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By George D. Dub

Owens Lake is at present a source of important nonmetallic minerals, sodium carbonate (soda ash, Na2CO3); sodium sesquicarbonate (trona, Na2CO3.NaHCO3.-2H2O) and borax, (Na2B4O7.10H2O). Owens Lake is a closed basin in the southern part of Inyo County, California, at the southern end of Owens Valley, east of the Sierra Nevada Mountains and west of the Coso and Inyo Mountains. Broadly considered, it is in the Great Basin area, but at no time was it a part of Lake Lahontan. Closed Basins Closed basins are phenomena of arid or semiarid regions where outflow is considerably less than inflow and where accordingly soluble salts concentrate in residual liquid. When inflowing waters originate in areas where the rocks are predominantly marine sediments, the residual basin liquid is primarily a chloride, brine; if the rocks are largely igneous, the residual brine tends to be alkaline containing carbonates, and sometimes borates as well. Since normally, both types of rocks occur in regions contiguous to closed basins, residual liquids do not often fit into the two broad divisions mentioned. No two residual brines are exactly alike, just as no two ore deposits are exactly alike. Even out of the same closed basin, it is possible to get widely different analyses of liquids since underground and surface flows, as well as local evaporation rates and other conditions might have a marked influence on residual-liquid compositions. At Searles Lake, The American Potash and Chemical Corporation is building a plant to process a lower-level brine which is considerably higher in sodium carbonate and borax, and lower in potassium chloride, than that company has processed for many years. F. W. Clarke1 has classified waters of closed basins as follows: I. Chloride type; largely sodium chloride (NaC1) and of oceanic type, such as Great Salt Lake. Related is -the Dead Sea, a bittern residue of magnesium, potassium and sodium chlorides. a. Sulphate type; largely sodium sulphate (Na2SO4) with considerable sodium chloride such as Sevier Lake, Utah; Laramie Lakes, Wyoming; Dale Lake, California. 3. Carbonate type; high in carbonates and fairly high in sulphates, such as Moses Lake, Eastern Washington; and the Nebraska Potash Lakes. 4. Carbonate—chloride type; lower in carbonates than type 3 and about equally rich in sulphates, such as Pyramid Lake, Nevada. 5. Sulphate—Carbonate type; quite high in sulphates and carbonates, such as Pelican Lake, Oregon. 6. Triple type; considerable quantities of carbonates, sulphates and chlorides

Jan 1, 1948

By D. J. McAdam

In a recent survey of the literature, the author has found evidence incompatible with prevalent views regarding the technical cohesive strength and yield strength of metals. Some of the evidence regarding the technical cohesive strength was presented in a previous paper,21 and views were presented that are believed to be in better accord with the evidence. In that paper, it was shown that the technical cohesion limit (contrary to prevalent views) varies with the combination of principal stresses,t and that the influence of plastic deformation on the technical cohesive strength is not in accordance with prevalent views. In this paper, consideration will be given to both the technical cohesion limit and the yield stress as affected by the combination of principal stresses. Evidence will be presented that the variations of the yield stress and the technical cohesion limit with the combination of principal stresses are not in accordance with prevalent views. Consideration will also be given to ductility and work-hardening capacity, which depend on the initial difference between the technical cohesion limit and the yield stress. By the "technical cohesion limit" is meant the mean stress at actual or poten- tial* fracture. An example of a rechnlcal cohesion limit is the "true" breaking stress of a tension-test specimen, the breaking load divided by the sectional area at fracture. The cohesion limit thus obtained for a ductile metal, however, has been greatly altered by the plastic deformation. It is important to consider the cohesion limit of the metal prior to this deformation, and the variation of the cohesive strength with plastic deformation. The relation between the cohesion limit and the corresponding flow stresst determines the ductility of a metal and the energy that it will absorb before fracture. In this paper, as in the previous paper,21 tensile stresses will be viewed as positive and compressive stresses as negative. The algebraically greatest principal stress will be designated S1, the least principal stress will be designated S3, and the intermediate principal stress will be designated Sz In the previous paper, attention was confined to stress combinations with two of the principal stresses equal (with the intermediate principal stress S2 equal eithel to S3 or 31). In this paper, consideration will also be given to the effect of stress combinations with all three principal stresses unequal. In the earlier sections of the paper (pp. 3I I to 324), attention will be confined to stress combinations with

