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By P. Campbell, E. M. Gardy, L. Hachkowski

Introduction The efficient process control of a uranium mill and its associated mining operation requires a nearly continuous knowledge of the uranium concentration in the ore. Generally, the approach is to use laboratory assays of grab samples from the mill feed belt. In some cases, elaborate and expensive systems have been used to ensure random sampling, but even with these systems, mass-balance discrepancies still exist. There is a requirement for an on-line instrument that can give a prompt, accurate analysis of a large portion of the feed stock. The authors have recently evaluated a laboratory system that achieves these goals using neutron activation and delayed neutron counting. The development of the on-line uranium ore analyzer is a consequence of previous work done at the Whiteshell Nuclear Research Establishment (WNRE) (Campbell et al., 1978 and 1981), and is based on the emission of delayed neutrons after the neutron irradiation of fissionable materials. The mechanism of delayed neutron emission has been described (Keepin, 1965), but briefly is as follows. The fission fragments resulting from irradiation are in an excited state. Certain of these delayed fission products, precursors, decay to a more stable state by the emission of a delayed neutron. The delayed neutrons can be divided into six groups with effective half-lives ranging from 0.2-55 sec. It is important to note that more than 50% of the delayed neutrons are emitted within the first 6 sec after irradiation; this has an influence on the design of the analysis instrument. Measurement Cycle The measurement cycle involves a neutron irradiation, a delay time, and a counting period. The laboratory demonstration model of the ore analyzer comprises a californium-252 neutron source, a source chamber, a pneumatic transfer facility, and a neutron irradiation and counting module [(Fig. 1)]. The ore sample is placed in the irradiation chamber. The neutron source is transferred to the center of the chamber for a 5-sec irradiation while the helium-3 thermal neutron detectors are switched off. As the source returns to the storage chamber, the detectors are switched on and the delayed neutrons counted. The total assay cycle is 16 sec. The current instrument design calls for the ore to be cut from the main belt every 16 sec by a mechanical arm and dropped into the analyzer directly, or via another belt. The cycle starts when there is a sufficient mass of ore in the irradiation chamber to depress a switch and stops when the delayed neutron count is finished. The ore is then dropped through the bottom of the analyzer chamber and returned to the main belt. The counts from each cycle are accumulated and the data reduced to give the desired information.

Jan 1, 1985

By L. Guttman

LESTER GUTTMAN* The question of whether classical thermodynamics alone imposes any lower limit to solid solubilities was raised during a discussion among various members of the Institute for the Study of Metals. Although our conclusions are not new,1 they may be worth reiterating, since confusion on this subject seems to persist. In a binary system (A,B), the conditions for equilibrium, at constant pressure and temperature, between a solid solution and a liquid solution are µAs = µAl [la] µBs = µbl [lb] Here µ denotes the chemical potential either of component A or B, as shown by the subscripts, and the superscripts s and l refer to the solid and liquid solutions, respectively. These two conditions are just sufficient to determine the compositions of both phases. Suppose, however, that the solid phase is found experimentally not to contain any of component B; then we cannot evaluate µbs, which is defined by Here nAs and nBs are the number of mols of A and B, respectively, in the solid, whose free energy is Fs. Eq lb cannot be used, but it is no longer needed: we stated at the outset that the solid has the composition "pure A," and only the liquid composition need now be determined, from Eq la. All the equilibrium properties of the system can be derived from Eq la rigorously and simply. Hence there is no purely thermodynamic reason to exclude phase diagrams which show precisely zero solid solubility. Marsh2 has objected to such diagrams as implying that even the transformation of pure A begins at the eutectic temperature. In any case, the user of the diagram must have a certain minimum knowledge of how to interpret it, and it would seem simpler to remember the general rule that eutectic temperatures have no significance for pure components, rather than to show, say, by broken lines, on a scale which will often be exaggerated, a hypothetical solid-solubility about which we know only that it is less than a certain amount. The case of an intermetallic compound in equilibrium with a binary liquid solution does not require a separate treatment since nothing in the foregoing was dependent on the nature of the components. Therefore, thermodynamics alone does not exclude the possibility that (intermetallic) compounds may exist with compositions always precisely those given by their chemical formulas. With the aid of extra-thermodynamic methods, one may conclude that the solubility of one solid in another is never exactly zero. On the other hand, one would generally expect that the solubility will be small unless the components resemble one another (as do the metals), or form weakly-bound crystals. Therefore, it is useful to know that when there is no direct evidence for solid-solubility, one may neglect it without violating thermodynamic principles.

Jan 1, 1950

By R. Schuhmann

PROBLEMS in expressing, interpreting, and using size-distribution data recur in many phases of mineral dressing; therefore it is desirable that size analyses be expressed in such a form, either numerically or graphically, that comparisons can be made readily. The size distribution of comminuted material is a result of a specific physical operation on the material and therefore should be governed by some definite physical principle, or cause-effect relation. Although this principle has not been discovered, it has been found that the size distribution of comminuted, homogeneous solids may be expressed by relatively simple mathematical equations. Two equations in particular, proposed by Gaudin1 and by Rosin and Rammler,2 respectively, have stood out, both having wide demonstrated ranges of application. The form and the method of application of the two appear almost irreconcilably different on the surface, but, as will be shown, they are asymptotic in the fine sizes when expressed in the same units. This paper gives the results of a further study of the Gaudin and Rosin-Rammler relations, made with the object of developing a form of size-distribution equation that will be more usable and more significant-particularly in comminution studies, both for interpreting the nature of size distribution and for evaluating surface. The work is summarized in the following: I. An equation relating cumulative per cent finer to particle size in the fine sizes is derived from Gaudin's original equation. That equation relates weight per cent retained on one size and through the next larger size in a geometric series of sizes to size (of the retaining screen, for example). 2. Comparisons with Gaudin's equation and with the Rosin-Rammler equation show that the cumulative equation applies to the fine sizes of all size distributions that fit either of these earlier relations, and, furthermore, has some advantages over each. 3. A simple graphical method of surface calculation based on the cumulative-equation plot is evolved and illustrated. 4. Additional confirmation of the cumulative equation was obtained in the more severe test of sizing down to the limit of a sedimentation balance a sample of jaw-crushed quartz. The equation was closely followed over a 20o-fold range in size, from approximately 0.4 mm. (35 mesh) down to 0.0023 mm. (2.3 microns), the lower limit of the size analysis made. There is no indication that the equation will not hold further to well below 2.3 microns. 5. Calculations of the surface produced in the jaw-crushing test mentioned above showed that at least 8o per cent of the total new surface was on particles finer than 35 mesh (about ½ 5 of the crusher setting). The minus 35-mesh portion was only

