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By R. S. Busk

A PREVIOUS paper1 has summarized the effect of A all metallic elements on the lattice parameters of magnesium. The present paper deals with the effect of temperature on the lattice parameters and the way in which that effect is modified by alloyed elements. The experimental techniques for sample preparation, exposure to the X-ray beam, measurement of line diameters, and calculation of the parameters were the same as reported in the previous paper. The temperature of the specimen was raised by surrounding the mounted compact with a conical Glas-Col,* the base of which was sealed with Saran film. In this way the temperature of the specimen could be raised to 200 °C, which was the maximum desired in the present study. Control of temperature was effected by maintaining a constant, desired voltage to the resistance coils of the Glas-Col through a Variac. Temperature was measured by an iron constantan thermocouple embedded in the body of the specimerl so that the thermocouple bead was within 1/16 in. of the surface exposed to the X-ray beam. Variation of temperature during an exposure was less than + 1/4 oF. A series of eight alloys, consisting of pure Mg, Mg+Ag, Mg+Al, and Mg+Sn were studied. All percentage values are in atomic percent. The results are given in Table I. The density values given in Table I were calculated from the lattice parameters, using an equation due to Foote and Jette:"

Jan 1, 1953

By Ralph Arnold

While much geological work and drilling have been done in Venezuela, the incompleteness of geological evidence obtainable and the restricted areas in which drilling has been done make any estimates of oil resources extremely speculative. Nevertheless, it is felt that the data given are justified, in the belief that the ideas of a geologist familiar with the region are better than those of a promoter. The oldest beds of interest to oil men are those of the Cretaceous, which are well developed in eastern Venezuela and about parts of Lake Mara-caibo. In both cases, a hard massive limestone is overlain by a great thickness, most of which is shale. In the west, in places this shale is more than 2000 ft. (610 m.) thick, while in the east it-mag attain 3500 ft. (1067 m.). The limestone and parts of the shale are oil producing. In the east, the Tertiary where present is almost entirely covered by the Quarternary, but in the west it is well developed. In probable unconformity over the Cretaceous shales is found a series of sandstones and shales, which may be called the lower coal series. It is well developed southwest of the lake, where it is the oil container, and is probably the lowest series exposed east of the lake. Above this is the Pauji shale, which is probably the point of origin of all the oil east of Lake Maracaibo and in the State of Falcon. It has a thickness of about 3000 ft. (914 m.) in the Mene Grande region. The upper coal series is developed as a great thickness of sands, shales, and clays, with some partly consolidated limestone. Its thickness may be as great as 10,000 ft. It is the oil container of the Falcon-El Mene region and a probable location for oil in some of the area about the lake. It may or may not include the so-called Maracaibo series in which the oil is found at Mene Grande. Over much of Venezuela, the Quarternary is present in unknown thickness, which probably is not great. It obscures oil structure in all cases and presents the greatest hazard of oil exploration for the fields which probably exist below.

Jan 1, 1923

By Ralph Arnold

While much geological work and drilling have been done in Venezuela, the incompleteness of geological evidence obtainable and the restricted areas in which drilling has been done make any estimates of oil resources extremely speculative. Nevertheless, it is felt that the data given are justified, in the belief that the ideas of a geologist familiar with the region are better than those of a promoter. The oldest beds of interest to oil men are those of the Cretaceous, which are well developed in eastern Venezuela and about parts of Lake Mara-caibo. In both cases, a hard massive limestone is overlain by a great thickness, most of which is shale. In the west, in places this shale is more than 2000 ft. (610 m.) thick, while in the east it-mag attain 3500 ft. (1067 m.). The limestone and parts of the shale are oil producing. In the east, the Tertiary where present is almost entirely covered by the Quarternary, but in the west it is well developed. In probable unconformity over the Cretaceous shales is found a series of sandstones and shales, which may be called the lower coal series. It is well developed southwest of the lake, where it is the oil container, and is probably the lowest series exposed east of the lake. Above this is the Pauji shale, which is probably the point of origin of all the oil east of Lake Maracaibo and in the State of Falcon. It has a thickness of about 3000 ft. (914 m.) in the Mene Grande region. The upper coal series is developed as a great thickness of sands, shales, and clays, with some partly consolidated limestone. Its thickness may be as great as 10,000 ft. It is the oil container of the Falcon-El Mene region and a probable location for oil in some of the area about the lake. It may or may not include the so-called Maracaibo series in which the oil is found at Mene Grande. Over much of Venezuela, the Quarternary is present in unknown thickness, which probably is not great. It obscures oil structure in all cases and presents the greatest hazard of oil exploration for the fields which probably exist below.

