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By H. R. Hendricks

The ore of the Iron King mine, being very hard and having a very fine crystalline structure, presents many problems in milling that are not present in ordinary lead-zinc ores. This very fine crystallization causes much interlocking of minerals and is most prevalent between the gold and pyrite and the sphalerite and pyrite. The recovery of gold is extremely important in the economic value of the ore and therefore constitutes a large part of the metallurgical problem to be solved. All of the gold is free but it has not been observed in any of the mill products except under very high magnification. Some of the gold is tarnished, and this condition also makes recovery difficult. The gold values have been determined to be highest in association with chalcopyrite and arsenopyrite. The silver occurs mostly in association with galena and tennantite. A small amount of native silver has also been observed occasionally. The galena, being softer than the other minerals, frees much better in grinding. Some oxidation of galena is present in the ore, as the non-sulphide lead content amounts to approximately 10 pct of the total lead. Some oxidation of sphalerite is also present in the ore. Microscopic examination of the zinc concentrate shows many grains in the minus 325-mesh to be half pyrite, together with the fact that some of the zinc is marmatite, makes production of a high-grade zinc concentrate very difficult unless recovery is sacrificed. In the grinding, it has been determined that the flotation feed must be maintained at 85 pet minus 200-mesh to free the minerals sufficiently to make the best economic recoveries. Table I illustrates the effect of fine grinding. It shows that the bulk of the recovered minerals is in the minus 325-mesh size and is the part of the minerals that is more readily freed. The pyrite concentrate is not given in the table because it contains practically no minus 325-mesh. Development of Present Flowsheet and Milling Practice The first milling operation started in 1938, with a mill capacity of 140 tons daily, producing a bulk gold-silver-lead concentrate. After several stages of expansion and changes of the mill flowsheet, the mill is currently treating 380 tons of ore daily and three separate concentrates are produced. In order of their importance and value, these are: (I) lead concentrate, (2) zinc concentrate and (3) pyrite concentrate. The present average grade of mill heads is: gold 0.11 oz, silver 4.00 oz, lead 2.50 pet, zinc 7.60 pct, copper 0.22 pct, iron 22 pct and 30 pct insolubles. Specific gravity of the ore is 3.8. During the period when bulk flotation mas practiced, the flowsheet was very simple. The ore was ground fine enough to liberate as much gold, silver and lead as possible. Necessary collectors and frothers were added in sufficient quantity

Jan 1, 1949

By D. W. Butner

As a part of the Bureau of Mines program for the discovery and development of critical and essential mineral deposits, the Darwin district, Inyo County, Calif., was chosen for investigation as a possible source of, tungsten ore. Two properties were active in the district at the time. One was producing tungsten concentrate and the other was preparing to produce tungsten concentrate as a byproduct in the milling of lead-zinc ore. The objective of the project set up by the Bureau was to investigate promising areas and to sample, thoroughly those which preliminary inspection indicated contained tungsten ore. Work was begun November 17) 1941, and completed January 19, 1942. In the course of the project, 18 localities were inspected and 9 were sampled. The following pages brief the history, note the location, and outline the physical features and geology of the district. The work performed by the Bureau of Mines is described and supplemented with assay maps.

Jan 1, 1949

By T. E. Newkirk

In the past few years there has been considerable controversy on the most suitable material to be used in ventilation doors which will have to serve also as fire doors. For this reason it was thought useful at Mclntyre to find out to what extent our zinc chloride treated spruce doors were fire resistant and what protection we could expect in case of an underground fire. The doors to be tested were constructed of the same material as that used in our regular ventilation doors underground, namely, 2-inch tongue-and-groove well seasoned spuce. It is to be noted, however, that whereas the dimensions of the ventilation doors are 5 ft. by 6 ft., those of the test doors were only 14 in. by 17½ in. The treated doors were made up from lumber that l1ad been treated with chromated zinc chloride, as described in Mr. Keeley's paper. Two kinds of tests were carried out: a heat test and a flame test. An electric muffle furnace for assaying, equipped with a pyrometer, was used for the beat test, the furnace door being replaced by the wooden door to be tested. In the flame test, the door to be tested formed one end of a rectangular steel box in which was placed a large blow-torch. Air was admitted at the rear of the box. A pyrometer recorded the temperature at the inside surface of the door where flame impinged. The results of the tests are presented in the accompanying table, in which the word 'failed' is to be interpreted as meaning that the door had been damaged to the extent that smoke, but not necessarily flame, issued from its outside surface.

