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By W. F. Boericke

THE character of the Wisconsin orebodies must be clearly understood to appreciate the difficulties encountered in sampling and estimating them. Unlike the western vein deposits, they do not lie between regular walls, but form large, often highly irregular bodies, on a single floor in the so-called Galena limestone. The bottom of the ore is nearly always on the clay bed, a flat generally well-defined clay seam a few inches thick, which marks the junction of the base of the Galena lime and the top of the next formation, the Trenton lime. Above the clay bed, the orebodies make the unique pitch and flat formation described by Chamberlin.1 Theoretically, we should find two pitches, one on either side of the top crevice, diverging as they go down, with the final flats a considerable distance apart on the clay bed. The area between, known locally as core ground, should thus constitute a single defined orebody, with the richer portions near the pitches and the leaner portion in the center of the core. Practically, we find few examples of this perfect orebody. Almost invariably one pitch is better defined than the other, and often the opposite pitch is marked only by a faint line of fracture showing stronger mineralization than the wall rock, or it may be totally lacking. But with an orebody of any size or richness, at least one strong pitch usually tends to follow an east-west trend, or a quartering course. Late mining developments in the rich New Diggings field indicate that these pitches often assume a horseshoe or crescent shape, the richer ore being in the heart of the crescent and the pitches showing less mineralization as they extend outwards. In the northern end of the field, on the contrary, the pitches maintain generally a consistent direction, and have been mined for thousands of feet from the shaft. The length of the orebody may be several thousand feet, the height varying from 10 to 60 ft. and the width from 60 to 250 ft. In fact, on the core side, there is no definite end to the ore. The mineralization simply gets thinner and thinner, and the limit of working depends on the cost sheet and the profit from the recovery.

Jan 2, 1922

By G. D. Tranter

THE importance of ladle and teeming practice and its relationship to the yield and quality of the product has focused considerable attention on this phase of open-hearth operation. Inherently bad steel cannot be perfected by manipulation during teeming, but through the proper handling of "off heats" serious losses, which in many cases arise, can be greatly minimized. Conversely, good steel can be ruined as a result of poor workmanship in the pit, and rendered unsuitable for application to orders where exacting surface and physical requirements are essential. Sufficient evidence has been secured to warrant careful investigation of pit operations when an abnormal number of basic defects develop in the semifinished product or are found on the inspection table. This paper is written from the standpoint of the problems involved in making satisfactory ingots for sheet rolling. Furthermore, since rimming steel is almost universally used for manufacture of sheet, the data presented herein refer particularly to that type of steel. Going back some 30 years in open-hearth history, the word "pit" properly described the hole into which the ladle was lowered to receive the metal tapped from the furnace. The working space was necessarily limited, and the disposal of slag and metal overflowing the ladle or lost because of breakouts presented a serious problem. The equipment, of course, was rather crude in the light of present-day development. Often no ladle stands were provided, the ladle being hung on the crane beneath the runner. When delayed taps occurred, the ladle was often held in position for many hours, thus tying tip pit operations and causing delays to adjacent furnaces. Because of crowded conditions, slag thimbles were not used and a network of cast-steel triangles and old tapping bars, placed before the heat was tapped, was the means for removing slag from the hole. Furnaces of different capacities were used

Jan 1, 1935

By Alfred F. Beasley

The slags derived from the smelting of lead and copper ores are composed essentially of silicates. The problems arising from the smelting of these ores consequently involve the study of silicate fusions. In the absence of specific knowledge concerning the compounds occurring in slags, it has been the practice of metallurgists to treat them in terms of their total chemical composition. Slags have been classified as monosilicates, bisilicates, sesquisilicates, and subsilicates according to the relative amounts of oxygen in the acid and basic radicles of the hypothetical salt produced. While this has been a useful guide to metallurgists in computing the furnace charges, it falls far short of a basis for the scientific investigation of slags. In order to understand the nature of a slag as a whole it is necessary to learn the physical and chemical properties of the mineral compounds which go to make up that slag. The efficiency with which an ore is smelted in a furnace depends primarily on the physical properties of the compounds formed during the operation. When the melting points and fields of stability of each of the mineral compounds in slags have been established, it will be possible to determine the properties of any given slag by examining the nature and amounts of its mineral constituents. Earlier Investigations In 1884 and succeeding years an extensive survey of lead and copper furnace slags was made in Europe by Vogt, and a large part of the present knowledge of the subject is due to his publications.' Vogt's interest in artificial fusions was directed primarily towards the explanation of problems in igneous and economic geology and his subsequent works have been confined wholly to that subject.

