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By J. C. McCarthy, J. B. Stone

AT Fresnillo a series of veins that has yielded very large quantities of silver and other metals has been developed over a length of 6500 ft. and to a depth of over 3000 ft. In the course of this work striking changes in the mineralogy and metal content of the principal veins have been disclosed and are well illustrated by assay data and ore reserve calculations. This paper describes those changes and suggests their bearing on theories of ore deposition. LOCATION The Fresnillo mines and the adjacent city of Fresnillo are in the central part of the state of Zacatecas, 35 miles northwest of the capitol city of Zacatecas, once a great silver-mining center whose total production is said to place it fourth in the world. Another famous old silver district, Sombrerete, lies northwest of Fresnillo. Fresnillo is connected by a 5-mile branch with Fresnillo station on the Mexican National Railways line from Juarez to Mexico City. Lying near the southwestern edge of the central plateau of Mexico, Fresnillo is surrounded by a semi-arid rolling plain from which rise widely scattered low hills. One of these is Proano hill, an inconspicuous elevation rising barely 300 ft. above the plain, but in and around Proa[n]o hill are the Fresnillo mines, and it is marked now by a great gash cut by gloryhole mining. HISTORY According to an account written in 1834 by Francisco García, then Governor of the state of Zacatecas, Fresnillo was discovered in 1554, and a presidio was established there in 1568 to protect travelers on the road between Zacatecas and Sombrerete. Little is known of the earliest mining at Fresnillo. which seemingly began soon after its discovery, but by piecing together various accounts, it appears that mines there were working in 1717 and were notable active in 1751 under an administrator named Murguía, who left detailed notes describing them. In 1757, however, the mines became inactive because of difficulties with creditors and no doubt also because of increasing water troubles, and they remained abandoned until 1830. In 1830 the mines at Fresnillo became the property of the State of Zacatecas because the interest of Governor García had been aroused by the long disregarded notes of Murguía. Under Garcia's energetic leadership the mines were reopened, and silver production began in 1832 in spite of serious difficulties, including abnormal rains and an epidemic of cholera. In 1833 English capital was sought to allow installation of steam-driven Cornish pumps at the two principal shafts and to build the great hacienda de benefacio (the present Proano hacienda). The company formed in 1835 (Cia. Zacatecano-Mexicana) operated successfully until 1872, when its career was brought to an end by the political and economic disturbances

Jan 1, 1942

By A. A. Jr. Smith, J. S. Smart

Numerous investigations of the effects of the various impurities common to commercial coppers have been published, and the data have found wide use in industry. Naturally, emphasis has been placed on the oxygen-bearing coppers of the tough-pitch variety, but need is now evident for data covering the oxygen-free condition as well. Furthermore, relatively few of the important reactions between oxygen and the various impurities are well understood, and often the data at hand are not applicable to the low concentrations encountered in commercial coppers. This paper presents the changes in conductivity and softening temperature derived from individual additions of the group VIII triad, iron, cobalt and nickel, to high-purity copper, and constitutes a portion of the copper-research program being conducted at the Central Research Laboratory of the American Smelting and Refining Co. The method of preparation of high-purity copper and its alloys in either the oxygen-free or oxygen-bearing form has been described previous1y.l The purity of this material, and the methods developed to produce it, provide the means for direct investigation of the very low concentration range in which the various impurities generally exist in commercial coppers. The usual necessity for extrapolation to this range using the results obtained from higher concentrations, and the interfering effects of extraneous impurities, are thus simultaneously circumvented. Since previously published data have adequately presented the effects of higher concentrations of iron, cobalt and nickel, the present investigation has been confined to the composition range o to 0.05 per cent. All alloys were initially prepared in the form of oxygen-free continuous cast rod, the weighed alloying additions in suitable form being charged with the pure copper to the continuous casting crucible. The best of the available high-purity alloying materials were used, and spectrographic analyses confirmed the absence of interfering elements from the alloys formed. Each series was compounded in the direction of increasing concentration, starting with pure copper in a new crucible and finishing with the 0.05 per cent alloy. Composition was closely controlled by adding the alloying element in the form of a carefully compounded and analyzed master alloy rendered uniform by continuous casting. The accuracy of this method of synihesis was proved by chemical analysis of the more concentrated alloys, and, accordingly, there is every reason to credit the stated composition of the very dilute alloys in the range where chemical analysis becomes highly uncertain. Oxygen-free samples for measurements of conductivity and softening temperature Were obtained by drawing the continuous cast rod to wire of 0.08I-in. diameter. employing a number of intermediate anneals to refine the grain structure. The