Jan 1, 1949

By E. V. Potter, H. C. Lukens

In general, during all electrowinning processes, large volumes of gas are liberated at the cathodes of the electrolytic cells. Most of this gas escapes from the electrolyte, but much of it may be absorbed or entrapped in the metal being plated. In many cases, this is not objectionable, but since it is liberated when the metal is heated, it is a potential source of trouble where the metal has to be heated or melted. Since this gas is principally hydrogen, in escaping from a melt it may cause minor explosions and scattering of the molten metal, while in other cases it has been reported to make the metal rise in the molds when cast. Therefore such metals would be more generally useful if all or most of the gas were removed before they were used. In an earlier investigation1 the gas content of electrolytic manganese was determined and methods found for removing most of it without detrimentally affecting the properties of the metal. In this paper results of a similar investigation on electrolytic chromium are presented. Apparatus The apparatus used in these investigations is essentially the same as that used in the manganese work1 and is shown in Fig I. It consists of sample chamber A, vacuum or pressure gauge B, volume gauge C, vacuum pump D, hydrogen container E, auxiliary gas chamber F, and furnace G. The pressure in the system is measured by the mercury manometer B. Changes in gas volume are measured by volume gauge C which is essentially another mercury manometer in which the volume of the mercury column is large and can be varied by raising or lowering the well H. In this apparatus the tube had a volume of 100 cc for a 60 cm length. Larger changes in gas volume can be accommodated by auxiliary containers attached to the system at F. The sample is heated by furnace G, the temperature of which is regulated within 5°C by an automatic controller. The only change in the apparatus was the sample container which was made smaller because of the chromium metal's containing more gas than the manganese metal and a smaller sample was used. Method The measurements were made as follows: The complete system when cold was evacuated and refilled with H2 until all the air was removed. It was then refilled with H2 to the desired pressure, as indicated on manometer B. The temperature was then raised to the desired value and as gas was either liberated from or absorbed by the sample, the height of well H was changed to alter the volume of the system so as to keep the pressure constant. In determining rates of liberation of gas, volume readings were taken at convenient tirne intervals depending on the rapidity with which the gas was liberated. In determining total volumes, the specimen was held at constant temperature and pressure until

Jan 1, 1949

By A. A. Collins

OF widespread interest to all lead metallurgists is a dross smelting process that will consistently give mattes and speiss of low lead and high copper contents. It is a problem that has intrigued operators through the years. In the past few years, with increased amounts of copper and overcrowded metallurgical furnaces, it has been almost imperative to evolve some method. Experience at El Paso has shown that high copper-to-lead ratios can be consistently attained by use of the new soda process. † Older Methods Three older methods, i.e., 1. Dross plus siliceous ore and limestone. 2. Dross plus one fourth blast-furnace speiss and siliceous ore, 3. Dross plus sulphides and siliceous ore and scrap iron, do not give consistent results and the products are higher in lead and lower in copper than those produced with the soda process. Any radical changes in the blast-furnace drosses will throw any of these older methods into a tailspin. For almost ten years, El Paso used the third method, with varying results. It was noted that as the arsenic and anti- mony contents of the dross increased, the speiss and matte steadily increased in lead and decreased in copper. A dross containing four parts of copper to one part of arsenic was considered to be ideal. At times this balance of copper to arsenic was almost impossible to obtain, especially when handling copper Cottrell dusts, which vary so greatly in copper and arsenic. Custom smelters like El Paso, handling a great variety of dusts, concentrates and ores, are apt to have anywhere from I to 15 pct copper on charge, with a variable in arsenic of from 0.5 to 2 pct. With such conditions, and drosses varying from 12 to 40 pct copper, an operator using the older dross reverberatory practices would find it just about impossible to control the reverberatory operations and to produce a good product low in lead and high in copper. These conditions prevailed at El Paso from the year 1940 to the present. In May 1942 it was decided to experiment with the soda process. The Soda Process After considerable work the soda method was finally evolved to the present state. It has never failed to produce a good product even under the most adverse conditions. It might be well to remember that many men of thought in the smelting industry have at one time or other sugge~ted the removal of all copper, arsenic and antimony from the lead blast-furnace charge as an aid to better metallurgy.