Jan 1, 1930

By C. D. Hunter, I. B. Browning, Ralph Thomas

Production of oil in Kentucky during 1938 was 5,566,154 bbl., showing a substantial increase over that of the year 1937; while in the gas areas development was somewhat retarded although deliveries were similar to those of 1937. In eastern Kentucky there was an oil increase during 1938 of over 150,000 bbl. above that of 1937. No new pools were developed and the increase was almost entirely due to the installation and expansion of secondary recovery methods and to the further development of the Maxon and Big Lime, Mississippian pools in Floyd and Martin Counties. With this development eastern Kentucky reached a level near that of 12 years previous, when the development of several large producing areas had just been completed. In eastern Kentucky during 1938 not over 50 gas wells were completed. Although the pipe-line deliveries were approximately the same as during 1937, development was retarded because of an excess of gas already developed and the intensive development of the Oriskany fields in West Virginia. In western Kentucky, in the part of the Illinois Basin lying in the state and known in Kentucky as the Western Coal Basin, the increase was due to the discovery of a number of new pools, the largest and the one producing the biggest wells being known as the Birk City pool. In fact, the discovery of this pool with the McClosky "sand" (St. Genevieve limestone) as the producing formation was in a large measure responsible for the increased activity, which in turn brought the discovery of the other areas. The discovery well in this pool is the C. T. Blackwell, which came in flowing 1000 bbl. daily from the McClosky sand at a depth of 1870 ft., drilled in April 22, 1938. This discovery, coming at a time when the major companies and many others were conducting an intensive leasing campaign and making extensive subsurface and geophysical surveys in the deeper part of the basin, resulted in renewed activity over the entire area and brought much new drilling. It also brought to Kentucky for

Jan 1, 1939

Discussion of the paper of L. A. DELANO, presented at the St. Louis meeting, October, 1917, and printed in Bulletin No. 129, September, 1917, pp. 1267 to 1286. L. A. DELANO.-Since this paper was written, our daily mill tonnage has increased, from 2100 tons to 2700 tons a clay. We now use 11,000 lb. of chrome-steel balls of 2 ½ -in. maximum diameter, for an average crushing capacity of 275 tons per 24 hr., this giving a much higher crushing efficiency. I wish to call attention to two typographical errors in my paper as published. On page 1271, Table 1, the screen analysis of ore at the Bonne Terre mill should be stated in inches instead of mesh. - On page 1272, "the rolls average 185 tons per 24 hr., and have shown by test that 250 tons per hour can be maintained" should, of course, read "250 tons per 24 hours." THE CHAIRMAN (0. M. BILHARZ, Miami, Okla.).-Is the increased mill capacity due to improvement in the re-crushing of middlings or to keeping middlings out of circulation and treating them separately? L. A. DELANO.-AS was shown in Mr. Rabling's paper (Bulletin No. 128, p. 1161) the tonnage was raised by about 400 tons by simply changing the arrangement of screen openings in the jig bed and using punched plate instead of woven-wire screen. Then, by adding a ball-mill, we are taking from the jigs a much larger tonnage of middlings than before or as much as the ball-mill will crush. By sending the ball-mill product to tables and flotation, the jig capacity for original ore is increased. CHAIRMAN BILHARZ.-In your opinion, the satisfactory operation of a Hancock jig on a given ore requires adjustment of the length of the different screens and careful selection of screen openings? L. A. DELANO.-I think it does. In order to decide what screen openings to use on the jigs, we did a great deal of experimental work, making screen analyses of the various hutch products, and then hand-sorting the galena, middlings and chat from each size screen to determine the percentage of galena in each. Then adjusting our jig screens, we gradually worked the process up to a very efficient basis. ERNEST GAYFORD, Salt Lake City, Utah.-I would like too ask Mr. Delano whether the work of the Allen cone has been quite as satisfactory as the manufacturers claim for it.