Jan 1, 1942

By Philip Davidson, Wm. B. Price

The object of this paper is to present, a survey of some of the physical properties of sheet nickel silver when subjected to cold rolling and when annealed at temperatures extending from 350" C. throughout the coininercial annealing range.. A number of alloys were taken from stock and analyzed. Four of these, representing low nickel and high nickel both with and without lead, were chosen for the tests. The analyses were as follows: Alloy Copper, Lead. Iron. Nickel, ManGanese, Zinc, PeR Cent. Per Cent. Per Cent. Per Cent. Per Cent. PEr Cent. A................ 64.68 Trace 0.194 6.73 0.06 remainder B................ 65.82 1.27 0.227 6.17 0.06 remainder C................ 65.44 Trace 0.345 17.83 0.08 remainder D................ 65.60 1.08 0.238 17.77 0.06 remainder These alloys were lolled in accordance with regular mill practice to approximately No. 6 R. & 6. gage. The bars (which were then coiled) were annealed together in a mill muffle and afterwards pickled in a 10 per cent. sulfuric-acid solution and cold rolled in stages of approximately one number (B. & S.) at a time to No. 18 B. & S. gage. After each stage, suitable samples were set aside for physical tests. Additional samples were chosen for tests after the bars were rolled to Nos. 8, 10, 12, and 14 gage. These samples were annealed together in the mill muffle and, after pickling, were finished on 0.050-in. gage, yielding a series of specimens 2, 4, 6, and 8 numbers hard for the subsequent annealing tests. Tensile test pieces 9 by 1 in. (22.9 by 2.5 cm.) were cut from both lots and the central portion was milled to give a test section 2 1/2 in. long by 1/2 in. wide. A distance of 2 in. was marked off on the edge of each specimen with light punch marks. In the case of the annealed alloys, Erichsen test pieces 4 by 235 in. were also taken.

Jan 1, 1921

By E. J. Horwood

The varying physical character and large extent of the Broken Hill lode necesarily involve the employment of a variety of underground methods. The lode had its origin in an extensive fault plane traversing metamorphosed schists conformably, as a rule, with their beds of stratification. The underground waters carrying minerals in solution deposited their contents in the original cavities formed by the faulting action, and in the enlargements of these cavities due to dynamic forces brought to bear on the rocks, more especially on the hanging-wall side of the fault. This deposition was supplemented by metasomatic replacement of a portion of the original rock contents by the argentiferous sulphides of lead and zinc which form the staple products of the district. Although the orebody is practically continuous throughout the mines, its width varies greatly, ranging from a few feet to about 350 ft. The widest portions occur in conjunction with huge folds in the inclosing country rock, almost exclusively on the hanging-wall side. The ore in these folds pitches to the south in the southern half of the field, and to the north in the northern half; there are, however, undulations in these ore channels evidently due to compression of the rocks in the direction of the channels. In the earlier days, before the orebodies had been opened up, vertical cross-sections across these bulges in the hanging wall gave the lode the appearance of the "saddle formation," so well exemplified at Bendigo in Victoria; but subsequent development of the ore channels has long since proved that the formations in the two districts are entirely different. The depth reached by the zone of complete oxidation of the sulphides varies from about 259 to 550 ft, from the original outcrop, while partial oxidation extends in places below the 1,000-ft. level. The result of oxidizing influences has been the production of ore of every grade of cohesive-ness, from that of dry sand to that of hard compact rock. The methods of mining followed in this field, and even in individual mines, are varied in accordance with the character of the ore. The general practice of the field is to sink vertical shafts, generally on the foot-

Jan 1, 1916

By P. W. George

This paper presents geological data regarding deposits of over 7,500,000 tons of fine-grained sulphide ore in barite gangue. A series of pyroclastics and arkoses was intruded by sills of quartz porphyry in connection with batholithic intrusions of granite 3 miles from Buchans. Gentle folding, with fracturing and shearing of tuff beds and intruded quartz porphyry, accompanied and followed the intrusion of granite and porphyry. The sulphide-barite mineralization is localized primarily in fractured and sheared tuff beds, but also in the sheared quartz porphyry. Two structures have proved particularly favorable for mineralization-one, the center of an anticline; the other, a series of weak, incompetent tuff beds intruded by thin sills of quartz porphyry and located between strong beds of arkose. During the folding the weak tuff with intruded quartz porphyry was intensely sheared and contorted and subsequently mineralized by sulphides and barite. INTRODUCTION Buchans is 5 miles north of Red Indian Lake, in Central Newfound- land, at an elevation of 900 ft. above sea level and 400 ft. above Red Indian Lake, and is reached by a 37-mile private railway from Miller- town Junction. The first ore body found at Buchans was discovered some 30 years ago in the bed of Buchans River, by Matty Mitchell, an Indian woodsman employed by the Anglo-Newfoundland Development Co. Ltd. In the years following the discovery the ore was developed by shaft-sinking and drifting, with results indicating a lenticular ore body in a shear zone, containing about 100,000 tons of extremely fine-grained sulphide ore in barite gangue averaging about 0.05 oz. Au, 4 oz. Ag, 1.5 per cent Ch, 10 per cent Pb, 18 per cent Zn. Because a profitable metallurgical treatment for the ore was not found, the mine was closed down in 1911. In 1915, Mr. H. A. Guess, Vice President of the American Smelting & Refining Co., learned of the Buchans River ore body, and after metallurgical tests had been carried on from time to time in his company's testing plant a satisfactory separation of the sulphides by selective flotation was finally effected in 1925. Prospecting for additional ore bodies was begun in 1926, utilizing geophysical methods under the direction of Hans Lund-

