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By John E. Riddell

This review is based on a questionnaire sent out in 1959 to 24 Canadian exploration companies, known to be active in the mineral exploration field, concerning their practice in the use of applied geo-chemical techniques in Precambrian terrains. Of the companies questioned, 21 returned completed questionnaires and of these, 15 companies stated that they had some experience in this type of work. Of the companies who stated that they had no experience in geochemical prospecting in glaciated Precambrian areas it is interesting to note that only three have reported that they do not use applied geochemistry at all. The others have reported that they make use of applied geochemistry in New Brunswick and in the Cordillera in areas of light glaciation. Their opinion is that it is a useful aid when accompanied by other techniques in evaluating mineralized areas. GEOCHEMICAL SURVEYS Table I reviews the type of geochemical surveys used by these companies, and the following points of interest should be noted.

Jan 11, 1960

By Howard Eavenson

THE mines of the United States Coal & Coke Co. are located in the Pocahontas coal field, in McDowell County, West Virginia. Twelve plants have been opened and equipped, of which, by reason of the present business depression, only nine are now in operation. These twelve plants serve eighteen mines; some of the plants having several openings, in either the same seam or in different seams, tributary to the same tipple. Only nine of these mines are now being worked at the plants in operation. Construction work was begun in 1902, and coal from the first three plants started was shipped in December of that year. An additional plant was opened in 1903, two in 1904, two in 1905, two in 1907 and the last two in 1908. On Jan. 1, 1915, the plants have been in operation an average period of 8.8 years. The rules of the company, which have been in operation since work was first begun, are similar to those in use by the H. C. Frick Coke Co. at its mines in Pennsylvania, in which "Safety First" was incorporated as early as 1907; and it has always been the policy of both companies to look carefully after the safety of their employees. Although the field is an extension of the old Pocahontas field, the mines, located upon a new branch railroad, are away by themselves, and, as is usual in such cases, the class of labor attracted at first was temporary in character. The rapid development of the mines required the employment of many new men and officials each year; and, while the accident record from 1904 to 1909 was as good as the average of the field, it was felt to be not as good as it could be made, and a determined campaign was inaugurated for preventing accidents of all kinds. Fig. .1 shows the tonnage of coal produced per fatality, both total and for inside accidents only, at our mines; the tonnage per fatality of all West Virginia mines, and of those in the Pocahontas field.

Jan 2, 1915

By Nathan C. Rockwood

LIME is a very general term applied to products of limestone, in popular treatises often incorrectly, including ground or pulverized limestone used in agriculture. When used without qualifying adjective, the term usually means burned or calcined limestone, or quicklime, or calcia. The corresponding calcined product of magnesite is magnesia. If the calcined rock is relatively pure calcium carbonate, the resulting lime is a "high-calcium" lime; if the rock is dolomite, or a combination of calcium and magnesium carbonates, the lime is a "dolomitic" or a "high- magnesium" lime. There are various grades in between, depending on the percentage of magnesium carbonate and other ingredients in the raw rock. For some industrial purposes only high-calcium limes are suitable; in other processing-as, for example, neutralizing acid wastes -the relative percentages of magnesium and calcium carbonates in the raw material are not important. Limestone is burned (roasted or calcined) in furnaces known as lime kilns by direct exposure to the hot gases and flames. When heated to incandescence, the carbon dioxide content of the rock is expelled, leaving calcium and magnesium oxides, according to the formula usually written: [ ] This thermal reaction is reversible, which requires that the lime be withdrawn from the "calcining zone" of the kiln about as fast as made. Pure calcium carbonate would lose 44 pct of its weight in the calcining process; dolomite, 48 pct. Generally, the lime will retain the approximate size and shape of the piece of rock from which it is made but some few limestones tend to break down into granules or dust. Hence lime (quicklime) comes from the kilns in lumps or granules and normally is packed in water- tight containers in these forms, or is pulverized after calcination.