Jan 1, 1940

By Fred C. Bond

THE process of comminution by grinding is properly classified as an art, rather than as a science. Like most other operations concerned in ore dressing, or in the treatment of nonmetallic minerals, the mathematical relationships of rock breaking are still obscure, and much disputed. If some basic formula exists, comparable for instance to E = RI in electricity, it has not yet been placed in, common use. The absence of such a comprehensive mathematical background undoubtedly is the cause of many inefficient designs and installations, as well as of much poorly oriented experimental work. On the other hand, the mathematical immaturity of the process constitutes an opportunity for a real contribution on the part of those who enjoy mental pioneering. Some of the statements made in any such exploratory work may be little more than opinions, and it is to be expected that some will later be proved untrue; but such an evolutionary development is necessary if we are ever to arrive at the fundamental laws, the discovery of which can change the empirical art into a mathematically exact science. The most significant principle of comminution that we possess is probably Rittinger's law, which states that the work done in crushing or grinding is proportional to the area of the new surface produced. There is an alternative theory, known as Kick's law, which formerly was preferred to Rittinger's but which now is rather discredited. According to Kick, the work done in comminution is proportional to the volume, or weight, of the pieces broken, or to the cube of their diameters, while according to Rittinger it is proportional to the square of the diameters. Careful determinations of size distribution below the limit of sieve sizes1 indicate strongly that under ordinary grinding conditions the amount of new surface area produced is directly proportional to the net power input. Changes in the size ratio between the balls and the particles cause important differences in the proportion of the product that passes a given mesh size, but do not affect the total amount of new surface area; they merely change the distribution of the surface among mesh sizes. The principle of constant surface area is the logical result of Rittinger's law, and it may well have an important bearing on the comminution theories of the future. FORMULA FOR BALL WEAR Grinding tests have been made in the Allis-Chalmers laboratory for more than 9 years, using a standard charge of mixed balls of different sizes; and a continuous record has been kept of the rates of wear of each size. These results have been combined and plotted in various ways in order to discover the relationship that exists

Jan 1, 1940

By F. S. Norcross

MANGANESE has long been considered one of the United States' most important strategic raw materials. Its indispensability in steel manufacture makes it vital to the nation's industrial life. Coupled with this indispensability is the fact that domestic producers have not been able as yet to supply more than a small part of the nation's manganese needs; consequently the United States has had to depend upon such distant foreign sources as the U.S.S.R., British India, the African Gold Coast and Brazil for the major portion of its ore. The great bulk of the mineral is consumed in the form of ferromanganese, an alloy that averages 8o per cent manganese and 20 per cent mainly iron and carbon. Manganese is used as a deoxidizer and desulphurizer in steel manufacture. By combination with the residual oxygen and sulphur of the bath, it helps to produce a clean, sound metal. Approximately 14 lb. of manganese, it is estimated, is used in making one ton of steel. For standard-grade ferromanganese, the ore should have a minimum of about 48 per cent Mn and preferably more. Beneficiation methods have not yet been developed for making the immense low-grade deposits in the United States salable at prevailing prices. Dependence upon foreign sources thrust the United States into serious difficulties during the World War. At the start of hostilities in 1914, the nation was obtaining about four-fifths of its ore from the Eastern Hemisphere. Domestic production had been negligible, amounting to only 3000 to 7000 tons a year. As the war continued, shipments from the East, endangered by sea attack and blockade, dwindled rapidly. Imports from Brazil mounted. In this grave emergency, the United States, starting almost from scratch and under the stimulus of high prices, succeeded by 1918 in producing a peak of 311,000 tons, about 3 5 per cent of its consumption that year. Since the war, the United States has depended upon foreign sources for more than nine-tenths of its high-grade ore. During 1936, 1937 and 1938, imports for consumption averaged 736,000 tons a year. During the same period, domestic shipments averaged 21,000 tons a year. Of the imports, the U.S.S.R. accounted for an annual average of 280,000 tons, the Gold Coast 208,000 tons, India 74,000 tons, and Brazil 72,500 tons. But India and Brazil, which in 1936 ranked third and fourth in the list of importers, were displaced in 1937 and 1938 by a new and much nearer source-Cuba. From less than 1 per cent of United States imports in 1931, the island republic increased its shipments to more than 27 per cent in 1938. Cuba's average for this three-year period was 97,000 tons, and its total for 1938 was 131,000 tons.

Jan 1, 1940

By Terkel Rosenqvist

Few subjects in iron metallurgy and steelmaking have caused so much discussion and have resulted in so many divergent and partly contradictory statements as the theory of the distribution of sulphur between molten iron and slag. An interesting survey of the different points of view is given by Grant and Chipman.1 Grant and Chipman themselves conclude in their paper that, at least in the open hearth and electric steel furnaces, the governing reaction is a simple distribution like: FeS steel ? FeS alas [Ia] and MnS steel ? MnS alas [Ib] They conclude that no significant reaction takes place according to the more classical scheme: FeS steel + CaO slag ? FeOslag + CaS slag [2a] This is the reaction which is regarded by most investigators as the primary sulphur reaction in the blast furnaces, where it is accompanied by the reaction with carbon: FeO slag + C ? Fe + CO [2b] The writer suggests an alternative to the reactions mentioned above. It is the opinion of the writer that the a priori assumption of defined molecular compounds being present in molten steel as well as in molten slag, has retarded rather than promoted a solution of the sulphur problem. As for the constituents of molten steel, it is taking too much for granted to mention the presence of FeS and MnS. If this infers that the melt contains diatomic molecules, the statement must be made that at least such molecules do not exist in the solid state at room temperature. Solid FeS crystallizes in the nickel-arsenide structure, and MnS (high temperature form) in the sodium-chloride structure. In these structures the sulphur is surrounded by six metal neighbors, and no molecular groups can be isolated. When sulphur is dissolved in molten steel, it is expected to maintain the same coordination number as in the solid state. We know that sulphur in steel has a certain chemical potential and a chemical activity which can be expressed by the reaction: S steel + H2 ? H2S where the activity A. is proportional to the ratio PH2S/Ph2. We know also that this activity is influenced by the presence of other constituents. For instance, there is reason to assume that carbon and silicon increase the sulphur activity and manganese decreases it. However, the exact relationship here is scarcely known, and further investigations are needed. Concerning the molten slags, if anything should be assumed a priori it would be that they are composed of ionic rather than molecular parts. There are indications of this in that solid silicates form ionic structures, and molten slags show high electric conductivity and can be electrolyzed.2 It has been claimed by Taylor and Chipman3 that their activity investigations