Jan 1, 1942

By A. A. Jr. Smith, J. S. Smart

Numerous investigations of the effects of the various impurities common to commercial coppers have been published, and the data have found wide use in industry. Naturally, emphasis has been placed on the oxygen-bearing coppers of the tough-pitch variety, but need is now evident for data covering the oxygen-free condition as well. Furthermore, relatively few of the important reactions between oxygen and the various impurities are well understood, and often the data at hand are not applicable to the low concentrations encountered in commercial coppers. This paper presents the changes in conductivity and softening temperature derived from individual additions of the group VIII triad, iron, cobalt and nickel, to high-purity copper, and constitutes a portion of the copper-research program being conducted at the Central Research Laboratory of the American Smelting and Refining Co. The method of preparation of high-purity copper and its alloys in either the oxygen-free or oxygen-bearing form has been described previous1y.l The purity of this material, and the methods developed to produce it, provide the means for direct investigation of the very low concentration range in which the various impurities generally exist in commercial coppers. The usual necessity for extrapolation to this range using the results obtained from higher concentrations, and the interfering effects of extraneous impurities, are thus simultaneously circumvented. Since previously published data have adequately presented the effects of higher concentrations of iron, cobalt and nickel, the present investigation has been confined to the composition range o to 0.05 per cent. All alloys were initially prepared in the form of oxygen-free continuous cast rod, the weighed alloying additions in suitable form being charged with the pure copper to the continuous casting crucible. The best of the available high-purity alloying materials were used, and spectrographic analyses confirmed the absence of interfering elements from the alloys formed. Each series was compounded in the direction of increasing concentration, starting with pure copper in a new crucible and finishing with the 0.05 per cent alloy. Composition was closely controlled by adding the alloying element in the form of a carefully compounded and analyzed master alloy rendered uniform by continuous casting. The accuracy of this method of synihesis was proved by chemical analysis of the more concentrated alloys, and, accordingly, there is every reason to credit the stated composition of the very dilute alloys in the range where chemical analysis becomes highly uncertain. Oxygen-free samples for measurements of conductivity and softening temperature Were obtained by drawing the continuous cast rod to wire of 0.08I-in. diameter. employing a number of intermediate anneals to refine the grain structure. The

Jan 1, 1942

By John H. Hollomon

THE interpretation of notched-bar impact results has been a matter of controversy since the introduction of more or less standard tests by Fremont,1 Charpy2 and others at the turn of the century. Many investigators3-l1 have contributed to the understanding of the significance of such tests. Several symposiums12,13 have been held with the express purpose of discussing the significance and interpretation of the impact tests. It is not the purpose of this paper to review all these contributions to the development of the knowledge on this subject, but rather to present an interpretation of notched-bar impact tests that appears to be in agreement with the available published data. Fundamentally, the interpretation discussed in this DaDer of the brittle failure of some steels in notched-bar impact test is similar to the qualitative analysis presented by Ludwik3 in the 1920's. Essentially, Ludwik stated that the function of the notch was to increase the tensile stress necessary for yielding, and that brittle failure occurred when the stress necessary for yielding exceeded the normal stress required for fracture (with no plastic de- formation). Ludwik, however, had no means of measuring this normal stress necessary for failure of a plastically* undeformed metal, therefore he could not extend his interpretation beyond the qualitative stage. Kuntze14 and recently MC- Adam15 attempted to determine the effect of deformation on the stress necessary for fracture by testing, in tension, specimens notched to various depths and angles. Because of the uncertainties in Kuntze's (and RlcAdam's) analysis of the stresses (which have been pointed out by SachsI6) and the inherent nonuniformity in the stress distribution in notched bars, the results of such investigations are difficult to interpret. A more direct method of measuring the effect of deformation on the fracture stress was suggested in previous papers17-l8 and has been utilized in the present investigation. The present paper concerns itself primarily with the behavior of notched impact specimens. The effects of strain rate, temperature, and stress distribution, which are discussed with reference to the impact specimen, apply just as well to the behavior of metal at the bases of notches in any engineering structure. All that is necessary is to determine the stress distribution in the structure at the base of the notches.

Jan 1, 1944

By R. J. Sullivan

DURING the past twelve months the Loudon pool of the Eastern Interior Coal Basin has become established as one of the two largest discoveries since the revival of exploitation in the Illinois region and now occupies a position among the important pools of the United States. On Aug. 1, 1939, production averaged 54,000 bbl. per day from 990 wells, having accumulated a total of 10,305,000 bbl. Its very active develop-ment promises to continue for fully another year. It is the site of one of the industry's most singular large-scale gas-injection programs. LOCATION AND RESERVOIR CHARACTERISTICS At present the pool embraces an area 15 miles long by 3 ½ miles wide in eastern Fayette County, along the western rim of the now well-known Illinois Basin's deeper part, through townships 6, 7 and 8 north and ranges 2 and 3 east of the Third Meridian. It is the northernmost, and most prominent, productive structure of a group extending over a 50-mile trend southward to the Dix pool of Jefferson County. Three horizons have been proved to carry production-namely, the Weiler sandstone at 1390 ft., the Paint Creek sand at 1460 ft. and the Benoist or Bethel, at 1500 ft. Geologically, they fall in the Chester series of the upper Mississippian system, where much of Illinois' early-day oil was found, having been laid down in shallow off-shore waters as a part of the vast delta deposit underlying central Illinois and parts of Indiana and Kentucky. As a result of this type of sedimentation, followed by repeated uplifts and warping along an anticlinal axis, differ-ential deposition of sands and shales contributing to great irregularity took place. This is further reflected by wide variations in permeability, porosity and thickness of all the sands and, accordingly, in the producing characteristics of wells. Permeabilities range as high as 550 millidarcies, with the average commonly lying between 60 and 100. For the three pay horizons the mean porosity is 18 per cent, at a calculated water saturation of roughly 30 per cent. The Weiler sand has an average thickness of 25 ft., the Benoist of about 23 ft. and the Paint Creek of approximately 18 ft. Of these three productive formations the Benoist

