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By Myron Rodgers

IN rock- or ore-crushing tests all data, in order to be valuable for study and comparison, should be obtained and tabulated under conditions as uniform as possible. The results of many such tests have no value for comparison, because of incomplete data and the lack of uniform conditions, screens and scale of sizing, etc. The purpose of this paper is to suggest a form of standard specifications which may be developed by the Milling Committee and approved by the A. I. M. E. for rock-crushing tests. It is not intended to discuss any theories of rock crushing. Reports of rock-crushing tests should include the following details: 1. Description of the machine employed (jaw or gyratory crushers, rolls, stamps, tube mills, Chilean mills, etc.). 2. Method and material (timber, concrete, etc.) of foundation. Much power is dissipated in the vibration of poor foundations. 3. Locality from which the rock or ore was obtained, and geological, mineralogical, and physical characters of the material. 4. The power consumption of the machine running with no load and with full load, the unit of power being 1 hp, per 24 hr. 5. The capacity of the machine in tons (of 2000 lb.) per 24 hr. 6. The duty in tons per horsepower-day. 7. The screen analysis of feed and product by the proposed A. I. M. E. standard screen scale. On the foregoing statement, the following remarks are offered.

Jan 9, 1915

By T. R. Cass, M. R. Achter

A technique for measuring the creep and rupture strength of nickel specimens bonded by sintered oxide layers has been developed for the investigation of the role of grain-boundary oxide in the oxidation strengthening of nickel during creep at 817 °C. Prom the inverse dependence of the rate of bonding on the thickness of the starting oxide, it is suggested that the rate of diffusion of Ni++ through NiO is the controlling process at the start of sintering. Cornparisons of the creep and rupture strength and of the microstructure of sintered couples with conventionul creep specimens suggest rupture in nickel is prevented by load-bearikzg, ikltergranular layers of oxide. A number of recent investigations have shown that at high temperature and low stress the creep-rupture strength of nickel and nickel-base alloys may be much greater in air than in an inert gas or vacuum.'-7 For example, at 817°C and 2500 psi, nickep failed in vacuum in 48 hr but not in air after 4000 hr, and a Ni-Cr-A1 alloy5 at 1038°C and 3000 psi failed in vacuum in 4 hr but not in air after 731 hr. Fig. 1 is a typical creep curve in air for materials experiencing this pronounced environmental strengthening. A common feature of these curves is the entry into a third stage of accelerating creep, concurrent with the for-

Jan 1, 1962

By R. C. Buehl, J. P. Morris

F. D. Richardson—The authors are to be congratulated on this further contribution to our knowledge of the thermodynamics of the interaction between sulphur and carbon and silicon in liquid iron. As the authors state, the influence of carbon and silicon on the activity coefficient of sulphur in liquid iron is clearly of great importance in the blast furnace, since it must cause a three to fourfold improvement in the partition of sulphur between slag and metal. The influence of increasing temperature in further increasing the activity coefficient of the sulphur in the metal in the blast furnace by increasing the carbon content is also of interest. This effect, however, is probably only part of the reason for the general observation in blast furnace practice, that the sulphur content of the metal is lowered by increasing temperature. Other contributing factors are the lowering of the oxygen potential in the presence of carbon by increasing temperature and the probable increase in the activity coefficient of the lime in the slag for the same reason. The former of these effects, which works via the (CaO) + [S] = (CaS) + [O] equilibrium, might possibly account for a 70 pct improvement in the sulphur partition and the latter might give a further 50 pct improvement. C. Sherman—I would like to compliment the authors on their very careful research. If I may, I would like to show results of calculations on the carbon-sulphur-iron system similar to the ones that were shown in our paper for the silicon-sulphur-iron system. For Fe-S-C ternary system k=PHgs/PH2 x 1/(f1°) (f2°) (%S) where fs = sulphur activity coefficient fs' = fs for Fe-S system of equal pct S f3° = f2/f2 for Fe-S-C ternary system This same analysis has been used on other systems, but the results shown in fie.- 7 are for carbon and silicon. L. S. Darken—I would like to make two brief comments in addition to complimenting the authors on an apparently very precise and accurate investigation. The first is that the present work is in agreement with a calculation by Larsen and myself." Our calculation (much less precise than the present work) was based on: (1) Unpublished work on the sulphur content of molten iron (1.5 pct at 1500°C) in equilibrium with graphite and an iron sulphide slag; (2) the distribution coefficient of sulphur between slag and carbon-free liquid iron. We expressed the result in a form equivalent to log 7. = 0.18 [%C] which gives an activity coefficient (?s.) of sulphur only slightly higher than the authors find and certainly within the precision of the earlier work. My second comment concerns the correlation of the thermodynamic findings with atomistics. A rough pic- ture of the atomic arrangement in the liquid solution is rather easily conceived for this particular liquid solution containing iron, carbon, and sulphur. Carbon has a very much stronger affinity for iron than for sulphur. Hence we may conclude that a sulphur atom will but seldom be adjacent to a carbon atom—since this would be a position of high energy. From the metallic radii of iron and carbon we know that six iron atoms pack neatly around one carbon atom. Thus each carbon atom in retaining this shell of iron atoms (which latter may not be replaced by sulphur on account of the high energy requirement) decreases the available positions for each sulphur atom by six. Hence each atomic percent of carbon decreases the equilibrium sulphur content by 6 pct (of itself). Or, at low concentration each atomic percent of carbon increases the activity coefficient of sulphur by 6 pct. This is in good agreement with the observed increase (6 or 7 pct at low carbon content). It is indeed gratifying to find a case where, by such simple reasoning, quantitative agreement is found between precise data and the modern picture of the atomistics of the metallic state. J. P. Morris (authors' reply)—We would like to point out that there is an error in the equation on p. 322 of the paper. The third equation should read: ½S2 (gas) + H2 (gas) = H2S (gas) The authors wish to thank everyone for the interest they have shown in the paper. In regard to the general observation in blast furnace practice, that the sulphur content of the metal is lowered by increasing the temperature, Dr. Richardson is correct in stating that the cause can be attributed only in part to the increase in activity coefficient of sulphur resulting from the rise in carbon plus silicon content of the metal with rise in temperature. However, this factor is probably an important one. The results of one experiment, performed since this report was written, indicate that at a constant temperature the addition of silicon to a melt saturated with carbon causes an increase in the activity coefficient of sulphur even though the carbon solubility is lowered. In this test, 2.5 pct silicon was added to a melt saturated with carbon and maintained at 1400°C. Although the carbon content dropped from 4.85 to 4.1 pct, the activity coefficient of sulphur was increased by about 20 pct.

Jan 1, 1951

By J. W. MacDonald

Review of equipment, performance, and shaft design plans including head-frames for sinking and forms for lining are given. Also included are comparative data for shafts of 11 x20 ft and 193/4 x 30 ft with 756 and 666-ft depths, respectively. Advantage from curtain wall installation with "slip forms" which warrant consideration for normal lining are outlined. A preliminary review is also made of "chemical grouting" to stop water inflow leakage with Halliburton equipment and service. Two coal mine shafts recently completed in Franklin County, Ill., afford experience and accomplishment to warrant review. These shafts provide access to a tract with a potential recovery of one hundred million tons. This work provided the initial sinking by other than a contractor's organization since 1918. Four additional shafts required during the interim were sunk under contract arrangement with others. Supervisory contact and inspection afforded experience to govern planning and conduct of the recent work. TEST DRILLING Core test holes were drilled within each of the two shaft locations. Four additional cable tool holes were drilled on a radius of 700 ft from the shaft bottom to determine the coal seam elevations and govern the underground layout plans for each shaft. A graphic log of the hoist shaft test drillhole is shown by Fig. 1. A very soft sandstone was found at a depth of 16 ft. Cored recovery of the sandstone proved much weaker than the overlying compacted sand secured with coring facilities for soil recovery. Gradation from the very soft sandstone to fairly hard material occurred with the next 14 ft, extending to a depth of 30 ft. The weak character within the upper portion entailed need of adequate protection when sinking as described later. A part of the sandstone provided sufficient porosity and permeability to warrant further investigation. Adjacent cable tool drilling for temporary water supply disclosed inflow was limited to 2 gpm with a depth of 40 ft. It was increased to 12 gpm during the next 50 ft of drilling and occasioned subsequent precaution when sinking. General arrangement of facilities for sinking the hoist shaft is shown by Fig. 2. An available steel frame building was used to accommodate the hoist, compressor, shop, and shower or change room. The stiff leg derrick with 90-ft length of boom permitted removal of forms from the wood shop to the storage area or delivery within the shaft. Reinforcing steel and other needs were similarly handled, in addition to headframe erection. The headframe was placed over one end of the shaft to enable passage of forms for lining through the adjacent half next to the stiff leg derrick. The steel headframe as shown by Fig. 3 was fabricated locally. Side anchored cables from the