Jan 1, 1937

By William H. Case

The Bertha zinc-mines of the Bertha Zinc and Mineral Company are in that part of the State of Virginia popularly known as Southwest Virginia, and are included in that division of the " Great Valley " named in recent years the " New River-Cripple Creek Mineral-Region of Virginia." The mines lie wholly in the southeast part of Wythe county, two miles from Pulaski county on the east, and three miles from Carroll county on the south. The only other zinc-mines that are opened in the region are those of Manning & Squier, which adjoin the Bertha zinc-mines on the northeast, and the mines of the Wythe Lend and Zinc Company, at Austinville, eight miles to the southwest. A little farther southwest of Austinville, near Ivanhoe, a new company is preparing to mine zinc-ores. The mining properties named are included within a distance of ten miles along a direct line, bearing north 55' east, on the south side of New river, which here flows northeastward in a tortuous course. The greatest distance of any of the mine-openings from the river is a little less than two miles. It is not the magnitude of the operations in zinc-mining in this particular region, as compared with that of other well-known districts, that merits especial notice. Aside from the exceptional purity of the spelter made directly from the Bertha ores, some peculiar features of the zinc-ore deposih, as shown at the Bertha mines, and

Jan 1, 1894

By H. W. Meyer, H. N. Lander, F. D. Delve

A method of calculating the changes in blast-furnace performance brought about by burden and/or blast modifications is presented. Essentially the method consists of three simultaneous equutions derived from materials and heat balances. These equations can be used not only to evaluate quantitatively the effect of changes in process operating variables on furnace performance, but also to provide a useful means of evaluating changes in process variables which cannot be measured directly. It has been customary for a number of years to use simple heat and materials balances as a basis for assessing blast-furnace practice. A good example of the method used to set up these balances is that proposed by Joseph and Neustatter.1 This approach to process assessment has limited utility, however, in that it cannot be used to predict the furnace coke rate or production under new operating conditions. Using an approach based on multiple correlation of blast-furnace variables, R V. Flint2 has developed an equation which may be used to predict the change in coke rate that will result from some changes in operating conditions with a reasonable degree of accuracy. Although this equation has useful applications in production planning, it cannot be used to study the relationships between the operating variables and the fundamental thermochemi-cal characteristics of the process. In attempting to analyze the blast-furnace process quantitatively, the idea of dividing the furnace into zones3 may at first appear attractive. In our present state of knowledge, however, it is not possible to define with any accuracy the physical limits of such zones in relationship to their temperatures or to the reactions which may occur in them. Although its application is restricted, the zonal approach to blast-furnace analysis is useful in some instances. For example, the change in the calculated flame temperature in the "combustion zone" caused by injecting steam constitutes information which is helpful in understanding why the addition of steam to the blast is best accompanied by an increase in blast temperature. The zonal approach cannot, at the present time, be used to establish the relationships between process variables and process performance if the whole process rather than part of it is to be considered. One of the earliest approaches to the problem of relating blast-furnace operating variables to pro- duction and coke rate was that developed by Marshall.4 Essentially Marshall's work showed that it was possible to estimate the performance of a furnace by solving three simultaneous equations which consisted of rudimentary carbon and heat balances plus a further equation relating the production, wind rate, and the carbon burned at the tuyeres. Although these equations did not include all of the chemical and thermal variables of the process, their derivation and application seems to be the earliest attempt which achieved any success in relating prior furnace operating data to the calculation of furnace performance under different blast conditions. Work carried out in Germany has been directed mainly towards prediction of coke rates using material and thermal balances rather than statistical methods. wesemann5 used prior furnace operating data as part of the basis for predicting the change in coke rate accompanying a change in burden composition. This author employed a method of successive approximations to estimate the secondary changes in slag volume and stone rate brought about by the change in coke rate. The most recent analysis, which seems to have been developed concurrently with the thermochemical model presented in this paper, has been described by Georgen.6 This author has succeeded in improving on Wesemann's approach by expressing the total changes in the slag volume and stone rate in terms of the change in coke rate itself. This is accomplished in a manner similar to that used in the thermochemical model described in this paper. Although Georgen makes use of a calculated furnace heat loss, he does not relate the heat loss per unit of hot metal to the production rate as is done in the present work. Georgen's approach may be used to calculate the changes in materials requirements accompanying changes in furnace operation; it cannot be used to assess the resulting changes in production. The fact that blast-furnace behavior can be interpreted by consideration of the heat requirements of the process was demonstrated by Dancy, Sadler, and Lander.7 In the analysis of blast-furnace operation with oxygen and steam injection these authors showed that it was possible to account for the changes in production and coke rate

