized to bridge the gap. After the necessary equations were derived for the iron-chromium system they were used to calculate AFff— Y for stainless steels by combining them with the previous results obtained for iron-nickel alloys. The results of these calculations offer a means of rationalizing the anomalous behavior of chromium in iron-base alloys. The latter depends on the regular solution approximation in contrast to the explanation offered by zener , which is based upon the "magnetic specific heat effect." THERMODYNAMICS OF THE IRON-CHROMIUM SYSTEM If y represents the atomic fraction of chromium in an iron-chromium alloy, then the free energy of the b.c.c. phase Fa, the f.c.c. phase FY, and the difference in free energy between the f.c.c. and b.c.c. phases of the same composition are given through [3] and Fmy represent the free energies of mixing of the a and y phases; while @era -' and uFea 4Y represent the difference in free energy between f.c.c. and b.c.c. chromium and iron respectively. The values of AEFe0ff4Y in the temperature range between 0" and 900K ( .

Jan 1, 1960

By S. LeFevre

, The solution of the housing and sanitation problem in mining communities, keeping in view both economic and humanitarian aspects, demands the best thought of the management of such enterprises. Upon the solution depend the health and comfort of the employee and his family, and as a consequence his contentment and efficiency. The following description covcrs briefly the methods installed for Witherbee, Sherman & Co. The operation of the system has been for about two years under the direct supervision of T. A. Hammond, Land Agent, and I am indebted to him for a compilation of the results. Sanitation The sanitary service covers 238 tenements. The installation cost was about $22 per tenement, as follows: Privy, $15; privy box, $3; proportion of incinerator cost, $2; garbage can, $2. The operating cost is $1 per month for each tenement. In considering the installation of an ordinary village system of water supply and sewerage two difficulties were encountered: First, lack of a sufficient quantity of water; there is no stream large enough, as Mineville is situated very close to the divide between the Lake Champlain and Hudson River watersheds. A storage reservoir was tried, but the only location available, among glacial hills, would not hold water. Second, the expense of digging ditches for the distribution of water and sewer pipes is so great as to be prohibitive; Mineville is built chiefly in a valley covered in the bottom with drift, in this case a combination of boulders and hard pan. Digging in this material costs 80c. to $1 per cubic yard, and pipes must be placed 41/2 ft. down to be below frost. The houses are scattered along 4 miles of streets, the oldest having been built in the original forest 65 years ago. The newer houses have gone up a few at a time wherevcr a clear space could be found and the slopes were not too steep. This has advantages, as it separates the dwellings into