Jan 1, 1949

By W. G. Pfann

THE problem of forming points on wires that were from 0.002 in. to 0.010 in. in diam arose in the recent appearance of silicon and germanium point contact rectifiers as elements in microwave radar.1 In these devices a pointed tungsten wire makes a delicate contact with the surface of a semiconductor. Because the quality of such rectifiers is largely determined by the nature of this contact, considerable attention has been given to the formation of the metal point. Reproducibility of contour and freedom from burrs and even microscopic irregularities are considered essential. The operations involved must be simple and rapid enough for mass production. Previously tungsten wires had been provided with conical points by a grinding method which involved rotating a wire about its axis and causing it to bear at an oblique angle against a spinning abrasive disc. This was later complemented with an electrolytic polishing treatment which removed irregularities from the ground point and also rounded it off at the tip-a desirable feature. Finally, these operations were supplanted by the electrolytic pointing method described here. MECHANISM OF POINT FORMATION The wires to be pointed are partially immersed- in an electrolyte in a .vertical position whereupon they are provided with smooth rounded points entirely by electrochemical action. The process is independent of the depths to which the wires penetrate the electrolyte because the points are formed at the surface level. It may be used to point a single wire or many of them simultaneously and is applicable to wires of different diameters within a limited range. Solution of the immersed portion of a tungsten wire is accomplished by making it the positive electrode in an electrolytic cell, as shown in Fig I. While the wire is maintained at constant voltage the current through the cell falls continuously, principally because of the decrease in wetted area which results from solution of the tungsten. This increase in resistance occurs in a reproducible manner and is used as a control in the process. As the immersed portion of the wire dissolves, it remains cylindrical in shape except for the short section which is wetted by the meniscus. Here the rate of solution is lower and decreases as the top of the meniscus is approached. The result is the formation of a tapered section inside the meniscus. As solution continues, the cylindrical section either dissolves entirely or drops off just before complete solution, leaving only the taper which is now sharply pointed. Further action causes this point to become rounded and then quite blunt, until finally the end of the wire is almost plane. Fig 2 shows the stages in this development. Contour PF is most representative of those used in point contact rectifiers. Thus a variety of point

Jan 1, 1947

By Terkel Rosenqvist

Few subjects in iron metallurgy and steelmaking have caused so much discussion and have resulted in so many divergent and partly contradictory statements as the theory of the distribution of sulphur between molten iron and slag. An interesting survey of the different points of view is given by Grant and Chipman.1 Grant and Chipman themselves conclude in their paper that, at least in the open hearth and electric steel furnaces, the governing reaction is a simple distribution like: FeS steel ? FeS alas [Ia] and MnS steel ? MnS alas [Ib] They conclude that no significant reaction takes place according to the more classical scheme: FeS steel + CaO slag ? FeOslag + CaS slag [2a] This is the reaction which is regarded by most investigators as the primary sulphur reaction in the blast furnaces, where it is accompanied by the reaction with carbon: FeO slag + C ? Fe + CO [2b] The writer suggests an alternative to the reactions mentioned above. It is the opinion of the writer that the a priori assumption of defined molecular compounds being present in molten steel as well as in molten slag, has retarded rather than promoted a solution of the sulphur problem. As for the constituents of molten steel, it is taking too much for granted to mention the presence of FeS and MnS. If this infers that the melt contains diatomic molecules, the statement must be made that at least such molecules do not exist in the solid state at room temperature. Solid FeS crystallizes in the nickel-arsenide structure, and MnS (high temperature form) in the sodium-chloride structure. In these structures the sulphur is surrounded by six metal neighbors, and no molecular groups can be isolated. When sulphur is dissolved in molten steel, it is expected to maintain the same coordination number as in the solid state. We know that sulphur in steel has a certain chemical potential and a chemical activity which can be expressed by the reaction: S steel + H2 ? H2S where the activity A. is proportional to the ratio PH2S/Ph2. We know also that this activity is influenced by the presence of other constituents. For instance, there is reason to assume that carbon and silicon increase the sulphur activity and manganese decreases it. However, the exact relationship here is scarcely known, and further investigations are needed. Concerning the molten slags, if anything should be assumed a priori it would be that they are composed of ionic rather than molecular parts. There are indications of this in that solid silicates form ionic structures, and molten slags show high electric conductivity and can be electrolyzed.2 It has been claimed by Taylor and Chipman3 that their activity investigations

Jan 1, 1949

By E. H. Crabtree, T. C. King

THE ferrosilicon grinding unit at the Central Mill of the Eagle-Picher Mining & Smelting Co. near Picher, Okla., was completed in March 1947. The object of the plant was to grind pigs of ferrosilicon, having a size of approximately 1 ½ by 4 by 12 in. and an analysis of 85 pct Fe and 15 pct Si, to approximately 100 mesh by crushing and wet grinding in a ball-mill classifier circuit. The minus-loo mesh ferrosilicon had formerly been purchased, and was used as medium in the heavy media process for treatment of lead-zinc ores. From the standpoint of dollars and cents, the cost of minus-100 mesh ferrosilicon was roughly $40.00 per ton more than pigs of the corresponding alloy delivered to the mill. Estimated wet grinding costs were considerably below this figure. The possible saving in cost of make-up medium for the cone plant was particularly attractive since, at that time, the plant losses averaged over four tons of minus-100 mesh ferrosilicon per day. Aside from the probable saving on the grinding charge, the need for such a unit was twofold. First, periodic fluctuations in the differential between the top and bottom densities of the cone were thought to be detrimental to metallurgy. There was little doubt that an intermediate differential existed at which the heavy media separation was most efficient, representing a compromise between concentrate grade and tailing loss. Furthermore, if a finer, but considerably cleaner, medium could be obtained, there was the possibility that the differential could be decreased without increasing the viscosity, or conversely, the viscosity of the medium decreased without increasing the differential. Preliminary to experimental work aimed at pegging the cone differential and circulating medium viscosity at whatever values were found to be most efficient, it was essential to have control over both the non-magnetic content of the medium and the particle size of the contained ferrosilicon. The latter was possible by means of a grinding unit at the mill. Second, whereas the ferrosilicon alloys are made by a number of manufacturers, only two, to our knowledge, grind it in any quantity for use in heavy media plants. For all practical purposes, the Central Mill was limited to one source of supply, a particularly vulnerable position. In case the supply was cut off, it was thought that pigs of the 85 pct Fe, 15 pct Si alloy could be obtained elsewhere, providing means of grinding them was available. There was no published information available covering wet grinding of ferrosilicon. Dry grinding in batch mills and in continuous mills with air classification was being practiced, but the cost of grinding was excessive and the dust condition objectionable. It appeared that primary and secondary crushing was no particular problem, providing heavy duty equipment was used;