Jan 1, 1949

By R. Hadfield

Last year this institute was good enough to accept some remarks by the writer regarding sound steel, entitle Plant for Hadfield Method of Producing Sound Steel Ingots…

Jan 1, 1915

By Joshua I. Tracey, Mackenzie Gordon

THE bauxite deposits in central Arkansas were formed by weather¬ing, in early Eocene time, of fresh or kaolinized nepheline syenite above the water table in a subtropical climate of fairly continuous moisture. Four types are recognized: (1) residual bauxite deposits on kaolinized nepheline syenite, (2) residual deposits on kaolin and bauxite detritus, (3) alluvial deposits of sorted pebbles and pisolites of early Wilcox age, and (4) rubble deposits of later Wilcox age. All were invaded by kaolin and siderite during subsequent burial by Eocene sediments. PART I-FIELD RELATIONS OF THE DEPOSITS The Arkansas bauxite region for years has produced about 95 pct of the domestic supply of bauxite of the United States and it contains today reserves of more than 40 million tons-nine tenths of the known domestic reserves. During World War II, the Arkansas bauxite deposits were explored jointly by the U. S. Geological Survey and the U. S. Bureau of Mines. Some 6000 holes were drilled by rotary rigs under contract to the Federal Govern¬ment, which resulted in the discovery and delimiting of approximately 20 million tons of ore in place. Detailed study of the many sections cored through the bauxite, of data from a much greater number of drill holes made available by all the operating companies, and of the extensive mine faces exposed in the strip pits, has yielded valuable geologic information about the origin of the deposits. It is now possible to piece together a paleotopographic picture of the deposits and the now-buried land surface on which they formed. The bauxite region in the center of Arkansas (Fig 1) contains two producing districts: one in Pulaski County, just south of Little Rock, and the other in Saline County, about 15 miles to the southwest. The deposits in both districts are associated with hills of nepheline syenite, bosses of a large batho¬lith that was intruded into Paleozoic rocks and later buried beneath coastal plain sediments deposited during Cenozoic time. The folded Paleozoic rocks are exposed in the northwest half of the state and are covered in the southeast half by the nearly flat-lying coastal plain sediments. The older rocks were intruded by nepheline syenite, probably in Late Cretaceous time, and sub¬sequently were eroded so as to expose the top of the batholith. In Paleocene time, marine limestone and laminated black silty clay of the Midway group were deposited in an advancing sea, and covered all but the most elevated parts of the batholith. Erosion of the islands of nepheline

Jan 1, 1952

By Henry Rawdon

THE structure of wrought iron as usually described by metallographists and workers in metal in general is that of a fairly pure iron. Impurities, if present, are usually considered as being in solid solution in the crystals of the ferrite matrix or as forming part of the ever present "slag streaks." Attention is herein directed to one type of these dissolved impurities; to the detection of such impurities, particularly in low-grade irons, and to their possible influence on the physical properties of the metal. The attention of the Bureau of Standards was first directed to wrought iron of the peculiar and unusual characteristics to be described later, in material which had failed in service and was submitted for test. The examination was extended to other grades of wrought iron, to see whether such features are of common occurrence in this class of material. MICROSTRUCTURE OF WROUGHT IRON Usual Structure The general microstructure of wrought iron is so well known that a detailed description of it here is needless. There is shown in Fig. 2 the structure of a sample of Swedish iron which illustrates well the matrix of ferrite crystals in which are embedded the slag threads so characteristic of the puddling process and the subsequent working of the material. The usual commercial product contains varying amounts of the impurities commonly associated with iron and steel, i.e., manganese, phosphorus, silicon, and sulphur together with small amounts of carbon. Of these, the larger part of the phosphorus and some of the silicon present are held in solid solution in the ferrite while most of the remaining impurities, other than carbon, exist in the inclosures of slag. Aside from the discontinuities introduced by the slag streaks the ferrite matrix has the microstructure and appearance of a pure metal. No definite orientation of the various crystals or grains is apparent; deep etching of the specimens reveals no intracrystalline features (i.e., within the grains) other than those noted in "pure" metals in general (etching pits).