Jan 1, 1923

By Edgar G. Tuttle

The accompanying figures show a jig arranged for separating the middle product or middlings obtained in the concentration of certain ores, minerals, coal, etc. In the preparation for sizing, prior to jigging, of ore-bearing minerals, coal, etc., with their impurities, it is generally desirable to avoid reducing them to unnecessarily small sizes. Where the impurities adhere to the mineral particles or are disseminated through its mass, it is generally necessary to crush to such a fineness as will unlock the minerals from the gangue. If a considerable proportion of the material consists of mixtures of the minerals and gangue, each in large size, and a lesser proportion consists of a mixture of the two, in smaller sizes of various dimensions, it is not always desirable to reduce the whole to such size as would be required to liberate completely the pieces of ore and gangue in the finest mixtures. In such a case it is preferable to reduce all the material to such an

Jan 1, 1897

By Theodore Holt

Outline of the Process THE Mines Operating Co.'s plant at Park City, Utah, was designed to treat the low-grade fillings in the old stopes of the Ontario mine. These fillings carry 6 to 14 oz. of silver, 1 to 2 lb. copper, 0.01 to 0.015 oz. of gold, and a small percentage of lead and zinc. The treatment consists in mixing the crushed ore with coal dust and salt and then roasting in a new type of furnace y combustion of the contained fuel. The roasted ore is leached with an acid salt solution to dissolve the silver, gold, copper, and lead. At present these metals are precipitated together on scrap iron and the product sold to a refinery. The only new feature of importance in this scheme of treatment is the roasting process, which makes possible the chloridizing of ores without any loss of the valuable metals, and at a very low cost. History of the Development of the Process The roasting process had its beginning in connection with research work at the Utah State School of Mines. N. C. Christensen, Jr., who held one of the School of Mines' fellowships, was engaged in some experimental work on blast roasting. At the same time I was doing some work on the insoluble gold in Mercur base ores. Mr. Christensen roasted some of this base gold ore in his pot furnace, and found that when he mixed a sufficiently low percentage of fuel with the ore, it did not sinter but roasted to a leachable product. We considered this a new application of blast roasting, and proceeded to test out a large number of ores in several different types of small roasters which we constructed. This experimental work continued well into the summer of 1911.

Jan 7, 1914

By H. F. Bishop, F. A. Brandt, W. S. Pellini

M. S. Fisher and D. R. F. West (Imperial College of Science and Technology, London, England)—It may be of value to compare certain features of the results recorded in this very interesting paper with those of an investigation11 involving determinations of temperature gradients and rates of solidification of steel blocks of 3 in. sq section, cast horizontally in sand molds, some of which contained chills. Castings in plain carbon steels containing up to 0.8 pct C were made and, throughout this range of compositions, both inverse-rate and direct cooling curves, obtained from thermocouples placed within the mold cavity, showed a major arrest, the temperature of which was taken as the "effective liquidus" for the particular steel in question, solidifying under the conditions of the investigation. At the completion of the arrest, the curves showed a comparatively rapid de- crease in temperature. Some of the steels investigated lay within the peritectic region, and a second arrest at 1485°C was detected. In Fig. 15, which shows the effective liquidus temperatures, the limit of the 8 + liquid field has been placed at approximately 0.5 pct C. In inverse-rate curves corresponding to the thermal center of each casting, there was always a discontinuity, viz: an abrupt change of slope, at a temperature about 10 to 20°C below the "liquidus arrest," after which there seemed to be no significant irregularity in the rate of cooling. Curves corresponding to other positions in the casting generally showed some irregularities below the main arrest, probably as a result of heat flow from the solidifying interior. Even the reproducible discontinuity in the curves for the central positions appeared to have no theoretical or practical significance, though it probably corresponded to a stage in the solidification process at which only a relatively small amount of interdendritic liquid remained. Its temperature, rela-