Jan 1, 1947

By H. W. Hitzrot

SEPARATING a mixture of particle sizes of material suspended in a liquid medium is by no means an exact science. Selecting machines for individual classifying operations is even more difficult. The plant operator's own background is of course invaluable, and considerable help may be obtained from technical articles, talks with sales engineers, and handbooks on ore dressing. These several sources of information, however, are difficult to marshal in proper perspective for the major decision on classification units that an operator may be called upon to make. To the present writer's knowledge this assembly of facts is not available in handbooks, and technical papers are scarce on classification equipment developed in the past five or six years. It is believed that this paper will be helpful to users of classification equipment at this particular period in the development of hydro-classification. For easy reference the following classification units now available and in use in metallurgical and industrial operations are listed below, with a brief description of each. Unit-type classifiers, bowl classifiers, and bowl desiltors are all rectangular tanks, slightly tilted, with reciprocating rakes or screws to remove settled sands. The unit-type classifier, Figs. 1 and 2, is available in widths from 14 in. to 20 ft and lengths up to 40 ft. The shallow bowl of the bowl classifier, Fig. 3, equipped with rotating rakes, is superimposed on the lower end of the tank. Reciprocating rake compartments for this design range from 18 in. to 20 ft wide. Bowl diameters vary from 4 to 28 ft. The flat-bottomed bowl of the desiltor, Fig. 4, of relatively large diameter, is equipped with rakes rotating outward and partly over a pit, which is created by extension of the rectangular tank into and under the bowl section. Bowl desiltors are available with reciprocating rake sections 4 to 20 ft wide and bowls from 20 to 50 ft in diam. The bowl desiltor is used for applications beyond the range of the bowl classifier. The hydroseparator, Figs. 5 and 6, is a circular tank equipped with slowly rotating rake arms, set on a slope, with interrupted rake or spiral blades to move the settled solids to a central discharge cone. Tank diameters vary from 4 to 250 ft. Tank depths at center are 2 to 3 ft for small units and up to 25 ft for larger units. Hydraulic classifiers of the sizer and super-sorter types, Figs. 7 and 8, are narrow, deep, rectangular tanks divided by vertical baffles into a series of pockets. Hydraulic water is added near the bottom of each pocket. Perforated constriction plates, spiral flow arrangements, or jets are used to disseminate the water under pressure (hydraulic water) throughout the bed of material in the pocket. Discharge valves on each pocket are operated automatically by a pneumatic mechanism, a pincer-type mechanism, or a pressure control and motor combination actuated by a hydrostatic tube within the pocket. Hydraulic classifiers are available in 4, 5, 6, and 8-pocket units of varying constriction plate areas to suit conditions. There is now a jet sizer of unit pocket design that can be made up in 1 to 25 sections or more to accommodate sizing requirements. The hydroscillator, Fig. 9, is a rectangular tank set on a slope of 3 to 4 in. per ft. A bowl is superimposed on the lower end. The bowl bottom is an oscillating rubber-covered disk, perforated to allow hydraulic water introduced beneath the disk to set up a teeter bed and thus produce an oversize or rake product exceptionally free of slimes, and material minus the mesh of separation. A shallow dam at the periphery of the bowl allows the coarse or oversize fraction to spill over and drop down into the tank compartment, where it is moved up the deck by reciprocating rakes. The material, minus the mesh of separation, overflows a circular and stationary weir which is several inches higher than the dam on the oscillating disk. It is carried off in a circular launder in the usual manner. These units are available in bowl diameters from 4 to 14 ft and with reciprocating rake compartment widths to suit the tons per hour to be handled. Centrifugal classifiers include the solid bowl centrifuge and the cyclone classifier. The solid bowl centrifuge, Fig. 10, consists of a truncated cone fixed to a horizontal shaft and rotating at high speed. An internal spiral rotating at slightly less speed continuously removes solids deposited on the inner surface of the cone, or bowl. Feed enters the cone by means of the hollow center shaft. Overflow leaves through ports at the large end of the cone, and oversize solids, moved by the spiral, exit through ports at the small end. Centrifugal classifiers of this type are available in cone diameters from 18 to 54 in. The cyclone classifier, Fig. 11, is a stationary cone having a cylindrical upper section and a lower cone section. Feed is introduced tangentially into the upper cylindrical section under pressure from a hydrostatic head or by means of a pump. Centrifugal force thus induced effects a classification within the cone, the fine sizes being carried off in the overflow through an opening at the top of the cylindrical section. Coarse solids at relatively high pulp density exit through a control valve at the apex of the lower cone section. Cyclone classifiers are available in 3, 6, 12, 14, 24, and 30-in. diam. The cone classifier, Fig. 12, is a steel-plate cone with sides usually about 60" from horizontal. It contains no rotating mechanism. Feed enters a feed-well at center and classification is effected by gravity and pulp density. Fines are carried off in the flow over a peripheral weir at the top of the cone. Settled solids exit through an opening at the apex. An apex valve actuated through levers and rods by the pulp density in the lower cone section is usually supplied. Diameters are usually maximum at 8 ft.