Jan 9, 1917

By A. L. Labbe

SOON after the Cottrell treater was placed in operation in the Murray plant in 1918 to treat combined lead sinter and Wedge roaster smoke, it was noticed that the power flowing through the treater did not remain constant. This was indicated by the varying milliamperes and also by the total amount of power consumed by the rectifiers. At times, for then unknown reasons, the treater current fluctuated through a wide range of from 40 to 300 milliamperes. Fortunately, these variations in power did not affect the treater's overall recovery, as this installation consisted of three independent units in series, a feature which made the Murray Cottrell an outstanding installation over many years of operation. Water conditioning of the smoke as a means of improving recovery was already known, but was not adaptable to this installation for reasons of excessive corrosion, which had been the case with other treaters using water as a conditioner exclusively. Tests conducted on the smoke had proved conclusively that water vapor played no part in the fluctuation of power, and the same was true of the SO2 contents and dust burden of the smoke. Neither did temperature variations of from 150 to 400°F have any effect on the power. Finally it was noticed that the variation in power taken by the treater, and the efficiency of recovery, had some definite relation to the number of Wedge roasters operating, and particularly to the sulphur contents of the charge. This observation soon led to the discovery that free sulphuric acid was the conditioner for Murray smoke and that variations in acid contents of smoke accounted for fluctuations of treater power. Analysis of recovered dust revealed that only a few hundredths of a per cent free acid was necessary to maintain a very efficient recovery. Once this knowledge was available, the roaster charge was adjusted as to sulphur contents to produce the necessary acid conditioner. For a number of years this practice was followed, but with improvements in the field of flotation, excess pyrites were eliminated from the smelting picture, so with a change in metallurgical practice we were confronted with the problem of providing the deficiency in acid by some other means. The situation was aggravated and our problem of acid conditioning made more difficult by the increase in the lead contents of concentrates roasted resulting from better flotation methods. The first step towards introducing acid vapors into the smoke stream by accessory means was accomplished by boiling sulphuric acid in cast iron pots placed in an open fire box. Acid fumes evaporated from the pots together with the combustion gases from the fire box were discharged into the flue through a cast iron pipe. Capacity of the iron pots was quite limited because of the comparatively slow rate of evaporation and the use of lump coal as a fuel, which did not lend itself to practical temperature control. This method of firing resulted in frequent pot failures due to overheating. In spite of these incidental difficulties, encountered in any new venture, the pot evaporator demonstrated the practicability of fuming sufficient acid to make up the deficiency in that naturally evolved in the Wedge furnace operation. As time progressed, less and less high sulphur material was available for Wedge roasting, and the necessity for accessory acid conditioning reached a climax when the supply of this material was entirely eliminated. Since the cast iron pots could only evaporate a few hundred pounds of acid per day, and were costly to replace, the logical thought presented itself of spraying the acid in a heated chamber. This new type of acid evaporator was simpler and more economical to operate and had sufficient capacity to fume several thousand pounds per day. The attached fig. 1 presents the outstanding features of the new furnace. A is the vaporizing drum, which is heated by the combustion gases of coal stoker fire box, B. Acid is sprayed into the hot gas stream within the drum through a number of air-acid atomizers, C, and the acid vapor, together with combustion gases, is introduced into the smoke stream through two cast iron pipes extending to different points in the flue. The operation of this type of furnace requires control of temperature inside the drum to prevent overheating, which dissociates the acid vapor to SO, and eventually to SO2, with resulting loss of acid. The dissociation to SO, at elevated temperatures is comparatively high, and in some cases accounts for as much as 60 pct of the acid sprayed into the drum. To reduce the loss of acid, through dissociation, some Cottrell plants, where the rate of acid sprayed reaches at times 9 ppm, use two evaporating units. Fuming the acid into the smoke stream is only

Jan 1, 1951

By R. W. Kobler, D. A. Dahlstrom

Since the development of the pelletizing process in the fifties for relatively high purity pellets from low grade taconites, there have been several distinctive changes in filter construction for the dewatering step preceding balling. In the original large pilot plant run by Reserve Mining Co. in Babbitt, MN, the disk type filter employing agitation (Agidisc filter) was selected as the most economical. Size was limited to 2.1 m (6 ft, 9 in.) in diameter. The second generation occurred in the sixties with a jump to 2.70 m (8 ft, 10 in.) in diameter to reduce capital costs per annual ton of pellet. In the early seventies, the third generation was the steam drying disk filter which became necessary to achieve balling moistures with specific surface areas of 2200 cm2/g to as high as 3800 cm2/g caused by finer grinding to obtain higher purities. Finally, the fourth generation has recently been developed to reduce maintenance costs and further improve hydraulics to ensure minimum moisture contents.