Jan 1, 1953

By H. R. Stahl

The ore of the Southeast Missouri lead district consists essentially of galena in dolomitic limestone. The galena usually is in a very finely disseminated condition. The experience of recent years in the concentrating plants of the St. Joseph Lead Co. has shown that dry crushing carried to a point where about 90 per cent of the ore passes a 14-mesh screen liberates the larger portion of the galena, and minimizes further necessary reduction in rod or ball mills. Concentration practice has attempted to carry size reduction to a point where economic unlocking of the galena is obtained, without causing undue sliming of the very friable galena. While 60 per cent of the tonnage of the district is treated by the flotation process, only 40 per cent of the total concentrates is produced by that means, as efforts have always been made to keep the grade and tonnage of flotation feed as low as possible. The milling practice of the district for some years has recognized the generally accepted idea of better selective grinding in rod mills as compared with ball mills, with the attendant decrease in production of fines. The function of rod and ball mills in the various concentrating plants is the grinding of table middlings and coarse tailings that carry lead values too high to be discarded. The simplified flowsheet in Fig. I illustrates the place of grinding mills in the circuit. The arrangement of classi- fiers, tables and mills constitutes a closed circuit within somewhat flexible limits. In 1937 experiments were conducted in replacing the ball charge in a 6 by 4-ft. ball mill with rods 2 in. in diameter and 51 in. long. This step was taken with some hesitancy, as it was easy to picture a mass of tangled and crossed rods. However, no difficulties developed, and after a short trial period, six more ball mills were converted to rod mills. These mills are of the spur gear and pinion, belt-driven type, using in most cases a 75-hp. motor and having a 9½-in. open trunnion discharge. As ball mills, they carried a charge of 16,000 to 20,000 lb. of balls, 2 in. in original size. The feed to the mills was dewatered by dewatering wheels and introduced at 70 to 75 per cent solids, with a single scoop feeder. One change in mill design was found advisable when rods were introduced. The existing end liners had an outward flare of 3½ in. in 12 in. These were replaced by linings having vertical interior faces, in order to prevent undue end travel of the rods. This change gave an effective working length of 52¾ in. inside the mill. No alterations were made in the shell liners, which are of the smooth type, made of manganese steel 2 in. in original thickness. The rods at present in use are 51 in. long and 1¾ in. in diameter. The composition has been varied at times; at present they contain 0.70 per cent carbon and 0.70 per cent manganese. Some breakage and etching occurs as the rods become small, but this is considered preferable to using softer rods, which may kink and cause tangling of the charge. The mill charges

Jan 1, 1947

By H. R. Stahl

The ore of the Southeast Missouri lead district consists essentially of galena in dolomitic limestone. The galena usually is in a very finely disseminated condition. The experience of recent years in the concentrating plants of the St. Joseph Lead Co. has shown that dry crushing carried to a point where about 90 per cent of the ore passes a 14-mesh screen liberates the larger portion of the galena, and minimizes further necessary reduction in rod or ball mills. Concentration practice has attempted to carry size reduction to a point where economic unlocking of the galena is obtained, without causing undue sliming of the very friable galena. While 60 per cent of the tonnage of the district is treated by the flotation process, only 40 per cent of the total concentrates is produced by that means, as efforts have always been made to keep the grade and tonnage of flotation feed as low as possible. The milling practice of the district for some years has recognized the generally accepted idea of better selective grinding in rod mills as compared with ball mills, with the attendant decrease in production of fines. The function of rod and ball mills in the various concentrating plants is the grinding of table middlings and coarse tailings that carry lead values too high to be discarded. The simplified flowsheet in Fig. I illustrates the place of grinding mills in the circuit. The arrangement of classi- fiers, tables and mills constitutes a closed circuit within somewhat flexible limits. In 1937 experiments were conducted in replacing the ball charge in a 6 by 4-ft. ball mill with rods 2 in. in diameter and 51 in. long. This step was taken with some hesitancy, as it was easy to picture a mass of tangled and crossed rods. However, no difficulties developed, and after a short trial period, six more ball mills were converted to rod mills. These mills are of the spur gear and pinion, belt-driven type, using in most cases a 75-hp. motor and having a 9½-in. open trunnion discharge. As ball mills, they carried a charge of 16,000 to 20,000 lb. of balls, 2 in. in original size. The feed to the mills was dewatered by dewatering wheels and introduced at 70 to 75 per cent solids, with a single scoop feeder. One change in mill design was found advisable when rods were introduced. The existing end liners had an outward flare of 3½ in. in 12 in. These were replaced by linings having vertical interior faces, in order to prevent undue end travel of the rods. This change gave an effective working length of 52¾ in. inside the mill. No alterations were made in the shell liners, which are of the smooth type, made of manganese steel 2 in. in original thickness. The rods at present in use are 51 in. long and 1¾ in. in diameter. The composition has been varied at times; at present they contain 0.70 per cent carbon and 0.70 per cent manganese. Some breakage and etching occurs as the rods become small, but this is considered preferable to using softer rods, which may kink and cause tangling of the charge. The mill charges