Jan 1, 1949

By J. F. Myers, S. D. Michaelson, F. C. Bond

This work was undertaken with the object of collecting plant data on rod milling, making laboratory tests on representative samples of the various ores, and arriving at a basis for comparing the relative efficiencies of different mills operating on widely different materials and under varying conditions. These efficiencies consider only power consumption and do not include the other elements of operating costs. The over-all economic efficiency may be quite different from the consumption efficiencies with which this paper is concerned. It is hoped that these collected data will be of value in indicating the most favorable conditions of rod-mill operation. Operating-plant data on 14 different installations from II plants throughout the United States and Canada were collected, and laboratory tests were made on samples of these ores. The plant rod mills were all operated wet, and either in open circuit or with circulating loads so small as to be negligible. Data were obtained on one roll-crushing operation at the Utah Copper Co., for comparison. The rolls were operated wet in closed circuit with 10-mesh vibrating screens, and with a high circulating load averaging about 600 pct. This particular roll operation can be considered as a model of efficiency. The collected plant data are necessarily incomplete in some instances. Each ore was assigned a code number, and the name and location corresponding to the code number are listed in Table I when permitted. The power data in every instance refer to motor input. Net power was reported by the operator for samples 1-30 and 1-45, and this was increased by 20 pct to compensate for motor losses and power factor. The plant rod-mill data are listed in Tables 2 and 3. In Table 2 the speeds are calculated on the basis of partly worn rod-mill liners 2 in. thick. Where D

Jan 1, 1949

By C. H. Riggs

This petroleum-engineering report of the Atlanta oil field in BO them Arkansas reviews reservoir conditions and pressure-production history of one of the important Smackover limestone fields in southern Arkansas. The study of the field was made and the report prepared to aid operators in this and similar fields to increase the efficiency of recovery through the application of better engineering practices. A similar study of the Magnolia field in Arkansas by Carpenter and Schroeder was the first of these reports prepared under a joint agreement between the Arkansas Oil and Can Commission and the Bureau of Mines. The increasing demands for petroleum and its products and the steadily increasing cost of discovering and developing oil field; emphasize the importance of conserving known reserves of oil and gas. In recent years vast improvements have been made in the technique of oil recovery, both by better utilization of the natural energy stored in the reservoir and by improved methods of applying external energy to the reservoir system. This report may aid in oil conservation by careful evaluation of the several factors that did or did not contribute the greater recovery in the Atlanta field and by suggesting improvements in production methods and applications.

Jan 1, 1949

By R. C. Gegg

The property of MacLeod-Cockshutt Gold Mines, Limited, is in the Little Long Lac area of the Thunder Bay Mining Division, Ontario, about four miles south of the Canadian National railway station of Geraldton. The town of Geraldton is about 170 miles east of Port Arthur and 22 miles west of Long Lac, which is on the main line of the Canadian National railway. The Trans-Canada highway passes through the property and during the winter season the road is maintained between Port Arthur and Long Lac. Electric power is supplied by the Hydro-Electric Commission of Ontario from their Cameron Falls plant and is distributed at 44,000 volts by a local transformer station. Early History The South orebody was discovered by diamond drilling in the spring of 1936. This is a porphyry contact orebody with major pyrite and minor arsenopyrite mineralization. Some gold is present in the 'free' state, but for the most part it is locked in the pyrite, with little or none in the arsenopyrite. Preliminary test work was based on comparing straight cyanidation with bulk flotation and cyanidation of the concentrates. The overall results by the two methods were found similar, but for economic reasons the flotation-cyanidation process was adopted. Before the mill came into operation, a new and entirely different type of orebody, consisting of massive sulphide lenses in banded iron formation, was discovered to the north of the existing workings. These bodies are an intricate mixture of pyrite and arsenopyrite, both carying gold.

