Search Documents

By E. E. Barker

Although flotation was known to be a successful process prior to 1312, Utah Copper Co.'s ores were not entirely treated by this process until 1923. Experiments had been conducted, of course, prior to that time, and in February, 1917, the first unit of the Arthur plant was changed over from gravity concentration to flotation. On May 16, 1918, the entire Arthur plant was on flotation, excepting the coarse concentrate which was tabled prior to fine grinding. The reasons for the delay in adopting flotation at these plants were, first, that it was decided to await the outcome of the litigation with the Butte & Superior Mining Co., which began with an injunction served on Oct. 3, 1911. This litigation continued for years. Then the demands on the plants in the war years 1916, 1917 and 1918, really precluded any further drastic changes during that period, as all efforts were bent on maximum tonnage, and the cry of tonnage, rather than of recovery, was the major issue. The postwar depression followed these years of intensive production, so that it was not until 1922 that the Magna plant was equipped to use this process. Excepting, then, the amount of coarse concentrate that was tabled, 1923 was the first year when the Utah Copper Co.'s ores were on a full flotation basis. Effect of Flotation on Ore Reserves The object of this paper is to point out the effect that flotation has had on the operations of the Utah Copper Co. In 1905, when D. C. Jackling began making history in copper, the ore reserves were 34,000,000 tons, with a grade of 2 per cent. copper and a mill capacity of 700 tons daily. The point of cut-off between nonprofitable and profitable grades was 1.2 per cent. A few years later the capacity was increased to 6000 tons, when 1.1 per cent. became the point of cut-off. Ore reserves had climbed to 80,000,000 tons, having a grade of 1.9 per cent. In 1911 there was an ore reserve of 300,000,000 tons, with a grade of 1.5 per cent. and a cut-off of 1 per cent.; in 1918 the ore reserve was approximately 400,000,000 tons, and a grade of 1.38 per cent., and point of cut-off 0.95 per cent., based on normal costs and prices. Up to this

Jan 1, 1928

By Frank A. Wardlaw

My subject covers the effect that recent metallurgy has had on operations at the Utah-Delaware mine. This mine is the old Highland Boy mine of Bingham Canyon, Utah, one that has now been in operation for about 30 years. It was first operated by Messrs. Weir and Newhouse as a gold mine, under direction of the Highland Boy Gold Mining Co. The ore was oxidized gold ore and it was treated on the property in a cyanida-tion plant. During the process of development, sulfide ore was found assaying about 5 per cent. copper. It proved to be the discovery of what is yet the largest sulfide copper orebody in the State. The first shipment was made in December, 1896. A smelter was built in Salt Lake Valley to treat this ore and not long after the property was sold to the Utah Consolidated Mining Co. From 1900 to 1924, the mine was a large producer of direct-smelting ore which averaged for many years 4.5 per cent. copper, then later dropped to 2 per cent. copper. During the years 1912 to 1920, direet-smelting lead ore was also mined. In 1918 and 1919 ore developments at depth showed a decided increase in silica content and a porphyry sill so closely associated with the ore that it was impossible to mine the ore clean and it was found necessary to build a copper flotation concentrator at Tooele. The mill had a capacity of 1000 tons a day and made a recovery of 90 per cent., and would have been responsible for a continuance of profitable operations had it not been for a slump in the copper market and the necessary high cost of mining ore on the lower levels. Results of Selective Flotation In 1924 the mine was on the point of a shutadown which had all the earmarks of being permanent; it was being operated at a heavy loss and preparations were made to pull the pumps and salvage the machinery. The plant at Tooele was remodelled for selective flotation by the International Smelting Co. in order to treat custom ores, and incidentally to clean up a possible 10,000 tons of lead-zinc ore which it was thought possible to get in the Utah-Delaware before shutting the mine down. Results were so encouraging that an effort was made to develop more lead-