Jan 1, 1915

By I. R. Kramer

The effect of the specimen diameter, d, on the flow stress, cra of polycrystalline aluminunz (99.997) was studied. The increase in the flow stress could be accountedfor by the increase in the surface layer stress, with decreasing specimen diameter. Both , and a, were found to be proportional to For the smaller-dianzeter specimen (< 0.033 in.) at strains less than aboul 0.1, the work hardening of the surface layer was greater than that associated with the bulk of the specimen. At higher strains the work hardening due to the bulk appears to be independent of the specimen diameter. THE increase in the strength of metals with decreasing diameter is well-known; however, an adequate explanation for the cause of the size effect is still lacking. The earliest systematic investigation of size effect appears to be that of Onol who reported that for aluminum monocrystals the resistance to slip at low strains increased as the specimen diameter decreased. A change in the stress-strain curve beyond 0.001 strain was not found. However, Suzuki et a1 .&apos; reported for monocrystals of a brass and copper having diameters in the range of 2 to 0.12 mm that the entire stress-strain curve was raised as the specimen diameter was decreased. The effect of size was most apparent when the diameter of the specimen was less than 0.5 mm. In the discussion of this paper Honey-combe reported a size effect in copper crystals as large as % in. diam. These results are in agreement with those of paterson3 and Garstone et al.4 While the majority of the investigations on size effects was conducted in terms of the variation in the diameter of the specimen, several investigators studied the influence of the specimen geometry. For example, Wu and smoluchowski 5 reported that in aluminum monocrystals the slip system was a function of the specimen dimension in the slip direction. King-man and Green 6 studied the influence of size on the compressive stress-strain relationship of aluminum monocrystals when the ratio of length to diameter was constant. Their specimen diameters ranged from to & in. For specimens oriented for single slip the critical resolved shear stress for the smaller-size specimens increased with decreasing diameter. No effect was observed in the large-size specimens. Specimens having an orientation near the corners of the stereographic triangle did not exhibit a size effect. Apparently, the increase in strength with decrease in the diameter of the specimen is a general phenomenon and has been observed in a brass |T and cadmium as well as in aluminum and copper.&apos; In a series of investigations (for example Ref. lo), it was shown that during deformation a surface layer was formed which imposes a back stress, a,, on the moving dislocations. It is reasonable to predict that this surface layer stress, as, should be a function of the specimen diameter and could possibly account for the flow stress size effect. In fact, experimental evidence will be presented to show that this is the case; i.e., the increase in flow stress with decreasing size is equal to the increase in the surface layer stress, as, with size. In addition, data will be presented on the variation with size of and a* where is the back stress associated with the generation of dislocation obstacles in the bulk of the specimen and a* is the net effective stress acting on the mobile dislocations. A limited investigation was carried out on gold specimens to determine the influence of an oxide film. EXPERIMENTAL PROCEDURE The aluminum specimens were prepared from -in. bar stock (99.997 pct purity). The 0.350- and 0.150-in.-diam specimens were machined directly from the bars while the specimens having a diameter of 0.033, 0.020, and 0.015 in. were prepared by swaging and drawing to 0.04 in. and electropolishing almost to final size. The specimens were prepared with a 2-in. gage length. The specimens were annealed in vacuum (-10-4 Torr) at 350°C for 8 hr. The grain diameter of the specimens in the various specimen diameter groups was 0.08 ± 0.02 mm. Gold specimens of two diameters, 0.14 and 0.03 in., were prepared in a similar way and annealed at 650°C for 8 hr. The grain diameter of the gold specimens was 0.2 mm. After annealing the specimens were electrochemically polished to the final size and tested in an Instron tensile machine at a strain rate, E&apos;, of 10- 3 per min. While it was possible to determine the surface layer stress, a,, in the larger-size specimens by measuring the difference, Aa, between the stress before unloading the specimens and the initial flow stress after removal of the surface layer as outlined in detail in Ref. 10, this method is not applicable for small wires because of the difficulty in obtaining a sufficiently accurate measure of the diameter. The values at the various strains were therefore determined by measuring after the specimen had been annealed at 35°C for 4 hr. It has previously been shown" that the two methods give the same results for a provided that the annealing temperature is low enough to affect only the surface layer and not the dislocation barriers in the bulk of the specimen. For the gold specimens a treatment at 150°C for 16 hr was found to be satisfactory for the determination of by the low-temperature annealing method. EXPERIMENTAL RESULTS Determination of a,, and a,. The stress-strain curves for the various diameter aluminum specimens, plotted in terms of the logarithms of the true stress, and true strain, are given in Fig. 1. These curves represent the average data taken from at least ten specimens at each size. Over the range of strains investigated the curves follow the empirical equation