Jan 1, 1962

By R. W. Heckel

The growth and shrinkage rates of second-phase particles of various size distributions are analyzed in terms of five mathematical models. These models apply to various forms of diffusion-controlled growth (one model) and interjace-controlled growth (four models). The expression for diffusion-controlled growth assumes local equilibrium at the interface between the second Phase and the matrix and through the choice of the proper flux area perrnits the definition of minimum diffusion-controlled growth kinetics. The interface-controlled kinetics models consider the situations where the rate-controlling step occurs at either the interlace undergoing solution or the interface undergoing deposition. In addition, each of these interface-controlled kinetics models is analyzed for interface reaction rates which are independent of the thermodynamic activity gradient across the interface or proportional to the gradient. THE variation of the size distribution of second-phase particles in a solute-saturated matrix has been a stimulating topic for many years. Research has been carried out in this area for situations where the matrix was either liquid or solid. It is well-understood that the over-all process results in an increase in the average particle size, a decrease in the total number of particles per unit volume, and a net decrease in surface area. Greenwood1 has presented a mathematical model based upon the mass balance existing between growing and shrinking particles. His analysis is stated in terms of Fick's First Law using concentration gradients set up between particles of different size as a result of the effect of surface curvature on solubility. Greenwood's analysis indicates that the mean particle size should increase approximately as the cube root of time. Asimow2 has shown that several of the simplifying assumptions employed by Greenwood may be eliminated from this diffusion-controlled model with little increase in complexity. Lifshits and Slyozov3-5 have also considered the growth-shrinkage problem involving diffusion-controlled kinetics. Their analysis was extended to include the effects of interacting particles ("kinetic approximation"), particle shape, and elastic strains. These authors also show that the mean particle size varies as the cube root of the reaction time. wagner6 has formulated expressions for both diffusion-controlled kinetics and interface-controlled kinetics in the growth-shrinkage problem. Wagner's diffusion-controlled kinetics is based upon the same form of mass balance using Fick's First Law as used by others and is shown to be applicable to situations where kr >> D (k is the interface reaction constant, r is the particle radius, and D is the diffusion coefficient). This analysis also yields a mean particle size which varies as the cube root of time, in agreement with other workers. Wagner also developed interface-controlled kinetics to handle kr « D. This model was based upon the fact that the interface reaction rate is proportional to the interface area. Wagner assumed that the rate was also proportional to the thermodynamic activity difference, ?a, existing across the interface. Thus, the specific reaction rate at the surface of radius r was defined as k • 4nr2 . ?a. It is the purpose of the present paper to reconsider the processes of diffusion-controlled growth and interface-controlled growth in the variation of particle-size distributions as a function of time at constant second-phase volume. The analyses are formulated in a manner which permits data obtained

Jan 1, 1965

By E. Horton Jones

INTRODUCTION WE have endeavored in the following "sheets" to give the unit construction costs derived from the building of the Arizona Copper Co.'s new smelter, Clifton, Ariz., starting in February, 1912, and completing February, 1914. 1 In Chapter I-Unit Costs-are to be found the most elementary total unit costs which the accounts provide for. They are usable in the Clifton district. Here too are found the percentages to be added to an estimate for Engineering and General Expense. In Chapter II-Comparative Costs-these elementary costs have been classified, averaged and reported as labor and material unit costs. In such form the labor unit costs, when properly applied to similar conditions as these under which they were derived, are usable anywhere. The material unit costs are better disregarded for more accurate estimates and replaced by a newly priced bill of material. In Chapter III Composite Costs are given. They are unit costs built-up from several elementary units, and likewise units of larger dimensions and simpler application, valuable for

Jan 7, 1914

By G. M. Schwartz, J. N. Gundersen

Utilization of magnetite-bearing taconite from the great Mesabi range of Minnesota is fast becoming a major industry in the state. The planning and early development stages of taconite processing projects have been de-scribed1-3 and an excellent discussion has appeared recently4 concerning the current mining, crushing, concentrating, and pelletizing techniques of the Reserve Mining Co., whose operations lie entirely within the Eastern Mesabi district. The present paper will deal only with some of the geological aspects of taconite utilization in the district, with emphasis placed upon some of the common mineralogical and textural features of the tacon-ites currently being mined or stripped and how these factors generally affect the potential recovery of the iron from the metamorphosed Biwabik iron-formation. The generalized geologic setting of the Eastern Mesabi district is presented in Fig. 1. The small circles on the geologic map, adapted from Grout and Broderick,3 mark the location of drill holes that were available for inspection during this study through the courtesy of Reserve Mining and Erie Mining Cos. The cores from these holes were divided into stratigraphic units and examined for their mineral content. Typical mineral associations and grain fabrics were also determined from hundreds of thin and polished sections. Available metallurgical tests made it possible to determine the probable response of a given variety of taconite to magnetic concentration. The longitudinal section of Fig. 1, projected along the strike of the iron formation, presents the simple stratigraphy of the area. It also shows the uniformly thick nature of the iron formation members in the district as well as a somewhat abrupt thickening of the formation to the west of the probable fault. The stratigraphy of these precambrian rocks has been described5-' elsewhere and need only be briefly reviewed. Historically, the oldest rocks unit, the Giants Range granite (Algoman) was covered by the Animikie group of sediments (Middle Huronian?), consisting of the Pokegama, Biwabik, and Virginia formations. Outcrops of the Pokegama quartzite are locally distributed along the northern limit of the Biwabik iron formation, but they are too limited to depict on the geologic map of Fig. 1. The effects of the intrusion of several small diabase sills within the Virginia and Biwabik formations are obscure because all of these rocks have been subsequently highly metamorphosed, mainly by the Duluth gabbro. The emplacement of the transgressing Duluth gabbro (Middle Keweenawan) and that of numerous peg-matitic veins in the iron formation8 has resulted in extensive contact and metasomatic metamorphism which has locally reconstituted the previously existing minerals of the Biwabik formation, in large part quartz and magnetite, into a wide variety of silicate mineral assemblages9 that have a profound effect upon the re coverability of the remaining magnetite. The entire region has subsequently been eroded and largely buried beneath Pleistocene glacial deposits. Structurally, the strike of the gently dipping Animikie group generally parallels the elongate outcrop pattern of the Biwabik iron formation on the map of Fig. 1. The Animikie group most commonly dips from 5" to 15" to the south-southeast. In this area, the Duluth gabbro pluton is somewhat sill-like, dipping about 30" to the southeast. Projecting these present structures upward, it seems probable that the gabbro once extended over the underlying Animikie group in the Eastern Mesabi district. SELECTION AND PRESENTATION OF DATA Because of the locally intense metamorphic activity in the district, mineralogical variations within the magnetite-bearing taconites are commonly abrupt, both vertically and laterally. Consequently it is difficult to select a single hole or small group of holes whose cores will be entirely representative of all of the district. Two holes have been selected, however, that in a general way represent mineralogical and textural variations that are likely to be encountered in both the eastern and western parts of the district. These recently drilled holes were chosen for discussion because it was possible to log the core and then to place the sample intervals at stratigraphic boundaries prior to preparation of the core for magnetic tube tests.