Jan 1, 1940

By H. S. McKnight

One phase of Canada's rapid development in mining and metallurgy during the past few years has been the creation of Canadian Copper Refiners Ltd., and the establishment of its modern electrolytic copper refinery at Montreal East, Quebec. Canadian Copper Refiners Ltd. is made up of three well-known metal firms whose combined activities cover the entire industry from the mine to the market. They are Noranda Mines Ltd., Toronto, Nichols Copper Co., New York, and British Metal Corporation, Ltd., London. Noranda Mines is the principal owner and provides the greater part of the raw material received by the refinery for treatment. Nichols Copper Co. designed and built the plant and is operating it under contract. British Metal Corporation acts as sales agent and markets the refinery's products. Montreal East was chosen as the site of the new refinery only after a careful survey had been made of the entire eastern part of Canada. The location chosen offers many advantages, chief of which arc the relatively low costs of power, labor, fuel and bupplies, and the fact that it is situated on Montreal harbor, the nearest seaport to the points of origin of the ram materials. All incoming material is received by rail. A majority of the plant's product is exported through the port of Montreal to European markets, being handled to the docks in railroad cars. A large proportion of the product is delivered by means of an industrial railway to the adjacent rolling mill of the Canada Wire & Cable Co., Limited. The refinery was designed to treat primarily the copper produced by Noranda Mines, Ltd. and Hudson Bay Mining &. Smelting Co., Ltd., and its rated capacity is 75,000 short tons per year. A general layout of the plant is shown in Fig. 1. Its capacity can be readily doubled, as shown by the dotted lines. Figs. 2 and 3 are sections taken through the furnace department and the tank house respectively. Except for the main office and laboratory and an auxiliary building containing the stores, warehouse, change rooms, compressor room and machine shop, the entire plant is under one roof. The type of building construction consists of reinforced concrete foundations supported on concrete piles, a steel structure covered with Robertson Protected Metal siding and a 3-in.

Jan 1, 1934

By W. R. Ingalls

In choosing the unusual title given to this paper, in which the term "cyclic " might be substituted for "intermittent," my idea has been simply in respect of precision. We might say old method vs. new method, but that antithesis implies something of substitution that has not occurred and perhaps is not in sight. We may say horizontal retorts vs. vertical, but none of the older retorts are truly horizontal, and there are three different types of vertical retorts. Other complexities might be suggested. There is, however, a broad distinction between distilling that is accomplished in a periodic cycle, generally daily; and distilling that is accomplished continuously, irrespective of whether it be done in the vertical retorts of the New Jersey process or in the retorts of the Trollhät-tan process. In offering a summary of the present state of the art with the intermittent process, I shall not have any idea of exemplifying it through descriptions of any single plants, but shall rather divide among the main steps of the process and outline what is best developed among them. In pursuing this thought, it will manifestly be impossible to make a composition that will exemplify the acme of good practice, for the latter will always be a function of the cost of labor, fuel, etc., which are widely variable. As of 1929, in respect of labor we might generalize the average plant wage in the Pittsburgh-Wheeling district of the United States at $5 per day; in Oklahoma at $3; in Great Britain at $3; in Germany at $1.75-$2; in Belgium and France at $1. With high labor costs the metallurgist will naturally use means of the nature of mechanization that with low labor would be uneconomical when capital charges are taken into account. Nor may we from records of performance safely make positive deductions in respect of the efficiency of different types of furnaces unless they have been tested comparatively under identical conditions in the same works. We shall find gas-fired regenerative furnaces showing coal consumption from 1.2 tons down to 0.7 per ton of roasted ore, which may be ascribable to difference in the thermal value of the coal that is used or to differences in proportions of the parts of the furnace, which is something quite distinct from the principle of the furnace. Thus, we have