Jan 1, 1949

By William Sharp

Discovery of new ore in the Eureka district, Nevada, as a result of bold and persistent exploration based on a geologic interpretation of structure has recently aroused wide comment in mining circles. It may therefore be of interest to the profession to have a summary of the historical background and an account of the recent program of development. The mines of Eureka, discovered in the late 1860's, were the first important silver-lead mines in the United States. They have a production record of more than fifty-one million dollars.' After the ore bodies were apparently bottomed in the 1880's, repeated attempts to find new ore were unsuccessful until 1940, when a diamond-drill hole cut a 40-ft width of sulphide ore 1409 ft below the 841-ft level of the Locan shaft. The next hole cut 30 ft, of ore and three additional holes drilled in 1943 and 1944 showed the ore body to be continuous, Shaft sinking to reach and open up the ore body has been in progress since 1945. Geology Stratigraphy The ore in the Eureka district occurs in a series of Cambrian sediments which is exposed through a thickness of over 6000 ft. The stratigraphic column is as follows:2 Thickness. Ft Ordovician : Pogonip limestone (upper part). Thin bedded limestone with some chert. Upper Cambrian: Pogonip limestone (lower part). Thin-bedded limestone (with shaly limestone and chert).... 425 + Dunderberg shale. Shale (clay and calcareous shale) and thin-bedded limestone............. 34°+ Hamburg dolomite. Dolomite and limestone.................... 900+ Middle Cambrian: Secret Canyon shale. Shale and thin-bedded limestone........ 1,035 Geddes limestone. Carbonaceous, flaggy limestone with some shale partings................ 335 Middle (?) and LoweR Cambrian: Eldorado dolomite. Dolomite and limestone.................... 2,000 Lower Cambrian: Prospect Mountain quartzite. Quartzite with occasional intercalated conglomerate and shale 1,660+ The Prospect Mountain quartzite has been subdivided into two formations: Upper part, Pioche shale. Shale, quartzite with some intercalated limestone (Lower Cambrian fossils)-200 ft. Lower part, Prospect Mountain quartzite proper- 1,460Sft. Total thickness................ 6,695 + Much the most productive of these formations has been the Eldorado dolomite, which yielded more than $42,000,000 at Ruby Hill and an additional amount

Jan 1, 1949

By Paul Averitt

The coal-bearing rocks in Iowa occur in an arcuate belt across the central and southern parts of the State (fig. 1) at the northern end of the Western Interior coal field. Coal beds of minable thickness are sporadic throughout this area and are mined principally in the three districts indicated by dark shading on the index map, although coal occurs also in the areas indicated by light shading. The coal is mostly of high-volatile C bituminous rank in all parts of the State. SURFACE FEATURES The coal-bearing rocks in Iowa are concealed for the most part by a covering of glacial drift ranging in thickness from a few feet to several hundred feet. This covering has been dissected to form a gently rolling upland, rising from an elevation of 466 feet at the mouth of the Des Moines River to about 1,200 feet in Calhoun County in the central part of the State. The local relief, however, rarely exceeds 150 to 200 feet.

