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By W. McA. Johnson

The work of J. E. Johnson, Jr., on the effect of small amounts of oxygen in cast iron in increasing its strength and resistance to shock, is of interest from the technical and scientific standpoints. The following exposition of the theory carries Mr. Johnson's explanation further and, in my opinion, will disclose certain related phenomena and perhaps shed further light on the subject. Some facts, so novel in character as to excite incredulity, have been accepted finally by many metallurgical authorities. One of these is that oxygen to the amount of 0.060 per cent. in cast iron gives it a breaking strength of 3,500 Ib. per square inch, compared with a breaking strength of 2,500 lb. per square inch in cast iron of the same composition with respect to elements other than oxygen, but having only 0.010 per cent, oxygen. The "oxygenated" cast iron, even if made in a coke furnace, has properties equivalent to the best charcoal pig iron. Mr. Johnson has also established the fact that the variations in combined carbon cannot account for the results. By microphotographic evidence he has shown that in oxygenated cast iron the graphite particles are dense and spherical, while in well-deoxidized cast iron the graphite particles are flaky and leaf-like. It is just a matter of common sense that iron of the latter structure is much weaker than the former. The analytical and microscopical work has been done with care and there is no doubt in my mind about the accuracy of the experimental results. In this paper then, we shall consider it as accepted that small amounts (0.060 per cent.) of oxygen increase the strength of cast iron by producing a structure in which the round and solid particles of graphite are surrounded firmly by a principal matrix of ferrite. The purpose of this paper is to show that the reason why the graphite particles are round is founded on the fact that the particles of any precipitate are made denser and harder by the presence of a reagent having a solvent action on the precipitate. This can be stated as a law, although I have never seen it so given in any textbook. Let us then, for purposes of illustration, give a concrete example and develop the theory of such mechanical and chemical actions as occur: It is well known that, by a soluble sulphate, a barium salt in solution is precipitated in a dense and easily filtered body when the solution is

Jan 1, 1916

By F. D. Weeks

The Salida smelting plant, owned by the Ohio & Colorado Smelting & Refining Co., is situated at an altitude of 7,000 ft., about 2 miles west of Salida, Colo., and 215 miles southwest of Denver. Salida is a division point on the Denver & Rio Grande R. R. where the narrow-gage line serving the San Juan region joins the through standard gage. This narrow-gage road also serves the coal and coke companies at Crested Butte and the lime quarries and mines at Monarch and Garfield, which are only 20 miles distant. These facts, together with the possibility of an appropriate site, doubtless influenced the builders to locate at this point. The smelting plant and yard occupy 80 acres of land on the northeast side of the Arkansas River. The slope of the ground is such as to permit of the ore being delivered by cars at a point 25 ft. above the feed floor; and in front of the furnaces there is a fall of 50 ft. for the slag dump. A recent purchase of river bottom has provided slag room for 20 years. As originally constructed, the plant consisted of four blast furnaces, 40 by 144 in. at tuyéres, two copper-matting furnaces, dust chamber, stack, power plant, blowers, sampling works, and one railroad trestle about 16 ft. from bed floor to rails. Later 12 hand roasters of excellent design were added and the practice of smelting leady matte in a copper-matting furnace was abandoned. Since the plant passed under the present management the equipment has been increased so that the following is a brief description of the same at this date: Two lead blast furnaces, 48 by 180 in. at tuyéres, 17 ft. 6 in. tuyéres to feed floor; four lead blast furnaces, 40 by 144 in. at tuyéres, 17 ft. 6 in. tuyéree to feed floor; two copper-matting furnaces, not used; blast-furnace flue, 900 ft.; area 245 sq. ft.; blast-furnace stack, 150 ft.; above furnace floor, 190 ft. When smelting a 500-ton charge per day, the gases approximate 63,700 cu. ft. per minute and travel at a speed of 260 ft. per minute. The slag is hauled by a 10-ton electric locomotive; the slag shells and settler refuse are elevated to a bin on the feed floor by an electric hoist. The matte is tapped in 1,600-lb. cakes and loaded into cars

