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By D. C. Seidel, E. F. Fitzhugh

The cementation of nickel from acidic solutions by metallic iron is discussed. The cementation is carried out in pressure vessels at temperatures above 100°C. The results from bench scale studies on variables such as retention time, temperature, and solution pH are presented. A continuous processing flowsheet is proposed. Cementation is one of the oldest known hydrometal-lurgical reactions. The cementation of copper on iron was recorded by 'Paracelsus the Great' in about 1500,1 and by 1600 the technique was being used to recover copper at the Rio Tinto operations in Spain.2 At least one early author felt that cementation may have been one of the primary reasons for the alchemists' belief in the transmutation of metals.' The early writings state that when iron was placed in the clear waters from some mountain springs, the iron disappears and copper is found in its place. To the alchemists this may well have been one of the most convincing proofs that transmutation could and did occur. Since these early times the recovery of copper from acidic solutions by cementation has been practiced in plants throughout the world. Probably the cementation technique in some form has been common to more copper mining and milling operations than any other single recovery process. During current hydrometallurgical extraction studies, which were sponsored by the Republic Steel Corp. at the Colorado School of Mines Research Foundation, Inc., it was found that under the proper conditions, metallic nickel could be cemented from acidic solutions by powdered iron. The reactions are apparently similar to those that occur during the cementation of copper, but the nickel cementation had not been anticipated because iron and nickel are nearly adjacent in the electromotive series. The potential difference between iron and copper is approximately 0.78 v, while the potential difference between iron and nickel is less than 0.21 v. The cementation of nickel with iron at room temperature is almost negligible, but when the reaction is carried out in a closed vessel at temperatures in excess of 100°C, the nickel can be cemented almost quantitatively. The part played by this discovery in a practical method of nickel recovery is set forth in a separate paper.3 The following paragraphs are a discussion of experimental studies that were made to investigate this cementation reaction. The technique has been designated the HTC or High Temperature Cementation procedure. The cementation work was part of a study on hydrometallurgical techniques for the extraction of nickel from the garnierite or silicate type nickel ores. A hydrothermal extraction procedure had been developed, and this technique produced an acidic pulp or solution that contained both nickel and appreciable amounts of magnesium.3 Small quantities of ferric and ferrous iron were also present along with cobalt, manganese, and chromium. The potential for recovering the nickel from these acidic solutions by ion exchange or solvent extraction did not appear to be promising because of the relatively high magnesium content. The Ni ++ and Mg++ have nearly identical ionic dimensions and tend to be co-absorbed or extracted during ion exchange or solvent extraction treatments. Preliminary tests indicated that the nickel could be precipitated as a sulfide when using the high pressure H2S precipitation technique developed for the Moa Bay operations of the Freeport Nickel CO.4 This technique gave good recoveries, but the nickel sulfide product requires considerable additional processing before a marketable form of nickel is realized. The process also requires clarified feed solutions, and the solids-liquid separations on the garnierite residues are difficult. A program was initiated to investigate alternate procedures that might shorten the route to a marketable nickel product, and hopefully also permit bypassing the difficult solids-liquid separation steps required for the H2S precipitation technique. It was during these studies that the nickel cementation reaction with iron was encountered. EXPERIMENTAL EQUIPMENT AND PROCEDURES The bench scale precipitation tests which were conducted during this experimental program were made in 2-liter stirred autoclaves.* A photograph showing the form and arrangement of the autoclave

Jan 1, 1968

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 bearing loads and elevated operating temperatures, the life of bearings made from tin-base 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 bearing 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 bearings 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 C.F. Pascoe, H.J. Roast

The object of this investigation was to compare the arsenical antimony-lead alloy with some of the regular bearing-metal alloys. With this end in view, the following tests were made: 1. Chemical analyses to show the composition. 2. Thermal analyses. 3. Macroscopic examination of fractures. 4.Microscopic examination. 5. Tensile tests. 6. Crushing tests. 7. Scleroscope tests.' 8. Brinell tests, at temperatures from 80' to 400' F. Tests 3 to 8 were conducted on both sand-cast and chill-cast specimens. For the purposes of this paper the alloys have been numbered from 1 to 6; their composition is as follows: Lead. ' Antimony, Tin, CoppeR ARSEnic, NumBeR PeR CENT. PER CENT. PeR Cent. PeR CeNt. PeR Cent. 1 82.64 12.95 4.40 0.01 none 2 27.54 9.04 59.87 2.65 none I 3 none- 0.13 ------------1.47 4 82.95 16.85 none 0.20 none 5 83.78 15.35 none 0.05 0.82 6 0.15 7.87 . 84.13 7.85 none The analysis, in the case of the antimony, was made by solution in sul-furic acid and titrated with permanganate. For tin, the ferric-chloride method was used; for copper, the alkaline separations, and the thio-cyanate methods; for arsenic, volatilization of arsenic and titration with iodine. The lead, with the exception of alloy 6, was taken by difference. With the exception of No. 5, the alloys were taken as received from the manufacturer. About 40 lb, were melted in a graphite crucible and recast into small ingots to insure uniformity of composition; samples

