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By J. H. Frye, C. P. Sun

A Frimary substitutional solid solution is a solution that has the same crystalline structure as the solvent metal, and in which solute atoms have replaced, solvent atoms at random on the host lattice. This replacement usually results in an increase in hardness as measured by conventional penetration tests. The factors controlling this hardness increase have been investigated in some detail for copper and silver alloys.1-5 It has been shown that at least four factors are involved: (1) concentration of solute, (2) row of the periodic table from which the solute is taken, (3) increase in lattice parameter of the solvent produced by the solute, and (4) work-hardening produced by penetration of the hardness indenter.2,4 The last factor is of the utmost importance. An investigation of the available data2,4 leads to the conclusion that the solutes that produce large hardness increases in copper or silver do so principally because they increase the rate of work-hardening. Furthermore, it is doubtful whether any solute produces appreciable hardening in the absence of plastic deformation and hence work-hardening. Thus it is clear that an understanding of the effect of solutes on rate of work-hardening is essential to an understanding of the hardness of solid solutions. This effect of solutes is also important in forming operations such as rolling or deep drawing. The presence of elements that confer rapid rates of work-hardening will increase power consumption and may well decrease the amount of deformation possible between anneals. This paper is the report of an investigation of the factors that control rates of work-hardening in copper and silver solid solutions. The silver and copper alloys considered are those in which all solutes are from the B subgroups or second short period of the periodic table. Experience with hardness, with structure of alloys, and with electrical conductivity has proved that results of a fundamental nature are most likely to be obtained if alloys are made up on the basis of the positions of the constituent elements in the periodic table. In previous work, Meyer analysis data on high-purity copper and silver solid solutions were obtained through careful measurements with fully calibrated equipment and these data were corrected to correspond to a standard grain size. From these corrected Meyer analysis data rates of work-hardening have now been calculated. Calculation of Rates as Work- HARDENING The Meyer analyses, on which this study of rates of work-hardening is based, consisted in determining impression

Jan 1, 1944

By J. H. Frye, C. P. Sun

A Frimary substitutional solid solution is a solution that has the same crystalline structure as the solvent metal, and in which solute atoms have replaced, solvent atoms at random on the host lattice. This replacement usually results in an increase in hardness as measured by conventional penetration tests. The factors controlling this hardness increase have been investigated in some detail for copper and silver alloys.1-5 It has been shown that at least four factors are involved: (1) concentration of solute, (2) row of the periodic table from which the solute is taken, (3) increase in lattice parameter of the solvent produced by the solute, and (4) work-hardening produced by penetration of the hardness indenter.2,4 The last factor is of the utmost importance. An investigation of the available data2,4 leads to the conclusion that the solutes that produce large hardness increases in copper or silver do so principally because they increase the rate of work-hardening. Furthermore, it is doubtful whether any solute produces appreciable hardening in the absence of plastic deformation and hence work-hardening. Thus it is clear that an understanding of the effect of solutes on rate of work-hardening is essential to an understanding of the hardness of solid solutions. This effect of solutes is also important in forming operations such as rolling or deep drawing. The presence of elements that confer rapid rates of work-hardening will increase power consumption and may well decrease the amount of deformation possible between anneals. This paper is the report of an investigation of the factors that control rates of work-hardening in copper and silver solid solutions. The silver and copper alloys considered are those in which all solutes are from the B subgroups or second short period of the periodic table. Experience with hardness, with structure of alloys, and with electrical conductivity has proved that results of a fundamental nature are most likely to be obtained if alloys are made up on the basis of the positions of the constituent elements in the periodic table. In previous work, Meyer analysis data on high-purity copper and silver solid solutions were obtained through careful measurements with fully calibrated equipment and these data were corrected to correspond to a standard grain size. From these corrected Meyer analysis data rates of work-hardening have now been calculated. Calculation of Rates as Work- HARDENING The Meyer analyses, on which this study of rates of work-hardening is based, consisted in determining impression