Jan 1, 1980

By C. F. Smart

SINCE their comparatively recent development, the alloys of cadmium with silver and copper or nickel, and of cadmium with nickel alone, have been used somewhat extensively as liners for connecting rod bearings and main bearings of internal-combustion engines. In applications where, owing to high hearing loads and elevated operating temperatures, the life of hearings made from tin-hale or lead-base alloys has been undesirably short, cadmium alloys have demonstrated considerable advantage. In applications where, because of deflection of the crankshaft or bearing supports, copper-lead bearing metals have resulted in scored shafts, cadmium alloys have been operated successfully. The somewhat serious drawback to these alloys in hearing service has been that they, in common with copper-lead and certain lead alloys, are susceptible to corrosion in the presence of lubricants containing organic acids. These acids may be fatty acids (such as oleic acid) derived from animal or vegetable oils that sometimes have been compounded with engine lubricants, or they may be compounds of an acid nature developed by the oxidation of some petroleum oils under operating conditions. The result of such acid attack on hearings is illustrated by the photograph, Fig. 1. The petroleum oils that are most readily oxidized to a corrosive condition, and the operating conditions under which such oxidation most readily occurs, are now rather well known among the technical organizations of both the petroleum and automotive industries. While means have been developed for inhibiting petroleum oils against such acid formation, the knowledge that this condition may be encountered has been a factor in restricting the wider use of cadmium alloys as engine bearings. It has appeared, therefore, that their value and use would be enhanced if they could be made corrosion-resistant to a degree compatible with the most severe operating requirements. Investigation, undertaken with this in view, indicated that the addition of fractional percentages of the metal indium to the cadmium bearing

Jan 1, 1938

By W. H. Moore, E. T. Powell

IN the early days of mining in the Anthracite field, only the thicker and better seams of coal were mined, because of the limited mining and coal-cleaning facilities, therefore many of the thinner and less productive seams were permitted to remain unmined. With improvements in mining methods, the introduction of modem preparation plants, and the increased market demand for small-size coal, the thinner and more laminated seams have now become assets. In many instances these seams overlie the mined-out beds of the thicker and better seams, from which all recoverable coal was removed by the conventional breast-and-pillar method of mining. Some collieries in certain sections of the Anthracite field are now relying on these thinner and laminated seams. For the purpose of illustration, a large area in the Hickory Swamp-Ridge Basin of the western middle coal field has been selected, where as early as 1898 the No. 4, or Little Buck Mountain, seam was mined and pillars removed. The No. 5, or Buck Mountain, seam lying on a comparatively light pitch and with intervening strata of but 55 ft. of rock was not mined until 1935. Farther east, in the same basin the Nos. 8 and 9 seams, which are two splits of the Mammoth seam, have been mined and robbed and the workings have been abandoned since 1896. The overlying No. 9½ seam, or the upper split of the Mam- moth, and the No. 9¾ or 4-ft. seam remained virgin, with only an interval of 50 ft. of rock separating them, and with 40 to so ft. of rock strata separating the No. 9½ seam from the completely mined area of the Nos. 8 and 9 seams. These seams lie on a pitch of 45° on the south side of the basin and on 75° on the north side of the basin. Mining has been carried on in the 9½ and 9¾ seams for 3 years on both dips of the basin, for a distance of 3500 ft. In some sections the overlying strata of the thicker seams previously mined have settled without any apparent breaks in the upper strata, as indications have shown in the top split of the Mammoth; while in other areas the roof and bottom of this top split of the Mammoth have settled unevenly, thereby causing the bottom and roof to break into blocks. In attempting to mine over a large area of the No. 5 or Buck Mountain seam, a similar settlement has been encountered; the roof of the seam has been broken and a considerable amount of timbering has been necessary in order to make the mining of the seam possible. The conventional breast-and-pillar method, using shaker conveyors, has been used in the No. 5 seam, and a slant-breast method in the mining of Nos. 9½ and 9¾ seams. Mining OF Light-pitch Seams The No. 4 seam was mined and robbed by the conventional breast-and-pillar system. In this area the seam pitched from 5° to 25° on the north dip, averaged 8 ft. in thickness and had a slate bottom and a