Jan 1, 1916

By John Blatchford, William T. Hall

The product obtained by roasting stibnite is likely to contain some unoxidized antimony trisulphide and a mixture of antimony trioxide and antimony tetroxide. It was desired to determine, as accurately as possible, the condition of the antimony as well as the total quantity present. Attempts were made to separate the trisulphide and the two oxides by methods based upon their varying solubilities in different solvents, but no satisfactory results were obtained in this way. It was found, however, that by determining the total antimony content, the antimony present as trioxide, and the antimony present as trisulphide, a good idea of the chemical composition of the roasted product could be obtained. The total antimony was determined by dissolving the sample in concentrated hydrochloric acid, reducing the antimony entirely to the trivalent condition by means of hydriodic acid and eventually titrating the antimony back to the pentavalent condition by means of iodine in the presence of sodium bicarbonate1* The details of the procedure are as follows: Weigh out 0.25 gram of the roasted product into a trapped flask (Fig. I), add 4 grams of powdered tartaric acid, 2 grams of potassium iodide and 40 c.c. of concentrated hydrochloric acid. Boil the solution gently f;r 5 min. and then cool to room temperature by shaking the flask while holding it under running water. Then, very carefully discharge the iodine color by the cautious addition of 0.05-normal sodium thiosulphate solution, adding a little starch toward the last. Nearly neutralize the acid with ammonia solution, but leave the solution distinctly acid. Pour the slightly acid solution into 200 c. c. of water containing an excess of sodium bicarbonate and titrate to a permanent blue color with standard 0.1-normal iodine solution.

Jan 1, 1916

By James F. Kemp

The iron ores of southeastern Cuba present a subject of exceptional geological interest. Their relations to the inclosing rocks are varied and in some cases unusual. The problem of their geological origin is not a simple one and for this reason has occasioned some divergence of views among previous observers. Profiting, however, by these earlier studies and by more detailed field and laboratory observations as well as by the experience of added years of mining, much more satisfactory descriptions and interpretations can now be prepared. Previous Work In 1884, interest in the southeastern coast of Cuba became active among the consumers of iron ore along the Atlantic seaboard of the United States. Active development was undertaken by the Juragua Iron Co., Ltd., a corporation affiliated with the Bethlehem Steel Co., and the Pennsylvania Steel Co. The property of the Juragua company, which is situated about 15 miles east of Santiago, was studied in June and July, 1884, by Dr. James P. Kimball, at the time Professor of Economic Geology in Lehigh University and one of the best known of American mining geologists. Dr. Kimball published on his return two papers, the first important descriptions which we have of the local geology.1 It is

Jan 1, 1916

By W. L. Honnold

Owing to a unique labor situation and other unusual circumstances, the mining methods of the Rand are hardly comparable with practice elsewhere. They are of considerable interest, however, and their importance is increased by the magnitude of the' industry with which they are identified, gold of $166,186,000 value, say 40 per cent. of the world's production, having been recovered from 25,701,954 tons of ore milled in 1914. In the sketchy remarks to follow, attention will first be drawn to some influential factors of general bearing, after which certain features at two mines of the writer's group will be referred to. It is hoped that engineers associated with similar or different conditions and practice will contribute to broaden the discussion. Before proceeding, recognition is due the generous assistance of C. E. Knecht and F. A. Unger. General Observations The principal mines are situated along the northern limit of an irregularly shaped geological basin, and extend in almost unbroken continuity over a distance of about 60 miles. As the boundaries are defined by vertical planes there are both "outcrop" and "deep-level" properties, all differing in size and in technical features. So much has been written in regard to the geology of the Rand that only a few pertinent features need here be recalled. Briefly, the ore deposits are profitably enriched portions of certain conglomerate beds or "reefs," rarely more than three, which are interposed between quartzite, or quartzite and slate, and lie so closely together that where more than one are mined they are dealt with as a working unit, although they may be stoped separately. Their collective extension is locally referred to as " the Reef." Along the outcrop the series dips more or less steeply, and in depth it flattehs, the latter tendency being most marked in the Far East Rand, where the bottom of the basin is nowhere so deep as to be inaccessible. Dikes and faults are a common feature throughout. They are,