Jan 1, 1968

By G. Ostberg

OXYGEN has frequently been considered responsible for phenomena in the metallurgy of cast iron which are not sufficiently explained by the presence of other elements. Oxygen and oxides have been regarded as critical factors, particularly for the solidification process, and specifically in the formation of graphite. However, the evidence for an influence of oxygen has mostly only been indirect or qualitative. Furthermore, quantitative data on the oxygen content of cast iron presented in the literature have often been less conclusive as being apparently determined with insufficient regard for the difficulties associated with the sampling and analysis. These problems, experimental and theoretical, have been studied in this investigation. An attempt is made to bring the views on the soluble amount of oxygen in cast iron on a level with the present understanding of the thermodynamics of these types of alloys. Although there are no experimental data on the complete multicomponent cast-iron alloys some conclusions have been drawn from investigations on simpler systems. The order of magnitude of the solubility of oxygen in cast iron is considered in the first part of the paper. The second part deals with the oxygen content actually present under practical conditions. First the possible sources of errors in sampling and analyzing methods are examined. Finally some observations on the oxygen and oxide content of malleable iron during the production process are made. THERMODYNAMICS OF OXYGEN SOLUBILITY When cast iron is saturated with oxygen the saturation phase is essentially carbon monoxide or silicon dioxide depending upon the composition of the iron, the temperature, and the rates of formation of the oxides.* The latter occurs accordiw to the O + C = CO, Kc0 - a COa J 20 + Si = SiO2, Activities are used here instead of weight pct concentrations, since activities are believed to show more sensitively the effects of the interaction of all the alloying elements. Calculations involving activities are, however, seriously hampered by lack of experimental data on the interactions between the components encountered in the multi component cast-iron system Fe-O-C-Si-Mn-P-S- and so on. In order to make an estimate of the oxygen solubility reason-

Jan 1, 1959

The authors have done a very thorough job in analyzing the factors affecting turndown sulfur performance at Inland&apos;s No. 4 B.O.F. While many of the results are not unexpected, it is helpful I am sure to have quantified many of the relationships that the operators rely upon daily. On the other hand, however, the results shown for ladle skimming are very surprising. I do not doubt the statistical relationship shown but I feel, based on our experience at Youngstown, that some other parameter has been ignored in the data on ladle skimming. It was our experience that by skimming our hot metal transfer ladle we could reduce our lime consumption by eleven percent while obtaining sulfur removal comparable to that shown by the authors. I would like the authors comment of similar adjustments on the lime charge that were made on skimmed heats at Inland. If not, it would seem that one could gain more desulfurization in the B.O.F. by raising the hot metal manganese than by installing a skimmer. The author&apos;s state that no comparisons are valid between the data sets for the two time periods involved because of the difference in the consumption of No. 2 Bundles. I feel this in fact is one of the important conclusions to be drawn from the paper, as there seems to be approximately a 10-20 percent difference in percent of heats that are under 0.020 percent sulfur for similar conditions, which is an extremely important difference with today&apos;s demands for low sulfur. I would like the author&apos;s comments on the difference in the amount of 82 bundles consumed. Along this same line I was quite surprised by the relatively low number of heats that had a first turn down of 0.020 percent sulfur or less. Also rather disconcerting are the conditions necessary to achieve a first turn of 0.020 percent sulfur, 50 - 60 percent of the time, as required by today&apos;s specifications. It is a problem which we are well aware of, but the numbers are still shocking.

Jan 1, 1972

By D. A. Woodford

In a constant-load creep test, if the density of nlobile dislocations is assumed constant for strains exceeding that corvesponding to the minimum creep rate, it is shown that the creep rate may be approximated by the equation where E is the nominal strain, o0. is the initial stress, and a&apos; the stress fov unit dislocation velocity. The constant C contains the mean density of mobile dislocations and n is a measure of the stress sensitivity of dislocation celocity. This equation is shoum to be applicable to published data on an alloy steel and on nickel. The degree of plastic instability and hence the creep ductility is dependent on the calue of n for these materials. When instability becomes gross and local macroscopic necking occurs, the assumptions are no longer valid. On the basis of this equation, and the strain-time relation derived from it, a method of predicting the creep behavior at different stresses is suggested. MOst analyses of creep data have been concerned with time-dependent transient creep or time-independ- ent secondary creep in constant-stress tests. The far more widely used constant-load test has been neglected in terms of a mechanistic approach. In general, since the strains involved are usually small, the same laws for transient creep may be applied to both types of tests. However, there is no steady-state creep in a constant-load test since the stress increases continuously. Any apparent steady state observed on a strain-time plot is either fortuitous or illusory, and is generally revealed as a continuously varying creep rate on a carefully constructed strain rate vs time plot.* Transient creep may be understood in terms of a simple differential equation representing the change in mobile dislocation density with time:&apos; where p is the density of mobile dislocations. The first term on the right is related to first-order kinetics of dislocation multiDlication and the second to second-order kinetics of exhaustion. On the basis of this equation, several authors2-4 have found good agreement with experimental transient creep curves for a number of materials. When k1p = k2p2, a steady-state creep