Jan 1, 1918

By Ellis Clark

THE greater portion of the brown ore in the vicinity of Easton, along the north slope of the Lehigh Mountain or Durham range of hills, is obtained from mines instead of from open cuts or quarries, as is the case in the ore-beds in the western part of Northampton County, in Lehigh County, and in Berks County. The reasons for this method of procedure are various, and the principal causes of the shaft and regular mine system of working being adopted instead of the open air cuts, are : First, the large amount of stripping that would be necessary, as the richer and more valuable portions of the ore do not appear to be found as near the surface as those in other portions of the Kittatinny valley; second, the occurrence of the ore in regular beds, interstratified with the clay, and, to a less extent, in pockets or erratic deposits, as is the case in the mines of Lehigh County; and third, on account of many of the mines being situated at the immediate base of the Lehigh Mountain, where the surface water would be in such large quantities as to be with difficulty handled, and where the subsequent caving in of the sides of the open cut would be almost certain to expose the gneiss rock, which is water-bearing, the large quantities of water coming from which would probably drown out the ore-bed. Shafts, and a regular system of gangways and mining are therefore required by the configuration of the surface, the depth of the deposits and their geological structure. The ordinary size of the shafts is about four feet square, but where an extensive plant of pumping machinery is required the size may be increased to eight feet by six, with two compartments for the hoisting and pump-shaft respectively. A man-way, consisting of ladders with rests at intervals of twenty or thirty feet, is comprised in the pump-shaft, the wire rope being used for purposes of ascent and descent only in exceptional cases, on account of accidents which

Jan 1, 1879

IT is customary to make an alloy of copper in the same way, not to increase its quantity as with gold or silver, but to corrupt it for the art of casting and to destroy a certain natural viscosity in it. For this result it is accompanied by tin and also, sometimes, brass. But its proper and true alloy is fine tin, except, however, when you wish to make hammered works. For these, copper must be pure and without any trace [of impurity], otherwise it cannot be hammered thin, gilded in the fire, drawn into a wire, or made into vessels for use as is customary; but in casting, tin is almost essential. When you come to this association, just as it changes nature and appearance according to the proportion of tin which you give it, so it changes name and is no longer called copper but bronze, or, for a greater distinction of the quantity of tin for every hundred or more of copper, it is called by skilled workmen metal of more or less fineness as it contains more or less tin. This is recognized by the whiteness and brittleness: changing from red, which is the color of copper, it becomes white; from soft and flexible it becomes hard and brittle as glass. This admixture removes it so far from its original nature that one who does not know that it is a compound material believes it to be one of the number of metals engendered by Nature. Now, so that you may understand well, I tell you that various kinds of this are made, all according to the proportions of the quantities and according to the kind of work. For one kind of bronze is desired by those who make statues; this is not desired by those who make artillery or by those who make bells, mortars, basins, and similar cast objects. Now in order to alloy it into the species of bronze, eight, nine, ten, up to twelve pounds of tin are put with every hundred pounds of copper. Those who want it to make bells put twenty-three, twenty-four, twenty-five, and twenty-six, depending on the tone, on whether they are large or small in form, and on whether they wish to make the tone deep or sharp and clear. From twelve upwards is used for all other works that need it, either for hardness or to give them fluidity in casting by surpassing the degree of

Jan 1, 1942

By Henry M. Howe

This paper is a discussion of the extremely valuable one of Mathewson and Phillips, The Recrystallization of Cold-Worked Alpha Brass on Annealing,1 which not only gives us a wealth of important data reached with great intelligence, but also shows both uncommon powers of imagination and a perfectly fair spirit. Let me try to analyze some of their results, and to enunciate some of the laws to which these results point. For brevity I refer to them as "The Authors." The page numbers in parentheses refer to their paper. 1. The Visible Aspects of Recrystallization.—Plastic deformation in a single direction, as in cold-rolling, draws the grains out in that direction, that is, inequiaxes them, without destroying their apparent individuality. Each grain seems to endure the drawing out without changing either its volume or its orientation materially, for the etching tint of each grain seems to be as uniform after as before the drawing out, save for the "etch bands,"2 shown in Figs. A to D of the Authors' Plate I. As the drawn-out metal is heated progressively, on reaching a certain "disintegration temperature" these old grains seem to break up, probably into submicroscopic fragments, and this apparent disintegration is followed by a coalescence of these fragments, so that their number decreases and their size increases, till they may become very much coarser than before the drawing out. Ruder3 coarsened the grains of silicon steel in this way till single grains reached an area of 50 cm. (7.75 sq. in.). The deformation itself probably breaks each grain up into many very small fragments, but the fact that each grain continues to have a uniform etching tint different from that of its neighbors indicates that these fragments retain the initial common orientation of the grain. The apparent disintegration which we observe, then, is rather an abandonment of community of orientation for individuality of orientation, by the fragments formed earlier during the deformation. Hence we