Jan 1, 1949

By The Staff

Introduction For many years the necessity of mechanization of all the underground services in the collieries of the Dominion Steel and Coal Corporation in the Sydney, Pictou, and Cumberland coalfields has been foremost in the minds of those concerned with the engineering and operating policies of the Corporation, and a gradual policy of the mechanization of these services has been undertaken over the years, until nearly all the underground services, except the actual loading of the coal at the face, have been mechanized. Early in 1928, experiments in mechanical loading were conducted which finally culminated in the installation of mechanical loading units in our No. 20 colliery in 1938. This attempt at mechanization of the final underground services proved unsuccessful because the workmen were not ready to accept complete mechanization, and the fruition of many years of study was delayed until the year 1947. During the ensuing years, the engineers and operators of the Corporation had studied the possibility of adapting to conditions prevailing in Nova Scotia mines the many types of loading equipment in use in the United States and Europe. The varied physical conditions which exist in the fields operated by the Corporation rendered it extremely difficult to apply the equipment and technique successfully employed in the United States, and more recently in Europe, and special techniques have had to be devised to meet the need of each individual colliery. Prevailing physical conditions in the Nova Scotia fields find no parallel for mechanization with those of any other part of the American continent.

Jan 1, 1949

By Richard Havens

Samples of low-grade oxide manganese ores were procured from two southern Utah properties by examining engineers of the Bureau of Mines during an investigation of manganese mineral deposits of continental United States. A 2-ton lot of ore was obtained from the Blackbird group of mining claims in Paiute County, 11 miles southeast of Marysvale, Utah. The deposit was described as an intrusive sheet or sill of altered brecciated rhyolite cemented together by manganese oxide minerals. A 2-ton sample from the Black John property, situated 32 miles northeast of Kanab, Kane County, Utah, was taken from open cuts on the vein. The blanket-type vein deposit consisted of nodules and streaks of manganese imbedded in compact gray clay.

Jan 1, 1949

By R. Havens

During the course of the investigation of possible domestic sources of manganese, engineers of the Bureau of Mines examined and sampled two manganese mineral deposits in northeastern Utah. The Daggett Chief claims are situated in the northern part of Daggett County,5 miles southwest of Manila, Utah, and approximately 57 miles from Green-River, Wyo. Owners of the property are J.E. Harris Of Salt Lake City, Utah, Nels Palleson of Manila, Utah, and J. P. Larson. The manganese occurs in a vein 6 to 8 feet wide in a shale and sandstone formation. Two distinct occurrences Of manganese are present Within the vein, and samples of each were subjected to testing. Lot A, representing .probably the major portion of the Manganese; was composed of nodules and veinlets of manganese oxides in a semicompact; gray, shaly formation. Lot B represented a 15-inch vein of dense iron-manganese ore. .Most of the laboratory work at Salt Lake City, Utah, was done on these two samples; but preliminary testing also was carried out on a small sample, physically similar to lot A, submitted by the owners.

Jan 1, 1949

By Harold E. Heide

Although the United States has been the largest consumer of tin, only a small quantity of tin was produced or. the North American Continent during peace times, and thy deficiency was made UP by foreign imports. Because the most important sources of tin wore cut off by the war with Japan, it became necessary to increase imports from South America and Africa and, if possible, develop domestic tin reserves. Most of the tin produced in North America has come from the Seward Peninsula, Alaska, and a substantial portion of this was produced from the Potato Mountain area. Tin was first identified as occurring in the Seward Peninsula by A. H. Brooks in 1900, when a part of heavy concentrates from Anikovik River and Buhna Creek, one of its tributaries, proved to be stream tin. 4/ His report was given considerable publicity by the daily press, and the first steamers north in 1901 carried a number of prospectors bound for the Seward Peninsula to search for tin. The deposits of stream tin on Buck Creek in the Potato Mountain area were discovered in the fall of 1901. 2/

Jan 1, 1949

By S. Benedict Levin

The Big mine has been a source of mica and high-grade felspar from separate zones in the same pegmatite dike since 1914. In 1943 it became the outstanding producer of strategic mica in New England. The urgent need for sheet mica led to core-drilling by the Bureau of Mines from February to May 1914. Nine diamond-drill hob's aggregating 1,474 feet in length were drilled by the Bureau of Mines with the objective of developing extensions of the known mica shoots. The use of large-diameter bits made it feasible to obtain reasonably reliable results in diamond-drilling this type of deposit.