Jan 1, 1961

By M. M. Singh, J. W. Brown

Three rocks of highly varied physical properties are selected for the first reported experimental study concerning microseismic activity of rock under tension to determine the effects of static fatigue, loading rate, repeated loading and rock anistropy compared to compression and bending loads. The microseismic method of detecting rock under deformation may be a great potential tool in surveillance and location in situ of dangerously stressed strata, as well as predicting rock failures. Noises emitted under load or the rate of energy (energy of noise pulses generated per unit time) are computed for varying conditions. The tests are conducted at constant load on a dead weight loading machine to eliminate all possible extraneous noises — an indispensable precaution — at various load levels. Experimental results are reported including the relation of total energy emitted to tensile strength of rock and the possibility of simulating a phenomenological model for microseismic activity. Plots of some of the experiments are included. INTRODUCTION Rock failure, whether on a large scale faulting in the earth's crust and rock bursts in mines, or on small scale ruptures of laboratory test specimens, generates transient vibrations in the rock mass. Mechanisms of intermittent deformation, such as twinning and cracking through and between mineral grains, also produce similar vibrations, but of very small amplitude. Such vibrations from both deformation and failure are referred to as microseismic activity, rock noise, or occasionally seismo-acoustical activity. An individual rock noise is sometimes

Jan 1, 1967

By Claudio Barsotti, Leonard C. Kitchener

INTRODUCTION The major mining operations of the Ontario Division of Inco Metals Company are located in the Sudbury Basin. Around the rim of this geo- logical formation and within thirty miles of the City of Sudbury are located eleven under- ground mines of the Company, two of which are on standby. In addition, there is one mine in Northwestern Ontario at Shebandowan, 1127 Km (700 miles) northwest of Sudbury. In 1980 ore production totalled 10.6 x 10 mt (11.7 million tons) from underground and open pit operations. Underground mining methods include blasthole, vertical retreat, sublevel caving, cut-and-fill and undercut-and-fill. Inco reduction operations in the Sudbury area include two concentrators, one smelter, one iron ore recovery plant, a copper refinery a nickel refinery and a high purity plate rolling mill. As well, an electrolytic nickel refinery is located on Lake Erie at Port Colborne, near Niagara Falls. In the Sudbury District, Inco employees total 14,278 of whom 6,490 are directly employed in the mining and milling operations. I T H DRILLING Background - The use of long blastholes in underground operations at Inco mines in the Sudbury district started more than 30 years ago at the Levack Mine. At that time, diamond drills were used to mine ore by open stoping and for pillar recovery. By the 1950's the concept had been extended to all mines. The heavy percussion drill was also developed by this time and led to the acceptance of the 11.43 cm (4% in.) piston diameter drifter for bar and arm and fan drill applications. The use however, of the nominal 5.08 cm (2 in.) diameter blastholes drilled radially from the level or sublevel causes overdrilling at the periphery. As well, the smaller holes are subject to hole deviation, are difficult to drill beyond 38.1 m (125 ft.) and require the need for several drilling horizons or sublevels. The need to improve blasthole stoping techniques and our experience with large diameter blastholes in open pit mining led to the trial of an I T H drill in 1962. The test was unsuccessful however, since button bits were not available at the time and the tungsten carbide cross bit experienced short life, gauge loss, and caused hole deviation and poor penetration rates. The successful application of the raise borer beginning in 1964 and the ability to drill 20.32 cm (8 in.) to 27.94 cm (11 in.) diameter pilot holes led to the development of smaller, more portable, rotary drills for blasthole mining. As a result, a mining method based on the use of 20 cm (7 718 in.) diameter, 60.1 m (200 ft.) long rotary blastholes was developed and tested successfully at our C.C.North Mine. Between 1972, when the method was first tested, and 1973 over 4572 m (15000 ft.) of large holes were drilled with rotary drills in the recovery of 175,960 mt (194000 tons) of primary stoping and pillar remnants. A new mining method was developed that made use of this large hole drilling equipment. Instead of longhole carbide drilling by bar-and-arm or fan drill set-ups at sublevels, the large diameter holes were drilled full stope height from the top sill, which substantially reduced development, drilling time and improved fragmentation. These drills however, were difficult to move and set-up, and had a high maintenance cost and low productivity. The capital outlay and extensive technical development required to improve these drills and reduce the high drilling costs however, was economically unattractive. New development in drills and button bits INTRODUCTION The major mining operations of the Ontario Division of Inco Metals Company are located in the Sudbury Basin. Around the rim of this geo- logical formation and within thirty miles of the City of Sudbury are located eleven under- ground mines of the Company, two of which are on standby. In addition, there is one mine in Northwestern Ontario at Shebandowan, 1127 Km (700 miles) northwest of Sudbury. In 1980 ore production totalled 10.6 x 10 mt (11.7 million tons) from underground and open pit operations. Underground mining methods include blasthole, vertical retreat, sublevel caving, cut-and-fill and undercut-and-fill. Inco reduction operations in the Sudbury area include two concentrators, one smelter, one iron ore recovery plant, a copper refinery a nickel refinery and a high purity plate rolling mill. As well, an electrolytic nickel refinery is located on Lake Erie at Port Colborne, near Niagara Falls. In the Sudbury District, Inco employees total 14,278 of whom 6,490 are directly employed in the mining and milling operations. I T H DRILLING Background - The use of long blastholes in underground operations at Inco mines in the Sudbury district started more than 30 years ago at the Levack Mine. At that time, diamond drills were used to mine ore by open stoping and for pillar recovery. By the 1950's the concept had been extended to all mines. The heavy percussion drill was also developed by this time and led to the acceptance of the 11.43 cm (4% in.) piston diameter drifter for bar and arm and fan drill applications. The use however, of the nominal 5.08 cm (2 in.) diameter blastholes drilled radially from the level or sublevel causes overdrilling at the periphery. As well, the smaller holes are subject to hole deviation, are difficult to drill beyond 38.1 m (125 ft.) and require the need for several drilling horizons or sublevels. The need to improve blasthole stoping techniques and our experience with large diameter blastholes in open pit mining led to the trial of an I T H drill in 1962. The test was unsuccessful however, since button bits were not available at the time and the tungsten carbide cross bit experienced short life, gauge loss, and caused hole deviation and poor penetration rates. The successful application of the raise borer beginning in 1964 and the ability to drill 20.32 cm (8 in.) to 27.94 cm (11 in.) diameter pilot holes led to the development of smaller, more portable, rotary drills for blasthole mining. As a result, a mining method based on the use of 20 cm (7 718 in.) diameter, 60.1 m (200 ft.) long rotary blastholes was developed and tested successfully at our C.C.North Mine. Between 1972, when the method was first tested, and 1973 over 4572 m (15000 ft.) of large holes were drilled with rotary drills in the recovery of 175,960 mt (194000 tons) of primary stoping and pillar remnants. A new mining method was developed that made use of this large hole drilling equipment. Instead of longhole carbide drilling by bar-and-arm or fan drill set-ups at sublevels, the large diameter holes were drilled full stope height from the top sill, which substantially reduced development, drilling time and improved fragmentation. These drills however, were difficult to move and set-up, and had a high maintenance cost and low productivity. The capital outlay and extensive technical development required to improve these drills and reduce the high drilling costs however, was economically unattractive. New development in drills and button bits