Jan 1, 1917

By Ben-Zion Weiss, Melvin R. Meyerson

FATIGUE crack propagation in some aspects can be viewed as being a result of localized plastic deformation concentrated near the tip of the crack.' Deformation is influenced by microstructure which in turn can affect the fatigue crack growth significantly. To study the effect of localized microstructural changes on fatigue crack propagation, cantilever-type notched specimens were subjected to cyclic bending at constant load amplitude. The extent of fatigue crack propagation was obtained by measuring the length of the crack by periodic observation with a microscope on the polished side of the sample,2"4 and these data were used in preparing the curves in the figures which follow. Figs. 1 (solid lines) and 2 show experimental results obtained in austenitic type 302 stainless steel containing higher than standard carbon (0.24 pct C) and 310 stainless steel. Both steels were homogenized for 8 hr at 1120°C and then water-quenched in order to prevent the precipitation of carbides. Fig. 3 shows experimental results obtained on 9440 low alloy steel in the normalized condition and in the quenched and tempered (580°C) condition. Indications of microstructural influence can be seen in scattering of the experimental data. The degree of scattering depends on both the microstructure and the amount of cyclic stress applied to the sample. For microstructures which are more easily deformed plastically, e.g., normalized or annealed as compared with hardened and tempered structures, the scattering is greater. Similarly, higher stresses produce greater scattering in the results, regardless of whether the cause of the stress increase is due to initial loading conditions or the growth in crack length. Scattering is believed to be a result of partial redistribution in the stress field around the tip caused by plastic deformation. This opinion is borne out by much smaller scattering of experimental results in the austenitic 302 high-carbon (0.24 pct C) steel. In this steel, plastic deformation in the neighborhood of the crack causes partial mar-tensitic transformation of metastable austenite,5 which in turn restrains further plastic deformation. As a result, the stress redistribution around The tip is reduced. Partial martensitic transformation was confirmed

Jan 1, 1969

By E. C. Pechin

Secretary's Note.—Mr. Bramwell was born in 1846, at Liverpool, England, his father being English, and his mother American, of Scotch ancestry. Brought to the United States in early childhood, he was educated at the public schools and graduated from the College of the City of New York. He subsequently studied his profession as a mining engineer at the Royal School of Mines, Berlin, Prussia, and was for a short time in the service of the Prussian government at Konigs hutte. Returning to this country in 1871, he was employed during succeeding years as mining engineer, chemist, and manager at various fnrnaces and iron works in Pennsylvania and the Southern States until, in 1883, he became actively engaged in the development of the Flat-Top coal-field of southwestern Virginia, in which region the town of Bramwell was named after him, in recognition of his services as a pioneer of the great industry which he did so much to promote. After prolonged and fluctuating ill-health, he died, July 13, 1894, at Paris, France. At the opening session of the Bridgeport meeting, the Secretary announced his death, and Mr. Pechin, being called upon, made the remarks reported below.— R. W. R. Widely as Joseph Herbert Bramwell was known, I. think I knew him somewhat better than any one else. After finishing his studies abroad, he came to this country, a very young man, seeking an engagement. This was about the beginning of 1871, as I remem ber, and I was then trying to make iron at Dunbar, at the foot of the Allegheny mountains, in Fayette county, Pa. His engaging manner and charming personnel, combined with his letters of intro duction, made me more than satisfied in securing his services as my chemist. At the outset we lived a sort of pioneer life, and for many months he was practically a member of my family, and the affection then formed remained unbroken to his dying day. I vividly recall the circular of April, 1871, signed by Eckley B. Coxe, R. P. Rothwell and Martin Coryell suggesting a meeting at Wilkes-Barre, Pa., for the purpose of forming the American Institute of Mining Engineers. At the last moment, prevented from going, I sent Bramwell; and I am more than glad that his name appears among those whose sagacity established an Institute that has become illustrious by reason of its numbers, the character of its members, its invaluable papers and its enormous influence for good