Jan 1, 1949

By Elton A., Youngberg

The Holden mine operated by the Chelan Division of the Howe Sound Co. is on the east slope of the Cascade Range in north central Washington on the south slope of Railroad Creek valley at an elevation of 3500 ft. The mine may be reached by a 40 mile boat trip from the town of Chelan which is at the southern tip of Lake Chelan, to Lucerne at the mouth of Railroad Creek and an 11 mile bus ride up Railroad Creek to Holden. All freight and concentrate is moved over this route to Chelan Falls on the Columbia River which is on the railroad four miles below the town of Chelan. The mine is now producing 2000 tons of gold, copper, and zinc ore per day which is treated in the Holden mill. Gold-copper and zinc concentrates are made, the first of which is shipped to Tacoma, Wash., and the latter to Kellogg, Idaho, for smelting.

Jan 1, 1949

By Walter C. Lawson

The primary objective in laying track in panel sections is to reduce the number of track laborers required. This is possible because the work is mechanized. Moreover, because the work is mechanized and each of the operations needs only a few men, and each is self-contained, the work can be carried on at night as well as in daylight; therefore, the method provides a means of preparing tracks on more than one shift. The laying of tracks in open pits and quarries in panel sections is not new but new methods have been made possible by the introduction of new types of equipment. The purpose of this paper is to describe the methods that are followed in the Morenci open pit. General Mining Operations Full-scale ore production at Morenci is about 50,000 tons daily. The normal yearly output of ore and waste is 32,000,000 tons, of which 30,000,000 tons is handled by rail haulage. This requires the building of approximately 47 miles of loading and other temporary tracks during a 12 months' period (Fig I). Broken ore and waste is loaded into 90-ton capacity dump cars with 5-cu-yd full-revolving electric shovels. Shovel banks are uniformly 50 ft high. They are blasted with churn drill holes and the broken material from a blast is loaded out with two shovel cuts, the first one being called (locally) the "splatter" cut and the second one the "clean-up" or "face-up" cut. When a shovel is making its clean-up cut, the loading track is about 70 ft from the toe of the bank and, under normal conditions, a bank is blasted against it without removing the track. It is thus evident that each track is used for two shovel cuts before relocation is necessary, inasmuch as a single position serves for both the clean-up cut prior to blasting and for the splatter cut following the blast. The normal advance of a solid bench by a single blast is 40 ft (Fig 2). Preparation OF Panel Grade In preparation of panel grades bulldozers and rooters are used for the bulk of the work and an auto-patrol road grader for the final stage. At Morenci, bulldozers have always been used for track-grade preparation, but before the introduction of the other two types of equipment grades were uneven and irregular at best, and always necessitated a big gang of track laborers to hand-block under the ties with rocks after a track was laid into place. High blocking also made poor track because of instability and caused frequent derailments. Under these conditions derailments were generally bad as re-railing became difficult. Much track was torn up and many tics broken in the process. The inadequacy of the hand-blocked track, together with the inefficiency of hand blocking, constantly pointed, to the desirability of reducing the amount of labor required, so various means were tried to improve conditions, such as wood

Jan 1, 1949

By Edward P. Barrett

The production of sponge iron is one of the earliest arts in the metal¬lurgy of iron and steel. The idea of producing iron direct from ore at temperatures lower than the melting point of, the iron has persisted throughout the centuries. The purpose of this paper is to present a general picture of the methods, proposed or tried within the past 100 years, for producing sponge iron by treating hot ore with heated reducing gases. The abundant literature has been abstracted, and quoted freely. Publications concerning more detailed information on the various methods are referred to in the footnotes and in the bibliography.