Jan 1, 1944

By Hans Lundberg

IN ORDER to comprehend the help and information that may be expected from electrical prospecting, it is necessary to have at least a general knowledge of the methods and principles involved in preparing and interpreting results obtained in the field. Electrical prospecting is rather complicated, both in theory and practice, and in order to make a brief description clear to the practical mining man it has been necessary to idealize and generalize the problems, at the risk of having such simplification criticized by the theoretical man. This simplification must not be misinterpreted as showing a lack of theoretical knowledge. FIG. 1.-FLOW OF CURRENT IN AND AROUND A CONDUCTING BODY. For its operation, electrical prospecting mainly, depends on the fact that most metallic ores conduct an electric current much more readily than barren rocks. If an electric current flows through a block of ground within which a body of chalcopyrite is enclosed in country rock of. schist, the current will find less resistance to its passage through the chalcopyrite than through the schist, even though the passage through the latter be much shorter. This is because the longer path through the chalcopyrite offers less resistance, owing to the superior electrical conductivity of the chalcopyrite (Fig. 1).

Jan 1, 1928

By Robert B. Johnson

The refraction seismic method has been used as an indirect means of exploring for fluorspar in southern Illinois by the Illinois State Geological Survey. Use of the method has been restricted to those areas where mineralization is associated with faults, a restriction necessitated by the operating principles of the refraction method. The seismic data enable the interpreter to locate fault zones in the subsurface and, under ideal circumstances, to make estimates of dip direction and magnitude and vertical displacement provided the local stratigraphic units are known and are geophysically recognizable. In this instance, refraction seismic exploration has been successful in the surface mapping of well-developed faults, both simple and complex with which fluorspar may or may not be associated. However, less clear results have been obtained where faulting has been less severe and where there are numerous accompanying mineralized joints.

Jan 1, 1964

By W. A. Boyer

Mine hoisting ropes can be loaded to capacity only when the strength of each component is exactly known. Characteristic curves provide this information. When load and rate of acceleration are specified for an individual rope, the characteristic curve assumes a predetermined shape. The factor of safety is the value derived by dividing the breaking strength of the rope by the total stress on the rope where it goes over the sheave wheel. Total stress is made up of the weight of skip and cage, weight of ore, weight of rope between sheave and rope clevise or attachment, friction load, and acceleration force. Bending stresses are neglected in this discussion, as sheaves and drum diameters are assumed to be large enough to reduce these stresses to a negligible value.

Mar 1, 1956

By W. S. Row

This mine is operated at 4000 to 4500 tons daily through a single shaft, with one rock hoist and one senice hoist. Latest shaft construction is concrete with wooden dividers. Economics of drifters and air-leg drills are compared, as well as costs of shrinkage and sub-level stoping. Economy of large haulage equipment and underground crushing is shown. THE property of Kerr-Addison Gold Mines is located in the Larder Lake district of Ontario about 3 miles from the Quebec boundary on the highway between Kirkland Lake and Noranda. It is roughly 320 miles due north of Toronto and 365 miles north of Buffalo, N. Y. The property was staked in 1906 and after several development efforts finally came into continuous production at 500 tons per day in 1937. In 1939 the daily tonnage -was increased to 1200 and in 1941 to 2100 tons. At the end of the war, because of the shortage of labor, tonnage was down to 1100 tons per day, rising to 2200 in February 1947. A further plant expansion inaugurated in 1946 was completed by December 1948, bringing the capacity of the mine and mill to a production rate of 4000 tons per day. Current production, with 880 men employed, is 4400 tons per day. Changes in mining methods and equipment, required by the recently completed expansion program, have yielded interesting comparative results, some of which are described in this paper. Hoisting of ore and waste from the mine, as well as all service facilities for handling men and supplies underground, are provided for in a single vertical five-compartment shaft. Ore and waste are hoisted in 12½-ton skips in balance by a 14-ft diam hoist with drums 115 in. wide, grooved for 2?-in. wire ropes. This hoist, with a total rope pull of 85,000 lb, was designed to operate at a speed of 2600 ft per min to a depth of 4000 ft, hoisting 15 tons of ore in each skip. It is amplidyne-controlled and powered with dual drive dc motors of 2250-hp each, through flexible couplings to two pinion shafts, one on each side of the main gear. Each motor has a peak load rating of 4500-hp at point of greatest duty. Direct current power to the hoist motors is supplied from a motor generator set consisting of a 5000-hp synchronous motor driving two 1750-kw dc generators. The loading pockets used so far are at depths of 1500 and 2725 ft respectively. Consequently the present hoisting speed is 2000 ft per min and the skips capacity only 12½ tons. Now that the shaft has been deepened the new pockets at 3934 ft will be increasingly used so that the speed and the skip size will be increased up to original specifications as required. Some difficulty in meeting the required hoisting cycle is expected because of the shape of the skips. As now in use the cross-sectional dimensions of the inside of the skip are 4 ft 6 in. x 3 ft 8 in. To hold 121/2 tons, therefore, they are 14 ft 10 in. deep. Since the skips are the conventional overturning or Kim-berley type, considerable hoisting time is lost in retardation in entering the headframe dump and in acceleration when leaving. It is expected that to obtain the specified tonnage from this hoist when the skip size is increased to 15 tons it may be necessary to resort to bottom dump skips. This type of skip could reduce considerably the present excessive acceleration and retardation times. Analyses of the skip-hoisting operation are made monthly, and every effort is made to reduce controllable delays. For a nine-month period in 1951 such an analysis shows that the hoist was in operation 92.5 pct of the available time, which was allotted as shown in Table I. Overall hoisting costs, cage and skip, for the year 1951, averaged $0.168 per ton of ore milled. The mine is serviced entirely by means of a single cage, capacity 35 men or 15,700 lb, operating in con-