Jan 1, 1895

By Rossiter W. Raymond

The lead-mining industry of Missouri, as of other parts of the Mississippi basin, appears to have been paralyzed by the shock of competition with the mines of the States and Territories further west. It is difficult for establishments working galena ores, too poor in silver to pay for extracting it, to compete with the producers of the argentiferous carbonates of Utah and Colorado, who look upon their lead as a by-product of little value, and are satisfied if it pays the charges of transportation, or the loss in smelting for silver. But simultaneously with this depression in the lead-mining of the Mississippi States, the development of new zinc deposits has caused a notable revival in the mining of zinc ores. There has recently been a sndden and strong rise in the price of' spelter, which, though probably to a large extent temporary in character, has given important encouragement to prospectors and miners. Apart from the general recovery of business in the United States from the prostration of the past six years, the cause of the advance in spelter has been the falling off in production, resulting from the shutting clown of several leading Western works,—the Martindale and Carondelet works, at St. Louis, and the Consolidated Company in Kansas. At the same time, the principal works at La Salle, Ill., were undergoing repairs, and produced less than usual, while the sheet-mill of Matthieson & Hegeler, at that place, was running as usual; so that this concern, ordinarily a large producer, entered the already depleted market as a buyer. These causes are in their nature transitory, and it is quite likely that spelter will be superabundant in a few months. Having lately received from a well-informed Missouri correspondent some samples of the zinc ores and a general account of the ore deposits of Southwest Missouri, I have thought that it might interest the Institute to examine these specimens, and to hear such brief statements concerning the region as I am able, upon good authority, though without the aid of any personal investigation 011 the spot, to give. The zinc-bearing territory in Southwest Missouri appears to be very extensive; that is to sap, zinc ores have been mined during the last year at various points throughout a district extending from the eastern edge of Kansas, west of Joplin, to a line some 50 miles east

Jan 1, 1880

By Robert Ignouf

MY REMARKS, which I feel highly honored in being invited to make, shall be limited to a consideration of -the mining and metallurgical problems involved in this question; in fact, these problems alone might well be considered as suffi¬cient to explain the existence of the Ruhr situation and to justify the steps taken by France in this affair. Comparing the France of pre-war days with the France of the period following the Versailles treaty indicates clearly how the Ruhr question is a direct consequence of the new economic conditions which have prevailed in Europe since peace was signed between the Allies and Germany. Before the war, in 1913 for instance, the French out¬put of iron ore was 22,000,000 tons. This output, which decreased considerably during and immediately after the war, has about regained its previous level. It should be noted that 96 per cent. of the production of iron in France is obtained from the eastern part of the country, in the districts of Nancy, Briey,. Longwy, Metz, and Thionville. This concentration of practi¬cally all the iron mines into one district creates a very serious problem for France, involving strategic and economic consequences. The strategic problem arises from the fact that the whole iron resources of France lie close to the frontier of Germany, at a distance of only a few miles; hence it is of the' utmost importance for France to be assured that her eastern frontier shall not be at the mercy of a sudden invasion by an army which could in a few hours deprive France of her iron resources and render her powerless to defend herself. Invaded by her Eastern neighbor three times during the past 100 years, France has had dire experience with what such invasions mean for her. Wherever the German armies have passed, they have left an inefface¬able trail of desolation. France, therefore, intends to protect herself by sufficient and efficient guarantees against renewed invasions of her territory.

Jan 5, 1923

By W. L. Zeigler

THE tailing from the gravity concentration plant of the Hecla Mining Co., Gem, Idaho, was former-ly loaded into railroad cars to be used for ballast, highway surfacing material, or concrete work, or dis-posed of by an aerial tramway, which dumped it into another gulch. The tailing averages about 1.2 per cent lead and 0.8 oz. silver to the ton. Further treatment for the recovery of the valuable metals had been investigat-ed several times with results that did not warrant the construction of a retreatment plant; but early in 1925 experimental work was again started with small tests, and later with a 10 tons per day flotation test plant. This showed that the tailing could be ground and treat-ed by flotation at a profit. It was decided to make the new flotation plant an addition to, rather than incor-porate it in the old concentrator, the original capacity of which had been increased many times, the present arrangement being by no means ideal. The old mill buildings were of timber and board contruction, creat-ing more or less fire hazard, so all new construction was made fireproof. Construction work on the new plant was started in the middle of July, 1925, and the plant was placed in operation early in December. The building is of concrete and steel with hollow tile walls, steel sash and corrugated asbestos roofing. The ball-mill floor has for its foundation a cellular reinforced concrete block about 14 ft. deep, extending to cemented river gravel, and the cells filled with tailing to give weight to the block. Overhead, built into the steel frame, is a 5-ton hand-operated traveling crane. With both ball-mills running fully loaded, there is not the slightest trace of vibra-tion in the foundation or building. The flotation floor is of reinforced concrete beam construction and spans Canyon Creek. Adjustable plunger feeders are arranged under the the tailing-bin, which feed by gravity to a 14-in. eleva-tor, discharging to a short 18-in. conveyor. From the conveyor the tailing is discharged into a dividing hop-per, from which point the plant is arranged in three separate units. Each unit consists of a ball-mill in closed circuit with a classifier, and an 8-cell flotation machine and has an approximate capacity of 300 tons per 24 hours.