Jan 1, 1981

By W. A. Klikoff, I. Fatt

The concept of fractional wet wattability is examined. Fractional water wettability of a reservoir rock is defined as the fraction of the internal surface urea that is in contact with water. Capillary pressure and relalive permeability of unconsolidated sand are shown to be functions of fractional wettability. INTRODUCTION The petroleum industry has long recognized that wettability of reservoir rock has an important effect on multiphase flow of oil, water and gas through reservoirs. As early as 1928 the American Petroleum Institute sponsored a study of wettability as part of API Project 27 at the U. of Michigan.' Despite 30 years of research, there is still little exact knowledge of the wettability of reservoir rocks. There are two parts to the wettability problem. After agreeing to a uniform nomenclature in regard to wettability,' the first question to be answered is, "What is the in situ wettability of a given reservoir rock?" If this can he answered the next question is, "What part does wettability play in determining the characteristics of multiphase fluid flow through the rock?" This paper represents an oblique attack on the problem of wettability. No attempt is made here to answer the basic question of wettability in situ. instead the consequences of the concept of fractional wettability are examined. Multiphase flow in sandpacks is shown to he highly influenced by fractional wettability. Jennings' has given 3 definition of wettability and the other terms used in discussing wettability. These terms must be applied to the physical situation existing In reservoir rock. A survey of the pertinent literature from 1928 to 1956 indicates that the concept of a contact angle was applied to reservoir rock in the same way it would be applied to a flat, homogeneous surface. Attempts were made to state quantitatively the wettability of a reservoir rock in terms of a contact angle which was presumably constant at all points on the very rough and heterogeneous interior surface of a porous rock. Calhoun, et al,3-6 prepared synthetic consolidated and un-consolidated porous media in which they claimed there was a known uniform contact angle. They then showed the effects on the capillary pressure and relative permeability characteristics of varying this angle. The API Project 47 at the U. of Texas' and others' have made extensive studies of an indirect approach to the contact angle through the use of heat of wetting data. Even if successful. however, this approach also states the wettability of porous rock in terms of a contact angle which is uniform over the entirc surface. If the angle varies from one part to another on the internal surface, there is no way of determining from the measurements the area distribution of contact angles. In 1956 Brown and Fatt5 suggested that the concept of a contact angle, as applied to reservoir rock, be abandoned. This suggestion was made because it is known that the internal surface of most reservoir rocks is composed of many different minerals, cach with a different surface chemistry and a different capacity to adsorh surface active materials from reservoir fluids. Furthermore, the operation of a contact angle in determining the form of a fluid-fluid interface is difficult to picture in the very complex geometry of a pore. Brown and Fatt proposed that the wettability of reservoir rock be stated in terms of the fractional internal surface area that is in contact with water or oil. All surfaces on which there is water are called water-wet; surfaces on which there is oil are called oil-wet. The fractional water wettability is then stated as a number which represents the fraction of the internal surface that is in contact with water. A symmetrical statement can be made for the fractional oil wettability. The concept of a fractional wettability as previously stated has in its favor the recognition of the heterogeneous mineral composition of most reservoir rocks. Another point in its favor is that fractional wettability can be measured quantitatively with relative ease. Hol-brook and Bernard10 se a simple dye adsorption test to obtain fractional wettability of reservoir rocks. Amott11 uses a combination of imbibition and displacement to arrive at a wettability index of reservoir rocks which seems related to fractional wettability in the range 0.25 to 0.75 fractional water wettability. Jennings" has shown the changes in relative permeability that take place when a porous material is changed from unity to zero fractional water wettability. He also shows that reservoir rocks in their natural state, but at room temperature and atmospheric pressure, have relative permeability characteristics which would indi-

By William T. Folwell

The Lucky Friday mine east of Mullan, Idaho, is an outstanding example of a property in the Coeur dlAlene district where a small and insignificant-appearing silver-lead-zinc vein at the surface has changed at depth into a large vein of great importance. The Lucky Friday vein has little if any surface expression and above the 1200 level the ore shoots are small and discontinuous. Between the 1200 level and 2450 level, the lowest developed level, the main ore shoot has shown remarkable improvement on each succeeding lower level, and today the mine is one of the major lead-silver producers in the Coeur d'Alene district. History: The Lucky Friday property is on the north side of the South Fork of the Coeur d'Alene River in sections 25, 26 and 35, T. 48 N., R. 5 E., Hunter mining district, Shoshone County, Idaho. This is about one mile east of the town of Mullan, which serves the eastern portion of the Coeur d'Alene district. The southern part of the property is crossed by a branch line of the Northern Pacific Ry. and by U. S. Highway 10. This highway is the principal road crossing the panhandle of Idaho and connects the district with Spokane, Wash., on the west and Missoula, Mont., on the east. The main portal and surface plant of the Lucky Friday mine is at an elevation of 3365 ft, only a short distance above the valley floor and a few hundred feet from U. S. Highway 10. so the mine is readily accessible for year-round operation. The property is comprised of six claims, known as the Lucky Friday group, owned outright by the Lucky Friday Silver-Lead Mines Co. There are four patented claims, Good Friday, Lucky Friday, Northern Light, and Lucky Friday Fraction No. 2 (Mineral Survey No. 3028), and two unpatented claims, Hunter and Creek. In addition, Lucky Friday owns an undivided one-half interest in the Hunter Creek property. which adjoins the Lucky Friday group on the north: a 90 pct interest in the mineral rights in the Jutila Ranch (160A), which adjoins the Lucky Friday group on the east; and a 60 pct interest in the Lucky Friday Extension claim group, which adjoins the Lucky Friday group on the west. The company also has a long-term mining lease on the Hunter Ranch, which adjoins the Lucky Friday group on the west. The claims of the Lucky Friday group were located between 1899 and 1906. The Lucky Friday Mines Co. was organized in 1906 and did considerable exploration work by surface trenching and shallow underground workings, only to see the property sold by the Shoshone county sheriff to satisfy labor claims totaling $2000 in 1912. Another firm, Lucky Friday Mining Co., bought the claims in 1914 and spent 12 years driving what is now known as the tunnel level crosscut. This tunnel intersected a vein previously exposed in a higher tunnel, but it was only a few inches wide. The vein was followed a short distance westerly but was so unpromising that the work was discontinued in favor of extending the main crosscut tunnel several hundred feet north. No ore was found and all work was discontinued. The property was held in such low esteem by the firm that it let taxes amounting to less than $15 a year go delinquent for nine years. Then the property lay idle for two more years until in 1938 John Sekulic, a Mullan service station operator, took a lease, with a $15,000 purchase option, on the advice of an old miner who had worked in the mine. Sekulic re-opened the tunnel level crosscut and explored the vein with an easterly drift for about 200 ft. The vein was too narrow to be of commercial value but was believed interesting enough to warrant further exploration at depth. Lacking funds to explore the vein at depth, Sekulic tried to get the district's larger operating companies to take over his lease and option. They were not interested because of the lean tunnel level showing and the fact that the property lay between the White Ledge fault on the north and Osburn fault on the south, an area which geologists always considered unworthy of exploration. Sekulic then organized the present company and assigned his lease and option to it for stock. This was in 1939. Enough stock was sold locally to finance sinking of a shaft 100 ft from the tunnel level east drift. The vein at this additional depth still was not commercial but showed some improvement. Treasury stock was offered at 5 to 10C a share

Jan 1, 1959

By D. R. Irving

Prior to this time, ample supplies of high grade mineral fillers, such as clay, have been available close to consuming centers. Now depletion of these accessible deposits, coupled with other factors such as increasing demand due to population expansion and standard of living rise, has made development of upgrading techniques attractive to producers. Various investigations being carried out by the U.S. Bureau of Mines are described. Factors making the development of improved methods for upgrading clays attractive to producers include: depletion of accessible deposits of high-grade clays, rising transportation and labor costs, more exacting specifications, and increased demand resulting from an expanding population and a constantly rising standard of living. Industry recognizes the need for upgrading presently unusable materials and has cooperated with the U.S. Bureau of Mines by providing samples from submarginal deposits and by evaluating the products that USBM upgrades. Following are some examples of recent clay beneficiation work by the Bureau. FIRE CLAY BENEFICIATION At the Rolla Metallurgical Center, Rolla, Mo., the Bureau is conducting research to develop technically and economically feasible mineral-dressing methods to remove quartz, pyrite, and other impurities from submarginal fire clays. The research has been spurred by the growing shortage of high-grade bur-ley, flint, and plastic clays in the Missouri fire-clay district resulting from increased consumption and the low rate of discovery of new deposits of adequate quality. There apparently is ample tonnage of submarginal flint and plastic clays to maintain the industry for many years if such material can be upgraded to commercial quality. Samples have been submitted to USBM by several major producers of refractories, including A. P. Green Fire Brick Co.; Harbison-Walker Refractories Co.; Mexico Refractories Div., Kaiser Aluminum and Chemical Corp.; and Wellsville Fire Brick Co. Research on some samples has been completed and still is in progress on others. The Bureau has found that specimens containing quartz or pyrite in grains coarser than the accompanying clay generally can be upgraded either with a gravity table or hydraulic cyclone. The improvement in quality is notable not only in chemical analysis but in an increase in pyrometric cone equivalent (PCE) of one or more cones. As an example, consider the results attained on a sandy flint—a fine-grained mixture of kaolin and halloysite with quartz and minor quantities of iron and manganese oxides. Chemical analysis was: 35.4 pct alumina (A12O3) and 47.8 pct silica (SiO2). The sample contained 9.2 pct free silica as quartz. The PCE was 33. The sample was crushed to -10-mesh, blunged with water, and then stage-ground to -48-mesh. The ground sample was tabled to yield a clay fraction containing 37.6 pct A12O3, and 45.3 pct SiO2 of which 2.4 pct was free silica. The PCE was 34. Recovery of alumina by this process was 73.6 pct. Concentration of the same clay in the hydraulic cyclone was even more effective in respect to quartz removal. The clay fraction analyzed 38.6 pct A12O3, 45.0 pct SiO2 of which 0.5 pct was free silica as quartz, and had a PCE of 34. Recovery of alumina was 71.9 pct. The hydraulic cyclone was also effective in reducing the pyrite content of a plastic clay which likewise was a mixture of kaolin and halloysite, plus some pyrite, quartz, and carbonaceous material. It analyzed 30.6 pct Al2O3, 52.3 pct SiO2, 2.3 pct Fe2O3, and 0.76 pct S. The sample was blunged and screened through 10-mesh to remove some coarse pyrite, then ground to -65-mesh and cycloned to remove most of the remainder. The final product recovered 81.8 pct of the alumina at a grade of 32.3 pct A12O3, 50.9 pct SiO2, 1.4 pct Fe2O3, and 0.1 pct S. The PCE was 32-1/2, compared with a PCE of 31-1/2 for the crude clay. The fine beneficiated clays, of course, must be pelletized and dried or fired to meet commercial acceptance. This operation has already been solved by industry. Little progress has been made on two remaining problems. One is that of fine-grained impurities, not separable by table or cyclone. Flotation is being investigated to solve this. Although fine pyrite is readily rejected, a satisfactory quartz-clay separation has not been attained. The second problem, a stubborn one, is that of alkali. Samples containing as much as 5 pct potas-