Jan 1, 1954

Andes Copper has an old camp and a new camp, both isolated until recent road improvements. The old camp is at Potrerillos and the new one at El Salvador, 25 road ,miles away. The company was incorporated in 1916 and the first mine, concentrator, leaching plant, smelter and town were built in the succeeding twelve years. The first blister copper was poured on January 24, 1927. Economic conditions determined the long development period but the end result was the design and construction of a mining complex that was all of a piece, unlike the piecemeal additions to El Teniente and Chuqui. The mining episode at Potrerillos came to an end in 1960 when the ore at La Mina became exhausted and the new mine, El Salvador was brought in. Both of these ore deposits come within four hpundred feet of the surface and had they been developed within the last five years might reasonably have been brought in as open pit mines. However, that is conjecture, and as it stands the operation at La Mina was so ideal for the mining method selected, blockcaving, it became a classic for the textbooks. On the other hand, El Salvador has turned out to be a blockcaving mine to test an engineer's skill. In mining the copper deposit at Indio Muerto mountain, there are ground conditions of every conceivable type. Tunneling in one section, the back can cave like sized pea gravel and only spiling can hold it; in another heading only the lifters are shot but often the rest of the round will sit down in one piece; elsewhere the bottom squeezes up and closes the drifts; and contrary to all other locations, there are sections where the block can be undercut without support. Today the scope of operations at Andes extends from mining to electrolytic copper. Ore is mined at El Salvador and the concentrator is also situated there. Concentrates pipelined 14 miles to Llanth, the water decanted, and then it is hauled by rail 40 miles and 5500 ft higher to the Potrerillos smelter and refinery. Potrerillos and El Salvador are respectively 90 and 70 miles from the port of

Jan 11, 1969

By Robert P. Roberts

In order to obtain data on the thermal effect of the blast-furnace jacket and on the water consumption in these jackets a series of tests were run on the 56 by 180 in. blast furnaces at the Great Falls Reduction Department of the Anaconda Copper Mining Co., at Great Falls, Mont. The arrangement of the jacket is shown in the transverse section of one of the furnaces (Fig. 1) on which the tests were made. The nose, spout, and breast jackets have independent water supplies and discharges. The end and side jackets are connected in pairs, as regards water supply, the feed water entering at the bottom of the lower jacket and overflowing into the bottom of the jacket immediately above it, and discharging into an independent overflow pipe for each pair of jackets. Tests were run on the main side and end jackets, the breast jacket, the spout, and the nose. Measurements were taken of the quantity of water discharged per minute and the rise in temperature of the water in its passage through the jacket. The product of these two results, expressed in the proper units, evidently gives the number of units of heat removed per minute by the water. Sufficient work was done on each type of jacket to determine its behavior under different conditions, and the results from each set of conditions were checked by repeating the work from three to six times. The method used was as follows: The overflow from the jacket to be tested was arranged to discharge at will into a barrel, mounted upon platform scales. The discharge temperature was read just before turning the water into the barrel, and just before turning it back to the sluice; if the test lasted more than a minute, additional temperature readings were taken at 1-min. intervals from start to finish. The elapsed time of running the water into the barrel was determined with a stop-watch, and this figure, combined with the in-