Jan 1, 1968

By J. D. Mooney, D. L. Darnell

The accompanying chart has been drawn for the convenicncc of engineers as&apos; a means of quickly determining the correct number of plies of conveyor belts operating under specific conditions. The calculations are based on the average safe strength (factor of safety, 15) of the various standard rubber conveyor belts. The calculations assume maximum loading conditions; that is, the belt is considered as carrying the greatest load that it will handle without spillage at ordinary belt speeds. This not only produces the most economical operating conditions, but also the maximum tension in the belt. The chart is a graphical representation of the formula: Where, p = kg W(L + 10H) p = the correct number of plies k = a constant, depending on the type of drive g = the weight &apos;in pounds per cubic foot of material handled W = the width of the belt in inches L = the length of the belt in feet (approximately twice the center distance). H = the difference in elevation between the head and tail pulleys, in feet. For a simple drive, with a bare pulley, k = ----1 For a simple drive, with a rubber-lagged pulley, k = --- 1 For a, tandem drivej with bare pulleys, k = 1/375.000 For a tandem drive, with rubber-lagged pulleys, k = ---1 The chart is drawn for a simple drive, with a bare pulley (k — =e)1 ; therefore, the number of plies obtained from the chart should be multi- &apos; plied by the factor 0.83 or 5/6 for simple, lagged drive; the factor 0.67 or 2/3 for tandem, bare; and the factor 0.55 or 11/20, for tandem, lagged. The formula p = kgW(L + 10H) has been developed mathematically

Jan 1, 1916

By Robin O. Williams

Robin 0. Williams (Oak Ridge National Laboratory)— The authors have questioned the degree to which the coherency strains between the iron-rich and chromium-rich phases are isotropic as proposed in Ref. 5 on the basis of the difference between the elastic properties of the two phases. The relative magnitude of the stresses is determined by the moduli as shown by Eqs. [2], [3], and [4] of Ref. 34. However, the moduli of the two phases have no direct bearing on the uniformity of either the stress or strain within either phase. The idea that the strains are isotropic within each phase (but normally of different magnitude and always of different sign) is based entirely upon the experimental observation that X-ray line broadening has not been detected even when the particles become rather large. It has not proven possible to grow the particles sufficiently large that they lose coherency. Based upon this lack of line broadening one can estimate an upper limit for the nonuniformity of the strains within each phase as follows. It is considered possible to detect line broadening if it is as great as 10 pct of the separation of the K, doublet for the (211) line using chromium radiation. The doublet separation would correspond to a total strain of 0.0017 such that the total variation of lattice parameter relative to the average lattice is now k0.05x0.0017 or something less than ± * For the present case the strain in each phase is roughly 0.002 such that the variation of strain within a phase will not exceed 5 pct. It is stated that the expression derived for strengthening for the hydrostatic straining as observed in this system would substantially overestimate the magnitude due to dislocation flexure. This is contrary to the conclusion reached in the original paper34 for the present range of particle sizes. What is the lowest temperature at which a has been observed to form in this alloy? M. J. Marcinkowski, R. M. Fisher, and A. Szirmae (nutlzors&apos; reply)— -Williams&apos; arguments based on X-ray findings for a chromium-rich precipitate and an iron-rich matrix strained to a common lattice parameter are certainly convincing. This being the case, there are no shear components of strain associated with the precipitate-matrix aggregate to interact with the shear components of the dislocation stress fields, contrary to the opinion expressed by the present authors. On the other hand, the present authors, in spite of this error, did not expect the shear interactions to be significant. The chief objection to Williams&apos; model in the present case is that the various segments of the dislocation line are assumed to pass from one potential valley to the next independently of neighboring segments. This is only true for a highly flexible dislocation line, i.e., one whose radius of curvature is something less than the center to center distance between precipitate particles which amounts to about 90A in the present alloy. In order to maintain this curvature, an externally applied shear stress of at least 230,000 lb per sq in. would be required or about four times the observed stress. It is therefore concluded that the dislocation lines move rather rigidly through the lattice. This being the case, the forces on the dislocation resulting from the hydrostatic interaction between the stress fields of the edge-dislocation components and the precipitate particles should average out to zero; that is particles above the below the slip plane produce forces on the dislocation of opposite sign and therefore will cancel when averaged over the entire length of the dislocation. On the other hand, since the dislocation is not perfectly rigid, Williams&apos; model may lead to some strengthening, but far less than that predicted. A second and equally serious objective to using Williams&apos; strengthening model for the present alloys is that profuse wavy slip due to the motion of screw dislocations played a predominant role not only in the unaged alloys but in the fully aged ones as well. Since the screw dislocation has associated with it only shear components of stress the hydrostatic strengthening model no longer applies. In view of these arguments the present authors must reject Williams&apos; model of strengthening as being pertinent to the present alloy system. The present authors have made no detailed study of the lowest temperature at which a forms in the quenched ferritic alloys. None was ever observed n the alloys aged at 500°C so that forma-tion must occur at temperatures higher than this and was therefore not a factor in the present study.