Jan 1, 1937

By R. F. Mehl, W. Batz, C. Wells

Diffusion coefficient values for carbon in austenite covering a wide range of temperature and composition have been determined employing statistical methods. In addition, the relation between concentration and each of the following, D (diffusion coefficient), Q (activation heat of diffusion), and A (the constant in the diffusion equation), has been obtained. A study of the rates of diffusion of austenite was published in Metals Technology in 1940.' The present investigation is an extension of the earlier work, undertaken (1) to obtain values of the coefficient of diffusion, D, over a much wider range of carbon concentration, and (2) to provide data in all ranges of concentration and temperature in such abundance that statistical methods could be employed to evaluate the accuracy of the data with unusual precision. Composition of Materials: Composition of materials used in this research program are listed in table I. These materials, all forged to about 1 1/4 in. rounds, consist of a high-purity iron, Armco iron, a large number of laboratory-made plain carbon steels containing amounts of impurities normally expected in high-quality commercial steels, and three silicon steels. Since it was known at the time these steels were being made that impurities in the amounts present were unlikely to affect carbon diffusion rates' significantly, the effort made to obtain high-purity steels was not as persistent as it otherwise would have been. Procedure: In general, the procedure followed by Wells and Mehl and described' in detail in an earlier publication' is the one used by the present authors. However, certain modifications have been made which it is believed allow diffusion data to be interpreted more objectively and correctly. These are described completely under appropriate subheadings which follow. Experimental Procedure: Experimental procedure includes (1) the preparation of specimens by welding steels of different carbon contents, (2) the heating of these for certain times at selected temperatures in order to diffuse carbon along the specimens across the welded "interface," and (3) the determination of the carbon distribution across the diffusion zone. Preparation of Specimens: Prior to welding steel of one carbon content to iron or steel of another carbon content, surfaces to be welded are machined perpendicular to the longitudinal axis of cylinders which are about 3/4 in. in diam and 1 1/2 in. long. The prepared, flat, parallel surfaces are brought together and butt-welded electrically. Evidence of a good weld is given in fig. 1. Welding temperature

Jan 1, 1951

By K. J. Rosengren

INTRODUCTION Mount Isa Mines Limited operates a major mining complex at Mount Isa in northwest Queensland, Australia. The location is arid and remote, some 600 miles distant from the nearest port of Townsville on the eastern coast (Fig. I). Current mine production is 6 million long tons per annum comprising 3.8 million tons of copper ore and 2.2 million tons of silver-lead-zinc ore. Copper ore production is scheduled to increase to 5.8 million long tons per annum by 1973 and a new silver-lead-zinc mine is being developed at Hilton, 13 miles north of Mount Isa. The Mount Isa and Hilton orebodies occur in a single stratigraphic member, termed the Urquhart Shale. This shale is conformable with several other sedimentary formations striking north-south and dipping approximately 65º west. The silver-lead-zinc orebodies occur in bands conformable with the shale bedding and tend to be of tabular shape, varying in width from a few feet to nearly 200 ft. The copper orebodies occur in breccia bodies, locally known as "silica-dolomite", within the Urquhart Shale and tend to be more irregular and massive with widths up to several hundred feet. The detailed geology is described by Bennett (1 2). At present, production comes almost entirely from underground operations with only about 25,000 tons per year of carbonate ore from open pit operations. The mining method used in the copper orebodies is sublevel open stoping with pillar recovery by pillar blast before stope filling or by sublevel caving after stope filling. A large primary sublevel caving operation was also carried out until recently in one of the orebodies. The wider lead orebodies are mined by sublevel open stoping, with or without pillar recovery, while cut and fill stoping is used in the narrower orebodies. Details of the mining methods are given by Davies (3). The company's interest in open pit mining stems from a large open pit which was mined out in 1965 and a further large open pit planned for future production. In addition, the company is currently evaluating several large ore deposits in Australia and New Guinea. All of these orebodies would be mined by open pit methods and rock mechanics studies of slope stability are now a standard feature of all open pit evaluations.

Jan 1, 1972

By D. J. McAdam, G. W. Geil, D. H. Woodard, W. D. Jenkins

INTRODUCTION IN a series of papers, the authors and their associates have shown that the resistance of a metal to fracture is a function of all three principal stresses. 10-18,20,21,23,[1] Consequently since a technical cohesion limit is defined as the technically determinable resistance to fracture, under a specific stress system, the technical cohesive strength comprises an infinite number of technical cohesion limits corresponding to the infinite number of possible combinations of the principal stresses. As shown in the previous papers, the technical cohesive strength increases with plastic deformation, with decrease in temperature, and with increase in the strain rate. The technical cohesion limit, therefore, is affected by the same factors that affect the flow stress,§ namely, the stress system, plastic deformation, temperature, and the strain rate. The effect of each of these factors on the fracture stress is qualitatively the same as its effect on the flow stress. The effect of any of the factors may be represented by a curve of cohesion limits, although only one point on the curve may represent actual fracture. Since the term "fracture stress" has come into general use with the same significance as technical cohesion limit; the authors are here using the' shorter term with the understanding that it, does not always refer to actual fracture. In the general investigation described. in the previously-mentioned papers,10-18,20,21,23 the influence of the stress 'system on the fracture stress was studied by means, of tension tests of notched specimens. When a circumferentially notched cylindrical' specimen is subjected to longitudinal ten-, sion, the minimum cross section is under " transverse radial tension. The mean radial stress increases with the depth and sharpness of the notch, but is always less than the longitudinal stress. The greatest principal stress (S,), therefore, is longitudinal, and since the radial stress is effectively equal to the circumferential stress, the transverse principal stresses are equiaxial in that plane (S2 = S3). By the use of notched specimens in tension tests it was shown that the fracture stress increases with increase in the radial stress ratio (S3/SI). The evidence is based not only on the results of tests to fracture at various constant temperatures, but also on. results of two-stage tests, in which each specimen was plastically deformed a predetermined