Jan 1, 1923

By F. R. Zacher

Unmined blocks in the Sewickley seam, surrounded by worked out areas, have been found to transmit overburden oads through the interval strata to the Pittsburgh seam workings 90 ft below. Operating experiences under these conditions are described in this paper. IT has always been understood, in northern West Virginia where both the Pittsburgh and Sewickley seams are mined, that pillaring or splitting in the lower Pittsburgh seam could break the interval strata and make mining in the overlying Sewickley difficult or impossible. In some instances, however, mining methods practiced in the overlying Sewickley have reversed the problem, resulting in a shifting of the intermediate layer and destroying valuable acreages of the Pittsburgh seam below. The Pittsburgh seam in this area is, on the average, 102 in. thick. About 18 in. of low-quality head coal are left for roof protection as it is mined. Immediately over the coal are two draw slates, two rider seams, and a good hard black shale, locally called black rock, which serves as a permanent roof where draw slates and riders are taken down on certain main haulroads. In most areas mined in this seam in the past the bottom has been satisfactory but soft. However, in areas now being mined by the Christopher Coal Co. in Monongalia County the bottom is very hard and unyielding. The Sewickley seam, lying 85 ft above the Pittsburgh, will average nearly 70 in. in thickness and is overlain by massive shales and sandstones. Much of the seam is mined only down to the "sheep skin," a hard band about 4 in. above the bottom, which is composed of relatively hard shale. Fig. 1 shows a typical cross-section through the Pittsburgh and Sewickley seams in the area being discussed. The thin Redstone seam is not deep-mined but is stripped in some locations where it outcrops. Most of the Sewickley seam in the Morgantown areas has been mined out from 10 to 30 years. There are at present only two large mines and five or six others of small production operating in that seam. Mining systems in the past were generally of the room and pillar type with some full pillar extraction. This seam in the Scott's Run area was operated by as many as thirty or forty different companies at one time, during an era beset with strikes, market failures, and depressions. Most of the barrier pillars along the many main haulroads were abandoned, and blocks of Sewickley coal that would be fully recovered with today's machinery were left unmined when conditions became anything less than favorable. Consequently the mined-out Sewickley is spotted throughout with solid blocks varying in size from 100x100 ft to 200x400 ft and larger. It is these unmined Sewickley blocks surrounded by worked-out portions that have caused transmissions of roof stresses to underlying Pittsburgh mining areas, see Fig. 2, resulting in untold difficulty. As the Christopher Coal Co. operates both the Pittsburgh and Sewickley seams, accurate maps of workings in both seams are readily available and are regularly consulted. The first Pittsburgh squeeze attributed to overlying Sewickley blocks occurred in the Arkwright mine, where a shuttle car crew was developing a section consisting of seven panel headings. These entries were being driven in solid coal 3400 ft from the nearest worked-out Pittsburgh area. As mining advanced it was noticed that certain ribs began to slough and certain areas to show signs

Jan 1, 1953

By Leonard D. Jaffe

In recent years, the advantages of tempered martensite as a microstructure for steel parts have been well established. For parts that must not fracture brittlely when loaded at high rates, at low temperatures, or under conditions of combined stress, tempered martensite is especially desirable. One of the factors that determine whether a steel part will have a martensitic structure after heat treatment is the hardenability of the steel. Despite the large amount of recent research, there are still many unsettled basic questions concerning hardenability and its relation to chemical composition, and, indeed, considerable disagreement among investigators in the field. Much of the uncertainty that beclouds the existing knowledge of hardenability arises from the fact that hardenability is still a wholly empirical concept, and has not been quantitatively related to the more fundamental knowledge of isothermal decomposition of austenite. Hardenability is, essentiaily, a measure of the time required for small percentages of austenite to decompose, during continuous cooling, to pearlite, bainite, proeutectoid ferrite, or proeutec-toid carbide. In the practical heat-treat- ment of steel, also, the decomposition of austenite is usually not isothermal but rather during cooling, or, in general, is not at a single constant temperature. Thus, both hardenability and practical heat treatment involve anisothermal decomposition of austenite. In a previous publication1 the author and his collaborators outlined the principles that govern the relations between isothermal and anisothermal decomposition of austenite, and reviewed the experimental information available in 1945. To avoid repetition, familiarity with that article (Paper I) will be assumed, and attention given primarily to work carried out or published since 1945. The method generally employed was explained in Paper I: anisothermal transformation is studied through investigation of the effect of holding austenite at one temperature upon its subsequent decomposition at another temperature. Both temperatures are below the stability range of the austenite. This paper is confined to studies in which only a small amount of transformation occurred at the first temperature. This is the only case of importance in practical heat treatment when an essentially martensitic structure is sought. Materials and Procedure It was necessary to use different steels for the studies involving proeutectoid ferrite and for those involving proeutectoid carbide, beeause the two constituents have not been found to form in the same steel. Several steels were tried and rejected