Jan 1, 1961

By H. Rees

This paper supplements and brings up to date a report prepared earlier this year and presented at the Eleventh Pan American Railway Congress in Mexico City. Most of the original report is contained in this paper in order to acquaint the reader with the project background. On December 1, 1961, construction of this locomotive was completed. Stationary testing started immediately, and a host of problems in various areas were uncovered. Many of these had no relation to the use of coal in a direct fired gas turbine. All of them emphasized the problems which plagued previous projects of this kind. In 1944 the Locomotive Development Committee of Bituminous Coal Research, Inc., was formed to study the problems of combustion, coal handling, and fly ash removal which would have to be solved before coal could be used to fire a gas turbine. By 1947 this committee was composed of the chief executive officers of nine eastern railroads and five eastern coal companies. All of the members of this committee were vitally concerned in a coal fired gas turbine because coal cost only one-fourth as much per heat unit as fuel oil. A number of research projects were started to develop and test components on a small scale. This included work at John Hopkins University, Battelle Memorial Institute, Institute of Gas Technology at Chicago, Purdue University, and Southern Research Institute. By 1946 a pilot plant had been installed at the Dunkirk, New York plant of Alco Products Div. of American Locomotive Co. The U. S. Bureau of Mines made available to the Locomotive Development Committee (hereafter called LDC) a Houdry process gas turbine plant, originally destined for a Russian oil refinery. The full scale pilot plant was used to evaluate the results obtained from the small scale studies under actual gas turbine conditions. Operation of this plant disclosed a number of problems which had not been discovered in the small scale tests. By 1951 four 250 hr tests had been conducted with the Houdry gas turbine to test out a variety of equipment. The major problem concerned the removal of abrasive fly ash from the products of combustion so that wear on turbine blades would not be excessive. The first 250 hr test showed that the original fly ash separating equipment was inadequate. Large ash particles passing through the turbine blading were nicking the leading edges of the first row blades. At the conclusion of the second 250 hr test, erosion was so severe that the first row of stator and the first row of rotor blades were replaced. At the end of the third 250 hr test, it was necessary to replace the first three rows of stator blading. The fourth test was used to test out a new design fly ash separator developed by LDC and named the Dunlab tube. This device gave more encouraging results than any of the others tested. Early enthusiasm for overcoming major difficulties had resulted in placing an order in 1946 with Allis-Chalmers Mfg. Co. for a 3750 hp regenerative gas turbine. The gas turbine, together with associated combustion and ash removal equipment, coal handling equipment, generators and gear reduction unit was installed at Dunkirk by 1951. A 750 hr test was made in 1952 with the LDC-Allis Chalmers gas turbine power plant. At the end of this test it was necessary to replace the first four rows of both rotor and stator blading. Rows 5 and 6 did not require replacement. In 1953 the American Locomotive Co. joined with LDC and assumed an active role in equipment development and manufacture. During 1954 LDC-Alco operated the rebladed gas turbine 300 hr to determine if a new fly ash separator was adequate. Performance of the improved separator was considerably better, but trouble was experienced with clogged blowdown lines from the Dunlab tubes. When these lines clogged it permitted the passage of abrasive particles. In 1955 a total of 1421 hr was run. This included 243 hr of cycle operation on Union Pacific coal from the D. O. Clark mine at Superior, Wyoming. During the 243 hr a total of 370 tons of coal was consumed. The purpose of this test was to determine if U.P. coal was suitable for use in a coal fired turbine. It was the unofficial opinion that it was as good or better than the eastern coals that had been tested.

Jan 1, 1964

By Hokuichiro Ohmachi

INTRODUCTION Metallic minerals have been formed only through complex geologic processes which took place at certain stages of the earth's histrory. Their concentration, abundance, and distribution are, therefore, restricted geologically, and very small in global scale. Ore is a mineral deposit or mineral concentration from which metals can be economically extracted by the contemporaneous technology. By this definition, several factors should be considered before any mineral deposit is regarded as an ore. The grade of mineral concentration and the scale of the reserve are most important. Demand for the metal and its price are the other factors. Advancement of technology in mining and extraction are also vital. Copper, for example, is now commonly extracted from the ores containing less than 0.3% of the metal. However, in the 1700's the common minerable grade was about 13%, and in the 1900's between 5 to 2.5%. This decrease in minerable grade is a function of not only the demand for copper, but also technological progress in mining and metal extraction in this case. Availability of ores has recently been subject to political factors, which were not the primary concern in the past. The petroleum of the Middle East is the prominent example. The metallic minerals are classified into three categories. The first one is the minerals free from these factors mentioned above. Iron and manganese belong to this category. The second category include aluminum, copper, nickel, cobalt, titanium, lead and zinc. These metals can be provided by the future progress in technology which enables the use of lower grade deposits. The third one represents the metals whose occurrence is geologically limited, and, thus, subject to the political factors. Niobium and tantalium are the example. In this article, I discuss these minerals in detail to give an outline of the factors which brought their concentration, distribution and availability. SELECTED MINERAL CONCENTRATIONS Iron Major iron ore deposits are the features of the relics of ancient continental crusts. They are found in all the continents where the Precambrian rocks expose. They do not occur in the oldest rocks (3,000my old) nor in the youngest ones (600my old). They are sedimentary rocks, normally, exhibiting alternations of bands of iron oxides (magnetite and hematite) or iron silicates (especially greenalite), and bands of silica, variously described as jasper, quartzite and chert. The best known banded iron ore deposits exist in the Lake Superior region of the North America, and their distribution extends at intervals to Labrador of Canada. Other examples occur in the U.S.S.R. Brazil, Venezela, India, the mainland China, and southern and western Africa (Liberia and Guines) (Table 1). Recently large deposits have been discovered in the Hammersley region of the western Australia, and northern and central region of Brazil. These ore deposits were originally formed in shallow seas where simple but abundant life existed. After the deposition, enrichment processes , related to tropical weathering, have brought about wholesale removal of silica to produce the deposits of best quality. Ores with 50-60% iron represent the best products in the Precambrian fields. Since local reserves of this high grade ores approach to exhaustion, benefication processes have been used effectively to upgrade much leaner ores (for example, taconite with 20-25% iron in Minesota, U.S.A.). There are two other types of iron deposits. Ironstone, a sedimentary rock containing goethite, chamosite and siderite, first appeared in early Paleozoic times in the stratigraphic record, and reached its Zenith in the Jurassic. Similar bedded iron deposits are found in the belt from the Cleveland Hills to Oxford in England, and in the "minette" oolitic ores of Alsace-Lorraine in Luxenburg and France. The iron content of these ironstones seldom exceed 30%, but they usually contain calcite and are self-fluxing. They also have phosphorous as undesirable impurity. Because of the grade and impurities, they require more fuel than the Precambrian ores. The third type of iron ore is associated with igneous activity and consists of magnetite and some hematite, and apatite. This deposit is believed to be a product of magmatic differenciation. It occurs in Kiruna of Arctic Sweden. The iron ore, therefore, present no global shortage problem. They extend to considerable depths. Their concentration is largest among the metals. If low grades were treated, the resource can be stepped up much substantially. Manganese Manganese is a ferro--alloy metal, thus, essential to the manufacture of sound steel. Manganese comes from manganese oxide and silicate ores. Many minerals contain manganese, but only a few oxides, silicates, and, in some places, carbonates (rhodecrosite) are mined as ore. Types of manganese deposits are bedded, massive,