Jan 1, 1948

By Frank J. Tolonen, Nicholas H. Manderfield, Paul Jasberg

EARLY in the study of the low grade iron formations of Michigan, wide variations in their structure and texture became evident. Because of these variations no simple method of concentration is possible, and those portions of the formations that can be exploited profitably under a given stage of metallurgical progress and existing economic conditions must be searched for carefully. Both structure and texture of the formations have an important bearing on their amenability to bene-ficiation. By structure is meant the banding planes of easy fracture, the porosity, and the degree of leaching. Texture, which includes grain sizes, shapes, and degree of interlocking, determines the amount of grinding necessary to liberate the mineral grains. Structure determines the liberation of portions richer in one mineral. Structure and texture may be termed the gross features of the formation. Earliest surveys of the formations were conducted to outline the parts that could be concentrated at % in. (3 mesh). Operators agreed on this size because anything finer required sintering or other agglomeration processes. At 3 mesh only the structural features, as defined above, would be liberated. With suitable ores concentration was easy, as any of the gravity methods could be used. Sink-and-float gave good results, but with this method only limited portions of the iron formations yielded a desilrable product.1,2 n general, the iroa content could be raised to shipping grade, but the silica content remained too high—sometimes 20 pct—instead of 9 or 10 pct. Grinding did not greatly increase the liberation ofsilica until 2010 mesh or finer was reached. To find the reason fpr this T. M. Broderick began micros~opic examin'ations of the formations in 1933.3 Procedure: The purpose of this discussion is to correlate the appearance of typical portions of the iron formations with their amenability to concentration. As liberation is essential before separation becomes possible, much of the work is based on measuring the liberation achieved by crushing and grinding. Both heavy media and magnetic tube tests were used in this work. On the basis of general structural features such as banding, degree of alteration, and leaching, a specimen was selected from each part of the iron formations studied. The specimen represents the gross features of the formation and not necessarily the particular formation quantitatively. One half of the specimen was polished for macro-study, and the remainder was used for analyses to obtain the grades represented by variations appearing in the macrosection. Specimens for microscopic examination were also prepared for each variation in gross appearance. For correlation of microstructure with liberation, the previous work by sink-and-float methods was supplemented by magnetizing roasting followed by magnetic separation. The results show the liberation of silica to be very important at the sizes used, because any silica particle with iron minerals occluded would be held in the concentrate. A magnetizing roast changes the crushing and grinding character of the sample. For partly leached soft samples the change is slight, but with harder types and those in which there is a considerable change in crystal structure of the minerals, the increase in liberation of the iron minerals may be 30 pct more than that of unroasted samples at the same crusheH size. Gogebic Range Iron Formations: The Gogebic Iron Range is located in the extreme western and southern parts of Michigan's Upper Peninsula, extending westward into Wisconsin. Samples for this study were obtained from the Ironwood formation, comprising three chief members—the Plymouth, Norrie, and Ahvil—with several intermediate slates and thick overlying slates. (The slates are not considered in this discussion because grinding finer than present practice is usually required to liberate the iron minerals.) The predominant iron mineral is hematite, about half of which is hydrated to goe-thite. Limonite occurs in some of the partly leached areas, and siderite in parts of the formation.

Jan 1, 1958

By W. G. Woolf, E. R. Crutcher

The electrolytic zinc plant of the Sullivan Mining Co. is in Government Gulch, at Silver King, Shoshone County, Idaho, in the Coeur d'Alene mining district, about one mile from the Bunker Hill smelter and about three miles from Kellogg. The Sullivan Mining Co., which owns both the zinc plant and the Star mine, is itself owned jointly by the Bunker Hill & Sullivan Mining & Concentrating Co, and the Hecla Mining Company. The Tainton high-density, high-acid process is used at this plant, therefore some parts of its flowsheet differ markedly from those of electrolytic plants using the low-density process. While it is a custom plant and has treated a variety of zinc concentrates originating from various sources in a number of states, it has obtained its main supply of zinc concentrates chiefly from the zinc-bearing ores of the Coeur d'A1ene mining district, of which the principal important sources have been the Bunker Hill and Star mines. Thus the main feed has been zinc concentrates obtained from complex lead-zinc-silver ores. For the year of 1935 the concentrates had the following composition: water, 9.5 per cent; Zn, 50.6; Pb, 3.3; Fe, 9.4; S, 31.0; Cu, 0.2; insoluble, 2.9. Flowsheet The concentrates are delivered to the plant in railroad cars, which discharge into storage bins directly under the track. Bottom-dump cars are dropped directly into the bins while a slip scraper, pulled by an air tugger hoist, is used to empty box cars. There are five storage bins for concentrates, each of approximately 300 tons capacity. At the bottom of each bin there are individual belt feeders that discharge onto a common conveyor belt, which in turn discharges to another belt that carries the concentrate onto a weightometer belt for weighing. Immediately after the weighing, the crushed oversize return from the roasters is added to the load on the belt and then the ground-up sweepings, lumps and roaster cleanings. An electromagnet at the end of this conveyor removes tramp iron from the feed. The belt discharges to a set of 30 by 12-in. rolls, which break up lumps or frozen chunks of concentrate, etc. Beneath the rolls there is a screw conveyor, which conveys the roll discharge to a