Jan 1, 1914

By Robert Roberts

(Butte Meeting, August, 1913.) In order to obtain data on the thermal effect of the blast-furnace jacket and on the water consumption in these jackets a series of tests were run on the 56 by 180 in. blast furnaces at the Great Falls Reduction Department of the Anaconda Copper Mining Co., at Great Falls, Mont. The arrangement of the jacket is shown in the transverse section of one of the furnaces (Fig. 1) on which the tests were made. The nose, spout, and breast jackets have independent water supplies and discharges. The end and side jackets are connected in pairs, as regards water supply, the feed water entering at the bottom of the lower jacket and overflowing into the bottom of the jacket immediately above it, and discharging into an independent overflow pipe for each pair of jackets. Tests were run on the main side and end jackets, the breast jacket, the spout, and the nose. Measurements were taken of the quantity of water discharged, per minute and the rise in temperature of the water in its passage through the jacket. The product of these two results, expressed in the proper units, evidently gives the number of units of heat removed per minute by the water. Sufficient work was done on each type of jacket to determine its behavior under different conditions, and the results from each set of conditions were checked by repeating the work from three to six times. The method used was as follows : The overflow from the jacket to be tested was arranged to discharge at will into a barrel, mounted upon platform scales. The discharge temperature was read just before turning the water into the barrel, and just before turning it back to the sluice; if the test lasted more than a minute, additional temperature readings were taken at 1-min. intervals from start to finish. The elapsed time of running the water into the barrel was, determined with a stop-watch, and this figure, combined with the in-

Jan 7, 1913

By G. W. Nabor, H. J. Ramey

A blotter-type electrolytic model was utilized to prepare flow diagrams for a field test of the in-situ combustion process. It is pointed out that the areal sweep of a combustion pattern is similar to sweep patterns that would be developed at an infinite mobility ratio, which exists (approximately) across a combustion front because of the complete removal of liquids from the sand behind it. The precision of the blotter method was tested by comparison with results obtained by other techniques and was found to be satisfactory. The blotter-type model will not furnish as much information as more elaborate and expensive potentiometric models, but its speed of operation and ease of construction make it a highly satisfactory tool to determine areal sweep patterns. A tabulation of sweep efficiency and mobility ratio is furnished for various well geometries. INTRODUCTION In connection with the first field experiment with the in-situ combustion method of oil recovery," flow diagrams for the unique well geometry and boundary conditions involved in the subject test were developed. From these diagrams, information concerning the areal sweep and distribution of flow was obtained. A survev of the literature on potentiometric models, utilized to solve two-dimensional flow problems,1,2,3,4,5,6,8,9,10 indicated that the information desired could be obtained by means of a blotter-type model by making certain modifications in the technique. The following discussion describes the application of the modified blotter-type electrolytic model to the solution of the problem of the flow of gases in an oil reservoir for an engineering appraisal of the progress of the combustion front during the first field experiment. Considerable information has been published on the theoretical aspects of electrical model techniques. One author stated that:' "The models are based on the observation that since the velocity of an ion in an electrolytic system is proportional to the potential gradient, just as the velocity of a fluid particle in a porous medium is proportional to the pressure gradient, the paths of the ions in the electrolytic systems must be similar to those of the fluid particles in a porous medium with the same geometry and equivalent boundary conditions." This suggestion of a real analogy has been established repeatedly by comparisons of analytic derivations with model results. The analogy between the movement of a combustion front and the model results is not as apparent, since the combustion process differs from water flooding and gas cycling in that the velocity of the burning front is different from the velocity of the injected fluid at the front in combustion. Analogy between electric current flow and movement of the burning front exists because the local velocity of the burning front is proportional to the air (mass) velocity at this point (assuming frontal temperatures are at least as high as the ignition temperature of the fuel). Most model studies fall into one of two categories: (1) electron conduction, or (2) ionic conduction models. Examples of these are models which employ as flow media either a thin metal sheet, or a porous blotter saturated with an electrolyte, respectively. Of the two, the porous blotter model, an ionic conduction model, appears to be the most adaptable to the flow problem that results when using in-situ combustion. This problem requires that the mobility of the particles in the model be controlled analogous to the mobility of gas flowing through burned sand of high permeability to gas and then through oil sand of lower permeability to gas. Although the blotter model has heretofore been