Jan 1, 1982

By P. A. Witherspoon, R. W. Donovan, T. D. Mueller

The use of petroleum-barren aquifers for underground storage has become extremely important to the natural-gas industry. A critical problem in assessing the feasibility of a specific aquifer for such use is the permeability determination of the caprock over the proposed storage project. The approach used here is to conduct both static and dynamic field tests on the aquifer being analyzed. Valuable information on the possibility of communication between the storage aquifer and any other aquifers above can be obtained by measuring hydrostatic water levels and water analyses. Significant differences in such data give evidence of the lack of communication between the intended storage reservoir and other horizons. The dynamic approach requires that one well be pumped in the storage aquifer, and changes in fluid levels recorded in both the aquifer and its caprock. The interpretation of the data from such pumping tests involves the solution of nonsteady radial flow in an infinite aquifer and the influence on such flow of a leaky caprock. A finite-difference model has been used to investigate this problem, and the transient behavior has been solved numerically with a digital computer. It has been found that the pressure transients in the storage aquifer are not affected significantly by moderate caprock leakage. The pressure behavior of the caprock is a much better indicator of the degree of leakage, and generalized solutions for this behavior are included. Field data are presented to demonstrate both the static and dynamic approach. If is concluded that appropriate investigation of the groundwater hydrology in an aquifer-type gas-storage project can provide much valuable information for determining the effectiveness of the caprock to hold gas. INTRODUCTION Underground storage of natural gas in the United States has been developing at a rapid rate over the past few years. In 1955, the total gas-storage capacity was about 1.6 trillion cu ft; by 1961, this figure was almost 3.2 trillion cu ft, an increase of 100 per cent in six years.' This trend un- doubtedly will continue because the economics favor the development of gas storage, as opposed to the construction of new pipelines, to meet the inherent cyclic demand for fuel in the metropolitan areas of this country.' About 15 per cent of the current underground gas storage has been developed in petroleum-barren aquifers, i.e., geological domes or anticlines in which no commercial quantities of oil or gas had been produced prior to the storage operations. The necessity for using barren aquifers outside many metropolitan areas of this country has been due to the lack of depleted oil or gas fields that were near enough and large enough to meet the demands of such consuming areas. Pipeline companies have developed aquifer storage along their transmission lines to meet the fluctuating needs of their complex systems. Considerable thought has also been given to the problem of storing gas in a structureless aquifer, both in this country' and in the Soviet Union outside the city of Leningrad.'," Conditions such as these have led to the development of aquifer gas-storage projects in many parts of the U. S. Most of these developments have centered in the Mid-Continent area, and the greatest amount of activity has been concentrated in Illinois.6 Thus, the use of petroleum-barren aquifers for gas-storage purposes has become extremely important to the natural-gas industry. There are three basic problems in developing aquifer-type storage: (1) finding an adequate geologic structure, (2) finding a suitable storage reservoir within the structure and (3) determining the tightness of the caprock over the intended storage zone. The first two problems can be solved by applying conventional methods of exploration geology, but once these problems are solved, the question arises as to why no oil or gas is present in an otherwise favorable setting. Two situations are possible: (1) an adequate source bed was never present, or (2) a source bed was present but the petroleum seeped away because of a leaky caprock. Determining the tightness of the caprock is one of the most critical problems in assessing the feasibility of a specific aquifer for storage purposes. In attacking this problem, one usually takes cores of the caprock and subjects them to a rigorous investigation. Such core data are desirable, but they only detail the matrix properties and cannot be expected to reveal the gross characteristics of the caprock. Several gas-storage projects in the U. S. have had considerable leakage where

By C. G. Albert

The clay particles of 2 microns or less required for modern paper coating are predominantly flat plates, lying smoothly on the sheet and producing a high gloss. Operating speeds of today's coating machines necessitate a clay composition of 60 and sometimes 70 pct solids as against the 35 to 40 pct required in the past. Since clays in suspension may solidify in flow, only those of intrinsically low viscosity can be used as high coating solids. THE literature of paper technology contains a number of articles having reference to developments in the field of coating and filler clays for use in paper manufacture. Much of this information has not been included in mining publications and has therefore not been readily available to all in the mineral industry. Recent developments in this field, including spray drying of clays, are presented here. U. S. Bureau of Mines figures for 1952 indicate that the paper industry consumes more than 50 pct of all kaolin produced and sold in the U. S. As most of the kaolin used by the industry comes from Georgia producers, the fraction of their output destined for paper use is thus appreciably higher than 50 pct. Small wonder that the kaolin industry, especially in Georgia, is highly sensitive to the quality requirements of paper mills and must respond promptly to technological developments in paper manufacturing. The paper industry itself is not the ultimate consumer. For the greater portion of the clay the end product is the printed page, and the demands of printing and publishing have sparked some of the technological advances in paper making which have, in turn, brought about methods employed in the kaolin industry. As compared with the product of 20 or 25 years ago, one of the most striking characteristics of the graphic arts today is the mass production of quality printing of fine pictorial work, much of it in full color. During this time periodicals with printing standards close to those of yesterday's slick-paper publications that were printed on a slow schedule and for a limited circulation have grown to the point where they go out to many millions of readers, often at weekly intervals. The complexity of technological improvements brought about by this increased circulation is probably not appreciated even by technical people, unless they have had reasonably close contact with the industry. The advance has come about through developments not only in the art of printing, but also in the field of paper making and even at the level of clay mining and processing. The smoothness required for faithful reproduction of the kind of printed matter under consideration is attained with a clay-coated paper. Since the distribution expense of the publication will depend to a great extent on its weight, the paper used must not be too heavy. This means a lower basis weight than was normal for conventional clay-coated papers some 25 years ago. And for this mass production market it becomes necessary to provide a paper having these and all the other required characteristics at a very moderate price—not the premium price conventional clay-coated papers formerly demanded. This challenge has been met by a new method of producing coated paper. In the past, application of clay coating to paper was a conversion operation, performed separately from the making of the base sheet. The newer development is called machine coating. Here application of the coating is an integral step in a continuous process by which wood pulp, clay, and other ingredients are manufactured into a sheet of coated paper. Many more problems are involved in this procedural change than are apparent at a casual glance. The coating operation, for example, must function at much higher linear speed than could be obtained with coating mechanisms previously employed. The application machinery developed to meet this requirement necessitated changes in composition of the coating color.* This created new requirements, summarized below, for the clay employed as coating pigment. In addition to smoothness, a relatively glossy printing surface is needed, and to a large measure it is the function of the coating clay to make possible the development of both these surface characteristics. Traditionally, pigments such as satin-white, prepared by reacting lime with paper-makers' alum, were used to assist in producing a high finish. However, economic considerations and others preclude large-scale use of such material in the new processes. In the 1930's Maloney' discovered that a certain particle size fraction of kaolin, consisting