Jan 1, 1937

By C. R. Fulton

The idea of making permanent magnets by sintering is not a new one, but until the introduction of the iron-nickel-aluminum permanent-magnet alloys there was little need for such a process. Permanent-magnet alloys in use up to the time of this development could be forged and machined to the exact size or shape required without too much difficulty. These new Fe-Ni-A1 alloys, which have become known as Alnico in this country, soon became very popular because of their superior magnetic properties, but when they were made by casting they did have some disadvantages when compared with the older type of permanent-magnet steels. Among these disadvantages were the following: 1. It was almost essential that they be cast to size, since they could not be readily forged or machined. The material could be ground but this was quite expensive, particularly for small magnets in large quantities. 2. The material had a large grain size and was mechanically weak. 3. Close tolerances could not be held and on small magnets this meant grinding almost the entire magnet. Because of the brittle nature of the material, there was considerable loss due to chipping and cracking during these grinding operations. 4. Because of its high aluminum content, the molten alloy had a high surface tension. It was difficult to prevent loss of aluminum by oxidation during the melting operation. In view of these difficulties it is not hard to understand why a process for sintering alnico was developed, or why sintered alnico has come into such widespread use in recent years. It might be well to qualify this by saying that it has come into widespread use for small magnets weighing 2 oz. and less, of the type shown on Fig. I. In this size range, sintered alnico has been able to compete in price with cast alnico. Magnets larger in size, of the type shown on Fig. 2, can be made but they do not compare favorably in price with the cast material. An idea of the range of sizes that have been produced is given by Fig. 3, which shows a small magnet weighing 0.006 oz. and a large magnet with a steel insert weighing 4 lb., both made from sintered alnico. Magnets weighing more than 2 oz., however, should not be made from sintered

Jan 1, 1945

Before hostilities between the colonies and Great Britain began in 1775, most of the coal used in the northern colonies undoubtedly came from England, with some supplies for New England coming from Nova Scotia; in New York and Philadelphia the supply came from England with some from the Virginia mines. Table 4 shows data obtained from the Public Records Office, London, for the years 1768-1772. This table supplements Table 3, which records only a few shipments from Richmond in 1768 and 1769; as there was no coal shown in imports coastwise in 1769 in Table 4, and only two small shipments in 1769 in Table 3, there may have been a cessation of shipments then, for some reason not now known. In Table 5 are incorporated all the data in the first two, with a few earlier years, and an attempt is made to estimate the amount used during the period by allowing for coal exported. With the small total tonnages a carryover from one year to another makes a large difference in the annual figures, but over the period the result may be fairly indicative of the amount of coal used before the Revolution. According to present day ideas the quantity of coal used during that war was very small, but apparently it was badly needed for the manufacture of the small amount of munitions made in those troublous days. As Philadelphia was the leading city along the Atlantic coast, and as the southeastern part of Pennsylvania was the leader in making iron, as well as making rifles, it is natural that the first call for coal came from there. At first, procuring supplies such as coal, powder and sulphur was handled by a Committee of Safety, which later became a Council of Safety and later this committee's functions were turned over to a Board of War, functioning with the military organization of the colony. . The organizations of the several colonies were slightly different, but Pennsylvania, Maryland and Virginia followed this scheme generally. Later in the war the Congress began to make arms for the troops, and it is difficult now, with the few records available, to tell when the state functions ceased and those of Congress began. On December 5th, 1775, the Committee of Safety