Jan 1, 1955

By Paul H. Johnson

Chemical mining may be defined as the use of chemicals in extracting metal values from in situ broken or unbroken ores within a mine. The present means for the generation and regeneration of sulfuric acid or sulfuric-ferric sulfate solutions are inadequate for most applications. This report describes a pressure oxidation process for the generation and regeneration of these solutions from pyrites and spent iron sulfate leach liquors. While recent mining developments such as solution injection, caving, and conventional and nuclear blasting, have provided new and better means for the per-colation of leach solutions through rock masses, little has been done to reduce cost and improve the quality of the leach solutions that are required. A leach solution for chemical mining should meet certain requirements. It should be 1) inexpensive on either a small or large tonnage basis, 2) capable of selectively leaching the desired ore minerals contained in a particular ore, 3) of the proper strength or quality to do the job of leaching in an underground environment and 4) amendable to simple and inexpensive metal recovery steps. It is also important that the capital investment be low for the solution generation and; or regeneration plant.

Jan 8, 1965

By John C. Calhoun, Frank T. Bethel

The application of a method for analyzing pressure build-up curves to determine the effective permeability in the Spraberry is presented. Sixteen upper Spraberry wells of the Driver area arc analyzed and show variations in the in-place effective permeability of 2 to 183 md. Since these value- are much larger than the .5 md. or less reported for the matrix rock. a large part of the vertical fractures noted in Spraberry cores- must be considered native to the formation. The effectiveness of the well fracture treatment in connecting the well bore to the native fracture system is analyzed by comparison of the effective permeability as determined from the PI test to that obtained from a pressure build-up curve. These results show that the fracture treatment is sufficiently effective in connecting tile well bore to the native fractrure system to yield a flow capacity within about 50 per cent of that dictated by the native fracture system About two-thirds of the wells show some degree of local blockage. INTRODUCTION Fractured reservoirs have been known in the past.' However, during the past two years the Spraberry formation has directed the attention of the petroleum industry to a study of the performance of a fractured reservoir. The large variation in the number of fractures noted in Spraberry cores rile large variation in the producing ability of offset well.. the dramatic effect of well treatment on well ,productivity. the unknown quantity of oil stored in the fractures and matrix rock. and the unknown oil to be recovered were part of the initial Spraberry puzzle. In this type reservoir where the well treatment is an important factor in determining the productivity. it i- particularly desirable to determine the in-place flow capacity of the drainage area at distance* far removed from the well bore. This value. when compared to the flow capacity obtained from a PI measurement, will show the effectiveness of the well treatment (completion. fracture treatment, acidization. etc. in establishing a flow capacity in the vicinity of the well bore equal to that in the more remote region. The object of this paper i.; to show the application of a method of pressure build-up analysis for this purpose in studying the Upper Spraberry formation in the Driver area. STATEMENT OF THEORY The theoretical basis for the application of a pressure buildup curve to show the effective permeability of the rock at distances far removed from the well bore has been presented in detail in earlier publications.', in the previous development it was shown that pressure build-up behavior of all wells could be reduced to the same dimensionless build-up curve of Fig. 1. This mean.; that the bottom hole pressure of a well after hut-in will rise in such a manner that a plot of pressure against the logaritllm of time yields a straight line. The slope of this linp can be rlzed to calculate the effective permeability as expressed by the following equation: where: effective oil permeability md oil production rate prior to shnt in. STB/I) In practice the flow which occurs into the well bore after shut in and the variation in flow capacity in the immediate vicinity of the well bore usually cause the early part of the build-up curve to be unsatisfactory in the determination of permeability. In general the effect of these factors on the straight line section will become negligible after a time when the of the following equation is equal to 0.1. and the preferred section Of the curve for use in defining the slope lie-