Jan 1, 1949

By Donald C. Barton

OF the four papers that comprise this Technical Publication, three are case histories of individual geophysical prospects, subsequently tested by drilling. The bibliography of Dr. Barton's published scientific paperst includes investigations in physiography and surface geological mapping, as well as his geophysical papers, and from this list it is to be observed that he early recognized the place of different types of geophysical prospecting methods in the search for geological structures of economic importance. Moreover, his viewpoint was not merely that the structures should be found, but that the geophysical data should be examined quantitatively to determine the geometry of the structures. In other words, that although these data are insufficient to deduce a unique form for the structure that caused the effects measured in the data, the engineering computations should be carried out to give as much of a definite configuration to the structure as the data permitted, in order that the testing of the structures could be conducted in the most economical manner and with the idea that the experience so gained could be later incorporated in the study of structures subsequently found. Dr. Barton continually emphasized the quantitative use of geophysical observations, whereas at the time that he began his specialization in geophysics it was considered sufficient to use observations only qualitatively. He, more than any one else, was responsible for the first volume on geophysical prospecting issued by the Institute in 1929, and his three articles in that volume stressed the quantitative use of geophysical prospecting. The fourth article in this present Technical Publication is a further contribution to this branch of geophysical interpretation, and is a supplement to the paper in the 1929 volume, "Calculations in the Interpretation of Observations with the Eötvös Torsion Balance." From the successful demonstration by Dr. Barton of the usefulness of quan-

Jan 1, 1946

By R. E. Anderson, R. S. Busk

The mechanical properties of many nonferrous alloys can be modified by heat-treatment. This is almost always effected by controlling the amount of alloy in solid solution and the amount and distribution of second-phase material. The latter often is accomplished through precipitation or age-hardening treatments. The procedure is to solution-heat-treat the metal to provide a supersaturated solid solution and then cause precipitation of the dissolved material at some temperature below the solvus for the alloy. The generally used magnesium sand-casting alloys arc based on the ternary alloy system Mg-Al-Zn. Solid diffusion reactions in the system are slow enough so that a substantially completely supersaturated solulion is maintained even though the metal is relatively slowly cooled from heat-treating temperatures. Consequently, it is general practice to air-cool magnesium-alloy castings from the heat-treatment temperature prior to aging at an elevated temperature. Since it is necessary with many other alloys to cool quite rapidly before aging if optimum benefit from the aging is to be obtained, this study was undertaken to determine the effects of quenching on magnesium alloys. The] alloys used are shown in Table I, being selected as typical of the presently used magnesium alloys. Cooling Rates Since quenching is a means of rapidly extracting heat, there are two factors of interest in its study: (I) the rates of cooling that can be obtained and (2) the effect of various rates on the as-cooled as well as subsequently aged metal. The cooling rates were measured by drilling a hole 1/16- in diameter longitudi~lally in the center of a cylinder % in. in diameter and 3 in. long. The drilling was stopped to place the end of the hole at the center of the long axis of the cylinder. An iron-constantan thermocouple was inserted in this hole and wedged in place. The temperature drop during rapid quenching was recorded by a special1y built General Electric potentiometer recorder capab1e of measuring ' 500°F. temperature drop in 1/10 sec. For the slower cooling rates in air or oil, the temperature drop was measured by a Leeds and Northrup potentiometer recorder with the chart speed increased to One inch per minute' The cylinder was heated to the desired temperature, quenched as desired, and the temperature-time curve recorded. The cooling rates in still air, in oil at 350°F., in water at 180°F. and in a cold, high-pressure water spray were determined in this manner. The results of this study are presented in Fig. I. There is a great difference between water quenching and either air or oil quenching. There is little difference between the hot water quench and the cold water spray. Cooling rates were also