Jan 8, 1927

THERE are many kinds and varieties of earth* that are used for the loam compositions for making the moulds for casting bronze, brass, or other metals. Since this is a very necessary thing, you must try to have the best kind and one that resists the fire well. It must be disposed to receive the metals well, must make a neat casting, and must not shrink or break with cracks on drying or baking. Aside from actual trial, I believe that there is little that can help you, since the clay in itself has no color or [77] visible sign that I know of to show how satisfactory it is. Understand that their colors are no less varied than their natures, for, as you see, one is white, one black, one yellow, and another red. Perhaps none will show by its color that it is good for this work, yet it might also happen that all colors are good. But all that I can say about it is to show you by their effects how the good ones should be in their nature; and so by trial you will choose among those that come into your hands. All earths are either sandy or tufaceous, either argillaceous and lean or pasty with unctuous viscosity. The lean ones make the loam powdery and without toughness and hold together poorly by themselves when they are dry. The unctuous and viscous ones shrink and break. Becoming smaller by shrinking, they often twist the forms and they stick together badly. They swell up again in the fire and produce works that are neither exact-a nor neat. Therefore, the good ones must be those that are neither unctuous nor lean, neither wholly tender nor rough, and with a grain that is fine, without pebbles or shells. When applied to a work they easily dry out without fractures and, when dry, hold their own shape well. Above all they must resist the fire well. These are usually of a yellow or red color, but whatever their color may be, the color does not force me to say that the yellow are really better than the black, and the red than the white or gray. Their quality depends on what trial shows. Thus, to summarize, whenever you have to do this work, you must try to have the best clay that you can, because it is to be the foundation of

Jan 1, 1942

By H. L. Talbot, R. S. Young, A. Golledge

THE differentiation of, oxidized forms of lead from lead sulphide in complex products by chemical analysis is of considerable importance to certain mining and metallurgical companies. A method for the separation of "oxide" lead, (non-sulphide), from sulphide lead, based on the solubility of the former and virtual insolubility of the latter in concentrated ammonium acetate solution, has been employed for many years in the industry.1,2 Unfortunately this solvent, while satisfactory for lead sulphate and carbonate, is without action on lead phosphates or vanadates, with the result that at Rhodesia Broken Hill Development Co. where these latter minerals are present, they have counted as sulphides when using this procedure. For the control of operations in the new concentrator it became necessary to have a closer check on sulphide lead than could be obtained by the only current method of differentiating this constituent from oxidized lead. For some time this problem has engaged our attention, and after a large number of solvents had been tried, the following procedure was evolved. Like all methods depending on selective solution in a complex material, it will not give 100 pct separation into oxide and sulphide portions. Because of the ease of oxidation of galena this is almost an impossibility. Our procedure will, however, give a recovery of 99-100 pct of the oxidized lead, contaminated with very small quantities of lead sulphide, and a corresponding recovery of 97-98 pct of the sulphide lead. The procedure is simple and reproducible, and gives results quite satisfactory for routine control purposes. EXPERIMENTAL For this work standard samples of the following pure lead minerals were obtained from Broken Hill, Northern Rhodesia. These were hand-picked specimens, and were carefully ground to - 200 mesh. Since it was realized that a fair-sized sample would be necessary for this work to thoroughly test the large number of procedures we had in mind, a compromise had to be effected between a few grams of very pure minerals and a large sample of lower grade material. For this work the presence of gangue was not deleterious, so long as the oxide lead minerals did not contain lead sulphide, and vice versa. Table I shows the lead minerals used, and their total lead content, both theoretical and actual. Apart from the pyromorphite the minerals are seen by analysis to be quite pure. Many solutions were tried in an effort to dissolve the oxidized lead minerals and