Jan 1, 1916

By Frank Fahrenwald

I. INTRODUCTORY METALLURGICAL research has discovered many an alloy possessing properties not combined in any single metal, and progress still consists chiefly in the investigation and utilization of alloys. In the case of iron, the demands of automobiles, high-speed machines and high-duty engines have led to the production of special iron alloys which will meet any reasonable specifications in that field. In like manner, the bronzes, brasses and other alloys of copper have been brought to remarkable perfection, and for nearly every industrial purpose some alloy has been found more suitable than the pure metal. Less complete success has attended the attempt to find substitutes for gold, platinum, and the other precious metals. Indeed, it is not likely that a material can be produced which will possess all the properties of any one of them; yet it is reasonable to hope that, for any given use, the properties required may, be found in some less expensive material. Thus, in incandescent electric lamps, the wire passing through the thick glass neck of the bulb was, until recently, almost universally made of platinum, for the single reason that no other known material, suitable as a conductor, had the same coefficient of expansion as glass. But a comparatively recent investigation of the iron-nickel series' has shown that alloys of those metals may be produced, the coefficient of expansion of which can be accurately controlled between that of iron or nickel and zero. Thus, in that particular industry, a substitute for platinum has been found. According to statistical reports, not only considerable quantities of gold and iridium, but also more than one-third of the annual supply of platinum, are used (and, in the nature of the case, irrevocably lost) by dentists. Platinum is thus employed in several forms. As a thin foil, it serves various purposes for which its high melting point, pliability, chemical resistance, and other properties-including the ease with which

Jan 1, 1916

By Hugh P. Tiemann

The desirability of employing the centigrade scale for the measurement of temperatures is becoming more and more recognized in this country, particularly in view of the fact that this scale is used almost exclusively in technical papers and publications—exclusively in those published in French or German. There are, however, a large number of factories and laboratories which still employ the Fahrenheit scale. It therefore frequently becomes necessary to convert one into the other and the accompanying chart (Fig. I) has been found very convenient for this purpose. The formulas for conversion follow: (1) °F. = 9o /5+ 32 (2) oC = 5 (oF - 32 ) (1) 100° C X 9 900/5 =180 .'. 180 + 32 = 212° F. 5 (2) 5 (212° F - 32) = 5 X 180 =100o C 9 9

Jan 1, 1916

By J. Norman Bulkley

As electrical power is used to a greater extent on the Rand than in any other mining center, it is thought that a short description of the methods used and results obtained may be of interest. In comparing Rand practice with that on other fields, several general factors should be borne in mind: First, the mines are all working on large tonnages of low-grade ore (average $6.50). The annual crushing rate is about 27,000,000 tons. Second, the mines are all grouped under the financial and engineering control of mining houses, so that the results of a group, rather than an individual mine, are open for study, and practice follows more uniform lines. Third, coal of fair quality (about 12,000 B.t.u.) can be obtained for $2 to $3 per ton. Good supplies of condensing water are, however, scarce, and have to be carefully stored. Power Supply Power is supplied by the Victoria Falls & Transvaal Power Co., Ltd., and its subsidiary, the Rand "Mincs Power Supply Co. Both companies, which are practically one from an engineering point of view, operate the following stations, all fecding into the same system. Capacity Installed Turbo Capacity Installed Steam-Tur- Station Alternators bine Dliven Centrifugal Compressors Brakpan............... Two 12,500-kw. sets Two 3,000-kw. sets Simmer Pan............... Six 3,000-kw. sets Rosherville................ Two 11,000-kw. sets Six 3,500-kw. machines Five 9,600-kw. sets Three 7,000-kw. machines Verckniging ............. Two 9,600-kw. sets Two 12,000-kw. sets 162,200-kw. sets Total steam-tkbine capacity installed, 204,200 kw. There are also at the Robinson Central air station six motor-driven centrifugal compressors, each of 3,500 kw., equal to 22,000 cu. ft. per minute at 120 Ib. capacity at 3,000 r.p.m. Power is supplied from these stations as 50-cycle, three-phase current generally at 40,000 volts,