Jan 1, 1955

By D. C. Hilty, W. Crafts

L. S. Darken—Laboratory investigation of deoxidizing and other steelmaking reactions is usually centered, at least first, on the determination of the equilibrium or equilibria involved. This seems a reasonable procedure since equilibrium, if attained, depends only on composition, temperature, and pressure; hence conclusions derived from data on small experimental quantities are applicable to a heat of steel providing eauilibrium is attained in both cases. A knowledge of equilibrium serves as a useful framework even though we may know that practical conditions do not correspond to complete equilibrium. On the other hand, nonequilibrium or rate phenomena depend on a wider variety of conditions and are more difficult to interpret; conclusions applicable to laboratory conditions may or may not apply to larger scale phenomena. Hence the attainment or nonattainment of true equilibrium in the experiments here reported is of critical importance in evaluating their significance. Since some of the statements in this paper and in the closely related preceding one (on aluminum deoxidation) imply some doubt on this matter, I should first like to ask the authors whether their conditions are intended and believed to represent equilibrium. I should like to point out three considerations which seem to cast considerable doubt on the achievement of equilibrium, at least of the particular equilibrium under consideration. 1. In the experiments on manganese deoxidation the authors point out that they could not maintain the manganese-oxide slag on top of the metal in their rotating crucible, and hence they substantially dispensed with this slag. This leads to serious trouble in the interpretation of the results, for any equilibrium is, of course, a particular specific equilibrium—in this case Mn + O = MnO The experimental deletion of the upper layer of manganese oxide means that if equilibrium is attained at all it is attained between the metal and the MnO which has soaked into or adhered to the crucible (under the metal) and has dissolved substantial amounts of the crucible material including impurities. These impurities may constitute a significant portion of the slag by virtue of the small total amount of slag, even though the crucible is relatively pure. Hence there would seem to be a strong presumption that the equilibrium (if attained) involves not a pure MnO (or MnO — FeO) slag but one saturated with alumina and containing perhaps considerable impurities which substantially lower the concentration and activity of MnO, causing the above reaction to proceed to the right further than it would in the absence of alumina and impurities. Hence it is not surprising that manganese here appears as a better deoxidizer than found by other investigators. The present results may represent equilibrium with a slag of unknown composition which seems unlikely to be particularly related to plant experience. 2. The curves representing the observed silicon deoxidation (figs. 3, 4, and 5) are all drawn with a discontinuity in slope at about 0.02 pct oxygen. This point is interpreted as corresponding to the three-phase equilibrium, metal, slag, solid silica. The type of construction shown in these figures (though apparently fitting the data) is contrary to a fundamental principle of heterogeneous equilibrium as pertains to the construction of phase diagrams. According to this principle, the two solubility curves (each of the two portions of the curves in figs. 3, 4, and 5) must intersect in such manner that their (metastable) extensions must lie outside the homogeneous field rather than inside as in these figures. In other words, the "point" in these curves should be aimed in the opposite direction, if it is to be interpreted as corresponding to the three-phase equilibrium. The construction adopted is in violation of the second law of thermodynamics. This matter is discussed in detail in several texts and also by Lipson and Wilson.&apos;" The same criticism applies to the later figures representing conditions for manganese additions. The occurrence of this discontinuity or break at 0.02 pct oxygen casts further doubt on its interpretation. The earlier investigation of this system by Korber and Oelsen is in substantial agreement with the several recent findings of Chipman and coworkers that the oxygen content of iron in equilibrium with silica and silica-saturated iron oxide slag is about one third to one half that (0.24 pct at 1600") of iron saturated with pure iron oxide; thus there seems reliable evidence that iron saturated with silica and iron silicate slag at 1600" contains about 0.1 pct oxygen, or certainly much more than the 0.02 pct proposed in this paper. 3. In the quarternary system iron-silicon-manganese-oxygen one of the equilibria involved may be written 2 Mn + SiO2 (solid) = 2 MnO <slag> + Si The activity of SiO, is constant (if equilibrium is attained) by virtue of its presence as a substantially pure solid. At not too low metallic manganese content, the activity of MnO in the slag is constant by virtue of the fact that the slag is substantially pure manganese silicate saturated with silica and hence of constant composition. Thus the equilibrium constant for the above reaction is asi/a2Mn. Barring unanticipated large changes in the activity coefficients, the equilibrium constant may be adequately approximated for the composition range covered as [% Si]/[% Mn]2. Thus a plot of log [% Mn] against log [% Si] would be expected to be linear with a slope of one half as found by Kijrber and Oelsen. In the present investigation the slope (shown in fig. 15) is found to be one. It is difficult to believe that this finding represents a correct equilibrium determination, since it is at odds both with prior experimental investigation and with. theory. In view of the above points it seems that, although this paper reports many interesting findings, there is room for considerable skepticism as to the attainment of equilibrium and as to the conclusions drawn. N. A. Gokcen—The authors consider that Si% x O2% product is constant. This product is a function of the asi X a0 activity of Si02. The true constant is -------------. If the asio2 slags of this investigation were always saturated with SiO2 then Si% X O2% product would be constant,

Jan 1, 1951

By R. F. Mehl, C. E. Birchenall

R. E. Hoffman and D. Turnbull—The authors have presented evidence which they have interpreted as indicating that the rate of self diffusion is not intrinsically more rapid at grain boundaries than within the grain. Grain-size effects which apparently exist are attributed rather to impurities concentrated at the grain boundaries. In view of our own experiments and the existing evidence, we believe that the support for this hypothesis is not convincing. We have in progress an investigation in which the rate of self diffusion of silver is being measured over an extended temperature range in both single-crystal and polycrystalline specimens. The results of the single-crystal experiments and some preliminary data on fine-grained polycrystalline specimens have already been reported:&apos; and it is anticipated that a complete report will be published in the near future. The self-diffusion coefficient of large-grained polycrystalline silver (1 grain per sq mm) has previously been measured by Johnson" between 730" and 940°C. The diffusion coefficients which we have measured in single crystals (210 plane normal to diffusion direction) agree within experimental error with values calculated from an extrapolation of Johnson&apos;s curve down to temperatures as low as 500 °C. However, it has been demonstrated that the overall self-diffusion rate in fine-grained polycrystalline specimens (initial grain size of 0.003 cm) becomes measurably larger than the overall rate in a single crystal at a temperature of 600°C, and the discrepancy between the two rates becomes greater as the temperature is further decreased. In fact, it has been possible to obtain satisfactory penetration curves for polycrystalline specimens using the sectioning technique at temperatures as low as 400°C. At this temperature, the penetration is 50 to 100 times greater in the polycrystalline specimens than in a single crystal. Fisher" has developed an analysis whereby the ratio of the rate of the unit diffusion process at the grain boundary to the corresponding rate within the grain can be calculated from the penetration curves and an assumption as to the width of a grain boundary. This analysis applied to our data indicates that the unit process at the grain boundary is faster by a factor of 10&apos; at 475°C when the grain boundary width is taken to be 5. The silver used in most of these experiments was obtained from the Handy and Harmon Co. and listed as 99.97 pct pure. Preliminary experiments on 99.999 pct silver from the Jarrel-Asch Co. indicate a grain-size effect of the same order of magnitude as in the less pure silver. Nominally, these purities are as good, at least, as that of the carbonyl iron used by the authors, but of course if an impurity effect does exist its magnitude might be very dependent upon the nature of both major and minor constituents. The authors have cited the work of other investigators who have found no grain-size effects. Neither Steigman, Shockley and Nix" nor Maier and Nelson8 were able to correlate self-diffusion coefficients of copper with grain size. However, all their measurements were performed at or above 750°C; and on the basis of our work with silver, no grain-size effect would be expected at temperatures above about 0.7 of the absolute melting temperature unless the grain size were exceedingly small. Likewise, in the investigation of the self diffusion of lead by Seith and Keil,14 the lowest temperature at which the diffusion coefficient was measured in polycrystalline specimens was 207°C, which is still sufficiently high so that the lack of a grain-size effect is not surprising. Finally, in those experiments on iron from which they concluded that there was no grain size effect, Drs. Birchenall and Mehl seem to have no information as to the actual grain sizes immediately prior to and following the diffusion anneal. Without this information, we believe that their own experiments offer little support for their hypothesis. F. S. Buffington, I. D. Bakalar, and M. Cohen—The results given in this paper agree in order of magnitude with those tentatively reported by us.27 However, significant differences exist in the two sets of data, and it may be well to make an explicit comparison. The diffusion studies at M.I.T. were conducted on somewhat higher purity iron (99.98 pct Fe) than the grades used by the authors, but this is undoubtedly not the answer. Fig. 4 shows the diffusion results of both laboratories for the gamma phase, omitting the authors&apos; data on the commercial steels, while fig. 5 presents a similar comparison for the alpha phase. The divergence is much more marked in the latter case than in the former. In connection with the M.I.T. determinations, all of the runs in the gamma range and those above 800 °C in the alpha range were conducted with specimens having a relatively thick (0.002 cm) electrodeposit of radioactive iron. This practice minimizes any possible error due to extraneous diffusion that may occur during the heating to and cooling from the operating temperature. An exact solution of Fick&apos;s law for these boundary conditions was used in calculating the diffusion coefficients. At a later time, three runs were made below 800°C, using very thin electrodeposits similar to those of the authors, and the points fell considerably below the values expected from the extrapolation of the results based on the specimens with the thick deposits (compare dash-dot line in fig. 5). However, in the runs with the thin deposits, deviations of 100 pct were found between the individual specimens, whereas the maximum deviation with the thick deposits was less than 25 pct. Accordingly, it is not known at the moment whether the M.I.T. points below 800 °C should be given as much weight as those above 800°C. If this were done, the frequency factor would be of the order of 400 cm2 per sec, which is quite high. In other metals, the frequency factor for self diffusion lies between about 0.1 and 10 cm2 per sec. As the points below