Jan 1, 1956

By T. V. Philip, Paul A. Beck

IN a previous note1 it was pointed out that the available information suggests a distinct correlation between the occurrence of the CsCl-type ordered structures formed in equi-atomic binary alloys of transition elements and the location of the components in the periodic table. A definite increase in the bond strength between unlike atoms, as compared with that between like atoms, is indicated when, in binary alloys of iron group elements, the other component is changed from chromium to vanadium or tantalum, to titanium. Laves and Wallbaum,2 eported a CsCl-type ordered structure for the stable intermediate phase at the equi-atomic composition in the Ti-Fe system (and also in the Ti-Ru and Ti-Os systems). On the other hand, Tagaya, Nenno, and Nishiyama8 found no ordering of this kind in the body-centered-cubic solid solutions in the Cr-Fe system. It was concluded' that the occurrence of a CsC1-type ordered phase of intermediate stability may be expected in the V-Fe system. Preliminary results were described: suggesting that at least some degree of order does exist in the equi-atomic alloy VFe, annealed at 1200° C and quenched in cold water. Using CrK radiation, at least one weak superlattice line was found in the X-ray diffraction pattern which could be indexed as (100). and which disappeared, as expected, when FeK radiation was used. It was found more recently that the impurity phase (presumably VN), previously giving rise to additional X-ray diffraction lines and thus preventing definite indentification of the (111) superlattice line,' can be eliminated by preparing the alloy by arc melting in a helium atmosphere. The alloy specimen melted in this manner had clean surfaces and showed very littte, if any, nitride phase in the microstructure after annealing in vacuum at 1250°C and quenching. The relative intensity of the (100). superlattice line with CrK radiation was greatly increased and, in addition, a fairly strong (Ill) , superlattice line and even a detectable (210) , super- lattice line was obtained, when the powder produced by filing from the 1250°C annealed and quenched specimen was reannealed in vacuum at 625°C for 1 hr. By using a cellulose acetate film as a filter in front of the photographic film in order to absorb the soft radiation resulting from air fluorescence, the background intensity could be reduced considerably in relation to the intensity of the diffraction lines. All three superlattice lines, (100),, (111)., and (210)a, disappeared, as expected, when a radiation other than CrK, was used.1 For the powder reannealed at 625ºC, the strongest lines of the s pattern also appeared weakly, indicating that during an: nealing at 625°C a small amount of transformation took place from the body-centered-cubic to the u phase. However, since all superlattice lines to be expected in the 0 angle range used were actually very clearly present, it was evident that the high temperature body-centered-cubic phase had transformed into a CsCl-type ordered structure before much s phase could form. Reannealing filings for 1 hr at 700°C resulted in complete transformation into the u phase. After 1 hr at 650°C the transformation into IT was not quite complete, the strongest body-centered-cubic diffraction lines were still present, although fairly weak, and only the (100), super-lattice line was detectable (extremely weak). These findings not only confirm the previous conclusionsL as to the existence of a CsCl-type ordered structure in the Fe-V system, but they also indicate that this metastable ordered structure forms at about 600°C preceding the formation of the stable s phase, while at 1250°C the disordered body-centered-cubic phase is stable. The faint (loo) , super-lattice line previously observed in specimens quenched from 1200°C was probably due to the fact that the quenching of the powder in an evacuated fuse:! quartz capsule was not fast enough to prevent completely the ordering reaction on cooling. The present results conform with the suggestion that the A-B bond, in relation to the A-A and B-B bonds, increases in strength from CrFe to VFe to TiFc. It is, therefore, of interest to compare the measured values of the average atomic diameter for these three alloys with the corresponding average atomic diameters calculated from the atomic radii of the elements concerned. ASSUming A-B contacts

Jan 1, 1958

By R. K. Datta

Eluropium -activated yttrium vanadate, excited by short- and long-wavelength ultraviolet radiations, shows enzission lines near 6100A, the principal ones correspondirzg to transitions of EU+3 ions from 5Do to the F2 energy level. The ~neclmlallistns of absorption and etnissiorz of this phosphor have been empirically studied. The absorption is caused by two separate processes in the matrix. The phosphor YVO4:Eu and tlie matrix YVO, haue essentially the same resectance spectra in the region 2500 to 4000if. There are two broat ah~o?~ption bands: natnely, around 2600 and 3600A regions. Whet? compared with the reflectance spectra of the constituent oxides, namely, Y2O3, V205, Eu2 03, the absorption bands of the tnatvix and the phosphor in the short- and long-wavelength ultraviolet regions can be interpreted as due to charge-transfer processes involving i) Y-O and ii) 17-0, respecti~~ely. The excitation spectrum' of the phosphor has two broad bands with peaks at about 2600 and 3600A regions. U can be postulated that enevgy absorbed under short- crud long-wavelength ultraziolet excitatzon, respectzcely, by Y-0 and V-O parts of the niatrix is transferred by radcatio~zless process to EU+3 ions. Thzs postulate can empirically be proved on the basis of the spectra of the systems (YVO4,-YPO4,):Eu. PO? ~ons do not have' any appreciable absorption in the short or long uv regions. When P0't3 ions are substitlited for v0i3 ions in YVO,:Eu phosphors, the absorption and excitation in the lo~zg uv region are gradually reduced. Ailoreover, the £«+3 emission under 3650A excitation is decreased zuitlz increase in PO^ content, lt~lzereas the emission under 253711 excitation remains esserztially constant with as IILUCIL as 60 mole pct of YPO, in YVO,:En. EUROPIUM-activated yttrium vanadate (YV04:Eu) is a well-known phosptor.'-3 When excited by cathode rays, 2537, or 3650A radiations, YV04:Eu emits typical narrow-band spectrum of EU'3 fluorescence in the region of 5900 to 7100A, a very strong maximum occurring at about 6190A. This emission is considered to be due to 5Do — 7F2 transition of excite: EU*3 in 4f state. Moreover, the emission under 3650A excitation increases with temperature markedly. Because of these properties, YV04:Eu phosphor finds wide application in industries, namely, in color television tube, laser, and high-pressure mercury-vapor lamps. It is the purpose of this paper 1) to present the empirical study of the mechanism of absorption and excitation of YV04:Eu phosphor by ultraviolet radiation, and 2) to offer an explanation for the process of emission. It should be mentioned that a theoretical discussion on the mechanism of absorption in terms of the crystal or ligand field theory is beyond the scope of this paper. Empirical studies presented here are based primarily on spectroscopic and crystal-chemical investigations. EXPERIMENTAL i) Preparation of Materials. The synthesis of vanadate phosphorinvolves preparation of (Yi-.tEuj^Oj, blending the mixed oxide with an oxidic compound of vanadium, such as V205 or NH4V03, and firing the mixture at about 1000°C for several hours when YVO4:Eu forms as a primary phase. (Y1-xEux)203 compositions are prepared by coprecipitation of oxalates, and igniting them at about 1000°C. The vanadium compound is added in excess of the stoichiometric proportion for fluxing purposes. The charge is cooled, ground, and the excess of vanadium is removed by a hot wash (-80°C) of alkali or ammonia. Y(VP)04:Eu samples are prepared by mixing (Yi-xEux)203 with stoichio-metric proportions of (NH4)2HP04 and the vanadium compound. The mixtures are prefired at 700" to 800°C for 1 or 2 hr, then ground and refired at 1150°C. The firings of the phosphors are done in silica crucibles, and the homogeneity of the final products checked by X-ray diffraction analysis. ii) Spectroscopic Measurements. Diffuse reflectance sqectra of the samples within the range 2500 to 7000A are obtained by using CaF2 as standard in Cary Recording Spectrophotometer, Model No. 14. The emission spectra from phosphor plaques, excited under different ultraviolet sources, are obtained by a direct-recording spectroradiometer4 adapted with a grating monochromator with nearly constant dispersion. Maximym sensitivity in the spectral region from 2500 to 7000A is obtained with a quartz-windowed, S-20 response p$otomultiplier. The total' band pass is kept about 30A. The spectra are calibrated against the light from a standard tungsten filament lamp reflected from a magnesium-oxide plaque. Brightness data given throughout this report $re measured from the height of the main peak (6190A) of the emission spectra. The excitation spectra are measured using sodium salicy-late as standard. RESULTS AND DISCUSSION The diffuse reflectance spectra of both pure and activated YV04, along with their constituent components, are shown in Fig. 1. YVO4 shows two broad absorption bands, namely, around 2500 and 3600A, respectively. Since there is little difference between the reflectance spectraoof YVO4 and YVO,:Eu in the uv region (2500 to 4000A), it can be postulated that excitation of the phosphor is mainly due to absorption by the matrix, i.e., YVO4. The question, now, is whether the matrix (Yv04) is excited as a whole, or different components are responsible for the two absorption bands. The characteristic absorption spectra of molecular ions involving transition metal ions, both in aqueous solutions and in crystals, are due to electronic transi-