Jan 1, 1942

By W. H. Moore, E. T. Powell

IN the early days of mining in the Anthracite field, only the thicker and better seams of coal were mined, because of the limited mining and coal-cleaning facilities, therefore many of the thinner and less productive seams were permitted to remain unmined. With improvements in mining methods, the introduction of modem preparation plants, and the increased market demand for small-size coal, the thinner and more laminated seams have now become assets. In many instances these seams overlie the mined-out beds of the thicker and better seams, from which all recoverable coal was removed by the conventional breast-and-pillar method of mining. Some collieries in certain sections of the Anthracite field are now relying on these thinner and laminated seams. For the purpose of illustration, a large area in the Hickory Swamp-Ridge Basin of the western middle coal field has been selected, where as early as 1898 the No. 4, or Little Buck Mountain, seam was mined and pillars removed. The No. 5, or Buck Mountain, seam lying on a comparatively light pitch and with intervening strata of but 55 ft. of rock was not mined until 1935. Farther east, in the same basin the Nos. 8 and 9 seams, which are two splits of the Mammoth seam, have been mined and robbed and the workings have been abandoned since 1896. The overlying No. 9½ seam, or the upper split of the Mam- moth, and the No. 9¾ or 4-ft. seam remained virgin, with only an interval of 50 ft. of rock separating them, and with 40 to so ft. of rock strata separating the No. 9½ seam from the completely mined area of the Nos. 8 and 9 seams. These seams lie on a pitch of 45° on the south side of the basin and on 75° on the north side of the basin. Mining has been carried on in the 9½ and 9¾ seams for 3 years on both dips of the basin, for a distance of 3500 ft. In some sections the overlying strata of the thicker seams previously mined have settled without any apparent breaks in the upper strata, as indications have shown in the top split of the Mammoth; while in other areas the roof and bottom of this top split of the Mammoth have settled unevenly, thereby causing the bottom and roof to break into blocks. In attempting to mine over a large area of the No. 5 or Buck Mountain seam, a similar settlement has been encountered; the roof of the seam has been broken and a considerable amount of timbering has been necessary in order to make the mining of the seam possible. The conventional breast-and-pillar method, using shaker conveyors, has been used in the No. 5 seam, and a slant-breast method in the mining of Nos. 9½ and 9¾ seams. Mining OF Light-pitch Seams The No. 4 seam was mined and robbed by the conventional breast-and-pillar system. In this area the seam pitched from 5° to 25° on the north dip, averaged 8 ft. in thickness and had a slate bottom and a

Jan 1, 1942

By J. B. NEALEY

MANY nonferrous smelting plants have recently adopted natural gas as fuel with resultant economies, both in cost and efficiency of utilization. Not only has this fuel been used for roasting, reduction, sintering, drossing, and smelting, but also for power. All of the three big smelters near Salt Lake City, Utah, have been changed over from coal and oil firing to gas. Among these is the plant of the International Smelting Co., Tooele, an Anaconda subsidiary, smelters of both copper and lead ores. In the roasting division at Tooele are 32 MacDougall six-hearth roasters, each 16 ft. in diameter and 19 ft. high. Two gas burners are installed on each furnace. These burners are of the single-valve, automatic gas-air proportioning type using high-pressure gas at 10 lb.