Jan 1, 1961

By Harmon F. Fisher

In the course of a comprehensive investigation for the development of the electrical process for the dehydration of oil-field emulsions, the writer has had an unusual opportunity to direct and execute a number of related, parallel studies concerning crude oil emulsions and extending over a period of nearly six years. Among these was an investigation utilizing the microscope for studying both the general characteristics of crude oil emulsions, the nature of their electrical charge and the action of an alternating current field in causing coalescence of the finely dispersed, more or less permanently suspended, emulsion particles into larger droplets of such mass that subsequent dehydration would follow by the ordinary process of settling, further agglomeration of the water droplets and, finally, separation into commercially dry, pipe line oil and a clean water discharge. It is not the purpose of this paper to enter into a detailed discussion of the highly specialized subject of microscope technique, nor into a theoretical discussion concerning the origin of the electrical charge on the emulsion particle or concerning the mechanism of the electrical phenomena occurring during the electrical treatment of oil-field emulsions. In a short paper of this character one can hope to touch on only a very limited discussion of the theoretical aspects of crude oil emulsions. The subject of emulsions and colloids is comprised in a comparatively young science in which, however, very rapid progress is being made and consequently it is undergoing many changes; a tentative conclusion valid today may be overthrown tomorrow. Nor would it be a simple science even if it were no longer undergoing this rapid progress; a thorough study of emulsions is based upon a knowledge of such fundamental subjects as molecular structure, ionization and other electrical phenomena, surface chemistry, surface and interfacial tension phenomena, viscosity and other factors too numerous to discuss here and the fundamental theories of which are still undergoing more or less change. Therefore this paper will confine itself largely to a generally descriptive discussion, a sufficiently detailed outline of the microscope technique and a brief comment on

Jan 1, 1931

THE refining period of an open-hearth heat blends imperceptibly into the finishing period, during which final adjustments are made in slag composition, in bath action and temperature, and in composition of the steel, in order to prepare it for casting into ingots. Operations must be governed by the consideration that alterations in the character of the heat during the finishing period must be held to a minimum. In a broad sense, the finishing and deoxidation period of a heat may be said to be- gin when the action of the last ore addition has ceased, and to end when the teeming of the ingots is completed. INGOT STRUCTLIRES A survey of the relative advantages and disadvantages of different types of ingot structure is required because deoxidation treatments during the finishing period control the amount of gas evolved during solidification and consequently determine the type of ingot structure. Four classes of ingots are recognized, de- pending upon the degree of deoxidation to which the steel has been subjected. These types are killed, semikilled, capped, and rimmed; typical examples of their structures are represented by ingots 1, 2, 5, and 7, respectively, of Fig 9-1. The actual ingots from which these cross sections were sketched were all poured to the same height (indicated by the dotted lines) in the same type and size of bottle-top molds. The eight drawings of the figure represent selected types of ingot structure in a continuous series from "dead killed" to "wild," resulting from variations in the total quantity of gas evolved, from none at all to copious amounts, during solidification of the steel in the mold. Killed and semikilled steels are adjacent at the "quiescent" end of the ingot structure series where only small amounts of gas are evolved. The term "quiescent" describes the action of the molten metal during and subsequent to the teeming of killed and semikilled steels. Capped and rimmed steels are adjacent at the