Jan 1, 1942

By C. W. Hoffman, W. R. Hinken

Probability sizing, a fairly recent development in the field of particle separation, is now under investigation to determine its value to the mining industry. The method employed is probability sizing through a vibrating network of decreasing screen openings. While probability sizing is not new, the possibility of using it for coarse, highly abrasive ores at high tonnage rates has only recently appeared to merit practical consideration. The advantages of probability sizing vs. conventional screening methods are: (1) reduced space requirements for a high capacity unit, (2) fewer moving parts, (3) more versatile product sizing selection, (4) reduced maintenance requirements, and (5) lower power consumption. Screening is a necessary part of virtually every crushing circuit. At some stage in the process, particle separation becomes a controlling factor in the operation. The basic type of screen in use today is the vibrating unit, which usually employs a mechanical vibrator of one type or another to obtain sizing and material transportation across the screen surface. The vibrating screen has its limitations and inherent problems. Operators must live with: (1) high maintenance — worn rods, screen cloth, spacers, cross members, chutes, etc., (2) low capacity in relation to floor space requirements, (3) extreme noise levels, (4) high power consumption, and (5) structural problems in the screens and supporting steelwork. All plant operators have become tolerant of these drawbacks through the years, and little innovation has been developed in screening. The vibrating screens in use today incorporate improved metals, more efficient power sources, better lubricants, and increased manufacturing skill, but the basic machine itself remains unchanged. Screening is a relatively small part of most plants' operating expenditures and gets relatively little attention from research and development. The HiProb sizer is a new approach to the problem of particle separation on a high tonnage basis. Developed by Dr. Fredrik Mogensen (president of Fred-rik Mogensen, A. B. of Djursholn, Sweden), this unit has found its primary acceptance in the sizing of wet or dry materials in the fine ranges. Climax has been working cooperatively with Mine and Smelter Supply Company, the United States licensee, on the adaptation of the sizer to obtain consistent recovery at a high tonnage rate on a — 3/8-in. separation. This was a new application for the sizer and opened up a whole new field of problems. Dr. Mogensen's machine utilizes the control of normal probability of falling particles and is designed to eliminate some of the undesirable features of vibrating screens mentioned earlier. The machine is light, compact, consumes less power than conventional screens and has a low noise level. PRINCIPLES Probability sizing is a process designed to allow solid particles to segregate with a high ratio of the materials reporting into their respective size fractions with a minimum amount of misplaced particles. Sizing is defined in this paper as segregation of dry solid particles into fractions of comparable sizes. Probability is usually estimated as the ratio of the number of ways an event might occur vs. the total number of allowable ways. The probability process described depends upon the fact that particles falling through a network will segregate across the sizing surface even though all openings are larger than the particles. The degree of segregation is dependent upon the particle size and the trajectory applied by the sizing network to each falling particle. Specifically, the process utilizes a maze made up of screen cloth to control the path of a particle in free fall as illustrated in Fig. 1. The usefulness of this probability sizing process is dependent upon the degree of efficiency of an effective maze. Primarily, a maze must be extremely flexible to meet the multi-faceted nature of solid particles and, at the same time, the degree of control must be reproducible so that a desired separation is predictable. Dr. Fredrik Mogensen has formulated a maze that is not only flexible and easily controlled, but, mote importantly, easily adaptable to a working machine.

Jan 1, 1970

By D. J. McAdam, G. W. Geil, D. H. Woodard, W. D. Jenkins

.In a series of papers, the authors and their associates have shown that the resistance of a metal to fracture is a function of all three principal stresses. Consequently since a technical cohesion limit is defined as the technically deter-minable resistance to fracture under a specific stress system, the technical cohesive strength comprises an infinite number of technical cohesion limits corresponding to the infinite number of possible combinations of the principal stresses. As shown in the previous papers, the technical cohesive strength increases with plastic deformation, with decrease in temperature, and with increase in the strain rate. The technical cohesion limit, therefore, is affected by the same factors that affect the flow stress,$ namely, the stress system, plastic deformation, temperature, and the strain rate. The effect of each of these factors on the fracture stress is qualitatively the same as its effect on the flow stress. The effect of any of the factors may be represented by a curve of cohesion limits, although only one point on the curve may represent actual fracture. Since the term "fracture stress" has come into general use with the same significance as technical cohesion limit, the authors are here using the shorter term with the understanding that it does not always refer to actual fracture. In the general investigation described in the previously-mentioned papers, 10-18,20,21,23 the influence of the stress system on the fracture stress was studied by means of tension tests of notched specimens. When a circumferentially notched cylindrical specimen is subjected to longitudinal tension, the minimum cross section is under transverse radial tension. The mean radial stress increases with the depth and sharpness of the notch, but is always less than the longitudinal stress. The greatest principal stress (S1), therefore, is longitudinal, and since the radial stress is effectively equal to the circumferential stress, the transverse principal stresses are equiaxial in that plane (Sz = S3). By the use of notched specimens in tension tests it was shown that the fracture stress increases with increase in the radial stress ratio (s3/s1) The evidence is based not only on the results of tests to fracture at various constant temperatures, but also On results of two-stage tests, in which each specimen was plastically deformed a predetermined