Jan 1, 1945

By Carl J. Christensen

This paper presents the development of diamond coring of the Weber sand section in the Rangely Field, Colorado. The description and operation of the diamond-coring equipment is included as well as the economics and results obtained by its use. Diamond coring is compared with the other methods used for drilling the Weber sand. Two types of diamond coring are discussed. The first and the one used most extensively is the regular 6 ?x 3¾-in. diamond bit and bottom hole 50-ft barrel. The second type, which has also been used successfully by Stanolind, consists of a 4 13/16 x 2 11/32-in. cutting bit and reverse circulation coring equipment. The first type will be referred to as regular or conventional diamond coring while the latter type will be referred to as reverse circulation diamond coring. Introduction Diamond coring has been used for many years in mines, quarries, surface coring at dam sites, laboratory work, and so on, but, until recently, had never been used very extensively in coring oil sands and other oil-bearing strata. It was demonstrated early in its development that hard forma-tions could be cut rapidly and that core recoveries far exceeded those obtained with conventional hard rock bits and quite often 100 pet recovery was obtained. However, the poorly constructed diamond-coring equipment and resulting high costs of early diamond coring slowed down its progress. Early in the development Of the Rangely Weber field, Rio Blanco County, Colorado, the Rangely Engineering Committee selected key wells that were to core the Weber section in its entirety. These cores were to be analyzed and various sections tested to gather sufficient information so that accurate reservoir studies could be made. The number of wells selected were kept at a minimum inasmuch as the productive Weber section, which is found at an approximate depth of 6000 ft, is thick, being as much as 600 to 700 ft on top of the structure. The Weber section is also comparatively hard to drill. For instance, one well took 60 days to core 490 ft of Weber sand with conventional coring equipment. It was noticed, however, that Core Labora-verse diamond plug cutter could quickly cut a plug from. one of these hard cores. This fact led to the serious consideration of diamond coring the Weber sand. The first diamond coring at Randy was performed by Stanolind in August 1946 andl although we experienced some difficulties, the results definitely indicated that diamond coring would be successful at Rangely for .coring the Weber sand. As more wells were cored and the equipment improved somewhat, diamond coring became less expensive than drilling the Weber sand with the conventional rock bit. In less than one Year diamond coring has definitely Proved its usefulness in oil-field drilling operations in the Rocky Mountain area where formations quite often drill very hard. The practice of spot coring in hard oil-bearing formations with conventional coring equipment is giving way diaimond coring of the entire sections, thus givng an operator the complete picture.

Jan 1, 1948

By Richard V. Hughes, Rudolf J. Pfister

The necessity for getting more water into sands of low permeability in any secondary-recovery water-flood operation in order to recover all the available oil always has been a major problem. In the early days of water-flooding the addition of soda ash and caustic soda to fresh input waters was recommended and tried without success in the Bradford field. The most frequent difficulty was sand plugging. The use of natural and artificial brines often has been suggested for water-flooding, in the belief that laboratory observations of decreasing fresh-water throughput rates in measuring water permeabilities of consolidated sand cores might be partially explained by swelling of clays. A review of available literature bearing on the subject and related laboratory experiments are presented herein, offering an explanation for failure in use of soda ash and caustic soda waters and supporting the recent concept that every water used for input purposes in the secondary recovery of oil should possess such physical and chemical characteristics as to keep the clay content of the producing sands in a permanently flocculated condition. Produced brines meet this requirement. Their use is recommended as a means of increasing intake rates of tight sands and oil recoveries, as a solution to brine-disposal problems, and as a means of conserving a natural resource. Introduction There is a general belief among those engaged in the secondary recovery of oil by water-flooding that one of the most important requirements for the success of any such project is an adequate supply of good, fresh water. Fresh waters have the advantages of being more easily stabilized than salt waters and of being usable about the lease for all purposes. The use of brines for water-flooding never has been given much consideration in the Pennsylvania Grade area. The use of brines in Mid-Continent operations has been considered largely as a substitute for fresh waters, which were economically unobtainable for many operations. Estimated recoveries for some water-drive fields indicate that less of the original oil in place will be left upon abandonment than with the above-average practices used in Pennsylvania Grade water-flood operations. Lower residual oil saturations than 25 pct of the pore space following the water-flooding of Third Bradford sand cores in the laboratory are unusual. In comparison, water-flushed sands surrounding any virgin oil pool are found to be practically devoid of oil. Such lower natural residual oil saturations may be the result of a time factor much longer than economic under artificial water-flood operations, but neither the producer of primary oil nor the producer of oil by secondary recovery can obtain maximum recoveries unless full advantage is taken of all natural forces and