Jan 1, 1968

By L. C. Brown, C. St. John, C. C. Sanderson

The diffusion rate in Cu-Te couples is very sensitive to compressive stress, with a load of 20 psi making a significant difference to the width of the diffusion zone. At zero stress, two phases appear in the diffusion zone (Cu4Te3 and CuTe). Under compressive loading the third stable phase (Cuz Te) also appears, and its thickness increases progressively with increasing stress. The results are explained on the basis of an incipient Kirkendall porosity which restricts the transfer of atoms from the copper into the diffusion zone. DURING a study of the Kirkendall effect in Cu-Te couples prepared by clamping together the two components, it was found that the diffusion-zone width and shape in the plane of contact were not reproducible. Although the stresses involved in clamping are not normally sufficiently high to affect diffusion rates, preliminary tests established that the Cu-Te system is particularly stress-sensitive. The phase diagram for the system Cu-Te given in Hanssen1 shows that there is practically no solid solubility at either end of the phase diagram. Many areas of the diagram are not fully substantiated, but there appear to be three intermediate phases: Cu,Te—hexagonal in structure, having a grey luster; Cu4Te3—a tetragonal defect structure, having a red-purple luster; CuTe—orthorhombic in structure and having a golden-green luster. The existence of a fourth phase, the X phase at 37 at. pct Te, is considered doubtful. The composition ranges of the three stable phases are small, and are not accurately known. The phase diagram changes little with temperature up to 305°C, at which temperature a polymorphic transformation takes place in Cu2Te. The nature of the Cu-Te phase diagram indicates that the diffusion zone in a Cu-Te couple would consist of a series of layers of intermediate phases. The relative thickness of any one phase will depend on its diffusion coefficient and composition range.' In this type of diffusion couple it is often found experimentally that some phases are not visible at all in the diffusion zone due either to a small diffusion coefficient or to a restricted composition range.3 Since the composition ranges of the phases in Cu-Te are not known, it is not possible to determine diffusion coefficients in this system from a knowledge of the phase thicknesses. Several investigations have been carried out to determine the effect of compressive stress on diffusion rates in multiphase systems. Diffusion couples of Ni-A1 have been investigated by Storchheim et al.4 and by Castleman and Seigle.5 Two phases (ß and ?) appear in the diffusion zone under zero stress and the thickness of both phases is progressively reduced with increasing stress. According to Storchheim et al.4 a stress of 25,000 psi reduces the thickness of the diffusion zone by 50 pct. In a-brass—?-brass couples the thickness of the 0 phase formed in the diffusion zone was reduced by 20 pct at a stress of 20,000 psi.6 In other investigations the compressive load has been observed to increase the width of the diffusion zone. In A1-U, several investigators3,8 have found the width of the whase UA13 to increase with stress. According to casileman,8 the rate of formation of UA13 at 520°C is 75 pct faster at a stress of 20,000 psi as compared with a stress of 2500 psi. In Cu-Sb the effect of stress is greater than in the other systems described. According to Heumann9,10 only one phase (y) appears in the diffusion zone at a stress of 500 psi, but at a stress of 850 psi two phases (y and k) are present. If a diffusion couple containing both y and k phases is annealed at a low stress level, the y phase grows at the expense of the k phase. EXPERIMENTAL The diffusion couples were prepared from electrolytic copper bar stock with a nominal purity of 99.92 pct and from tellurium of 99.7 pct purity. The tellurium proved difficult to machine because of its brittleness and a technique was developed for casting the tellurium into a graphite slab mold and spark-machining specimens from this slab. Both the copper and tellurium were produced in the form of discs 2 in. diam by approximately 1/4 in. thick with surfaces ground flat to 3/0 emery paper. The diffusion apparatus is shown in Fig. 1. Auni-axial compressive stress was applied to the system through a simple lever system. A stainless-steel rod actuated by the lever arm lay inside a stainless-steel tube. The diffusion couple lay on top of the steel rod, and pressure was applied to the couple between the rod and a plug welded into the center of the tube. To ensure a uniform stress across the couple, a hemispherical boss and cup were used to transmit the load to the diffusion couple. A 400-w tube furnace with a uniform hot zone 3 in. long slid around the stainless-steel tube and maintained the assembly at temperature. A thermocouple situated 3 in. from the specimen operated a proportional temperature controller which maintained the specimen temperature constant to ±2°C. Most diffusion runs were carried out at 250C although a few tests were made at other temperatures in the range 235° to 300°C. The specimens were inserted and removed with the furnace at operating temperature, and took only 2 min to reach diffusion temperature—a time small compared with the total diffusion time. All the diffusion experiments were carried out in a hydrogen atmosphere, since consistent results were obtained in hydrogen and nitrogen atmospheres and in

Jan 1, 1967

By S. A. Mersol, C. T. Lynch, F. W. Vahldiek

The room-temperature microhardness of as-grown and annealed MoaC single crystals was measured on the (0001), {2110), and1012) planes using Knoop and Vickevs indenters at loads ranging front 25 to 1000 g. The orientatimz dependence of hardness with respect to crystal axes was also studied. The average random hardness of as-grown crystals was determined to be 1520 kg per sq nm. Annealing to 2000°C decreased. the average hardness by 150 units. An increase in hardness after annealing at 2200 ;C was noted. Optical and electron microscopy revealed slip and twin traces on all planes studied, as produced by mi-cvohavdness indentations. Basal (0001)(2i10) slip was determined to be the primary slip system and was substantiated by electron transmission microscopy. A secondary {1010)(2110) slip was produced by mi-crohardness indentations. The lattev also produced twinning- of the {10i2)[0001] type, as proven by electron diffraction. Electrical resistivity and elastic-modulus anisotropy were found and correlated with hardness anisotropy and Mo2C crystal structure. Elastic-modulus values were obtained by microhard-uess and ullrasonic methods. Bonding mechanism of Mo2C is discussed. ROOM-TEMPERATURE microhardness indentations are useful for studying hardness anisotropy, slip, and twinning in brittle materials. Slip has previously been produced in this manner in and WS~~.~ Recently, the authors5 reported slip of the {10i0)(11~0) type produced by high pressure and microhardness indentations on hexagonal TiBz (c/a = 1.066) single and polycrystals. This slip system was also reported by French and ~homas' and Taka-hashi and ~reise~ for hexagonal WC (c/a = 0.976) crystals. These results suggest that prismatic rather than basal slip is favored in hexagonal nonmetallic materials having a c/a ratio considerably less than the ideal (1.633). Buerger precession and cylindrical X-ray rotation patterns were previously' taken on cleaved sections of the Mo2C single crystals studied in this work. Th~y were found to be hexagonal MoaC with a. = 3.0233A, co = 4.7344A, and C/O = 1.5660. The latter ratio is close to that of the beryllium metal (c/a = 1.57), which slips primarily on the (0001) plane: but also slips on the (1010) planes.'0 ~irconium" (c/u = 1.59) and titanium12 (c/o = 1.59) deform mainly by slip on nonbasal planes which contain a close-packed direction. This is due to the fact that for these two metals the initial resolved shear stress for slip on the (10i0) prismatic planes is lower than that on the (0001) plane. The prominence of basal rather than prismatic slip in metals of high c/o ratios is shown by cadmium (c/a = 1.89), zinc (c/a = 1.86), and magnesium (c/a = 1.62) which deform mainly by basal slip. However, in case of the latter, by stressing magnesium crystals in tension or compression parallel to the basal plane, slip on (10i0) planes can also be produced.13 Several hardness values for polycrystalline Mo2C are reported in the literature: Biickle ' and Samsonov'~ give a value of 1800 and 1479 kg per sq mm, respectively, at a 100-g load; and Kieffer and Benesovsk~'~ report a value of 1950 kg per sq mm at a 50-g load. ~ott'~ reports a Vickers hardness value of 2000 kg per sq mm, with the load unspecified. A Rockwell A hardness value of RA = 88 has also been reported.'' In the present work, for comparison with single-crystal Mo2C hardness values, a Khnloo value of 1600 It 150 kg per sq mm was found on 99.6 pct pure and 99 pct dense hot-pressed Mo2C. This work was undertaken partly to explain the considerable differences in hardness values reported for polycrystalline Mo2C. EXPERIMENTAL The Mo2C single crystals investigated were prepared by a Verneuil-type process using an electric arc by the Linde Co. of the Union Carbide Corp.lg The largest specimens grown were boules 7 mm in diam by 40 mm in length. The crystals had an average density of 9.04 g per cu cm, with a Mo + C content of 99.8 wt pct. The major impurities were: 100 ppm each Na, Zr, and Ca; 85 ppm 0, 55 ppm Fe, and 10 ppm each Cr and Ta. The crystals were found to be carbon-poor, the average carbon content being 5.73 wt pct (stoichiometric value is 5.89 pct). The molybdenum content was found to be 94.08 pct, which is nearly stoichiometric. Electron-microprobe traverses of selected specimens were done with a Phillips-AMR microanalyzer. Thin-film and carbon replicas were used to prepare electron micrographs. This work was done with a JEM-6A electron microscope. Prior to optical and electron-optical studies, specimens were mounted in Lucite and polished on a vibratory polisher using diamond-paste grades ranging from 9-1 p and Linde A powder for up to 48 hr. Dilute nitric acid was used for thin-section polishing and chemical etching for 1-15 min. Electrical-resistivity measurements at room temperature were taken with a Rubicon bridge, using gold contacts. For hardness measurements, a Tukon Microhardness Tester Type FB with Knoop and Vickers indenters was used. Measurements were taken at loads ranging from 25 to 1000 g; however, the 100-g load was chosen as the standard load. All measurements were taken at room temperature. Indentations of cracking classes 1 and 2 only were considered for hardness determinations.20)21 (There are six cracking classes, ranging from "class 1" for a perfect inden-