Jan 1, 1964

By F. C. Bond

Since the Third Theory of Comminution was presented eight years ago (I) it has found increasing use in crushing and grinding problems. The practical utility of its wok index equation is quite generally acknowledged (2). However, its theoretical basis has been questioned in at least three technical articles (3) (4) ('). The purpose of this paper is to present experimental proof that it is scientifically correct. Particles under compressive stress are strained and deformed. They absorb strain energy, and when this locally exceeds the breaking strength, a crack tip forms. The surrounding strain energy flows to the crack tip, which rapidly extends and splits the rock, releasing the strain energy as heat. The initial energy flow causes additional crack tips in highly strained areas. If the compression is rapidly applied by impact, crack tips may form before the strain energy has reached equilibrium in the particle, thus decreasing the total work input required for breakage. The energy necessary to break is essentially the energy necessary to produce crack tips, since the energy necessary to extend the cracks to breakage is already present as strain energy in the deformed particles. After breakage nearly all of this energy appears as heat. The crack length cannot be measured directly. However, in particles of regular and similar shape the crack tip length is considered as equal to the crack depth, or crack extension necessary to break, so that the crack length equals the square root of one-half of the surface area. The Third Theory states that the useful work done in crushing and grinding is directly proportional to the total length of the new cracks formed. It can be confirmed by showing that a constant work input produces a constant length of new cracks when reducing the same material to different product sizes. This is done in the present paper on a wide variety of material. The constant work input was supplied by one revolution of the 12" x 12" laboratory ball mill used in making grindability tests by the Allis-Chalmers method (12) (13) The new crack lengths produced per mill revolution were measured from all available grindability test results at 28, 35, 48, 65, and 100 mesh on fifteen different ores, and were found to remain substantially constant for each ore at all mesh sizes. A new technique is used for the measurement of crack lengths. Size distribution plots of the mill feed and product are made by the Third Theory method (9) and the crack lengths are obtained from these plots by the procedure described in the present paper. The energy input required to produce a unit length is of fundamental importance in the size reduction of brittle solids. The crack length Cr is expressed in centimeters per cubic centimeter of solid material. It bears a definite relationship to the external surface area of the crushed or ground solid. A uniform particle shape must be assumed before the surface area and crack length can be evaluated. In this paper it is assumed that the relationship between the surface area and the particle volume of a particle d microns in diameter is the same as that of cube d microns on a side. The external surface areas of particles with a cubical breakage probably agree approximately with this rule, and correction factors can be applied when physical measurements of the surface areas are available for comparison. However, the assumption of equivalent cubes has been found satisfactory for most calculation purposes. Assuming equivalent cubes, one cubic centimeter of particles d microns in diameter will have a crack length Cr of v30.000/d centimeters, and a surface area of 60,000/d square centimeters. The specific crack length is thus equal to the square root of one-half the specific surface area. Where Sa is the surface area in square centimeters per gram and Sg is the specific gravity of the ground solid, then Cr = vSg . Sa/2 = 173.2/ vd (1)

Jan 1, 1961

By C. J. Haigh

Stockpiled and currently produced zinc plant residues containing zinc ferrite are to be treated at a rate of 350 tonne per day in a plant that is being constructed at the Risdon works of the Electrolytic Zinc Company of Australasia Limited (E.Z.). Start-up of the plant is scheduled for December, 1970. The residue is to be treated by a hydrometallurgical process that was developed by E.Z. in Australia, by Det Norske Zinkkompani in Norway, and by Compagnie Royale Asturienne des Mines (Asturiana de Zinc) in Spain. The process, as it is to be used at Risdon, is simple, consisting essentially of two stages : 1. A leaching stage at atmospheric pressure in which the zinc ferrite contained in zinc plant residue is dissolved in spent electrolyte, fortified with make up sulphuric acid, at a temperature in the range 70 to 95ºC. Undissolved solids containing lead and precious metals are separated for sale to a lead smelter. 2. An iron precipitation stage at atmospheric pressure in which ferric iron is precipitated from a slightly acidic solution at 90 to 95ºC as ammonium jarosite, a crystalline double sulphate which is relatively coarse, and is easily filtered and washed free of zinc-bearing solution. Calcine is used to neutralise the acid that is liberated during precipitation of the iron. Also included in the Risdon plant are a separate stage for oxidation of ferrous iron by injection of air, and a pre-neutralisation stage, using recycled oxidation stage pulp, to neutralise excess acidity from the leaching stage before the jarosite precipitation stage. The residue treatment plant consists of conventional equipment such as covered agitated vessels, thickeners and filters. Operation will be continuous and product solution will join the main zinc plant circuit for further purifications and electrolysis. Marketable zinc production will be increased by 70 tonne per operating day by the treatment of residue.

Jan 1, 1970

By T. F. McCormick

THIS year, 1950, is the golden anniversary of the construction and operation in this country of a tube mill for the sole purpose of fabricating aluminum alloy tubing. For a short period prior to the beginning of this mill, aluminum alloy tubing was obtained from foreign sources or produced in mills fabricating other metals. Indeed, it was a very short period because aluminum itself was a newcomer in the field of metals. The first sheet rolling mill had been in operation just eight years and a rod rolling mill for the production of wire and cable had been functioning one year. The first tube mill for aluminum was built by the Aluminum Co. of America in May 1900 and started production with four draw benches and five men, including the mill superintendent. The building itself was a lean-to attached to a sheet-rolling mill at New Kensington, Pa. The cost of equipment, including the draw benches, dies, mandrels, and other tools, was about $5000. This figure seems absurd in the present day but nevertheless it was questioned at that time whether or not such a large sum should be spent to embark on a new and uncertain venture. The initial production system for the new mill consisted of casting round hollow ingots in a tilting-type iron mold, reducing the ingots to bloom size outside the plant, and finally drawing to size on the draw benches. This method of producing aluminum alloy tubing survived just two years, as blooms of satisfactory quality were not obtained consistently. It was superseded by the cupping method in which a 24-in. diam circle cut from a rolled plate was formed into a tube bloom using first a small cupping press and then a push bench. It was stated that tubing so produced was limited to a maximum diameter of 2 3/4 in. and to wall thicknesses varying from 12 to 24 gage. However, for the first time all the operations were performed at one plant and the result was a quality product. Later larger cupping presses and push benches were obtained and the method continued to be the principal one for aluminum alloy tubing for almost a quarter of a century. During the interim the use of a Alan Mannesmann -type billet piercer was explored and it was used to some extent, but failed to produce consistently a smooth inside surface suitable for drawing with aluminum alloys. Experiments started during the latter stages of World War I in the production of tube blooms with a hydraulic extrusion press, led to the use of this type of equipment to make tube blooms for further drawing during a change-over period extending from 1925 until 1930. At this time the cupping method was discarded completely. From these humble beginnings, the aluminum alloy tubing business expanded until well over six million pounds were produced in a single month in the United States during World War II. The major portion of this production was in strong alloys which did not appear in the tubing picture until 1022, when 17S alloy tubing was introduced for aircraft construction.