Jan 1, 1947

By Lawrence B. Wright

The occurrence of gold, gold-silver, silver-lead-zinc ores in the post-Cambrian sediments in the Black Hills of South Dakota, and their genetic relationship to the Tertiary intrusives, is well known and accepted, but the occurrence of gold in the pre-Cambrian rocks at the Homestake mine, which underlie this Tertiary gold mineralization, has given rise to differences of opinion as to the period of major gold introduction into the older rocks. Fig. 1 illustrates the distribution of mineralization along the axis of the Tertiary uplift in both pre-Cambrian and post-Cambrian rocks. The following possibilities have been presented: 1. The Homestake gold deposit was formed in pre-Cambrian time and the deposits in the overlying Cambrian sediments formed during Tertiary time. 2. The Homestake lode existed at the beginning of Cambrian time, but was enriched by the passage of Tertiary auriferous solutions. 3. The Homestake lode was represented by pre-Cambrian quartz-chlorite sulfide mineralization, very low in gold content and the gold superimposed on the older nearly barren mineralization during the Tertiary metallogenic epoch. 4. The sulfides and sulfarsenides as well as gold were deposited in both pre-Cambrian and post-Cambrian rocks during the Tertiary period. Those favoring the first proposition feel that the question is pretty well settled, while the proponents of the second theory are more flexible in their view, feeling that the amount of gold introduced in either period is still an open question. The third and fourth theories, with some modifications, are supported by the proponents of the Tertiary theory, who feel that the evidence strongly favors that period. The following pages are intended to present the evidence and to give the reaction of various observers with my own comments and conclusions; also, the accumulated facts of observation gathered over a period of close association with the area under discussion, beginning in 1919.

Jan 1, 1937

By Ramesh Malhotra, Hubert E. (Deceased) Risser

THE WORLD Coal, as a source of energy and as a source of coke for the smelting of iron ore, has contributed significantly to the development of every major industrial nation of the world A number of other materials-oil, natural gas, wood, waste materials, water, and, in some applications, nuclear materials -can be used satisfactorily for the generation of heat and energy, but no generally acceptable and economic substitute has been found to supplant coke in iron ore smelting in conventional blast furnaces. In the development of many nations, coal provided the initial source of energy for industrialization Although coal provided more than 60% of the world's energy each year between 1902 and 1929, oil and natural gas have surpassed it as a source of energy in more recent years (Ref. 38, pp. 439, 441). Despite the rapid rise in importance of other fuels, a considerable amount of the total energy consumed is still supplied by coal. In 1972, coal supplied 32.1 % of the total world energy (Ref. 49, pp. 2-3). To produce this energy, 2.4 billion metric tons of coal were consumed-a 16.0% increase over the amount consumed in 1962. A further worldwide increase in the amount of coal used to supply energy is projected. According to Dr. Karlheinz Bund (Ref. 37, p 145), by the year 2000 world coal consumption could amount to 5 billion mt. Coal Resources and Production The world's coal resources are estimated to be about 10,872 billion mt. Of this amount, 1,297 billion mt are classified as measured reserves (Ref. 50, p. 57). In Table 13.A.1 the estimates of measured reserves and total resources, recent production, and the ratio of measured reserves to production of major coal-producing countries are shown. About 90% of the world's total coal resources are estimated to be in the USSR, the United States, and China. The United States ranks second in the world in coal resources, but in measured reserves, it ranks first. Because of the difference in rate at which available coal resources are being exploited, the ratio of measured reserves to actual production varies widely among areas and countries. France and East Germany have the lowest reserve to production ratios, 49.5 and 60.1, respectively. In Mexico, where currently less than 5 million mtpy of coal are produced, the ratio is 1257 For all countries producing more than 200 million mtpy of coal (except East Germany), the ratio ranges from 350 to 670. For the United States, the ratio is 669.7. In Table 13.A.2, distribution of the world's coal resources by area is shown. More than 97% of the total resources lie in the northern hemisphere, with Asia, including European USSR, having 63.4%. Only 1.9% of the total coal resources are estimat-