Jan 1, 1947

Discussion of the paper of Edward Keller, presented at the Butte meeting, August, 1913, and printed in Bulletin No. 80, August, 1913, pp. 2093 to 2115. GEORGE L. HEATH, Hubbell, Mich.:-In the admirable paper quoted, is mentioned the date when the 5-g. assay was first proposed is a standard for refined metal. The author omitted, however, to mention a later paper,1 which contains a proposal for the use of a single " stock solution " of mixed acids, which has been adopted by the American Brass Co. and others. The suggestion is made that such a scheme would save time and manipulation with converter metal. Unless the low current of 0.5 ampere and the large excess of nitric acid, recommended by Mr. Keller, have been proved necessary to hold up antimony-and selenium, the use of the mixed solution which we have adopted will permit the use of 1 ampere of current per square decimeter of cathode surface, until the solutions have become colorless. At this point, we wash the covers and reduce to 0.5 ampere to prevent oxidation or contamination. The ordinary split electrodes have been used. With strong current, at least, the use of an excess of nitric acid tends to the neutralization of the solution before the end of the period, from the formation of too much ammonia., and this is detrimental. Antimony deposits more easily than arsenic, and tends to drag the latter dorm with it. There must, be some such explanation for the fact. that we have made hundreds of deposits of pure cathodes from Lake material containing no antimony, but as. high as 0.5 per cent. of arsenic. By placing an assay beaker in the Frary solenoid, or rotary device, a cathode deposit with less than 0.01 per cent. of impurity is easily obtained in one operation from 3 g. of whitneyite, carrying from 7 to 10 per cent. of arsenic. In such a case, it is necessary to test several times for end point and remove the plate as soon as no brown tint appears (at once) when hydrogen sulphide water is added to the test portion on the spot plate. Edward Keller's proportion,, of solvent figure out as 14 cc. of strong nitric to 11.4 cc. of sulphuric acid (1.84).

Jan 11, 1913

By William P. Blake

THE occurrence of enormous masses of pure copper has given the mining district of Lake Superior worldwide reputation. The first masses brought from there excited great attention, and directed the notice of the mining world to the few particular mines from which they were taken. It may not now be generally known that nearly all the veins which are worked, and which cut across the trap ridge, contain mass copper, and that large masses are continually being raised from them. The largest continuous mass which has been taken out was probably that from the Minnesota, in 1857, which is variously stated as weighing 420 tons and 470 tons. Its length was about 45 feet, its breadth or height 22 feet, and its greatest thickness 8 feet. All such masses are very irregular and ragged in their form and thickness, thinning out generally from a foot to a few inches, and straggling through the vein until they connect with other large masses. This was the character of a mass found in the Phoenix Mine, one of the oldest on the lake, which mass altogether weighed some 600 tons. But this was really a series of masses more or less connected by strings of metal, yet no one large part of it weighed, singly, over 200 tons. A similar series of masses, weighing about 600 tons, was extracted from the Minnesota. Some of the Phoenix masses were four to five feet thick of solid copper. The Cliff Mine has yielded masses weighing from 100 to 150 tons in one piece. One of 40 tons was taken out this year, besides numerous blocks weighing from 1 to 8 tons. This mine and the Central are now yielding mass copper in abundance. It is, of course, impossible to pick, or drill, or to break out such huge masses of solid metal when they are found, by drifting upon the course of the vein. The method of extraction is as follows The miner picks out or excavates a narrow passage or chamber upon one side of the mass, laying it bare as far as possible over its whole surface. It is usually firmly held by its close union with the vein-stuff, or by its irregular projections above, below, and at the end. If it cannot then be dislodged by levers, the excavation of a chamber is commenced behind the mass, and this excavation is made large enough to receive from 5 to 20 or more kegs of powder. In one

Jan 1, 1876

By H. F. Seifert

THE classification and preparation of non-ferrous scrap metals is a subject of interest to every individual and corporation that employs in its processes of manufacture non-ferrous metals and alloys and accumulates in such manufacture scrap from these materials for disposition in the form of turnings, borings, short ends, punchings, clippings, grindings, washings, etc. All too frequently the disposition of such scrap is left to individuals who have no knowledge of the subject either from a metallurgical or a commercial standpoint, or both. Much money is thus lost which with a little more intelligent effort might be saved, and in this day of fierce competition and small profits such an opportunity cannot be ignored. The Mineral Resources of the United States for the year 1925, the latest issue available (Part 1, page 221), gives the secondary non-ferrous metals of certain classes recovered in the United States for the year 1924 to 1925 as shown in Table 1. The figures are tabulated simply to give the reader some idea of the scope of the secondary non-ferrous metal industry. Approximately 900,000 tons of an annual value of $244,000,000 were recovered in the year cited. In 1916 (again using the Mineral Resources of the United States as authority) the tonnage reclaimed was approximately 632,000 tons and the value, because of the high prices for metals prevailing at that time, amounted to $265,000,000. It is not the purpose of this paper to discuss smelting and refining operations or the feasibility of an organization, primarily engaged in some other activity, conducting smelting and refining operations to convert its own non-ferrous scrap metals and alloys. The classification and preparation of these non-ferrous materials is a preliminary step, regardless of what subsequent operations the material may pass through or where these operations may be conducted. For that reason the subject can be readily discussed without any reference to subsequent activities. Space must be provided, with suitable handling facilities, scales and the necessary storage bins where the material may be accumulated. The scope of the activity, the nature of the industry and the tonnage handled will determine, in a large measure, what preparation, if any, it may be desirable to give these scrap materials to put them in the best marketable condition. By preparation is meant largely what, for the want of a