Jan 1, 1949

By J. C. Motz

Introduction The variations in crushing plant design seen during the past 30 years provide a valuable source of information and experience from which one can select the optimum arrangement for a specific requirement. The purpose of this paper is to examine some specific equipment configurations as they relate to operating efficiency, plant availability and final product size in the form of mill feed. Due to the varied ore properties, physical space limitations, size of ore body and financial resources, there is no single crushing or grinding flowsheet that is best. Since energy costs have risen dramatically during the recent past, it became obvious that significant savings are possible by selecting equipment and a circuit which offers low power requirements balanced with low wear metal and installed equipment costs. Relatively clear and measurable are the metal and maintenance costs and the ever increasing cost of energy. Metal costs in the form of crusher liners have not quite doubled in the last 20 years, while power costs have increased by a factor of four or five, depending on the region considered. Estimates made only a few years ago on the increased cost of energy are obsolete in view of the recent unstable conditions. Since power is therefore becoming the largest cost area, it behooves us to investigate potential cost saving methods in the form of lowering the KWH/Ton of crude. Crushing To reduce run-of-mine to mill feed requires reduction in a series of stages. The physical size and power requirements of a crusher capable of reducing hard rock vary depending on the application. Extended experience in most crushing plants show that a minus 19MM (3/4 in.) rod mill feed can be and is being produced in three stages of crushing. In the majority of cases, the last crushing stage is closed circuited using anywhere from 19MM (3/4 in.) to 14MM (9/16 in.) slotted openings on the sizing deck. In some copper concentrators, the mill feed is all passing 13MM (1/2 in.) using three stages of crushing and 14MM (9/16 in.) square openings on the closed circuit screens. It was felt by some that crushing to this fine size is only possible if the ore is friable and is relatively soft in comparison to the harder ores. Our results show that some of the copper ores are as hard as taconite and are crush-

Jan 1, 1982

By A. W. Schillinger

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

Jan 1, 1964

By T. N. McVay, E. E. Creitz

Rather large amounts of titanium minerals and some zircon and monazite are being recovered from dune sands about I. miles west of Jacksonville Beach, Fla. The Mining Branch of the Bureau of Mines, Southern Experiment station, has been interested in developing additional large deposits of similar sands in other areas of Florida. The results of a field study have been published1 concerning the sands occurring on Trail Ridge which runs north and south about 30 miles southwest of Jacksonville. This investigation was undertaken in order to identify the opaque titanium minerals occurring in Trail Ridge and because of the economic importance of Florida titanium minerals. Nature of the Problem The minerals present in the heavy fractions (sp gr > 2.94) of the sands are principally opaque minerals; staurolite, zircon, sillimanite, tourmaline, kyanite, rutile, a small amount of corundum, traces of the zinc spinel gahnite and garnet. Epidote, monazite, hornblende and ti-tanite are present in other dune sands but these minerals were not noted in the samples collected on Trail Ridge. All of the minerals except those which are opaque can be readily identified by the usual petrographic methods. However, Opaque minerals cannot be examined satisfactorily by transmitted light. Polished sections of the opaque-mineral grains were examined but these sections did not furnish useful information. Consequently, it was necessary to use means other than the usual petrographic methods to identify them. Previous Investigations of Heavy Minerals in Florida Sands Martens2 has described the heavy minerals in the Florida beach sands and has classified the opaque minerals as ilmenite with minor amounts of leucoxene. Miller3 studied a number of sands collected by Ross and Mertie in Florida. He states that the material from the Florida beach and dune sands that commonly has been called ilmenite has essentially the chemical composition of the material that has been called arizonite.4 The x ray diffraction patterns of several samples indicated that the supposed ilmenite is essentially amorphous, with only weak lines corresponding to ilmenite, and hence the material is unlike the type of arizonite from Arizona. Opaque Titanium Minerals Rutile is essentially titanium dioxide. However, as pointed out by Dana,5 important amounts of Fe+2 and Fe+3 as well as Cb and Ta are reported in some analyses.

Jan 1, 1949

By Leslie G. McMillion

A field project to develop and demonstrate a method for alleviating problems of highly mineralized ground water where they occur as isolated zones or pockets in fresh-water aquifers is being conducted in the Ogallala aquifer of eastern New Mexico by the Robert S. Kerr Water Research Center. Removal of contaminated water is by a large-capacity well and the pumped poor-quality water is used locally for secondary recovery of oil. Benefits of the operation are threefold in that the aquifer is being reclaimed for future fresh-water production, contaminated water is being used beneficially, and fresh water which otherwise would have been used for waterflooding is available for other purposes. Ground-water quality in the site has improved markedly after two years of operation. Use of water in the arid and semiarid Southwest has grown so rapidly that demands in many localities far exceed available supplies. In the High Plains region of New Mexico, Oklahoma, and Texas, the only large sources of water are underground reservoirs, being depleted at an alarming rate. In some parts of this region, the aquifers contain pockets of highly mineralized waters; and the mineralized waters in some of them have started spreading in response to the drawdown effects of pumping wells. The aquifers of this region are primarily composed of sand and gravel beds; any cleanup by natural processes will require many tens or even hundreds of years. Most of these pockets of highly mineralized water were caused by oil-field brines seeping downward from earthen disposal pits. Thousands of these pits were in use in the region until only a few years ago when they were prohibited by the oil regulatory agencies of the respective states. The resulting areas of poor quality water are serious enough to warrant much corrective attention. Correction will involve one or more of three approaches: 1) Restrict the pumping of fresh water in the zones where the drawdown of the pumping well will significantly influence the movement of the contaminant. 2) Remove the highly mineralized water entirely from the affected area. 3) Develop and pump the poor-quality water in such a manner that its movement is held in check, even though only a part of the contaminant is actually removed. The principal hurdle to overcome in applying methods 2 or 3 is financing of the operations, since there are very

Jan 1, 1972