Jan 1, 1947

By C. C. Henning

In any line of human activity logical handling requires an amount of attention to each phase that is in keeping with the importance of that phase. In the complex and rapidly changing field of modern ferrous metallurgy it is difficult to allot the proper degree of attention to the various subjects, and it is not surprising that one or more subjects of major importance should for a time be neglected. Bessemer steel is a good example of a neglected subject. It is largely taken for granted, inferior physical quality is assumed for it, and there is a • lack of interest in its technical development and commercial application. Availability of supply is limited. In spite of this technical indifference, limited source and lack of commercial exploitation, bessemer steel constitutes a large volume of tonnage in daily use. There is a fairly prevalent idea even among technical people that bessemer steel survives mostly because of low cost and by reason of a dwindling residue of applications sufficiently common to permit of its use. The difference in cost between bessemer and open-hearth steel is governed largely by the relative cost of manufacture of pig iron and the market price of steel scrap. During the past four years there have been long periods of time during which the price of steel scrap was low enough so that open-hearth steel was considerably lower in manufacturing cost than bessemer. As to application, it may be pointed out that thousands of tons of this material in the form of bessemer screw steel are applied for moving parts in automotive and other equipment. Some forging work and considerable heat-treating, particularly carburizing, is performed on this grade. From an economic standpoint the most effective contribution of bessemer steel lies in its free machining qualities. It is by far the freest

Jan 1, 1935

By C. C. Henning

In any line of human activity logical handling requires an amount of attention to each phase that is in keeping with the importance of that phase. In the complex and rapidly changing field of modern ferrous metallurgy it is difficult to allot the proper degree of attention to the various subjects, and it is not surprising that one or more subjects of major importance should for a time be neglected. Bessemer steel is a good example of a neglected subject. It is largely taken for granted, inferior physical quality is assumed for it, and there is a • lack of interest in its technical development and commercial application. Availability of supply is limited. In spite of this technical indifference, limited source and lack of commercial exploitation, bessemer steel constitutes a large volume of tonnage in daily use. There is a fairly prevalent idea even among technical people that bessemer steel survives mostly because of low cost and by reason of a dwindling residue of applications sufficiently common to permit of its use. The difference in cost between bessemer and open-hearth steel is governed largely by the relative cost of manufacture of pig iron and the market price of steel scrap. During the past four years there have been long periods of time during which the price of steel scrap was low enough so that open-hearth steel was considerably lower in manufacturing cost than bessemer. As to application, it may be pointed out that thousands of tons of this material in the form of bessemer screw steel are applied for moving parts in automotive and other equipment. Some forging work and considerable heat-treating, particularly carburizing, is performed on this grade. From an economic standpoint the most effective contribution of bessemer steel lies in its free machining qualities. It is by far the freest

Jan 1, 1935

By W. L. Hill, W. H. Armiger, S. D. Gooch

The physical properties, chemical behavior under thermal treatment, and fertilizer value of fluorine-containing pseudowavellite (hydrous calcium aluminum phosphate) that occurs as phosphate clay admixed with quartz sand in an extensive deposit near Bartow, Fla., were determined by experiments in laboratory, pilot-plant, and greenhouse. When the mineral was heated at 500°C, 80 pct of the phosphorus become citrate soluble. The heated product, however, deteriorates markedly under moist conditions such as obtain in the soil. EXTENSIVE beds of a soft aluminiferous phosphate lie in the Bartow-Pembroke region of Florida. Although the extent of the deposit is unknown, it is thought that the equivalent of at least five million tons of material containing 25 pct P,Os exists on one large mine property alone, and soft phosphates similar in nature are known to occur elsewhere in the state. This aluminum phosphate, unlike fluorapatite, is rendered fairly soluble in neutral ammonium citrate solution merely by heating it at a temperature below 600 °C, a property made the basis of a process patent by the junior author.' The possibility of producing an available form of phosphorus from what has hitherto been a worthless natural phosphate prompted the Bureau of Plant Industry, Soils, and Agricultural Engineering to make a brief study of the Pembroke phos- phate during the war years with a view towards determining the optimal temperature for heat treatment and the reactions involved in the process. The results of this study supplemented by data and information furnished by the junior author are presented in this paper. Description of Natural Phosphate: The phosphate with admixed quartz sand and siliceous matter occurs as a low-grade (8 to 9 pct P2O5) clay-like material that can be beneficiated cheaply by making a water separation of the very fine material from the sand, floating off the fine material, settling the suspension to a thick slurry, and finally de-watering the slurry by vacuum filtration followed by drying. The product dried at 100 °C contains about 16 pct water and 25 pct P2O5, about 1/12 of which is citrate soluble. Partial analyses of representative samples of the crude ore and corresponding concentrate are given in table I. The concentrate is light yellow in color and even after being dried at 100°C sticks very noticeably to laboratory grinding and screening equipment. Under the microscope the ground material is predominantly light brown crystalline aggregates with index of refraction near 1.61. The principal phosphate mineral constituent was identified as pseudo-wavellite with admixed phosphosiderite.2 Pseudo-