Jan 1, 1916

Jan 1, 1916

By Frederic K. Bunton

The smelting plant of the British Columbia Copper Co. at Greenwood, B. C., now closed because of the decline in the price of copper due.to the European war, is of special interest to metallurgists for several reasons. It was successfully smelting in blast furnaces the lowest grade copper ore of all plants in America. In order to do so, it had to run at very hig~; efficiency, which necessarily required a large tonnage per square foot of hearth area, together with the minimum amount of labor and other costs. The furnaces smelted daily 2,250 tons of ore (6.62 tons per sq. ft. of hearth area), carrying 0.85 per cent. of copper, at a smelting cost of 81.18 per ton. The entire plant required 130 men to operate it and keep up repairs, showing a labor efficiency of about 17.5 Dons per man per day. In the present paper, the method of obtaining these results is shown. Most of the information contained in this description I obtained as Assistant Superintendent, during the two years preceding the present shutdown. Other sources of information to which I am indebted are: The British Columbia Copper Co., Ltd., by ALfred W. G. Wilson, in The Copper Smelting Industry of Canada; Greenwood Smelting Worlrs, by J. E. McAllister, Engineering and Mining Journal, vol. xci, pp. 1011 to 1015 (May 20, 1911); and Description of the Copper Smelter of The British Columbia Copper Co., by W. L. Bell, in Transactions of the Canadian Mining Institute, vol. xvi, pp. 151 to 154 (1913). II. Location and Connections The smelter of this company is about half a mile south of GreenGood, in the Boundary district of British Columbia, on the Columbia & Westerg branch of the Canadian Pacific Railway. It was originally built in 1900 by Paul Johnson, to handle 600 tons of ore a day in two small furnaces of 300 tons capacity; but in 1907 the old furnaces were torn out, the plant was remodeled, and three blast furnaces, 48 by 240 in. at) the tuyéres, were installed. In 1910, two of these were enlarged to 51 by

Jan 1, 1916

By Walter O. Snelling

Continuation of the discussion of the paper of WaltER o. SnelLIng, presented at the New York and San Francisco meetings, February and September, 1915, respectively, and printed in Bulletin No. 100, April, 1915, pp. 695 to 704 and Trans., li, pp. 657 to 671. A. C. McLaUghlin, San Francisco, Ca1.—It seems to me that Mr. Snelling's work, if it is what it appears to be, is one of the most important contributions to petroleum technology that has been brought out in the last 25 years. If his end products are what they seem to be, he has made. a discovery, which, in my judgment, will revolutionize the petroleum industry. As to industrially carrying out these processes I do not think there is the slightest difficulty, for I have had some work recently where crude oils were handled successfully at a temperature of 750" F, and pressure of 750 lb. per square inch. William A. WillIams, San francisco, Ca1.—What metal did you use, Mr. McLaughlin? A. C. McLauGHILIN.—Ordinaly seamless tubing, extra strong. William A. Williams.—Tests were made in Dr. Rittman's experimental plant at Pittsburgh, upon samples of residuum furnished by Mr. Bell, the residuum being the remainder of the crude oil after the lighter fractions had been removed. After the material had been run through Dr. Bittman's plant, yields of from 43 to 58 per cent. of material, boiling under 150" C. were obtained. The distillate had a peculiar odor, due to the presence of benzene, toluene, and other unsaturated hydrocarbons. W. N. Best, New York, N. Y.—What is the difference between the Rittman and the Snelling process or system? William A. Williams.—he fundamental difference is that reactions take place under relatively higher temperatures and pressures and with none of the oil present in its original or liquid state, while the reactions secured by Mr. Snelling and other investigators have been from the liquids in the presence of some vapors. Under such conditions as Mr. Snelling operates, the temperatures and pressures used are limited and the flexibility in the control of these variables is extremely limited when compared with the Rittman process.