Jan 1, 1928

By A. M. Gaudin

A great number of hypotheses has been advanced to explain the complex phenomena that are encountered in flotation. In the days of bulk-oil flotation, when a large quantity of oil was employed, it was observed that the oil-mineral aggregate floated above the ore-pulp because of its lesser density. With the advent of froth-flotation this phase of the theory had to be modified to take into account the presence of air bubbles. The reason why the mineral-bearing froth of today rises to the top of flotation cells is, of course, because the air-mineral aggregate, with whatever reagent may be associated with it, is lighter than the pulp. In an attempt to explain froth-flotation the fact has been emphasized that certain minerals are wetted and therefore remain in water; while others are not wetted and therefore become associated in preference with air. Such an "explanation" does not explain facts, it merely states them; for instance, it does not explain why minerals are wetted under some circumstances and not wetted under others. Going somewhat further, the wetting or nonwetting of minerals by water has been ascribed' to differences in the angles of contact between oil droplets and mineral surface at the surface of the minerals. This explanation, however, was abandoned when soluble flotation reagents came to replace the so-called "insoluble" reagents. Another theory, sponsored about 10 years ago by Ralston,² involves the effects of electric charges at the surface of the mineral particles and air bubbles. This theory, unfortunately, was not developed as fully as it seems desirable. The belief that adsorption phenomena might play a preponderant role in flotation3 has been rather widespread for the last 10 years. Unfor-

Jan 1, 1928

By W. J. McKenna

The International Smelting Co. concentrator at Tooele, Utah, first operated on a custom basis for the treatment of lead-zinc-iron ores on Nov. 1, 1924, with a capacity of 500 tons per day. On May 1, 1925, the capacity was increased to 1000 tons per day by the addition of another 500-ton section. The concentrator has since operated on the two-section basis, each treating 500 tons per day entirely by flotation. Laboratory and Test Mill In connection with the concentrator, a well-equipped laboratory and small continuous testing plant are maintained. The flotation equipment in the test plant is of both the Callow and Mineral Separations Sub-"A" types, so that actual concentrator conditions may be duplicated as closely as possible. Purchasing of Ore A new shipper usually sends to the plant a 25 to 50-lb. samplc, representative of the ore to be shipped. Batch flotation tests are made on this sample to determine its amenability to the concentrator treatment flowsheet and reagent combinations. The results of these tests, which are accurately correlated with those obtained at the concentrator, both as to grades of products and recoveries, determine the actual value of the ore as submitted in the ore purchasing department's treatment schedule. Hearty cooperation has existed at all times between the ore purchasing and operating departments in the selection of ores as conccntrator feed. Ores Treated The basic tonnages for mixtures on the two sections come from the Park City and Bingham mining districts. The Park City tonnages arc received by rail and consist of crude ores and gravity mill products. Those from the Bingham district arc received by rail and aerial tramway,

Jan 1, 1928

By A. W. Hahn

Although enormous tonnages of sulfide lead and lead-silver ores are treated by flotation, the products of flotation mills treating oxidized ores of lead and silver are almost negligible. However, each year there is a slight tonnage increase, and it is possible, should a sizeable oxidized lead ore tonnage be developed, the field will become an important one. In 1918, the Shattuck-Arizona Copper Co. erected a plant at Warren, Ariz., for the treatment of oxidized gold-silver-lead ores. Glenn L. Allen worked out the metallurgy for this plant, employing tabling, followed by flotation concentration. The Shattuck-Denn Mining Co., successors to the Shattuck-Arizona Co., now employs straight flotation in the same mill, treating about 200 tons daily. In 1922, the San Diego mill of the American Smelting & Refining Co., in the Parral district, Chihuahua, Mexico, started operating a 600-ton mill on a mixed sulfide-oxide gold-silver-lead ore. Tables and flotation are used here, and about 200 tons of oxidized slime treated by flotation. In 1923, the Eureka Holly Mining Co. started operating a 90-ton mill at Eureka, Nev., using tables and flotation, on an oxidized lead-silver ore. In 1924, the American Smelting & Refining Co, started a 250-ton mill in northern Peru, to treat a high-grade oxidized gold-silver ore with very low lead values. In 1924, the Grand Central Mining Co. started a 200-ton flotation mill on old oxidized gold-silver-lead tailings at Fairbank, Ariz. In 1926, the Compania Internacional Minera, a subsidiary of the Union Corporation of London, started a 250-ton flotation mill on oxidized lead-silver ores, and during April of this year, the Phelps-Dodge Corpn. will probably start a 250-ton mill at Warren, Ariz., on oxidized lead-silver ores, similar to those being treated in the Shattuck-Denn mill. Methods All of these mills use sodium sulfide for filming the oxide minerals, with the possible exception of the North Peru mill, and all use oily reagents. The total tonnage of strictly oxidized lead-silver ores concentrated by flotation is probably less than 2500 tons daily. The following figures give the various types of ores being treated:

Jan 1, 1928

By A. W. Fahrenwald

Flotation practice in Idaho is now about 13 years old. The advance has been steady during these 113 years. The operators have been alert to take advantage of the newest developments and they have themselves made a constant effort to improve their work. The metallurgical improvements are due to the development of (1) better equipment, (2) better reagents, and (3) better understanding and control of the process. During the eight years the writer has been in this district, there has been continuous experimentation by several companies to improve the ore dressing by flotation methods. Gravity concentration has fought hard to hold its place, and although it is true that gravity processes have been improved during this period, nevertheless, in several of the important mills of the district the few tables that managed to hang on have finally given way. On the straight-lead ores, tables and jigs are still used; and, due to better classification, their hold on this job has been strengthened. Ores of the District The ores of the district are of two general classes: (1) clean lead ores, such as those from the Bunker Hill and Hecla mines, and (2) complex lead-zinc ores. The general impression prevails that the clean lead ores, class 1, are simple and that it is easy to obtain high recoveries and good grade products. This, in a way, is a false impression; there is quite a high percentage of locked middling in these ores at 100 and' 150-mesh sizes. Improved classification preparatory to table treatment of this type of ore has made it possible to segregate effectively a high-grade middling which is further comminuted and floated. The complex lead-zinc ores of the district, of which the ore from the Morning mine of the Federal Mining & Smelting Co. is a typical example, always have presented a difficult problem. It always has been comparatively easy to make a clean lead concentrate, but it never has been possible to concentrate sphalerite with a specific gravity of 4.10 from siderite of gravity 3.88. Also, a large percentage of the galena and sphalerite of this class of ore refuses to separate out until the crushing has been carried

Jan 1, 1928

By Arthur F. Taggart

In 1900, Elmore found that if an acidulated pulp was stirred up with an oil which was relatively insoluble in and lighter than water, and the mixture was allowed to stratify, much of the sulfide would be found hanging to the lower surface of the floating body of oil and could be skimmed off or overflowed with it. The action of the flotation agent here is twofold, first to select the sulfide from the gangue—what today is called collection—and second, to buoy the collected sulfide away from the gangue. The action in collecting is illustrated by Fig. 1: which contrasts the behavior of drops of oil brought, under water, to the surfaces of gangue and sulfide particles, respectively. The oil displaces water from the surface of the sulfide; the water displaces oil from the gangue surface. If, instead of considering the solid particles the fixed parts of the system of oil, water and solids. we consider the oil and water fixed and the solids presented to the surface of contact, the condition will be as shown in Fig. 2; the sulfide tends to be drawn into the interface and the gangue to be excluded. The sulfide particles are held in the oil-water surface by the tension of this surface. The buoyant effect is supplied by the oil droplets in bringing the sulfides to the water surface, but not in holding them there. Gas as Buoyant Medium of Flotation In 1906 Sulman. Picard and Ballot discovered a practical and substantially fool-proof method of doing what Everson, in 1886, and Fro-nlcnt. DeBavay Bessel and Elmore had been doing in the intervening years, viz.: using gas in place of oil as the buoyant medium for flotation, but retaining oil as the agent. To all of these inventors "oil," with no qualification as to its kind and no true understanding of its function, was the important thing. Hut they placed on the oil an additional burden viz.: froth-making, and it soon became apparent that not all oils were suitable for the dual purpose. As a matter of fact, it is probable that each of these inventors knew only one suitable oil at the time he wrote his patent, but used the generic term in the patent with the usual idea

Jan 1, 1928

By W. L. Zeigler

The Galena mill of the Callahan Zinc-Lead Co. is of 150 tons daily capacity and is situated at Lake Gulch, 2 miles west of Wallace, Idaho. (Fig. 1. It was constructed during the summer of 1926, work having been started late in June and the plant placed in operation Oct. 15, following.