Jan 1, 1965

By Y. Nakada, U. F. Kocks, B. Chalrners

THE orientation dependence of work hardening has previously been studied over the entire range, i.e., including special orientations of high symmetry, in aluminum1-3 and silver.* The differences between various orientations are substantial, and the trend is the same in all fcc crystals. However, there are quantitative differences in behavior between aluminum and silver at room temperature, particularly in the (100) orientation. While many experiments on gold have been reported,5&apos;9 none included the special orientations. We therefore undertook tension tests on gold crystals of the axial orientations (100),(110), (Ill), (211), and, as a representative of single slip, one whose Schmid factor was 0.5 (hereafter referred to as m = 0.5). Single crystals of dimensions 1/4 by .1/4 by 6 in. were grown from the melt1&apos; at a rate of 4 in. per hr, using gold of 99.99 pct purity obtained from Handy and Harman. A growth rate of 1/8 in. per hr, or a purity of 99.999 pct,ll produced no difference in results. The crystals were annealed at 1000°C in air for 24 hr and furnace-cooled down to room temperature, after which they were electro-polished in a solution of potassium cyanide (40 g), potassium ferrocyanide (10 g), soda (20 g), and enough distilled water to make 1 liter of solution, at a current density of 0.02 amp per sq mm. After 2 or 3 hr, a very smooth surface was obtained by this method. Nine m = 0.5, two (loo), one (110), three (Ill), and one (211; crystals were tested at room temperature in a floor-model Instron machine at a tensile strain rate of 0.5 pct per min. The accuracy of the stress measurement was k10 g per sq mm up to 500 g per sq mm, k2 pct for higher stresses. The strain measurement was accurate to t2 pct. The scatter of stress at a given strain among the crystals of the same orientation was small, *5 pct being the largest. The representative stress-strain curves for various orientations are shown in Fig. 1. Table I summarizes the work-hardening parameters as used by seeger.12 Results of other investigators are also included in this table for comparison. There are no previous data for the corner orientations. Values for m = 0.5 crystals agree fairly well with those of Berner. Tm is defined as the stress at which the stress-strain curve begins to deviate from linearity of Stage 11. However, in practice this is a very difficult value to estimate because each investigator has a different idea of linearity. Therefore, the comparison with the values of other investigators may not be valid. In the present experiment, (100) crystals had the highest 111, followed by (111) crystals. The work-hardening rate in Stage I1 was highest for the (111) crystal followed by the (100) crystals. Our value for 0x1 of m = 0.5 orientation agrees very well with those of other investigators. 1) Single-Slip orientation (m = 0.5). These crystals were oriented so that the primary-slip vector was contained in one side face. The dimension perpendicular to this side should then not change if single slip indeed takes place. Within the accuracy of measurement, this dimension did not change during the deformation. Since the accuracy is k0.2 pct, the amount of secondary slip is less than 0.7 pct of the slip on the primary-slip system at 30 pct tensile strain. This is in conformity with the results obtained by Kocks" for aluminum crystals. The tensile axis moved toward (211) during the deformation. Slip bands, Fig. 2(a), were very fine and closely spaced. Some deformation bands were observed. There were no clear-cut cross-slip traces such as the ones observed on aluminum m = 0.5 crystals. 2) (111) Crystals. The orientation of the tensile axis was stable during the deformation. From this observation, one can deduce that at least three slip systems must have operated, and probably all six because the remaining three all have cross-slip relation to one of the first three.&apos; It was very difficult to observe the slip markings. Consequently, we could not confirm by this method that six systems were operative during the deformation. At high strains (above 50 pct shear strain), this orientation had the highest stress-strain curve. At 80