Jan 1, 1976

By P. F. McIntyre

Electrolytic refining of lead was first carried out on a commercial scale at Trail, B. C., in 1903. The capacity at that time was small, and cell-changing operations were managed with chain block and monorail. Development up to 1922, utilizing crane installations and cell additions, attained a production of 100 tons per day. Successive additions from 1922 to 1926 further increased the rated capacity to 396 tons per day. Other improvements in metallurgical practice have gradually increased production, until for the year 1935 it was 448 tons per day. During the first five months of 1936, an average daily production of 504 tons was maintained without increasing the size of the plant. Lead Refinery Buildings.—The tank room and the lead-melting and casting equipment are housed in two buildings, which parallel each other and which for the greater part of their lengths are joined by a connecting aisle. All basements, walls, sumps and floors are of concrete. Where exposed to the electrolyte, the concrete has a protective coating of asphalt (220" melting point) over a coat of asphalt paint as a primer. Buildings erected before 1924 are of wood. More recent additions are made of structural steel with gunite siding. A brick building, adjoining the older section of the tank room, houses the slimes-treatment plant. The power house, containing motor generators, air compressors and rectifiers, is constructed of tile. Cell Construction.—The electrolytic cells are of reinforced concrete lined with asphalt. They are cast in pairs, the outside dimensions of a double cell being 113 in. long by 73 in. wide by 48 in. deep. The bottom is 6 in. thick and the sides, ends and center wall are 4 in. thick. A ledge 3 in. wide and 2½ in. deep is left along the top of the side walls to carry the insulators and bus bars. Reinforcing consists of ½-in. round iron rods. Cell inspection and repair work is done in conjunction with the cell-changing operations. The repair work necessary is usually confined to renewing chipped asphalt, unless a tank has been accidentally broken. Minor concrete repairs can be easily made in place. The broken concrete

Jan 1, 1937

By Robert W. Thomas

During the past 20 years the advantages of reinforced concrete as a substitute for timbering in so-called permanent mine openings have been fully recognized, and its use has become almost general practice in the larger mines. The term "permanent mine openings" is used advisedly and is assumed to embrace such openings as ventilation shafts, material shafts and main hoisting shafts, with their adjacent stations, ore pockets, etc. In the concreting of such openings, the actual cost as compared with that of timbering has been of secondary consideration, the support of large openings, elimination of fire hazards, and prevention of interruptions to production being the primary objectives. The ultimate economy, however, of lining such openings with concrete when they are to remain in use during the entire life of the mine, and especially when they are in bad ground, will seldom be questioned. During this 20-year period there have appeared in engineering publications a number of interesting articles on the use of concrete in mines, nearly always describing some specific project of major proportions, such as the concreting of ore pockets, large stations or main hoisting shafts. Such articles are unquestionably of interest, since the projects require considerable engineering thought, not only in the method of excavating large openings, but in their engineering design as well. Concreting of main hoisting shafts, particularly when this is done with the least possible interruption to production, provides some interesting problems. The concreting of haulage drifts of the dimensions commonly found in mines, on the other hand, soon develops into a routine operation, therefore it is with some hesitancy that the author attempts to prepare an article on this subject. The writer can recall only one article on the concreting of mine haulageways, although several articles have been written concerning the concreting of tunnels considerably larger in dimensions than those required in general mine practice.' From the scarcity of articles on the subject, it would seem that the concreting of haulage drifts is still far from being considered general

Jan 1, 1937

[Cyanidation: Fresnillo, Mexico. 112. 734, 739. Abs.. YB84. 34 gold ore: before or after amalgamation. respective merita. 118, 538 high residue avoided by milling in water, 118, 542 Homestake mill, 112,607. 614 impoverishing gold ore before cyanidation justified. 118, 539 Kirkland Lake district, 112, 648, 649. Abs., YB84.32 literature prior to 1932. See Source Book. McIntyre Porcupine mines, 118, 638 mercury as an aid to extraction of gold, 118, 53 8]