Jan 1, 1916

By Arthur F. Taggart

The following conclusions on the work of the Hardinge mill are based on data furnished to the writer by the Hardinge Conical Mill Co. in the form of the mesh cards hereto appended. Energy units (E. U.) and relative mechanical efficiencies (R. M. E.) are computed by the "volume method" of Stadler.' Screen apertures used are the average apertures of testing screeps of the meshes given. i . Card 122. June 28, 1912. Vipond Porcupine Mines Co., Ltd., Schumacher, Ont., Canada. Ore from mill bin. Gangue, quartz and basalt. 4.5 ft. by 13 in. ball mill, Capacity, 48 tons per 24 hr. Charge, 4,000 lb. balls. Speed, 33 rev. per minute. Horsepower, 15 to 17. Water, 100 per cent. by weight. (50 per cent. ?) Product deslimed and oversise reground in pebble mill, see Card 113. Feed to mill through 2-in. mesh. Mesh Feed, Per Cent. Discharge, Per Cent.. + 1in. 5.50 ....... + 3/4. 28.00 + in. 30.00 ...... + 1/4in. 19.72 ....... + 10 10.87 2.10 + 20 2.42 8.00 -20 3.51 .....,. + 40 ....... 22.68 + 60 ...___ 10.50 4-80 ....... 11.80 + 100 ....... 3.90 -100 ....... 40.15

Jan 1, 1916

By James A. Barr

When phosphate mining operations first commenced in Tennessee the loss of both high- and low-grade material was large, because of the crude hand methods employed. Practically all rock smaller than 2 in. was thrown back into the pits and covered with the overburden of subsequent openings. With the advent of mechanical washers and other mining machinery, together with the abolishing of the contract system, the waste has been steadily diminished, until at present only the finest sand is being rejected; and even this is not lost, for in the majority of cases the water, slime, and sand from the washers are run into large settling ponds where practically all of the phosphate-bearing material is caught and stored for future use, or until it is profitable to ship lower grades. The latest equipment used to recover the fine and lower-grade phosphate sand includes Dorr thickeners, settling tanks, Dorr classifiers and settling ponds served by locomotive cranes with clam-shell buckets, making it possible to save material averaging as fine as 200 mesh. TO attempt saving finer material would necessitate a different type of drier, which might be patterned after the multiple-deck roasters. The present tailing runs from 15 to 25 per ccnt. of available phosphate, which, howrever, as mentioned before, does not represent a total loss to possible future operations. The demand of the consumer is the greatest factor in regulating the recovery at present. In the mining of blue rock, which does not require washing treatment, the main waste material is the capping of low-grade kidney formation, generally used for backfilling of the rooms. By reason of the introduction of improved methods in the early stages and the re-mining of territory skimmed over at the start, at the present time the net result of all operations is a comparatively small loss. One of the largest temporary losses in the industry occurs in the Florida pebble district where the recovery of phosphate runs about 25 per cent. The loss is classed as temporary because most of the tailing has been collected in ponds and piles, where it may be reclaimed in the future should developments or improved processes warrant. In cleaning the Florida pebble phosphate, aside from the elutriatisn