Jan 1, 1951

By G. H., Wigton

THE present milling operations of the Chief Consolidated Mining Co. consist- of flotation of oxidized minerals from siliceous ores of the Tintic Mining district, Utah. As the lead, silver, and gold minerals that give the ores economic value are almost entirely oxidized; their successful flotation has necessitated the solution of many new problems. Laboratory effort resulted in the selection of a, natural float for the oxidized minerals. 'By "natural float," I mean that no reagent was added to produce a sulfide film on the osidized minerals before or while subjecting them to flotation operations. As compared to methods involving sulfide filming, the natural float was selected because it promised successful treatment of a larger variety, of ores without change in quantity or kind of reagents, as well as a better economic balance between metal recovery and concentration ratio. In explanation, I can, state that with sulfide filming, the efficiency of the method decreases rapidly with increase of hydrated, iron osides in the ore, whereas with the natural float the same effect has not been noted. Evidence shows that although sulfidizing can produce greater metal recoveries, it yields a concentrate so much lower in grade that no advantage results.

Jan 1, 1928

By A. S. Odeh

This paper analyzes the effect of limited entry to flow at the wellbore on the steady-state productivity of a well. Wells that have been opened to flow along a fraction of their productive interval are termed wells with limited entry. Previous work treated the cases of a partially penetrating well, a well producing from the central portion of the productive interval and a well in which several intervals equally spaced were open to flow. In this paper the open interval can be located anywhere within the productive interval. Thus, in a sense, it generalizes previous work. The finite cosine transform was used to arrive at a solution for steady-state flow of a slightly compressible fluid. The solution was programmed for a CDC 1604 computer. Numerical vaIues for rd = 660 ft, r, = 1/4 ft, and range of sand thickness of 20 to 200 ft are presented in graphical form. The effect of rd and r, values on the result is shown in a table. The correct calculation of skin and damage ratio in the presence of limited entry to flow is explained and illustrated by examples. Moreover, the paper shows how to calculate the net decrease in productivity due to the combined effect of limited entry and perforations. INTRODUCTION In some wells only a fraction of the productive interval is open to flow. Location of this fraction is usually dictated by formation characteristics and reservoir behavior. For instance, if a gas cap exists, the open interval is located away from the gas-oil contact to prevent any possible gas coning. Wells that intentionally have been opened to flow along a fraction of their productive formation are tened wells with limited entry. Obviously, unintentional completions of this type also exist. Limited entry to flow decreases well productivity. Magnitude of the loss depends on the fraction of the formation open to flow, on the thickness of the sand, on the location of the open interval and on the ratio of rd /r, where r , is well radius and rd is the drainage radius of the well. The use of pressure buildup data on producing wells to calculate the condition of the formation around the wellbore is an accepted practice. van Everdingenl and Hurst2 introduced the concept of she skin factor s considered to be due to a thin layer of different permeability immediately around the wellbore. These authors dealt with the case of a well of complete radial geometry, i.e., a well with open-hole completion that completely penetrates the formation. The presence of a low-permeability skin results in a loss of productivity, as does limited entry. Therefore, if pressure buildup data obtained on a well with limited entry are used to establish the presence or absence of skin (i.e., formation damage), and a correction is not made for this loss of productivity, the calculations would result in an erroneous skin value. They might indicate the presence of formation damage when in reality there is none, or they might indicate a value larger than the true value. This could lead to an incorrect basis for planning remedial measures. Muskat3 studied the problem of partially penetrating wells for the case of incompressible flow. He presented equations and figures which allow estimation of loss in productivity. Brons and Marting,4 using equations based on Nisle's work,5 studied the loss of productivity for three cases. The first was for a partially penetrating well; the second was for a well producing from only the central portion of a productive interval; and the third was for a well in which several intervals equally spaced were open to flow. Their work was for steady-state depletion-type reservoirs wherein the well radius of drainage is established and the fluid is considered to be slightly compressible. Considered in this paper is the problem of wells with limited entry in which the open intervals are located anywhere within the productive sand. The finite cosine transform is used to arrive at a

Jan 1, 1969

By LEWIS F. THOMAS

MINING geologists and mining engineer, rarely give due thought to the geography of mining deposits. They realize, it is true that what may be ore in one place would be only worthless rock in another because f the in accessibility of a smelter r a market; and they may consider the lack of a supply of labor, or water, or power to carry on operations. But they do not think (of these things in terms of geography. Geography may be defined as the science of relationships between earth and man, so geographers are vitally interested in the how and why of mining as it is carried on at different place, having entirely different environment-. Such environments are geographic complexes or composites of natural and cultural elements, which are listed and briefly discussed in the following paragraphs.