Jan 1, 1933

By Kenric C. Owen

INTRODUCTION Situated 23 miles east of Pretoria the Premier Mine started diamond production in 1903. Two years later it produced the largest diamond yet discovered, the 3 106 carat Cullinan stone. For the period to 1932 when the mine was closed down due to economic circumstances, open cast mining was practised using inclined rope haulage traction and by this time had produced 48.5 million carats of predominantly industrial quality diamonds from 148 million tons of ore. In 1945 it was decided to re-open the mine using underground mining methods. A system using long hole benching was devised (Hodgson and Sewel 1960). This system was later modified but still forms the basis for approximately fifty percent of production. Following the successful introduction of block caving on associated diamond mines in the Kimberley area (Gallagher and Loftus 1960) this mining method was introduced to the Premier Mine in the late 60's and now accounts for nearly 50% of production. This paper will discuss block caving at Premier Mine in the light of our experience and discussion will be directed to the main geological and structural features of the orebody and host rocks and the related constraints imposed on the mining methods. Reference is made to the long hole benching mining method for comparative purposes. GENERAL DESCRIPTION OF OPERATIONS The ore from the block caving and bench mining areas gravitates via ore passes to a twin haulage, 500m below surface. Electric 13 tonne locomotives, each hauling a train of 10 Granby type cars, deliver the ore to two 42" x 48" (1.06m x 1.22m) jaw crushers. The ore is reduced in size to minus 0.15m before being hoisted to the treatment plant on surface. Hoisting is done with 12.5 tonne bottom discharge skips in a five compartment rectangular shaft using two 3 240 H.P. semi automatic Ward Leonard winders. A small single drum service winder operates in the other compartment. Men and materials travel in a separate shaft in a 5,4m x 2,8m cage operated by a Koepi winder. These two shafts are also the main intake airways to the mine. The two main extraction fans capable of 250 m3/s at 3,2 kPa are situated on surface and are connected to the underground workings by an incline and a network of air passes and return airways. GEOLOGY Although there are numerous occurrences of kimberlite bodies in the district, the Premier Mine kimberlite pipe is the only economic orebody. It is roughly oval in shape with surface dimensions of 900m on the long axis and 450monthe short axis. The surface area is 320 000 m2. The kimberlite has intruded a massive body of felsite and norite which is intersected by a number of faults. The contact between the kimberlite and host rock is clear cut and dips inward at an average angle of 80°. Cutting across the pipe and the country rock is a 75 metre thick gabbro sill, the top of which intersects the pipe below the 347m level and dips at 15' to the north west. The kimberlite has been metamorphosed a distance of some 20m both above and below the sill contacts. Age measurements on biotite from the gabbro date the sill as 1 115 million years, thereby providing a minimum age of the pipe. (Pre Cambrian). All other pipes in South Africa are of Cretaceous age (60 million years). A simplified geological plan and section of the Premier Mine orebody is shown in Figure 1. It is thought that the kimberlite intruded in at least three distinct phases. The kimberlites of these different phases can be distinguished most easily by their characteristic colour. BROWN KIMBERLITE. This is the oldest kimberlite and forms a crescent shape in plan on the south eastern side of the pipe. In depth it increases in relative area. Although the brown kimberlite carries the highest diamond grade in terms of carats per ton all the various kimberlites in the Premier pipe are economic to mine.

Jan 1, 1981

By S. B. Christy

The fact that many millions of gold have been extracted by the cyanide process, during the last five or six years, from South African tailings which could not be profitably worked by any other method previously tried upon them, lends a peculiar practical interest to this branch' of metallurgy. Numerous writers have recently made valuable contributions on the process without exhausting the many complexities of the subject. Among the first American writers on the subject in its modern aspect were two former students of the University of California, Mr. Louis Janin, Jr., and Mr. Charles Butters. Young Janin was one of the first to enter this field, and performed much useful experimental work in the early stages of the history of the process. At about the same time, Mr. Charles Butters, who had left California to erect chlorination works for the Robinson mine at Johannesburg, became interested in the cyanide process, and, after getting the chlorination plant in working order, put up, for working the tailings from the Robinson mine, a cyanide plant which achieved the first large-scale success ever made with this process. The papers which Charles Butters, together with John E. Clennell and Edgar Smart, have contributed on the chemistry of the process, were among the first to give any adequate idea of its nature.* Another important contribution to the literature of the subject is the paper of Dr. A. Scheidell, published in 1894 as a " Bulletin of the California State Mining Bureau." This contains an admirable summary of the state of the art at the time of its publication.

Jan 1, 1897

By J. C. Reed, P. J. Shenon

The Elk City district is in north-central Idaho about 60 miles east of Grangeville and near the headwaters of the South Fork of the Clearwater River (Fig. 1). At the height of its boom in the early sixties, Elk City had a population of more than 2000 people, Published estimates of the placer production from the district range from $10,000,000 to $18,500,000. Although the first quartz-vein location was made as early as 1870, very little gold was extracted from quartz ores until after the completion of the American Eagle mill in 1902. Between 1902 and 1932 a total of from $750,000 to $1,000,000 was taken from about six properties. General Geology Distribution of Rocks Banded gneiss (Fig. 2) is the most aburldant bedrock in the Elk City district. It is interbedded with schist and quartzite and ill some places these rocks have been intruded by sills and dikes. Banded gneiss prevails in the western part of the area whereas schist prevails to the southeast. Augen gneiss occurs for the most part as irregular bodies, which lie transverse to the trend of the banded gneiss and schist. It is largely confined to two areas; one of which extends from east of the American Eagle mine westward as far as the Black Pine mine, a total distance of about five miles, and another area of about two square miles in the vicinity of the California-Idaho placer mine. Quartzite is fairly abundant. In addition to the bands of the latter interbedded with gneiss and schist, some large lenses crop out in the southeastern part of the dik-trict near the southern contact of the area underlain by granodiorite. Granodiorite crops out over a large area and extends northeastward