Jan 1, 1953

By A. E. Harnsberger

I is obvious that details such as the physical and chemical properties and methods of heat-treating of the materials mentioned must be omitted in a paper on the subject of materials used in oil-refinery pumps1. Only sufficient identification is given to enable the reader to find additional information from already published data. Also, the subject covers many operating conditions, but only the most important services can be discussed here. To the oil refiner, the pumping equipment is of vital interest. Practically every refinery operation, from the charging of crude oil to a distillation unit to the shipping of refined products, involves some form of pumping machinery. Intermediate refining stages involve a great variety of pumping machinery to carry out successfully the aims of the refinery technologist. Temperatures at which oils are handled vary from — 50" F. to 900" F. Pressure may vary from subatmospheric to 2000 1b. per sq. inch. The selection of the proper materials involves a careful consideration of the conditions under which the particular pump must operate. The four most important considerations are temperature, pressure, corrosion, and abrasion or wear. Consideration of temperature may be divided into three parts: (1) extremely low temperatures of the order of —50" F.; (2) temperatures at atmospheric or slightly above; (3) temperatures from 200" to 900" F. Pressures fall into two groups: (1) pressures reached by ordinary trade pumps of the order of 200 lb. per sq. in., (2) extreme pressure conditions involving pressures as high as 2000 lb. Noncorrosive and corrosive oils can be treated differently. Under the latter falls corrosion in the presence of water, which will occur up to 250" F. The second type of corrosion occurs in high-temperature pumping equipment, with temperatures from 450" to 900" F., largely caused by sulfur compounds. There are also other corrosive conditions met with in refinery operations, particularly in the handling of acids, acid sludges and other specialized processes. These problems, although difficult to

Jan 1, 1935

By A. E. Harnsberger

I is obvious that details such as the physical and chemical properties and methods of heat-treating of the materials mentioned must be omitted in a paper on the subject of materials used in oil-refinery pumps1. Only sufficient identification is given to enable the reader to find additional information from already published data. Also, the subject covers many operating conditions, but only the most important services can be discussed here. To the oil refiner, the pumping equipment is of vital interest. Practically every refinery operation, from the charging of crude oil to a distillation unit to the shipping of refined products, involves some form of pumping machinery. Intermediate refining stages involve a great variety of pumping machinery to carry out successfully the aims of the refinery technologist. Temperatures at which oils are handled vary from — 50" F. to 900" F. Pressure may vary from subatmospheric to 2000 1b. per sq. inch. The selection of the proper materials involves a careful consideration of the conditions under which the particular pump must operate. The four most important considerations are temperature, pressure, corrosion, and abrasion or wear. Consideration of temperature may be divided into three parts: (1) extremely low temperatures of the order of —50" F.; (2) temperatures at atmospheric or slightly above; (3) temperatures from 200" to 900" F. Pressures fall into two groups: (1) pressures reached by ordinary trade pumps of the order of 200 lb. per sq. in., (2) extreme pressure conditions involving pressures as high as 2000 lb. Noncorrosive and corrosive oils can be treated differently. Under the latter falls corrosion in the presence of water, which will occur up to 250" F. The second type of corrosion occurs in high-temperature pumping equipment, with temperatures from 450" to 900" F., largely caused by sulfur compounds. There are also other corrosive conditions met with in refinery operations, particularly in the handling of acids, acid sludges and other specialized processes. These problems, although difficult to