Jan 1, 1928

By Julian Glasser, Arthur Fleischer

THIS country was faced with the possible necessity of utilizing nonbauxitic ores for producing aluminum during World War 11. Construction of four experimental plants to treat such ores by four different processes was authorized1-4 including one at Salt Lake City, for the conversion of Marysvale alunite to metal-grade alumina and potassium sulphate by the Kalunite process.' An aluminum reduction plant in Tacoma, Wash., with a capacity of 50 tons of aluminum per day, was constructed to handle the Kalunite alumina produced at the Salt Lake City plant. The alumina and aluminum plants were Defense Plant Corp. projects operated by Kalunite, Inc. and Olin Industries, Inc., respectively. Kalunite alumina differs from Bayer alumina in several respects, particularly in that the former contains potash instead of soda. The presence of potassium in the alumina feed to the cryolite bath was stated by aluminum technologists to be undesirable under the usual operating conditions of aluminum furnaces. In 1940, about 20 tons of Kalunite alumina produced in the Kalunite pilot plant, were reduced electrolytically in a 10,000 amp furnace at Battelle Memorial Institute and at the aluminum reduction plant at Neuhausen, Switzerland. These relatively small scale tests indicated that satisfactory commercial aluminum could be produced in a practical manner. The Tacoma plant reduced about 2,500,000 lb of Kalunite alumina to produce more than 1,200,000 lb of aluminum, in accordance with the accepted gractice for reducing Bayer alumina. It was desired to determine whether commercial aluminum could be produced on a large scale and to establish the technological differences in commercial reduction practice between Kalunite and Bayer alumina. In the course of collecting data, the effects of the abnormal chemical composition and physical properties of Kalunite alumina were observed wherever possible. Tacoma Reduction Plant The reduction plant had two pot lines, each with 120 cells in series, operated at approximately 32,000 amp. Each cell was provided with a single rectangular Soederberg continuous anode, with side-pin electrical contacts. Each cell was hooded and ventilated through a stack connected to 14 other cells, with 25,000 cfm blower. When the plant started operation in 1942, all of the cathode linings were of block construction. Rebuilt cells were of rammed paste construction.

Jan 1, 1952

By Charles Himrod

IN presenting this paper it is not intended to enter into any discussion of the theory of the blast-furnace, but simply to give the results of a number of determinations of CO and CO2 in furnace gases, and some conclusions arrived at, and the result of an experiment made in hopes of increasing the production of CO2. In the first table, on page 198, is given the averages of a number of analyses of gases of different furnaces, and in the second, of two stacks of different sizes under varying circumstances. The conclusions arrived at as to what influences the production of CO2 are as follows: 1st. Kind of fuel. Bituminous coal not permitting the formation of as much CO2 as coke. 2d. Kind of ores. Rolling-mill cinder being much less favorable to formation of CO2 than hematite ores, magnetites probably less than hematites, but more than cinder. 3d. Length of time of contact of gases with ore. Small furnaces and too rapid driving both waste fuel by not allowing enough time for conversion of CO into CO2. 4th. The temperature of blast. The hotter the greater the quantity of CO2. 5th. Regularity of passage of stock through the furnace. The hanging and dropping of the stock, however slight, seems to pack it more solidly in the hearth, increases the resistance to the blast, and decreases the quantity of CO,. These are the most common causes noticed which influence the production of CO2, unless we except the one that might be considered the greatest of all, the one that has spoiled so many tons of iron, and the amiability of so many furnacemen, the one most dreaded, the great unknown. It would seem a self-evident proposition that the better the contact of the gases with the ore and the less with the fuel, the greater would be the amount of CO2 produced. To prolong the contact of gas with the ore furnaces are enlarged with advantage, and pains are taken to charge the stock in such a, manner as will best promote it. Still, with furnaces of most approved construction and most skilful manipulation, the gases leave the furnace so rich in CO, that not half the heating power of the fuel is obtained, and the highest

Jan 1, 1877