Jan 1, 1928

A general discussion on the subject of Crushing and Grinding was held on August 24, 1927, during the meeting of the Institute at Salt, Lake City. G. H. Clevenger, Chairman of the Milling Methods Committee of the Institute, made the following introductory remarks "Several years ago, a number of us felt that the time was ripe for a fundamental investigation of crushing and grinding. At the outset, it appeared that the first thing we needed was a yardstick whereby grinding efficiency could be measured. "The U. S. Bureau of Mines has been most generous in cooperating with the Milling Methods Committee in getting this work well undrr way. John Gross, who served the committee for a year as secretary, was placed in charge of the investigation and has been ably assisted by S. R. Zimmerley. "Messrs. Gross and Zimmerley will now present the Progrcss Report covering their work to date." Progress Report J. Gross, Salt Lake City, Utah.—-This has not been a simple problem, and when I tell you that me can give you a verbal report of what we have done during three years in five or ten minutes you will realize that we have necessarily done a great deal of work that dors riot appear in the final results. We have been in a virtually new field. Therefore, we have had to feel our way at each step so as to be fairly sure we were correet. The first thing that was realized in this problem ass that unless we had some method of measuring what happened when we crushed an ore, we might as well quit and not go further. we realized that if we tried to measure the results of crushing or grinding by the ordiuary sizing test we would not get anywhere. We could sieve down to say 200 mesh fairly accurately; we could even go one sieve size further with a little less accuracy, and then we were absolutely lost, unless we went into microscopic measurement. Even this latter method would not have been satisfactory. The question was: what were we going to measure? If we were going to measure size, we might measure one particular dimension of a particle and be off 200 per cent. or more. If we tried to measure surfaee, we would have something tangible; the surface of a particle, no matter how irregular, is something definite, far more definite than any linear dimension would be. Moreover, if we could measure surface, we would immediately have a method of checking the Rittinger law as to whether it was or was not correct. So the first thing that we sought. was a method whereby we could measure surface on crushed particles or, we will sag., on particles; it would not make any difference whether crushed or not. It then became a question of what kind of particles we were going to use in obtaining a method for surface measurement. We realizcd that it would be rather a corn-

Jan 1, 1928

By Paul Billingsley

When I was an undergraduate at the Columbia School of Mines, the mining curriculum was subdivided into two major branches's known respectively as the Metallurgical and the Geological Options, which diverged rather widely as the course progressed. When I duly elected the Geological Option, I believed, I am sure, that this divergence would continue throughout the years. The facts have proved otherwise. The mining industry is a closely knit entity in which the search for its raw matcrials, called mining geology, cannot be separated from the conversion of these raw materials into marketable form, called metallurgy. The mining geologist searches for materials which the metallurgist can utilize, and only such; and whenever an advance in metallurgy opens the gates for new matcrials, the geologist's problem is correspondingly modified. No better instance of this relationship can be given than that seen in the Salt Lake region in recent years. Our mineral industry here is of long standing. The three principal tributary districts, Bingham, Park City and Tintic, were discovered in the 1860's and have been vigorously and intelligently developed for many decades. Their production has made the Salt Lake valley one of the world's great smelting centers. The maintenance of this industry has for years depended upon the continual development of new orebodies to take the place of those being extracted. Intensive exploration in the old districts has been supplemented by prospecting in outlying districts of lesser prominence. Many new orebodies have been discovered, such as those of the Daly West, Silver King, and Judge mines in Park City, the Utah Apex in Bingham, and the Sioux, Colorado Chief Consolidated and Tintic Standard in Tintic. New mines in other districts have been discovered also, but without affecting the dominánce of the three major districts. In all cases discoveries, to be of value, had to be of ore amenable to treatment in the available metallurgical plants. This meant that, for the most part, zincy ores, both oxide and sulfide, must remain untouched, and also that ores high in iron and ores high in silica had a fictitious value, either high or low, set by the balance of these

Jan 1, 1928

By R. W. Diamond

Three mills are now operated by the Consolidated Mining & Smelting Co. of Canada, Ltd.: (1) The Sullivan Concentrator, Chapman Camp, B. C. (near Kimberley, B. C.), (2) the St. Eugene Concentrator, Moyie, B. C., and (3) the Tadanac (Customs) Concentrator, Trail, B. C. Differential flotation is the only concentration method employed in all three plants. Tables are used for pilot work only. Values in all ores treated are as minerals of some or all of the metals —lead, zinc, gold or silver. Occasional customs ores contain small amounts of copper. Sulfide iron content of ores treated varies from a low figure in some ores to 38 per cent. iron in Sullivan ore. Gangue minerals vary considerably. A few customs ores are somewhat oxidized and as a result present difficulties. Flotation equipment and control are similar in the three mills, and the flow sheets at the St. Eugene and Tadanac mills, while much simpler, are modeled after that of the Sullivan Concentrator. In this paper I will not detail treatment nor equipment in any of our plants but will describe special features in, or in connection with, all of them. I will then deal with flotation reagents, describing our use of each of them, particularly where differing from general practice. I will conclude with comments on special subjects related to differential flotation.