Jan 1, 1951

By Michael Perch, Charles C. Russell

Nearly 7 5 per cent of the total coke production in the United States in 1940 was consumed in blast furnaces. In 1939 the percentage was 69.9, and in 1938 it was 61.3. To produce a net ton of pig iron 1757 lb. of coke was required in 1940 and 1760 lb. in 1939. These figures indicate how dependent the production of pig iron is upon the production of coke. The rate of iron production is in some measure influenced by the physical and chemical characteristics of the coke. Blast-furnace operators believe that the physical properties of coke profoundly influence the operation of the furnace. Although it is true that coke of a wide variety of physical properties is used in American blast furnaces, the periodic fluctuation of these properties is one of the causes of irregularity of blast-furnace operation. The purpose of this paper is to present a broad picture of the physical properties of by-product coke that are generally considered to be important by coke-oven and blast-furnace operators. Effects of the kind of coal used, the preparation of the coal, the rate of coking, and other factors will be considered. Of all these variables, the kind of coal used is by far the most important. In this discussion use will be made of the method of coal classification by rank1 standardized by the American Society for Testing Materials. Although this method of classification has been an American standard for several years, it has not been given the attention it deserves, especially in connection with the selection of coal for production of coke. Despite the fact that many producers of blast-furnace coke are limited to some particular source or sources of coal, a good understanding of the principles involved in setting up the standard method of classification may lead to a better appreciation of the causes of changes in coke characteristics. Furthermore, where it is necessary to select new coals as a substitute for the original supply, this method of classification can be of great assistance. The data presented have been gathered from many sources, but no attempt has been made to cover the vast literature. Instead, some of the more important data have been selected to show how coals and their carbonization affect some of the physical properties of coke. It must be emphasized that generalizations concerning the behavior of coking coals are presented and that individual coals may vary widely from the specific behavior depicted. Nevertheless, it is believed that the data illustrate certain typical behaviors. Neither was it possible to cover completely all the factors involved; for example, the macroscopic constituents of coal are believed to have important effects on the physical properties of coke, but since there is little information available for such study, and since vitrain, clarain, and durain are neither

Jan 1, 1943

By Michael Perch, Charles C. Russell

Nearly 7 5 per cent of the total coke production in the United States in 1940 was consumed in blast furnaces. In 1939 the percentage was 69.9, and in 1938 it was 61.3. To produce a net ton of pig iron 1757 lb. of coke was required in 1940 and 1760 lb. in 1939. These figures indicate how dependent the production of pig iron is upon the production of coke. The rate of iron production is in some measure influenced by the physical and chemical characteristics of the coke. Blast-furnace operators believe that the physical properties of coke profoundly influence the operation of the furnace. Although it is true that coke of a wide variety of physical properties is used in American blast furnaces, the periodic fluctuation of these properties is one of the causes of irregularity of blast-furnace operation. The purpose of this paper is to present a broad picture of the physical properties of by-product coke that are generally considered to be important by coke-oven and blast-furnace operators. Effects of the kind of coal used, the preparation of the coal, the rate of coking, and other factors will be considered. Of all these variables, the kind of coal used is by far the most important. In this discussion use will be made of the method of coal classification by rank1 standardized by the American Society for Testing Materials. Although this method of classification has been an American standard for several years, it has not been given the attention it deserves, especially in connection with the selection of coal for production of coke. Despite the fact that many producers of blast-furnace coke are limited to some particular source or sources of coal, a good understanding of the principles involved in setting up the standard method of classification may lead to a better appreciation of the causes of changes in coke characteristics. Furthermore, where it is necessary to select new coals as a substitute for the original supply, this method of classification can be of great assistance. The data presented have been gathered from many sources, but no attempt has been made to cover the vast literature. Instead, some of the more important data have been selected to show how coals and their carbonization affect some of the physical properties of coke. It must be emphasized that generalizations concerning the behavior of coking coals are presented and that individual coals may vary widely from the specific behavior depicted. Nevertheless, it is believed that the data illustrate certain typical behaviors. Neither was it possible to cover completely all the factors involved; for example, the macroscopic constituents of coal are believed to have important effects on the physical properties of coke, but since there is little information available for such study, and since vitrain, clarain, and durain are neither