Jan 1, 1949

By Carl H. Unruh, D. L. Katz

Experimental data are presented for hydrate formation conditions for gas mixtures of carbon dioxide and methane. Equilibrium constants for carbon dioxide, defined as the mole fraction of carbon dioxide in the gas phase (dry basis) divided by the mole fraction of carhon dioxitle in the hydrate (dry basis)... were calculated usine the equilibrium constants for methane developed in an earlier paper. INTRODUCTION A series of papers have been written on gas hydrates, giving temperatures and pressures at which hydrates would form.2. 3, 4. 5, 6. 9. 12, 13, 15 Vapor-solid equilibrium constants have been developed for methane, ethane, propane and iso-butane' by which it is possible to compute for a specific gas mixture the temperature and pressure at which hydrates will form. Carbon dioxide is present in many natural gases in concentrations usually from 0.1 to 2.0 per cent. It also forms a gas hydrate with water and presumably enters the solid solution when present with a natural gas which is forming a hydrate. The equilibrium constant giving the ratio of the concentration of carbon dioxide in the gas phase to the concentration in the solid phase is required for including the effect of the carbon dioxide when computing the conditions for hydrate formation. Direct determination of the concentration of carbon dioxide in solid hydrate requires a difficult separation to be made. Therefore, this research resorted to the measurement of the temperature and pressure at which specific mixtures of methane and carbon diox- ide would form hydrates. From the equilibrium constants previously mentioned for methane, it is possible to compute the equilibrium constant for carbon dioxide, without measurement of the composition of the solid phase. Fig. 1 shows the conditions at which methane and carbon dioxide form hydrates. The methane curve FG extents to some high pressure while the carbon dioxide curve AB intersects the vapor-liquid water-liquid carbon dioxide

Jan 1, 1949

By David Turnbull

In most reactions involving solids the transformation kinetics may be represented by the combination of two processes —those of nucleation and of growth. For example, Mehl and his coiq-orker in their studies on the transformation of austenite to pearlite in steels have been able to describe the kinetics in terms of the rates of nucleation and of growth. By micrographic examination of their specimens they determined the nucleation rate (rate of appearance of pearlite colonies) as a function of time, of percent transformation, and of temperature. They observed that after an incubation period in which no nuclei are observed, the number of nuclei appearing per unit time per unit of untransformed grain boundary area, increases at an increasng rate with time. When recrystallization rates have been interpreted in terms of rates of nucleation and growth and the rate of nucleation determined, an incubation period followed by an increasing rate of nucleation has also been reported. Among the studies in which this phenomenon has been observed are those of Kornfeld and coworkers3 on cold-drawn aluminum wire, of Anderson and Mehl4 n the recrystallization of aluminum sheet, and of Stanlcy and Meh15 on the recrystallization of silicon ferrite sheet. The existence of an incubation period during which no nuclei are observed does not appear to have been reconciled with the generally accepted Volmer-Becker theory of nucleation in condensed systems. This theory ostensibly requires the rate of appearance of nuclei to be independent of time. Volmer-Becker Theory In order to form a stable nuclcus of a new phase B in a mother phase A, the molecules or atoms of the mother phase must surmount a free energy barrier whose height is determined by the volume free energy change accompanying the transformation and the energy of the surface separating the two phases. The free energy difference between nuclei of various sizes and the molecules or atoms of the mother phase is plotted schematically in Fig I as a function of the radius of the nucleus r. This free energy difference is plotted for two values of the change of free energy per atom, (Jb — ja) accompanying the transformation when the volumes of the two phases are infinite. One of these values of (Jb—Ja) corresponds to a temperature above the temperature To at which the phascs A and B are in equilibrium, and the other to a temperature below To. In order for a nucleus of phase B to be stable at a temperature T² < TO, its radius must be equal to or greater than the critical value r* for which the free energy of formatioil of the nucleus is a maximum, the of phase B having radii equal to or less than r* will be called embryos hfehl and Jetter," in discussing the kinetics of nucleation in solid systems, suppose that embryos of critical size are formed

Jan 1, 1949