Jan 1, 1951

By W. L. Hill, W. H. Armiger, S. D. Gooch

The physical properties, chemical behavior under thermal treatment, and fertilizer value of fluorine-containing pseudowavellite (hydrous calcium aluminum phosphate) that occurs as phosphate clay admixed with quartz sand in an extensive deposit near Bartow, Fla., were determined by experiments in laboratory, pilot-plant, and greenhouse. When the mineral was heated at 500°C, 80 pct of the phosphorus become citrate soluble. The heated product, however, deteriorates markedly under moist conditions such as obtain in the soil. EXTENSIVE beds of a soft aluminiferous phosphate lie in the Bartow-Pembroke region of Florida. Although the extent of the deposit is unknown, it is thought that the equivalent of at least five million tons of material containing 25 pct P,Os exists on one large mine property alone, and soft phosphates similar in nature are known to occur elsewhere in the state. This aluminum phosphate, unlike fluorapatite, is rendered fairly soluble in neutral ammonium citrate solution merely by heating it at a temperature below 600 °C, a property made the basis of a process patent by the junior author.' The possibility of producing an available form of phosphorus from what has hitherto been a worthless natural phosphate prompted the Bureau of Plant Industry, Soils, and Agricultural Engineering to make a brief study of the Pembroke phos- phate during the war years with a view towards determining the optimal temperature for heat treatment and the reactions involved in the process. The results of this study supplemented by data and information furnished by the junior author are presented in this paper. Description of Natural Phosphate: The phosphate with admixed quartz sand and siliceous matter occurs as a low-grade (8 to 9 pct P2O5) clay-like material that can be beneficiated cheaply by making a water separation of the very fine material from the sand, floating off the fine material, settling the suspension to a thick slurry, and finally de-watering the slurry by vacuum filtration followed by drying. The product dried at 100 °C contains about 16 pct water and 25 pct P2O5, about 1/12 of which is citrate soluble. Partial analyses of representative samples of the crude ore and corresponding concentrate are given in table I. The concentrate is light yellow in color and even after being dried at 100°C sticks very noticeably to laboratory grinding and screening equipment. Under the microscope the ground material is predominantly light brown crystalline aggregates with index of refraction near 1.61. The principal phosphate mineral constituent was identified as pseudo-wavellite with admixed phosphosiderite.2 Pseudo-

Jan 1, 1951

By J. H. Frye, R. M. Treco

ONE of the chief hindrances to an understanding of the hardness of solid solutions is the sparsity of suitable hardness data. There is great need of a large body of hardness data obtained from many different alloys of high purity by identical testing methods and corrected to correspond to a single grain size. A considerable amount of such data has been obtained by means of the Meyer analysis on copper-rich and silver-rich solid solutions. There are enough of these data to permit comparison of numerous alloys at equal solute concentration in these two solvents, but only in a few cases is it possible to construct hardness-concentration curves for any given alloy. Therefore additional hardness work on most of these alloys is desirable. The primary solid solutions of silver with cadmium, indium, tin, or antimony as solute are of special interest, since all elements are from the same period of the periodic table and since, as shown previously,1 the hardening effect of equal percentages of these solutes is proportional to the lattice expansion produced by them. This paper presents the relation between hardness and concentration in the silver-antimony system from very dilute to nearly saturated solutions of antimony in silver. EXPERIMENTAL PROCEDURE Alloys were made from silver, 99.99 per cent pure, and antimony, 99.975 per cent pure. Silver and antimony were placed in graphite crucibles that had an ash content of less than 0.75 per cent. These metals were melted and allowed to solidify in a vacuum, which, after melting was complete, was usually equivalent to 3 mm. Of mercury. A high-frequency furnace was used. The ingots so obtained were worked and annealed to lessen segregation and to control grain size. Annealing was done in evacuated and sealed glass tubes. Hardness measurements on specimens from these ingots were by means of the Meyer analysis.3,4 Using a 4-mm.. ball, impressions were made under seven loads from 25 to 400 kg. The diameter of impression made under each load was measured. This enabled calculation of the constants in the well-known formula of Meyer. L=ad¬n where L = load, d = diameter of impression, a = a constant, n = a constant The Meyer hardness is defined as P = L/(1/4pd2) or P = 4adn-2/p Meyer hardnesses corresponding to d = 1 and to d = 4 are of especial interest, since the first corresponds to comparatively slight plastic deformation and the second to the maximum plastic deformation producible by penetration of the ball. These