Jan 1, 1967

By L. D. Jaffe, R. K. Pitler

IT would clearly be advantageous to produce molten titanium, suitable for alloying and casting, directly from the relatively inexpensive tetra-chloride, without using a metallic reducing agent. Accordingly, a preliminary investigation has been made of the production of titanium by hydrogen reduction of the tetrachloride in an electric arc furnace. For the reaction: TiCl4(g) + 2H2(g) ? Ti(s) + 4HCl(g), Lockhart and his colleagues' have calculated equilibrium constants at 800" to 1500°K, based on tables of Brewer.' Extrapolating these data on the assumption that the heat of reaction varies linearly with temperature, and using Brewer's value for the heat of fusion, there is obtained for the reaction: TiCl4(g) + 2H2(g) ? Ti(l) + 4HCl(g) an equilibrium constant at 2100°K (100" above the melting point) Kp = (PHCl)4/pTlcl4 (Ph2)2 = 0.167 atm. If this is correct, starting with hydrogen saturated with TiC14 at 25°C, where the vapor pressure of the latter3 is 12.6 mm, and maintaining the overall pressure at 1 atm, 99 1/4 pct of the TiC14 would be reduced to Ti at equilibrium. An arc furnace designed for the melting of titanium and lined with water-cooled copper was used for the preliminary experimental check. Its inside diameter was 2 7/16 in. and height 6 in. The water-cooled electrode was tipped with 3/4 in. diam tung-sten. In the one run carried out, the furnace was charged with a starting batch of 293.35 g of scrap titanium,' analyzing 99.24 pct Ti. After the system was twice pumped out and flushed with argon to remove air, the arc was struck in argon and kept at 400 amp dc, electrode negative, for 1 min to melt the starting batch. Commercial tank hydrogen, dried by passing over silica gel and bubbled through technical grade TiC14 at room temperature, was then admitted to the furnace. The inlet tube directed this gas downward toward the arc and melt; the gas outlet was at the top of the furnace. The quantity of gas passed through was not measured accurately but is estimated to be about 50 liters per min at a few mm of mercury above atmospheric pressure. The arc was held in the H2-TiC14 mixture at 400 amp with little difficulty. After 10 min the arc burned through the copper lining above the melt, admitting water and ending the run. (This had happened frequently in straight melting runs under argon, and is not thought due to the atmosphere used.) In the furnace were found an ingot, several small pieces, and a little metallic powder, evidently formed when water entered the hot furnace. Some titanium also clung to the electrode. The portions were weighed and separately analyzed for titanium. Their total weight was 321.76 g of 94.82 pct overall purity. Much of the impurity consisted of tungsten broken off when the titanium was removed from the electrode. No detailed impurity analysis was made because the starting materials were impure and flooding the furnace introduced further contamination. Thus, the metal recovered contained 305.08 g Ti, as compared to 291.12 g in the starting batch. Titanium could not have been carried over from previous melting runs in the furnace, as the interior was sandblasted and the electrode ground clean before the run. Evidently at least 13.96 g of Ti were produced from TiCl4. Since this would require complete reduction of 430 liters of TiCl, gas at 12.6 mm of Hg, the efficiency of the process must have been rather high. Experimental work is continuing. No reason is seen why the process could not be scaled up to the size of a 30-ton steel-melting arc furnace. In a large unit there would be a large molten pool from which liquid metal could be withdrawn for alloying and casting. The problems of electrodes and power supply are similar to those arising in arc furnaces for melting titanium. Unreacted TiC14 and the HC1 in the exhaust gases could be recovered; unreacted H2 could be dried and re-used. The melt would be saturated with hydrogen, whose solubility in liquid titanium is unknown. Small amounts of hydrogen seem to have little effect on the properties of titanium.5 If the amount introduced was so large as to impair properties or cause casting difficulties, it could be reduced by holding the melt under an inert gas or by pumping on it, prior to casting. Acknowledgment The authors would like to thank I. Preble and M. Goldman for their valuable aid in the experimental work and the thermodynamic calculations, respectively. References 1R. J. Lockhart, J. J. Ward, M. J. Hussey, and J. W. Clegg: In unpublished report, "U. S. Air Force Project Rand. Titanium and Titanium Base Alloys." Battelle Memorial Institute (1949) 37-44. 2 L. Brewer et al: Chemistry and Metallurgy of Miscellaneous Materials—Thermodynamics. Ed. by L. L. Quill. pp 13-39, 60-192. (1950). McGraw-Hill Book Co. 3 K. Arii: Scientific Reports. Tohoku Imp. Univ. (1933) 22, 182. 4 R. S. Dean, J. R. Long, F. S. Wartman, and E. L. Anderson: Transactions AIME (1946) 166, 369-381; Metals Technology (Feb. 1946) TP 1961E. * R. I. Jaffee and I. E. Campbell: Transactions AIME (1949) 185, 646-654; Journal of Metals (Sept. 1949) TP 26813.

Jan 1, 1951

By C. R. Wohrman

A MELT of pure electrolytic iron with about 0.4 per cent. sulfur and 7 per cent. manganese was prepared in connection with a study of inclusions in iron. The alloy darkened rapidly when etched with a 10 per cent. alcoholic solution of nitric acid, and microscopic examination revealed the presence of a semimartensitic, semi-Widmanstättian structure (Figs. 1 and 2) closely resembling the structures found by Benedicks in his artificially prepared "meteoric iron." 1 The structures found by Benedicks were readily explained by the presence in his alloy of the two well known iron-nickel solid solutions, taenite and kamacite, and of their eutectoid (?) aggregate, plessite. By analogy, one would infer the existence of similar phases in the iron-manganese alloy just described, such as inference being strengthened by the similarity displayed in many ways by nickel and manganese in their effects on iron. IRON-MANGANESE ALLOTS ARE REGARDED AS HOMOGENEOUS Heterogeneity of iron-manganese alloys has not been recognized heretofore, judging from the data of standard metallographic books and those of original papers. In fact, iron and manganese are quite generally regarded as completely miscible in the solid state. Dr. Rosenhain,2 for example, writes: Alloys of group (a), of which those of gold-silver, iron-manganese, and copper-nickel are examples, when allowed to crystallize sufficiently slowly to attain their condition of final equilibrium, present a microstructure exactly similar to that of a pure metal-the homogeneous character of the molten solution is in that case so completely maintained that the microscope cannot detect the presence of any second constituent.

Jan 1, 1927

By Waldemar Lindgeen

PART I.—General Features. PAGE Purpose and Extent of Inquiry ; Definitions (Fissure-Veins, Metamorphism, Metasomatism, Impregnation, Cementation, Weathering); Metasomatism in Connection with Mineral Deposits, Especially Fissure-Veins; Calcu lation of Analyses ; Criteria of Metasomatism; Crystallization of Sec ondary Minerals in Other Bodies; Secondary Alteration of Veins; Structure and Composition of Metasomatic Vein-Rocks and Their Rela tion to General Metamorphism,.759 PART 11.—Minerals Developed by Metasomatic Processes in Fissure-Veins. Quartz, including Chalcedonite and Opal; Rutile and Anatase; Fluorite; Calcite ; Magnesite and Dolomite ; Siderite; Muscovite and Sericite ; Biotite; Chlorite; Pyroxene and A mphibole; Garnet; Epidote ; Ortho clase ; Albite ; Tourmaline ; Topaz ; Kaolinite ; Zeolites ; Pyrite ; Mar casite ; Chalcopyrite ; Arsenopyrite; Pyrrhotite ; Galena; Zinc-blende and Other Sulphides; Tellurides; Native Copper ; Gold and Silver,. 601 Resistant Minerals,........... 619 PART III.—The Fissure-Veins Classified According to Metasomatic Processes. 1. Topaz-Cassiterite Veins (Altenberg and Zinnwald, Saxony ; Mt. Bischoff, Tasmania),..........620 2. Scapolite-Apatite Veins,.........625 3. Tourmalinic Gold-Copper Veins,.. 626 4. Biotitic Gold-Copper Veins,........ 641 5. Propylitic Gold and Silver Veins (General Features; Comstock Lode; The Veins of Nagyág, Hungary; Pachuca, Mexico ; New Zealand; Alaska ; Silver Cliff, Colorado; Silver City, Idaho),. .. .645 6. Fluoritic Gold-Tellurium Veins (Cripple Creek ; Other Occurrences),. 654 7. Sericitic and Kaolinitic Gold- and Silver-Veins (General Remarks; The Freiberg Veins ; De Lamar, Idaho; Summit District, Colorado),. . 6.58 8. Sericitic and Calcitic Gold-Silver Veins (The Gold-Quartz Veins of Cali fornia ; Idaho Types ; San Juan, Colorado ; Treadwell Mine, Alaska ; Gold Veins of Ontario, Canada; Schwarzwald Fissure-Veins),. . 664

Jan 1, 1901