Jan 1, 1943

By C. W. Volek, G. Mandl

Steam-injection tests in the field have shoum that heat transport into the oil/water region, ahead of the steam zone, may have a significant effect on the production process. Earlier theoretical work on reservoir heating by hot-fluid injection can be used to describe the growth of the steam zone, but only if heat transport from steam zone into oil/water region is neglected. It therefore became necessary to re-examine and extend the earlier theory. It was found that this theory ceases to be consistent with the physical model of the process at a certain critical time, tc, which depends on reservoir thickness, temperature, and quality of the steam. The critical time marks an important change in the heat flow across the condensation frovt: the heat flow, which is purely conductive during 0 < t < tc&apos; becomes predominantly convective at t = tc. Accordingly, at t = tc, the equation which governs expansion of the steam zone changes. On the basis of the new equation, an approximate description for the steam-zone growth during t > tc is developed by making use of upper and lower bounds for the exact solution of the problem. The boundary conditions for heat and mass flow inside the oil/water region are established, and a simple method of detennining the saturation at the downstream side of the condensation front is presented. Experimental results show that the theory gives an accurate description of steam-zone expansion before the critical time. At times near critical time, and also after critical time, experimental data show theoretical steam-zone volumes to be high. Even so, the accuracy of the calculated steam-zone volumes is sufficient for practical purposes. THEORY OF STEAM-ZONE GROWTH INTRODUCTION The engineering evaluation of steam-drive processes is based mainly on a mathematical description of reservoir heating by hot fluid injection presented several years ago by Marx and Langenheim.l This theory can be used to determine the growth of the steam zone,2 provided the flow of heat from the steam zone into the liquid zone ahead of the condensation front (Fig. 1) is neglected. Since we must expect, however, that heat transport into and inside the liquid region (the region between condensation front and production wells) will affect both the water/oil flow and the growth of the steam zone, it is necessary to re-examine and to generalize the Marx and Langenheim theory as applied to steam drive, which will hereafter be termed the "old" theory. Observations made in several field tests emphasize the need to study the heat flow across the condensation front (CF). On several occasions it was discovered that heat was transported in the liquid zone far ahead of the advancing CF, resulting in an early breakthrough of warm water. This seems to occur most often when high injection pressures are applied, when "wet" kONINKLIJKE/SHELL EXPLORATIE EN PRODUKTIE LABORATORIUM RIJSWIJK, THE NETHERLANDS SHELL DEVELOPMENT CO. HOUSTON, TEX.

Jan 1, 1970

By Axel Hultgren

IN ingots of silicon-killed carbon steel made without addition of - aluminum, transparent spherical or nearly spherical inclusions, up to about 0.15-mm diameter, are generally present. They may be glassy or crystallized, wholly or partially. In steels of high carbon contents they consist of silica,4,12 in steels of lower carbon contents arid, in consequence, higher oxygen contents, of iron-manganese-silicate with higher iron and manganese contents the lower the carbon content. Small silicate inclusions, about 0.01 mm or smaller, may occur, just as sulphide inclusions do, in portions that have been the last or nearly the last to freeze,10,12,17 this location being verified by primary etching.10 Hence, it may be concluded that they have separated from liquid steel that has become enriched in oxygen through segregation during freezing. Silicate inclusions with a diameter larger than about 0.02 mm appear, at least in ingots of moderate size, to be distributed at random in relation to the primary structure and therefore may be assumed to have existed as droplets in the liquid steel before freezing started at the point in question. The average size of such large inclusions has been found to increase toward the axis17 and their amount to increase from the top downward in the axial region .2,10 It has been assumed that this accumulation is a result of the droplets being brought down by the slowly settling metal crystals.2 It has been suggested that large silicate inclusions generally form,6 or may form," by coagulation of smaller ones, or by small drops being caught by larger ones during the rising of the latter.9 The occasional occurrence of a large inclusion surrounded by a swarm of small ones has been taken as a proof of this mechanism but, as suggested later, may be explained in another way. It has been pointed out that an inclusion droplet in the liquid steel may grow at the expense of an adjacent one, which implies diffusion of soluble inclusion-forming elements in the liquid steel.15 Further, it has been pointed out that, on cooling in ladle or mold, the metal-silicate equilibrium is displaced in the direction of separation of silicate. Thus inclusions may grow in size and amount.2,3,8,12,13,17,19,20 The exposure of the liquid steel to the oxygen of the air during tapping and pouring, and the consequent reaction, has been mentioned as a source of the formation of inclusions19,20 but this possibility has not received particular attention. Further, the admixture of furnace slag during tapping&apos; or of refractory material from various sources1,20 has been considered. The inclusions due to the latter may be recognized, however, through their resistance to hydrofluoric acid, imparted by the alumina content. Silicate inclusions, are often referred to as. the natural and inevitable product of reaction on adding silicon or silicon

Jan 1, 1948

By Claude P. Heiner, Carl S. Westerberg

OBSERVATION of the clinkering action of coal from the Castle Gate D seam in underfeed stokers over a period of years has given rise to the present investigation of the effect of bony coal on clinkering. Some very costly examples of slagging clinkers in particular cases have been of especial interest to the authors. In such cases coal analyses and ash-softening and fluid temperatures have been found of little value. In general, it was noted that slagging was accompanied by a somewhat higher ash content of the stoker slack, and the suspicion naturally arose that foreign matter was the troublemaker. One would naturally suspect bony coal and rock impurities because of their occurrence and prevalence within the seam and the relative absence of sulphur or other impurities in the coal mined. The test results given herein indicate that the concentration of the bony coal and rock impurities in the coal determines the clinkering tendencies of the ash during combustion. CASTLE GATE D SEAM Extending from the Utah-Colorado line near Grand Junction, Colo., westward to Greenriver, Utah and thence in semi-circular form north and westward to Castle Gate, and from there southward to Mt. Hilgard in central Utah, is a line of nearly perpendicular cliffs, composed of interbedded massive sandstone and shales, making up the coal-bearing horizon of the Mesaverde formation of Upper Cretaceous age. The cliffs face roughly concentrically toward the south and southeast and the strata, with conformably enclosed coal beds, gently dip away from the face of the cliffs and at right angles to their front line. Back of these cliffs the mountain masses rise to an altitude of from 8000 to 10,000 ft. above sea level. Stretching away from their foot is a great shale plain, on which the towns are situated and over which the railroad lines and highways run. The Castle Gate mines enter the coal beds at their outcrop; the workings extending within a particular coal bed away from the outcrop and under the mountain mass back from the cliff line (Fig. 1). Throughout the Mesaverde formation in the Castle Gate general area, there are from two to five workable coal beds. Usually not all of these are coincidentally of workable size and merchantable quality throughout any given large tract. In the Castle Gate mine area, the base of the Mesaverde horizon and coal measures is marked by a massive, cliffmaking, gray sandstone bed 6o to 8o ft. thick. Locally this is referred to as the "Castle Gate sandstone floor." Frank R. Clark, in U. S. Geol. Survey Bull. 793, designates it as the "Aberdeen sandstone." At Castle Gate the various coal beds are designated by the letters A, B, C, etc., commencing with the lowest bed. Fig. 2 is a graphic generalized vertical section, compiled from physical exposures