Jan 1, 1935

By P. J. Shenon, J. C. Reed

The Elk City district is in north-central Idaho about 60 miles east of Grangeville and near the headwaters of the South Fork of the Clearwater River (Fig. 1). At the height of its boom in the early sixties, Elk City had a population of more than 2000 people, Published estimates of the placer production from the district range from $10,000,000 to $18,500,000. Although the first quartz-vein location was made as early as 1870, very little gold was extracted from quartz ores until after the completion of the American Eagle mill in 1902. Between 1902 and 1932 a total of from $750,000 to $1,000,000 was taken from about six properties. General Geology Distribution of Rocks Banded gneiss (Fig. 2) is the most aburldant bedrock in the Elk City district. It is interbedded with schist and quartzite and ill some places these rocks have been intruded by sills and dikes. Banded gneiss prevails in the western part of the area whereas schist prevails to the southeast. Augen gneiss occurs for the most part as irregular bodies, which lie transverse to the trend of the banded gneiss and schist. It is largely confined to two areas; one of which extends from east of the American Eagle mine westward as far as the Black Pine mine, a total distance of about five miles, and another area of about two square miles in the vicinity of the California-Idaho placer mine. Quartzite is fairly abundant. In addition to the bands of the latter interbedded with gneiss and schist, some large lenses crop out in the southeastern part of the dik-trict near the southern contact of the area underlain by granodiorite. Granodiorite crops out over a large area and extends northeastward

Jan 1, 1935

The Elk City district is in north-central Idaho about 60 miles east of Grangeville and near the headwaters of the South Fork of the Clearwater River (Fig. 1). At the height of its boom in the early sixties, Elk City had a population of more than 2000 people, Published estimates of the placer production from the district range from $10,000,000 to $18,500,000. Although the first quartz-vein location was made as early as 1870, very little gold was extracted from quartz ores until after the completion of the American Eagle mill in 1902. Between 1902 and 1932 a total of from $750,000 to $1,000,000 was taken from about six properties. General Geology Distribution of Rocks Banded gneiss (Fig. 2) is the most aburldant bedrock in the Elk City district. It is interbedded with schist and quartzite and ill some places these rocks have been intruded by sills and dikes. Banded gneiss prevails in the western part of the area whereas schist prevails to the southeast. Augen gneiss occurs for the most part as irregular bodies, which lie transverse to the trend of the banded gneiss and schist. It is largely confined to two areas; one of which extends from east of the American Eagle mine westward as far as the Black Pine mine, a total distance of about five miles, and another area of about two square miles in the vicinity of the California-Idaho placer mine. Quartzite is fairly abundant. In addition to the bands of the latter interbedded with gneiss and schist, some large lenses crop out in the southeastern part of the dik-trict near the southern contact of the area underlain by granodiorite. Granodiorite crops out over a large area and extends northeastward

Jan 1, 1935

By P. J. Shenon, J. C. Reed

The Elk City district is in north-central Idaho about 60 miles east of Grangeville and near the headwaters of the South Fork of the Clearwater River (Fig. 1). At the height of its boom in the early sixties, Elk City had a population of more than 2000 people, Published estimates of the placer production from the district range from $10,000,000 to $18,500,000. Although the first quartz-vein location was made as early as 1870, very little gold was extracted from quartz ores until after the completion of the American Eagle mill in 1902. Between 1902 and 1932 a total of from $750,000 to $1,000,000 was taken from about six properties. General Geology Distribution of Rocks Banded gneiss (Fig. 2) is the most aburldant bedrock in the Elk City district. It is interbedded with schist and quartzite and ill some places these rocks have been intruded by sills and dikes. Banded gneiss prevails in the western part of the area whereas schist prevails to the southeast. Augen gneiss occurs for the most part as irregular bodies, which lie transverse to the trend of the banded gneiss and schist. It is largely confined to two areas; one of which extends from east of the American Eagle mine westward as far as the Black Pine mine, a total distance of about five miles, and another area of about two square miles in the vicinity of the California-Idaho placer mine. Quartzite is fairly abundant. In addition to the bands of the latter interbedded with gneiss and schist, some large lenses crop out in the southeastern part of the dik-trict near the southern contact of the area underlain by granodiorite. Granodiorite crops out over a large area and extends northeastward