Jan 1, 1936

By R. L. Perrine

A stability theory is developed for miscible liquid-liquid displacement within a porous medium. In the usual case considered, a high-density high-viscosity "oil" is displaced downdip by a low-density low-viscosity "solvent'! Perturbation methods are used to find the conditions under which the spreading mechanism changes from the stable dispersion process to unstable viscous fingering. We find that instability is conditional, that there is a dependence on the shape of a disturbance leading to a"diameter" effect and that very difficult experimental scaling problems may result. A useful consequence is the definition of a minimum "slug size" for stable miscible displacement. This should make possible optimum use of the solvent process for oil recovery. The results may apply to many situations in which one fluid displaces another of somewhat different fluid properties within a porous medium. INTRODUCTION The process of miscible liquid-liquid displacement looked very favorable when it first received serious consideration as a means to increase petroleum recovery. Some early laboratory results were interpreted to mean that only a small "slug" of solvent was needed, perhaps 2 to 3 per cent of a hydrocarbon pore volume. This "slug" could recover the oil in an entire reservoir provided the fluids remained miscible.1,2 Only a slow growth of the "mixing zone", or gradual transition from oil to solvent, should occur. This would follow naturally from mixing by a dispersion mechanism such as that described by Scheidegger.3 And, indeed, laboratory cores have shown this kind of behavior provided solvent viscosity was at least as great as that of the oi1.2,4-6 The same kind of behavior has also been observed with solvent viscosity lower than that of the oil after the distance traversed became large. Not all laboratory results have been this favorable, however.7-9 Mixing zone growth may become proportional to the distance traveled rather than the square root of this quantity. Such behavior accompanies high rates in short systems of large diameter, provided the viscosity ratio is adverse. If this more rapid spreading were to persist in a reservoir, the solvent requirement for miscible "slug" displacement could exceed 30 per cent of a pore volume. There is considerable economic importance in the difference between 3 and 30 per cent solvent. A 1ogical conclusion is that two different kinds of flow behavior are possible, with each kind leading to a different result. One kind gives efficient displacement, the other does not. In the efficient case, the solvent bank spreads out only by the mechanism we have termed dispersion. Under these conditions, displacement is as near piston-like as possible. In the second kind of flow, found only with adverse viscosity ratios, viscous fingering occurs. That is, permeability variations cause a small finger of low-viscosity solvent-rich fluid to move ahead of its average position within the mixing zone. This creates a path of low resistance to flow, and an even greater amount of solvent-rich fluid follows. Thus, the process is autocatalytic. Once started, the fingering mechanism rapidly becomes dominant. The question to be answered is this. Under what conditions will each of these two different kinds of flow occur? In particular, what conditions denote the transition from dispersion to viscous fingering as the solvent spreading mechanism? The answer may tell us whether or not the desirable, near piston-like miscible displacement process is practical. An answer to this problem can be obtained from theory by the use of perturbation methods. The procedure is as follows. We first formulate a mathematical representation of the system. Then a small disturbance (or perturbation) in solvent concentration profile is introduced and observed to see what happens. Unstable flow is indicated when a small disturbance will grow larger. This will lead to eventual viscous fingering. If on the other hand the disturbance dies out, the displacement is stable. In this latter case, with dispersion as the spreading mechanism, near piston-like displacement is possible. Thus, this paper presents the development of a stability theory for miscible liquid-liquid displace-

By A. G. H. Andersen

A FEW years ago, while the writers worked on the constitution of ternary and quaternary alloys consisting mainly of iron, chromium and silicon1,2 some information on microstructures and hardness, not incor-porated in the published work, was obtained. It has not been possible to do any further work on the properties of the alloys. The information possessed is hereby presented for the interest it. may have in connection with the constitution of the alloys.

Jan 1, 1937

By D. W. Fuerstenau, G. D. Gumtz

Use of the discretized batch-grinding model for the simulation of locked-cycle grinding tests from batch-grinding data is illustrated. The simulated results were compared with actual locked-cycle experimental data for the dry grinding of dolomite. To illustrate the utility of this simulation method, a number of relationships between operating and performance parameters were generated, including relations between batch grinding and the Bond test, the production rate and the circulating load, and the production rate and the product size. A large amount of testing is done each year in the mineral industry for use with grinding unit cperations, and much of the experimental test data is obtained by the tedious locked-cycle testing procedure. In addition to being experimentally complex and requiring large amounts of sample, the very nature of the locked-cycle experiment yields results that cannot readily be interpreted. The attempts that have been made to model comminution systems during the last few years may provide a better and simpler way to obtain and analyze grinding test data; these models involve elements of both fitting to "model equations" and deterministic representations.1-1 The traditional engineering design techniques for comminution systems have relied mainly upon large amounts of industrial operating data that have been correlated with locked-cycle grinding data.',' Bond's empirical mill-design procedure involving locked cycle testing remains the most generally available method to this day. The purpose of the present paper is to show how much more useful data can be obtained from a simple batch-grinding experiment that provides information which subsequently can be utilized with population balance models of the grinding process to simulate locked-cycle behavior. These simulation results can, of course, be used in Bond's general design scheme. However, the combination of batch-grinding tests and the population balance model allows many more conditions to be considered than do single laboratory tests under specific locked-cycle conditions. The use of these simulation techniques for prediction of locked-cycle behavior represents one of the first practical applications for the simulation method in comminution.