Jan 1, 1916

Jan 1, 1916

Jan 1, 1916

By William Hall

THE product obtained by roasting stibnite is likely to contain some unoxidized antimony trisulphide and a mixture of antimony trioxide and antimony tetroxide. It was desired to determine, as accurately as possible, the condition of the antimony as well as the total quantity present. Attempts were made to separate the trisulphide and the two oxides by methods based upon their varying solubilities in different solvents, but no satisfactory results were obtained in this way. It was found, however, that by determining the total antimony content, the antimony present as trioxide, and the antimony present as trisulphide, a good idea of the chemical composition of the roasted product could be obtained. The total antimony was determined by dissolving the sample in concentrated hydrochloric acid, reducing the antimony entirely to the trivalent condition by means of hydriodic acid and eventually titrating the antimony back to the pentavalent condition by means of iodine in the presence of sodium bicarbonate.' The details of the procedure are as follows: Weigh out 0.25 gram of the roasted product into a trapped flask (Fig. 1), add 4 grams of powdered tartaric acid, 2 grams of potassium iodide and 40 c.c. of concentrated hydrochloric acid. Boil the solution gently for 5 min. and then cool to room temperature by shaking the flask while holding it under running water. Then, very carefully discharge the iodine color by the cautious addition of 0.05-normal sodium thiosulphate solution, adding a little starch toward the last. Nearly neutralize the acid with ammonia solution, but leave the solution distinctly acid. Pour the slightly acid solution into 200 c.c. of water con-

Jan 1, 1916

By J. R. Campbell

In order to discuss the subject intelligently, it will be necessary to touch briefly on the forms in which sulphur is supposed to exist in coking coal to be carbonized in beehive or byproduct ovens. Sulphur is known to exist in coal as sulphides and sulphates, as can be determined experimentally. Then there is the so-called organic sulphur, i.e., sulphur in combination with the carbon, 'hydrogen, and oxygen of the coal, about which much has been written, but nothing definitely proven in an experimental way. In fact, so far. as we know, there never has been developed a satisfactory and conclusive method in the laboratory for the direct determination of organic sulphur. For most coking coals, it can be safely assumed that the preponderance of sulphur is in the form of pyrite (FeSz). When exposed to comparatively low temperatures during the coking process, it loses its sulphur according to the following chemical reaction: 7FeS2 = Fe7Ss + 6S. It will be seen, therefore, that six out of the 14 atoms of sulphur (or 42.8 per cent.) are expelled, or volatilized. FelSa remains in the coke as pyrrhotite, or magnetic sulphide of iron; this accounts for the highly magnetic properties of the powdered coke. A more commonly accepted reaction is FeS2 = FeS f S, in which the straight sulphide of iron is produced with 50 per cent. volatilization of sulphur. In the beehive methods of coke making, air is introduced in sufficient amounts to carry on the distilling and the coking processes, and the sulphur is oxidized along with the other volatile products: First, into sulphur dioxide (SO2), known by the pungent andsuffocating odoremitted from the trunnel head; second, into sulphuric anhydride (SOa); and, finally, into sulphuric acid, as it comes into contact with the air and moisture. In a properly regulated draft on a beehive oven, there is never complete combustion of the gases. In other words, the coking process should be carried on in a reducing atmosphere, and a low-grade producer gas kept issuing from the trunnel head. For coals of about 30 per cent. volatile matter, the ratio of air to gas is 3½ to 1. In complete combustion, the ratio is 6 to 1, producing an extremely high temperature

Jan 1, 1916

By D. B. Dowling

The interest which has been aroused in prospecting for oil in the foothills of southern Alberta, and in the oil possibilities of the known gas fields situated in the less-disturbed areas, called for a much closer examination of the structure, thickness, and composition of the underlying rocks of the region than had hitherto been made. The areal geology of the larger part of the great plains was outlined by Dawson, McConnell, and Tyrrell, between 1881 and 1885. The foothill area was not critically examined at that time, owing to the time which would have been required for its proper study and the difficulty that was found in recognizing in the foothills the divisions which had been adopted in the mapping of the formations of the plains. This was due in great measure to the paucity of exposures in continuous sections of the lower divisions of the Upper Cretaceous. Since the pioneer work on the plains was published, the beds which form continuations south into Montana have been critically examined, The Canadian beds arch over the end of a flat anticline which rises to the south; and two sections in Montana, one on each side of the anticline, help to explain some of the difficulties previously encountered in the mapping of the measures found in the foothills. The index map, Fig. 1, shows the area studied, and its relation to southeastern Alberta, and to Montana. Problems of Correlation The principal difficulty experienced in correlation was in connection with the measures known as the Montana group. The beds of this group did not seem to agree in the two Montana sections. They showed the same differences to some extent as are encountered in comparing the foothill measures with those of the plains. The Montana group in Dakota and Manitoba consists of a series of shales of marine origin. In the eastern Montana section two sandstone members appear, leaving a shale member at the top and another