Jan 1, 1935

J. M. Wadsworth,' Okmulgee, Okla. (written discussion).—It seems to me that thc young man must first be absolutely sure that the work he has taken up is to his liking and then I would advise that he get into the practical end of it and learn as much about it as he can as fast as he can. For instance, in the refinery business the college man, graduated in chemical engineering, should ground himself by going through the various departments in a refinery and working in them until he is thoroughly familiar with the operating problems. In a short time he would be able to apply many things which appear in the abstract form in his college work and this would opcn up an opportunity for him to capitalize on them, promoting his own interest. I do not criticize the present-day college men as a class, but I have found that many of them come into your plant with the idea that mere possession of a diploma should insure rapid promotion; in a short time they are impatient and usually leave. We have found in employing men that the single attribute most difficult to find in them is a lot of energy and whole-hearted interest in the work at hand, and it seems to me that if a man combines these with a careful technical training he should be fairly well equipped to gamble on his future. They must learn that the biggest satisfaction to be gained in life is in their work. Procedure for Attaining Objective of Young Petroleum Engineer H. B. Fell,+ Ardmore, Okla. (written discussion).—Of course, the first thing to be ronsidered in any proposition is the ultimate objective that is desired to be attained. 1 presume that the desired objective of most of the young men graduating in petroleum engineering will be that of chief executive of one of the companies in the petroleum industry, or one of the allied industries, or to become an independent oil producer, or oil refiner, or possibly an oil distributor. The next phase then to be considered is the procedure which should most logically be followed to attain the objeckive of the indivitlual, since in these cases these young men will have been trained in a course in petroleum engineering and it would logically be presumed that the attainment of their objective should be through the operating end of the branch of the business with which they are connected. The operating end is, of course, the department of each branch of the business that develops the resources, or refines the products, or distributes the products, from which the profits are derived, and is a part of any business which is filled with enough interesting work to appease the desires of the most ambitious young men. In order to get the best possible training it would appear to me that the place for a young man to start after graduation would be at the bottom of that branch of the industry which he expects to follow. For example, if he is going in the producing end he can gct no better training than is secured by starting as a roustabout in the ficltl, where he romes in constant rontact with the character of work that is necessary

Jan 1, 1928

By John L. Lamont, Walter Crafts

SECONDARY hardening in tempering has long been recognized as a typical characteristic of steels containing large amounts of carbide-forming alloys. These steels, when quenched and tempered, tend to soften somewhat after tempering at low temperatures, and to reharden at intermediate tempering temperatures before finally softening to low hardness. This behavior in tempering has been studied by many investigators, chiefly with respect to the secondary hardening of high-speed steels. The initial softening has usually been ascribed to decomposition of martensite and growth of iron carbide particles. The secondary hardening has been explained by formation of fresh martensite from residual austenite, formation of very fine alloy carbide particles and by precipitation hardening of metallic compounds in the ferritic matrix. However, factors derived for the calculation of tempered hardness in low-alloy steels1 indicated only a tendency toward retarded softening rather than a discontinuous rehardening. A study was, therefore, made to determine whether fresh martensite from residual austenite and precipitation hardening are essential to rehardening and the degree to which tempered martensite could be rehardened by precipitation of alloy carbides. The mechanism of carbide rehardening was also investigated to determine the nature of the process. The causes of secondary hardening in high-speed steel have received serious consideration for about thirty years. Among the earlier workers Bain and Jeffries2 in 1923 recognized residual austenite as a significant factor, but in addition, emphasized the effect of the formation of alloy carbide particles of "critical" size. They conceived that freshly quenched martensite forms low-alloy iron carbide at low tempering temperatures; with further tempering the iron carbide is then modified by alloying elements that can diffuse most readily at intermediate temperatures (850°F); and that the iron carbides grow in size by coalescence so that the steel tends to lose hardness. At higher tempering temperatures carbide stability rather than alloy availability becomes the controlling influence and in high-speed steel "iron-tungsten carbide is the most stable one and forms to the elimination of the earlier formed carbides. . . . The iron-tungsten carbide particles reach approximately the size for critical dispersion after a short reheat at 1100°F." This concept, the formation of alloy carbide nuclei, redissolving of the iron carbide, and diffusion of carbon to the alloy carbide, was restated for vanadium steel in 1932 by Houdremont, Bennek, and Schrader.3 Subsequent studies have confirmed that iron carbide changes to alloy carbide at 1000 to 1100°F, but the mechanism of the process has received little consideration .

Jan 1, 1948