Jan 1, 1944

By F. S. McNicholas

IN the first big blast at Climax, a "loop back" (three-wire system) was used (Fig. 1), with the idea of securing a wiring system that would give to all series the same amount of current. Single-phase, 60-cycle, 13,800-volt FIG. 1.-Loop BACK. A THREE-WIRE SYSTEM WITH BRANCHES VARYING IN LENGTH. power was furnished to the blasting transformers, each of which has a rated capacity of 600 amp. at 480 volts. Two blasting leads were taken from these transformers, giving a blasting current of single phase, 60 cycle, 480 volts, with an instantaneous capacity of more than 600 amp. The blasting switch was on the secondary leads from the transformers, to avoid lag in energizing them.

Jan 1, 1938

By Samuel Storchheim

IT was noted that variations existed in the magnetic properties, namely, coercive force, He, and rema-nence, Br, of 18/8 stainless steel wires of the same analysis when these wires were given the same percentage of reductiori of area at different rates of reduction. In order to determine the extent of these variations, a program was initiated whereby commercial 302 and 304 stainless steels were drawn with a wide variation of speeds. Ultimate tensile strengths and specific resistances, as well as the Hc-BT measurements, were determined. If variations of considerable magnitude were found, it was thought that the knowledge accrued from the investigation could probably aid in determining the particular drawing speed required to obtain desired properties for a given specification for a particular analysis of wire. Drawing Procedure Two specimens (set I, No. 302 and No. 304), the chemical analyses of which are given in Table I, were drawn at different speeds, namely, 45 to 3900 ft per min, directly from 0.016 in. diam wire to 0.004 in. diam wire, a 93.7 pct reduction of area. To obtain this high percentage of reduction, the wire at 0.016 in. was softened by annealing at 1150°C. It was then strung in a No. 16CF variable speed Vaughn drawing unit and passed through 13 B.&S. gage dies of successively smaller diameter, as shown in the accompanying table. Following this step, several hundred feet of the wire was drawn, at a particular speed, through the dies and wound upon a spool. Once the sample of wire was drawn at the desired speed, the spool with the finished sample was removed, tagged for later identification, and an empty spool was placed in the machine. Another wire sample was drawn at a different speed, after making the necessary adjustments upon the Vaughn unit. At the greater speeds, several thousand feet of sample were drawn to insure that the wire to be used for testing would be uniform. This was done because the desired speed was not reached instantly; the wire taking a short while to attain the required drawing rate. All speed measurements were made by placing a tachometer against one of the machine's pulleys during the actual drawing of the specimens. 84.4 pct reduction of area. The 0.072 in. diameter wire was also annealed at 1150°C. The process of obtaining samples was the same as described above. The only difference between the two was that the rates involved were not as extensive, set II covering a range of 625 to 1150 ft per min. Tensile tests were carried out for the 0.004 in. diam wires on a horizontal Scott Tensile Tester, while tensile tests for the 0.028 in. diam wires were made on a vertical Scott Tensile Tester. Each value listed in the tables represented the average of five samples with a maximum deviation of approximately 3 pct among the samples. Resistance per foot of the wires was determined by use of standard bridges accurate to the nearest hundreth of an ohm for the 0.004 in. wire, and to the nearest thousandth of an ohm for the 0.028 in. wire. The resistances per foot were then converted to resistivities by use of the following equation: P = RxCM where R is the resistance per ft in ohms per ft; CM, the area in circular mils; and P, the resistivity in ohms—CM. The values listed in the tables likewise