Jan 1, 1971

By Theodore Simons

(Butte Meeting, August, 1913.) DURING the last half century a great amount of ingenuity and energy has been devoted to the invention of appliances for the recovery of valuable minerals from very fine sands and slimes. The reason for this is that in almost every dressing plant the greatest losses of values, considered relatively as well as absolutely, occur in the treatment of slimes. The natural aim of mill operators to minimize these losses has in recent years received another impetus, from the fact that the gradually diminishing occurrence of high-grade ores makes a more intense recovery of values from existing resources absolutely necessary for profitable operation. The results of such attempts are found in the appearance of a great number of machines and appliances which make it profitable to-clay to rework old dumps, containing the tailings of older processes. A constantly growing difficulty of obtaining skilled labor and the increase in wages made it prohibitory to employ those types of machines which required much attention, and in the middle of the nineteenth century the various concentrating tables of the percussion type, which automatically discharge their products, began to take the place of the intermittent tables, on which, when the bed had reached a certain thickness, operation had to be stopped in order to allow the skimming by hand of the separate layers formed on the table. In the course of years, two general types of machines evolved for the treatment of fine sands and slimes : The so-called concentrators and vanners, on the one hand, and the buddies and round tables on the other.1 For the profitable treatment of the very finest slimes, round tables have as a last resort proved more satisfactory and economical than other machines.

Jan 7, 1913

In response to Dr. Ledoux's paper, a large number of metallurgical establishments and individual assayers expressed their willingness to co-operate in the plan he proposed. The necessary samples were prepared as follows : Matte.—A car-load of matte was put through a Blake crusher, then automatically subdivided into tenths by a Taylor and Brunton sampler, and one-tenth passed through Cornish rolls and through a 12-mesh screen. This was again subdivided by an automatic sampler into tenths, and a final tenth, representing 1 per cent. of the original lot, was finished by hand-sampling on an iron floor until 100 pounds remained. This was pulverized to pass a 40-mesh screen and then thoroughly mixed and divided in the presence of my assistant for distribution to the parties co-operating. Copper Borings.—These were taken from a lot of anodes, the dip-samples from several batches having been united, remelted and cast into a plate, which was proved by assay to be of uniform quality in its different parts. Borings from this plate were intimately mixed and divided in the presence of my assistant. Samples were sent out November 1,1894, with the following letter: New York City, November 1, 1894. Samples of matte and copper (about I pound each) carefully prepared in the presence of my assistant and sealed with his seal as below, are sent you by express for assay and report according to the plan of Dr. Ledoux. These samples are constituted about as follows: Matte. Copper. From. To. From. To. Copper, per cent.,.....50 60 ..........., Gold, oz. per ton,.....2 3 0.20 0.50 Silver, oz. per ton,.....100 150 140 180

Jan 1, 1896

By W. R. Atkins

THE Negaunee mine is at the east end of the Marquette Range, in the city of Negaunee, on the Upper Peninsula of Michigan. Iron ore was first discovered on this property in 1883 by diamond drilling. A shaft was sunk soon afterward, and the mine has been worked continuously ever since. The total shipments from 1887 to' Jan. 1, 1945, were 20,417,587 long tons. The mine was taken over by the present operators, The Cleveland-Cliffs Iron Co., in 1903. During the past four years, 1941 to 1944 inclusive, the property has produced an average of nearly a million tons a year, and has employed more than 400 men. GEOLOGY The ore body at the Negaunee mine lies in a large syncline dipping about 15° to the west, the ore coming to ledge at the east end of the trough. The ore occurs in the Negaunee formation of the middle Huronian age, which consists of ferruginous cherts and slates through which intrusions of diorite have occurred. Although these intrusions of diorite have an important bearing on the formation of ore on the range, they are not important in this ore body. Some faulting has taken place along the diorite intrusions, or dikes as they are called locally, and also other faults are known to exist across the formation. These faults do not complicate either mining or development of the ore body. The footwall consists of jasper underlain with Siamo slate; the hanging wall is jasper and ferruginous chert. Fig. I shows a typical cross section of the ore body. The ore body is quite regular, about 120 ft. thick in the middle of the syncline and thinning out toward both sides of the fold. Its width ranges from 400 to 800 ft., the average being about 600 ft. The footwall is uniform in dip and strike, but irregular lenses of ore often extend 50 to 75 ft. up into the overlying jasper capping. METHOD OF MINING Ore in the Negaunee mine has been mined by top slicing almost exclusively since 1903. This method of mining has been employed because it is especially adapted to the size and shape of the Negaunee ore body, and to the soft, sticky nature of the ore as well as the generally soft, broken jasper capping. Development The ore body has been developed from six different levels with intervals varying between l00 and 125 ft. On the average, the dip causes a horizontal offset of 350 ft. per level, and under these circumstances it has been necessary to carry on considerable development in rock to make available the ore lying along the north and east footwalls. This is not as disadvantageous as might be supposed, because of reduced pressure and the accompanying saving in maintenance costs. The development of the

Jan 1, 1946