Jan 1, 1964

By Carl Gatlin, S. Srini-Vasan

It is shown thar k k ratios cornputed on the as.sutnption that the average sandstone reservoir has a pore-size distribution characrerized by a linear 1/P saturatioN plo are in good agreewient with correspotlding k,,/k,, ratios obfairzed by Wahl, Mul-/ins and Elfrink by correlatiotz of field measurements. Methods of calculating relative permeability data have recently been reviewed and extended.&apos; It was shown that for sandstone reservoirs where &apos; is approximately a linear function of effective saturation (as suggested first by Corey&apos;) it is possible to express relative permeabilities in very simple terms. For systems wetted by connate water and producing oil and gas (gas saturation growing at the expense of oil) the expression for the The accuracy of Eq. I has been frequently confirmed in our laboratories; because of its simplicity we make routine use of it in the analytical solution of proble~ns in reservoir engineering. Nevertheless, comparisons of the k,,,/k,, ratios yielded by Eq. 1 with appropriate reservoir data have been few. The publica- tion by Wahl, Mullins and ElfrinkV of an empirical equation for k,,/k,, ratios is thus of interest since this equation is specificalljl stated to be Vased on the correlation of field measurements made on sandstone reservoirs. Not surprisingly these authors found that it was necessary to

By M. C. Fuerstenau, W. F. Cummins

The flotation response of quartz in the presence of calcium chloride and various fatty acids is presented. Flotation is not effected with lauric acid until calcium laurate is precipitated in solution, while flotation is obtained with the 9, 10, and 11-carbon members before precipitation of the calcium carboxylate. The active species responsible for flotation is suggested to be a basic aqueous complex, e.g. Ca(OH) (RCOO) (aq), and a possible mechanism of adsorption is given. Most of the previous work on activation in non-metallic flotation systems has involved the use of high molecular weight collectors, such as oleic acid or alkyl aryl Sulfonate.1-7 In these studies the levels of addition of collector and activator required for flotation were much above the concentrations necessary for precipitation of the metal-collector salt. In fact, Clemmer, Clemmons, Rampacek, Williams, and Stacy1 have stated, "Our experiments on a variety of materials have demonstrated that the metal salts which react with anionic collecting agents, as soaps and alkyl sulfates, to form relatively insoluble compounds may be employed as activating agents provided an optimum quantity is used at a proper pH." On the other hand, Gaudin and chang8 obtained flotation with additions of barium and lauric acid below the level required for barium lautate precipitation.

Jan 1, 1968