Jan 1, 1952

By C. H. Benedict, H. C. Kenny

A 2000-ton ammonia leaching plant has been operated by the Calumet & Hecla Mining Co., at Lake Linden, Mich., continuousl~ since February, 1917, except from April, 1921, to April, 1922, during the period of depression following the European war. A description of this plant has been published elsewhere.1 The present article describes the operation of the plant since 1917, together with the changes in the process that have been found to be advisable. In spite of the fact that this plant and that of the Kennecott Copper Co. are the only ones operating on a large scale, much experimental work is being done on ammonia leaching, as applied to oxidized ores, and it is to aid these and other investigators that this detailed account is given. For the benefit of those unfamiliar with the original article, the following short abstract is given, as an introduction to a more detailed account of the process and the reasons for present practice. Rêsumê of Previous Article Leaching on a commercial scale began in July, 1916, and by Februsry, 1917, the plant of 2000 tons capacity was in commission. The ore treated is a conglomerate containing native copper and no sulfides or oxidized copper minerals. Before leaching, the ore is crushed by steam stamps and Hardinge pebble mills to pass a 28-mesh screen. Free copper is removed, by gravity concentration, on Woodbury jigs and Wilfley tables, and colloidal slime for oil flotation is removed by quadruplex Dorr classifiers. The leaching sand remaining contains about 2 per cent. + 28-mesh material and 15 per cent. — 200-mesh and carries from 8 to 14 lb. of copper per ton, depending on the grade of ore worked. The solvent employed in leaching is a water solution of ammonium carbonate,

Jan 1, 1924

A. Cook K.A. Grace's article advocates usage of the terms describing resources as defined by the US Bureau of Mines (USBM) and the US Geological Survey (USGS). Originally recommended in 1943, the terms were revised in 1976 and again in 1980. Mr. Grace further recommends that the term reserves should be severely restricted and that it should encompass mining and processing factors. The list of references quoted in the article does not, however, mention the study concerning reserves that was made over a period of almost four years by a joint Committee on Definitions. This committee was sponsored by three societies - the American Institute of Mining, Metallurgical, and Petroleum Engineers (AIME), the Society of Economic Geologists (SEG), and the American Institute of Professional Geologists (AIPG). Most of the detailed work was done by a working party appointed by the Society of Mining Engineers (SME) of AIME. After studying previous definitions in the US as well as in the United Kingdom, Canada, Australia, France, and other European countries, the following definitions were proposed and accepted in 1974 by the SME board of directors. (Ore reserve definitions in Rhodesia, South Africa, or New Zealand were not mentioned as having been reviewed. Table 1 lists definitions from MINING ENGINEERING, July 1975. Notice in Table 1 it was mentioned, under Resources, that the definitions of reserves were recommended for the purposes of commercial evaluation. It was further stated that the terminology adopted by the USBM and the USGS, as released to the public, covering resources, was recognized as such. Mr. Grace also recommends that reserve calculations should allow for future losses in ore tonnage, ore grade, milling, and smelting. Unfortunately, it might take years to develop a prospect after it has been discovered. Mining methods have to be tested before being finalized in order to know what tonnage of ore is going to be lost, or how much dilution is likely to occur. The same problems are present in concentrating or in smelting. Operating mines usually know how mined grade compares with ore reserve grade. They can usually predict fairly closely what their final metal recoveries are probably going to be. It is, however, expecting too much to have to delay quoting reserves from say a recently drilled prospect. Any responsible mining geologist or mining engineer (or company for that matter) should state several things about an ore reserve calculation: that it is or is not a preliminary estimate only; that estimated grades have not been diluted to allow for a future, as yet unknown, mining method; that metal recoveries are or are not only tentative, based perhaps on drill hole samples only or on no metallurgical testing at all. [ ] Reserves of a partially explored prospect are always quoted when it becomes necessary to raise capital. If they are described as preliminary only (really just tonnage and grade estimates and not true "reserves") in addition to being accompanied by some of the items just mentioned, nobody is going to be misled. If, on the other hand, a report by a member of the SME or SEG is intentionally misleading, the individual must be reported and eventually disbarred from his or her society. To summarize, let us stay with the definitions of ore, ore reserves, and related terms as recommended and accepted by SME. I might add that on occasion I have used a fourth category in my ore reserve calculations, namely Geologically Possible Ore. This category is more or less the same as Inferred ore defined by the USBM and USGS. I have never given a grade to this category, which may be based only on geological mapping, geochemical sampling, or geophysical traversing. ?