Jan 1, 1941

By J. B. Newkirk

ABOUT twenty-seven years ago W. bergl discovered that interesting detail could be seen in an X-ray diffraction spot made with a rock-salt crystal if the recording photographic film were held very close to the reflecting crystal. He also showed that much of this detail could be explained in terms of the deformation structure of the specimen. In the same year, 1931, Fox and parr2 reported that X-ray reflections from a quartz crystal were much more intense while the crystal was undergoing piezoelectric oscillation than when it was at rest. This observation was soon confirmed (and explained in terms of X-ray extinction) by Barrett3 in the United States and by Nishikawa et a1.,4 in Japan. During the next few years Barrett refined Berg's method for making X-ray diffraction micrographs and applied it to the study of metallic crystals. He reported this work in the 1945 annual lecture of the AIME.5 Since 1945 relatively little use has been made of the Berg-Barrett (B-B) method for making diffraction micrographs, though considerable interest in investigating subcrystal structures by other methods has developed. Some of these methods depend upon evaporation6 or chemical etching7 and, therefore, suffer from the occasional intrusion of subtle artifacts such as those recently reported by Hooker for leada and Welsh for aluminum.9 Other methods require the introduction of foreign elements which "decorate" the substructure in some way, thereby rendering it detectable by direct" or indirect1'. visual examination. Still others depend upon complex and often highly questionable analyses of X-ray diffraction line profiles. There have also been a number of related X-ray techniques, such as the Lambot,12 Schulz13 and Merlini and Guinier,14 which have proved to be quite useful for specific applications. Some of these methods give a picture of the substructure which is more or less directly related to the shape and the topography of the specimen. The Berg-Barrett method carries this desirable feature to the limit. Closely related to the B-B method are three photographic X-ray techniques recently described by Lang.15 These methods give beautiful pictures of the internal structure of crystals but are limited by the facts that the required apparatus is rather complex, exposure times are long, and the specimen crystals must be thin. An experimental appraisal which I have recently made of the original B-B method has shown that this simple technique is not only inexpensive and convenient, but can give more information than was apparently realized by its inventors. For example, first-order subgrains in iron-silicon crystals were clearly delineated and their relative orientation subject to description. Also defect structures could be seen within the subgrains and in some materials slip bands could be clearly seen with respect to the topography of the rest of the specimen. This much had already been described by Barrett. However, in the present study it was found that the actual sites of individual dislocations which emerge upon the crystal surface could be located. Furthermore, by a simple procedure, the direction (but not sign) of the dislocation Burgers' vector could be defined as well.

Jan 1, 1960

By C. G. Goetzel

Until recently porous bronzes have found many applications for self-lubricating bearings in the automotive, electrical, household appliance and general machine industries. The bulk of an annual production of several million pounds of copper powder had gone into the manufacture of these articles, of which by now more than a billion are in service. However, the war has brought severe curtailment of these bearings because of the scarcity of the two main constituents, copper and tin. The development of substitute materials of equal or superior properties has led to a large-scale replacement by products made from iron powder. In spite of the diminished interest in powdered copper-tin alloys caused by this trend, it was believed worth while to investigate some mechanical properties of these materials as a reference base for substitutes. In addition, experiments using the hot-press method were incorporated to investigate possibilities of saving vital equipment or opening new fields of applications. It must be assumed that by this time the powder metallurgy process of making porous bronzes, a development that found its first industrial applications more than 10 years ago, is well known in engineering. The patent literature and bibliography is volu- minous, most publications referring to porous, self-lubricating bronzes, containing about 10 per cent tin and I to 5 per cent graphite, with porosities in the range of 25 to 35 per cent.l-7 The excellent antifriction properties of these bearings, caused by the uniform impregnation with oil, are dependent on the interconnection of the pores and the uniform distribution of the graphite inclusions. The sintering mechanism of porous bronze, with its structural changes during the transformation of the metal powder into the sintered bearing has been shown in colored photomicrographs by Hall.* Diffusion between copper and tin was found to begin within one minute, when tin-rich transitory phases were formed. After 25 min. a homogeneous matrix of alpha was reached for an alloy containing 9.5 per cent tin and 6 per cent graphite. Price, Williams and Garrand,9 studying certain metal powder combinations in which sintering is accompanied by the cementing action of a liquid phase, including a 90-10 bronze sintered at 925°C., have found that if the treatment is carried out within the liquidus and solidus temperature range of the system, a peculiar process of alternate solution and precipitation occurs. The particles of the major solid constituent are dissolved in the minor liquid phase, but thereafter are reprecipi-tated on certain nuclei, which develop into large grains, usually rounded.

Jan 1, 1945