Jan 1, 1935

By Robert H. Hafner, Irvin R. Kramer

In recent years considerable interest has been shown in surface-alloyed metals, particnlarly those of chromium (chromized steels), which have excellent corrosion resistance under a variety of severe conditions, and may undergo bending and flanging operations without spalling. An early method1 employed to produce a high-chromium case on steel was developed by Kelley in 1921. It consisted of heating a low-carbon steel in intimate contact with powdered chromium in the presence of hydrogen at temperatures sufficiently high to permit diffusion of chromium into the steel. The usefulness of this process was limited because the extremely high tem- peratures of 1200º to 1400ºC. (2200° to 2550°F.) caused excessive grain growth and warpage. Louenstein and Ulmer2 packed iron castings in ferrochromium and com- mon salt, and heated for approximately three hours at 930º to 1040°C. (1700º to 1900ºF.). A recent and more practical method of depositing chromium on steel surfaces has been developed in Germany by Daeves, Becker and Steinberg3-8 and in Russia by Izgaryshev and Sarkisov.9 This process consists of passing gaseous chromium chloride over the steel to be chromized at temperatures of 900º to 1000ºC., or packing them in a porous, ceramic material that previously had been saturated with chromium chloride. They reported obtaining a Chromized layer, 0.004 in. thick, on low-carbon steels. Later, Daeves4 established a relationship between the carbon content

Jan 1, 1943

By Robert H. Hafner, Irvin R. Kramer

IN recent years considerable interest has been shown in surface-alloyed metals, particularly those of chromium (chromized steels), which have excellent corrosion [ ] resistance under a variety of severe conditions, and may undergo bending and flanging operations without spalling. An early method1 employed to produce a high-chromium case on steel was developed by Kelley in 1921. It consisted of heating a low-carbon steel in intimate contact with powdered chromium in the presence of hydrogen at temperatures sufficiently high to permit diffusion of chromium into the steel. The usefulness of this process was limited because the extremely high temperatures of 1200° to 1400°C. (2200° to 2550°°F,) caused excessive grain growth and warpage, Louenstein and Ulmer2 packed iron castings in ferrochromium and com- [ ] mon salt, and heated for approximately three hours at 930° to 1040°C. (1700° to 1900°F,). A recent and more practical method of depositing chromium on steel surfaces has been developed in Germany by Daeves, Becker and Steinberg3-8 and in Russia by Izgaryshev and Sarkisov.9 This process consists of passing gaseous chromium chloride over the steel to be chromized at temperatures of 900° to 1000°C., or packing them in a porous, ceramic material that previously had been saturated with chromium chloride. They reported obtaining a chromized layer, 0.004 in, thick, on low-carbon steels. Later, Daeves4 established a relationship between the carbon content

Jan 1, 1942

By Robert H. Hafner, Irvin R. Kramer

In recent years considerable interest has been shown in surface-alloyed metals, particnlarly those of chromium (chromized steels), which have excellent corrosion resistance under a variety of severe conditions, and may undergo bending and flanging operations without spalling. An early method1 employed to produce a high-chromium case on steel was developed by Kelley in 1921. It consisted of heating a low-carbon steel in intimate contact with powdered chromium in the presence of hydrogen at temperatures sufficiently high to permit diffusion of chromium into the steel. The usefulness of this process was limited because the extremely high tem- peratures of 1200º to 1400ºC. (2200° to 2550°F.) caused excessive grain growth and warpage. Louenstein and Ulmer2 packed iron castings in ferrochromium and com- mon salt, and heated for approximately three hours at 930º to 1040°C. (1700º to 1900ºF.). A recent and more practical method of depositing chromium on steel surfaces has been developed in Germany by Daeves, Becker and Steinberg3-8 and in Russia by Izgaryshev and Sarkisov.9 This process consists of passing gaseous chromium chloride over the steel to be chromized at temperatures of 900º to 1000ºC., or packing them in a porous, ceramic material that previously had been saturated with chromium chloride. They reported obtaining a Chromized layer, 0.004 in. thick, on low-carbon steels. Later, Daeves4 established a relationship between the carbon content