Jan 1, 1986

By A. A. Burr, D. Caplan

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

Jan 1, 1956

By C. D. Hall, F. E. Dollarhide

Conventional static tests of fluid-loss agents do not realistically simulate conditions in a fracturing treatment. The dynamic tests reported here show that fluid-loss volume is better represented as proportional to time, rather than as the square root of time. This leads to a different equation for fracture area. The leak-off rate increases with increasing shear rate at the fracture wall, but appears to approach a limiting value. Pressure effects are minor. Spurt loss ordinarily is not affected by the flow velocity in the fracture and is inversely proportional to concentration of agent. The filter cake, once it is well established, is resistant to damage by the flow of plain fracturing liquid (without fluid-loss agent). The latter two findings indicate that a treatment employing a high-concentration spearhead followed by plain fluid can offer a more economical treatment under suitable conditions. INTRODUCTION The successful design of hydraulic fracturing treatments depends on accurate knowledge of the fluid-loss properties of the fracturing fluid. Howard and Fast,' in giving the basic equation relating fracture area to fluid and treating parameters, described three mechanisms which might control the rate of fluid leak-off from the fracture. One mechanism usually is dominant in a given well treatment. For each mechanism, the leak-off velocity is inversely proportional to the square root of time, and the proportionality constant is designated as the fracturing-fluid coefficient. For the wall-building type of fluid-loss agent, the coefficient is determined by a filtration test in a pressure cell, usually with a rock wafer or core as the filter medium. In these static tests, the cumulative volume generally is proportional to the square root of time, after an initial spurt volume. The static-fluid-loss test is not representative of the con,-&tions under which a fluid-loss agent performs in a fratturing treatment. The marked difference between the dynamic- and static-fluid-loss behavior of drilling fluids reported in the literature2,3 implies that dynamic testing is also needed with fracturing fluids. We have therefore undertaken a study of the dynamic-fluid-loss behavior of fracturing fluids. The testing apparatus has also afforded opportunity to evaluate the resistance of the filter cake to removal or damage by flowing fluid containing no fluid-loss agent, with and without sand. The results of these studies offer a means for more accurate evaluation of fluid-loss agent performance, and point the way to a "spearhead" fracturing technique which may offer more economical treatment for some wells. EXPERIMENTAL METHODS The dynamic-fluid-loss testing method is applicable to any type of wall-building fracturing fluid. The present study aimed first at finding what phenomena are involved, and therefore has been limited in the number of materials tested. All of the results specifically reported herein are for kerosene containing a commercial solid fluid-loss agent, which is commonly used at 50 lb/1,000 gal of oil. Another agent in liquid form, used usually at 20 ga1/1,000 gal oil, has shown all the same phenomena in dynamic tests, and generally the same level of fluid-loss control as the solid agent. The dynamic-fluid-loss core cell used in all tests is shown in Fig. 1. The fracture was simulated by the an-nulus between a 2.03 in. OD sandstone core and the surrounding pipe. Annulus widths of 0.234 and 0.117 in. were used, and the core was 3.5 in. long. The annular geometry provides a uniform fluid velocity and a well-defined shear rate over the entire filtering surface, and permits a large filter area (144 sq cm) in a reasonably compact cell. The leak-off fluid passed into a 0.5 in. diameter axial hole in the core. A hollow steel rod through this hole was threaded into a rounded "streamliner" upstream of the core, and into a mounting stud downstream. The streamliner and the stud had the same outside diameter as the core. In all tests except those where sand was circulated, the mounting stud had protruding rings which constricted the annulus, to minimize any tendency for channeling of the fluid to the side exit port. The ends of the core were sealed by Neoprene, steel and Teflon washers. The leak-off fluid was conducted from the hollow rod to an exit tube, through a metering valve (a fine-pitched needle valve) and a quick-opening toggle valve in series, and into graduated cylinders for volume measurement; Two separate circulating systems were used in the experimental program. The extensive initial testing was done at 50 to 150 Psi. The fluid was circulated by a variable speed Moyno pump, and the flow rate was read by a rota-meter flow meter. The filtration Pressure was supplied by holding a back-pressure with a throttling valve. The discharge streamcould be diverted into any of four sections

Jan 1, 1965

The Institute conducts jointly with the American Society of Civil Engineers, American Society of Mechanical Engineers and American Institute of Electrical Engineers, certain activities as listed below, and is joint owner with them of certain properties. United Engineering Trustees, Inc. When Mr. Carnegie made his gift which made possible a home for the engineering societies, a separate corporation was organized to act as a holding company and to manage the common property. This was long known as the United Engineering Society, but the name has 'now been changed to one which more nearly reflects the real function of the corporation. It holds title to the Engineering Societies Building and to various funds placed in its charge and is managed by a Board of Trustees, of which three are chosen by each of the four societies, American Institute of Mining and Metallurgical Engineers, American Society of Mechanical Engineers, American Institute

Jan 1, 1934