Jan 1, 1941

By Edmund M. Warner

Any city dweller who has walked alone along a remote mine passageway has to be impressed by the eerie silence-the total absence of noise except for one&apos;s own breathing and scuffing of boots on the bottom. An underpound mine can have tremendous contrasts, from complete silence to the ear piercing clatter of percussion drills working in the mine roof. Noise in underground mining machines is both useful and detrimental. Without characteristic noises that are recognizable to miners, operators would have a more difficult job operating their machines. Noise serves as a warning of approaching machinery, and changes in noise give indication that breakdowns may occur. But excessive noise is a health hazard. In recognition of this fact, our congress included noise exposure as a part of the "Coal Mine Health and Safety Act of 1969." As [Fig. 1] indicates, allowable noise exposure is established by specifying a time duration permitted at different noise levels. This is the well known "Walsh-Healey" criteria which govern noise exposure in many industries. In most industries noise exposure is not a constant value. Noise emitted from coal mining machinery varies widely, and therefore interpretation of compliance is not a simple matter. Time, environment, instrumentation and machine productivity, as well as the designed machine characteristics affect the results. It is the purpose of this paper to present the underground coal face machinery noise problem from a manufacturer&apos;s point of view. Manufacturers are very much aware of their responsibility to alleviate the noise problem. Not infrequently, orders are received from coal operators that a given new machine must comply with "The Coal Mine Health & Safety Act" noise regulations. Such a request is indicative of the misunderstanding which exists with machine users, while indicating their desire to comply with the law. Obviously, if all machinery could be designed and used at some arbitrary low noise level, it would be possible to simplify the regulations. However, there still remains the pervasive problem of existing machines. Extensive investigation to create practical retrofit kits for existing machines, to substantially reduce noise have been only marginally successful. Certain machine mechanisms are inherent noise makers. The only really good solution is to substitute a quieter mechanism. Such overall new designs have been undertaken, and several examples are mentioned later in this paper. Before describing machine noise problems, it is necessary to define noise values. Noise levels are measured in decibels. The designation "DBA" is decibels measured on the "A" scale, or one in which the frequency sensitivity of the instrumentation is about the same as the human ear. The numbers are often misinterpreted because it is a logarithmic, rather than an arithmetic scale. In comprehending the numbers, it should be remembered that any increase of 3 dB doubles the sound power. In practical terms this means that two noises of equal power at the same distance away increases a dB reading by three. Consider [Fig. 2] showing a simplified case of a worker on a machine subject to

Jan 1, 1980

By C. E. Sims, H. M. Banta, S. L. Hoyt

OVER the past few years, metallic arc welding has been extended to steels of the hardenable type. As compared with other methods of fabrication, production has been facilitated, service performance frequently has been improved, and the overall results have been so favorable that engineers and production men have placed great emphasis on this application of welding. The metallurgist and welding engineer, however, have been confronted with new problems, since, for one reason, the harder steels are more apt to crack and special precautions must be taken if sound joints are to be procured. Some of the factors that are pertinent to this problem are metallurgical, and it is an object of this contribution to discuss steel manufacture and processing methods and to show how their control assists in reducing the hazard of cracking. By this time, the problem of cracking specifically underbead cracking, or "hard cracks," or parent-metal cracks-is well known, and a large amount of work has been published, both here and abroad, dealing with this subject.1 The tests described are made under restraint to facilitate cracking and, are so designed that a semiquantitative, or at least a relative, value of the cracking tendency is obtained. As one would expect, it has been reported that the incidence of cracking increases with the carbon and alloy content of the steel. More specifically, the cracking tendency is said to increase with the hardenability of the steel, though it is now recognized that the situation is not so simple. The phenomenon itself, on the other hand, has been but loosely described, and it is in only broad generalizations that this cracking has been related to steel metallurgy. Thus, while it has been abundantly demonstrated that the hardenable aircraft steel, S.A.E. 4130, must tend to crack when arc-welded, the problem still remains of why two lots of the same composition and the same hardenability can vary so greatly in cracking tendency, or what it is about the prior metallurgical history that so strongly affects the results. Further, while the efficacy of preheating and postheating in preventing cracking is recognized and utilized in production, the mechanism by which they function seems not to have been described. These points are thought to be significant in the production of steel for fabrication by metallic arc welding and, hence, of interest to mill and process metallurgists and welding engineers. Consequently, it is another object of this paper to discuss the mechanism of parent-metal cracking, including the influence of the condition of the steel and the control of cracking by manipulation of the time-temperature conditions during and after welding. A new method of rating the cracking. tendency is also discussed, and the results of the test are related to steel composition and structure or prior history. The work to be described was the outgrowth of work in 1940-41 at Battelle Memorial Institute 0n two projects, one

Jan 1, 1945

By J. L. Bruce

AFTER six years of war-enforced idleness, Cyprus copper mines are operating again. This relatively long shutdown seems infinitesimal when compared with something like seventeen hundred years of inactivity that followed the important operations of the ancient Romans, and continued until the twentieth century. The historical records of the ancient operators are scarce but full of interest to those who realize that Cyprus copper production was important more than two thousand years ago. This paper, however, will give only a brief review of the ancient history and the evidence of ancient mining which have been quite fully covered in other publications.1-6 The purpose of this paper is to give a general sketch of the Cyprus Mines Corporation&apos;s enterprise in the Island of Cyprus, with special emphasis on those conditions and practices which are unusual. The picture is not intended to be complete, and many interesting details will be omitted. Some of these may be described in subsequent papers covering departmental practices. The principal operations of Cyprus Mines Corp. are in Cyprus, where it has leasehold concessions from the Cyprus Government covering about 50 square miles. It holds additional areas under "Prospecting Permits." The consideration for the concessions is a royalty on production, coupled with minimum annual rentals. Cyprus is a British Colony in the Eastern end of the Mediterranean with an area of about 3000 square miles and a population of about 400,000, of which 65 pct may be classed as Greeks on the basis of language, religious training, and traditions. About 25 pct are Moslems, principally of Turkish extraction. Prevailing languages are Greek and Turkish. A large proportion of the supervisory and administrative staff and practically all of the skilled and unskilled craftsmen and laborers are drawn from the native population. The "overseas" staff constitutes not more than 2 pct of the employees. The majority of these are from the United Kingdom, the British Dominions and Colonies. About one-third are from the United States of America. The topography of the Island is rugged, with mountains and foothills covering more than two-thirds of the area. Between the Kyrenea range rising to 3000-ft elevation, along the north coast, and the Troodos range (maximum elevation 6400 ft) occupying the southwest half of the Island, lies a cultivated rolling plain, the "Mesaoria," dry in summer and watered only by the winter rains and by scanty irrigation from small mountain streams and wells in the spring and summer. Climate and natural vegetation are characteristic of semi-arid conditions. The "urban" population, about one-fourth of the total, lives in the five port towns, Famagusta, Larnaca, Limassol, Paphos and Kyrenia, and in the capitol Nicosia, near the center of the Island. The suburban population lives almost entirely in villages or small towns, with very few people residing in isolated country homes.

Jan 1, 1947