Jan 1, 1916

By Guy W. Crane

The geology of the Tintie mining district, fully treated, would occupy an elaborate monograph. This less comprehensive paper is devoted primarily to the occurrence and origin of the orebodies of the district, while space is given to those phases of the general geology only which seem necessary to a proper setting of such a discussion. The first geological report on the Tintic district to be issued was that by Tower and Smith, published in 1899, in the form of a paper in Part III of the Nineteenth Annual Report of the U. S. Geological Survey, and also as Folio No. 65, Geologic Atlas of the United States. About four years ago a second investigation was begun by Waldemar Lindgren and G. F. Loughlin, of the U. S. Geological Survey. With the exception of a paper on the oxidized zinc ores of the Tintic district by Mr. Loughlin, the results of their work are still unpublished. The writer's acquaintance with the district began in the fall of 1913, and has extended over a period of about two years. 1. Location The Tintic district lies about 65 miles due south of Salt Lake City, and on the west central slope of the Tintic Mountains, a short range, which forms the connecting link between the Oquirrh Mountains on the north and the Canyon Range on the south—the group constituting the first of the Basin ranges to the west of the Wasatch Mountains. The productive portion of the district embraces an area about 6 miles long and 2 miles wide, which is divided between Juab and Utah Counties. Within this area are the towns of Eureka, Mammoth, Robinson, Silver City, and Knightsville. All these points are tapped by the Denver & Rio Grande and S. P., L. A. & S. L. Railways. 2. Topography As in most of the Basin ranges, topographic relief is strong, dropping rapidly from peaks of 8,100 ft. above sea level to elevations of 5,600 ft. in Tintic Valley on the west and 4,500 ft. in Goshen Valley on the east.

Jan 1, 1916

By R. C. Patterson

In view of the circumstance that very few important changes have been made within the last 15 or 20 years in the equipment of rod and wire mills, the description of a new process introduced by Grenville Mellen, of LleweIIyn Park, N. J., to take the place of the present laborious system of producing rods of lead, zinc, brass, copper, aluminum, etc., may be interesting to members of the Institute. This new process consists in the production of cast rods at one operation in a small continuous casting machine. The hot liquid metal is transferred from the melting crucibles directly into an endless chain of mold blocks in the machine, where solidification takes place, and the rod comes out continuously in a solid form at one end as long as the molten metal is supplied. The operation of these mold blocks so as to produce a solid rod of uniform structure constitutes an important part of Mr. Mellen's invention. The machine (Figs. 1 and 2), 12 ft. in height over all, 2 by 3 ft. in horizontal cross-sectional area, and 6,000 Ib. in weight, has a framework of cast iron, holding in position two endless chains of mold blocks, which are made in sections and join in center alignment. The mold orifice is made up of these mold sections, into which the liquid metal flows. The molding chains are composed of steel blocks, grooved on one face to form the molding cavity, and linked together with flexible joints. Each block is carried on four rollers which guide the chain around the end sprockets and carry it in its course through the machine. The accuracy of alignment requisite to the production of a perfect rod in the mold groove formed by these sections is secured by careful machine work, and by building the four ways, down which the alignment takes place, so that two of the sides are fixed permanently, whiIe the opposite sides are held to their position by spring pressure. The guides carrying the molds, while in casting position, are water-cooled square tubes. The length of this machine is somewhat indeterminate. A certain amount of both time and cooling surface is required to solidify the metal jn the mold; and one may either use a long molding chain and run it fast, or a shorter chain and run it correspondingly slower. We have found,

Jan 1, 1916