Jan 1, 1943

By Robert H. Hafner, Irvin R. Kramer

IN recent years considerable interest has been shown in surface-alloyed metals, particularly those of chromium (chromized steels), which have excellent corrosion [ ] resistance under a variety of severe conditions, and may undergo bending and flanging operations without spalling. An early method' employed to produce a high-chromium case on steel was developed by Kelley in 1921. It consisted of heating a low-carbon steel in intimate contact with powdered chromium in the presence of hydrogen at temperatures sufficiently high to permit diffusion of chromium into the steel. The usefulness of this process was limited because the extremely high temperatures of 1200° to 1400°C. (2200° to 2550 F.) caused excessive grain growth and warpage. Louenstein and Ulmer2 packed iron castings in ferrochromium and com- [ ] mon salt, and heated for approximately three hours at 930° to 1040°C. (1700° to 1900°F.). A recent and more practical method of depositing chromium on steel surfaces has been developed in Germany by Daeves, Becker and Steinberg3-8 and in Russia by Izgaryshev and Sarkisov.9 This process consists of passing gaseous chromium chloride over the steel to be chromized at temperatures of 900° to 1000°C., or packing them in a porous, ceramic material that previously had been saturated with chromium chloride. They reported obtaining a chromized layer, 0.004 in. thick, on low-carbon steels. Later, Daeves4 established a relationship between the carbon content

Jan 1, 1942

By Nobuhiro Takahashi

PROLOGUE The Yaguki mine, which is located near the southern end of the Tohoku region, Japan, near the coast of the Pacific Ocean, had had a long history as a copper mine, but had not been operated since 1925. About 30 years later it was reopened by Nittetsu Mining Company, and in 1958 a small flotation and magnetic separation plant, a so-called test plant, started production. The ball mill the author first met, a 8x4x4x7 ft tricone mill, called No.1 mill, had started running at that time. With his experience in initial, running of this mill he realized that the ball mill was not simply a preparatory process to flotation, but a determinant factor for flotation performance, an energy-consuming and costly process, and that its operation control was an important task in the plant (1). The working part of an operating ball mill is the ball load inside, whose properties govern the grinding performance. Its size distribution affects the efficiency greatly and the capacity is nearly proportional to its weight. When a ball mill operates for a long period of time with a constant and consistent addition of renewal balls, wear and renewal of balls are balanced, and an equilibrium ball load is formed. A stable equilibrium load is the primary requirement for a stable ball mill operation. However, renewal ball size would be the most influential factor affecting efficiency and capacity of operating ball mills. To improve efficiency, to increase capacity, or to get finer product, attempts to find out the optimum size assortment of renewal balls should be made. When re-

Jan 1, 1976

By Ivan Simon, R. K. McConnell

Anticipating the availability of tiltmeters suitable for use in boreholes, this article proposes a method of using one or more of these meters to determine in situ stress distribution before and after mining operations. Several ways of using relief techniques to measure in situ stresses in rocks have been developed. These techniques, summarized by Merrill,1 involve measuring deformations which take place when a specimen of rock is wholly or partially isolated from the stress field in the surrounding rock. Measurements of rock deformation caused by stress relief from mining operations can, in principle, also be used to estimate in situ stress distributions. Müller2 has mentioned experiments being carried on in Austria with a "deformation indicator" to determine the tilting of holes drilled before a mine opening is made. Similar work using borehole extensometers has been reported by Potts? and Duvall4 has called attention to the possible use of tiltmeters in making these measurements. Considerable attention is now being paid to the use of tiltmeters in detecting rock deformation associated with fault movements, earthquakes, earth tides, etc. The instruments used for these measurements have usually been of the mercury tube, water tube, or quartz fiber type, but because of their large size, and/or fragile nature, such instruments would not be suitable for use in mining stress measurements. Recent instrumentation developments, however, give promise that rugged, sensitive tiltmeters small enough for use in boreholes may soon be available.' In anticipation of these developments, this article presents a number of calculations of the possibility of using one or more tiltmeters emplaced in a drill hole prior to mining operations to determine the in situ stress distribution both before and after a mining operation. Consider point p (x, z) near a tunnel of radius a

Jan 1, 1969