Jan 1, 1928

By Henry F. Adams, Guy H. Ruggles

In this paper the authors aim to chronicle the experience and salient points brought out in changing flotation reagents at a concentrator which had probably been using a minimum amount of oil at a minimum cost per ton, to a more complicated system requiring a partial chemical control. The original oils were coal tar and pine oil. The new reagents are lime, xanthate and pine oil. It was found that there were no insurmountable difficulties and eventually the cost of the new reagents per ton of ore proved to be very little more than that of the old. Lime Plant When the question of handling lime came up there were two propositions open: (1) to handle the lime dry; (2) to handle it wet in the form of an emulsion. Granting that a dry lime feeder would function properly, delivering lime uniformly, the chief objection was the dust that would be raised when these feeders were being charged or operated, as the only possible location was on a gallery beneath which men were working. With this objection the handling of lime as an emulsion seemed to be the better scheme. The lime plant which was first installed was for experimental work on one section. This was gradually changed and developed until it was of sufficient size to accommodate 20 sections. The various changes will not be discussed and the following description deals only with the fully developed plant. Crude as it is, and perhaps somewhat expensive to operate, the lime-handling plant has been quite successful. Much has been learned which will pave the way for a better installation in the future. Burnt lime is received loose in standard gage box cars. This is the product of the kilns and has had no screening or sorting. A grab sample is taken from each car as it is unloaded. An average of the analyses of samples from 44 cars gave the following: Available CaO Total CaO, Oxide and Hy- MgO, Fe²O³ + Al²O³, Insoluble, Per Cent. drate, Per Cent. Per Cent. Per Cent. Pek Cent. 88.3 79.2 1.6 ' 1.3 2.2

Jan 1, 1928

By E. H. Mohr

At the McGill concentrator of the Nevada Consolidated Copper Co., all flotation operations have been carried out in Forrester cells since November, 1926. In respect to cost of operation, the new cell has proved to be superior to the Callow, which it supplanted, and its incorporation in the flow sheet has resulted in a marked improvement in the metallurgy. Before citing the advantages of the new machine over the old, it should be explained that the Forrester has not only replaced the Callow, but it has made possible the elimination of a number of tables as well. When Callow cells were in use, table concentration preceded flotation. The ball mill pulp was elevated, by means of air lifts, to roughing Wilfley tables situated two floors directly above the flotation floor. On an intermediate floor were additional tables which cleaned the concentrate from those above. The finished table concentrate constituted approximately 60 per cent. of the total mill production, and contained about 40 per cent, of the recovered copper. The table tailings were returned to the classifiers, the overflow from which went to the flotation cells and the oversize to the ball mills. As soon as the Forrester cells were installed, this tabling operation was discontinued. The ball mill pulp now discharges into the classifiers direct, without the intermediate tabling. It was impracticable to ernploy the present flow sheet when Callow cells were in use, for the reason that unless the great,er part of the heavy mineral was first removed, it accumulated on the mats and adversely affected the flotation operation. By tabling ahead of flotation, it was practicable to maintain the concentrator at 10 to 20 per cent. greater capacity (depending up on the pyrite contnt of the ore) than was possible by flotation alone. This was due to the necessity for finer grinding in the latter case. Even though the table conecntrates contained only 6 to 8 per cent'. copper content, the low smelting costs made it advantageous to operate at high mill capacity, at the cost of low grade of product. Standard Section equipment With the old flow sheet, a standard section consisted of four ball mills and classifiers, one air. lift, 20 roughing tables, 10 cleaner tables, 12 primary

Jan 1, 1928

By R. A. Pallanch

Sodium sulfite as a zinc retarding agent in the selective Hotation of complex lead-zinc sulfide ores was discovered by the writer in the summer of 1919 in the testing laboratory of the U. S. Smelting, Refining & Mining Co.'s gravity concentrator at Midvalc. Utah. It proved to be the first reagent to give consistently good results in the laboratory on the cornpanj's Bingham ore. In December, 1919, C. A Lemke, concentrator superintendent, recommended that a flotation unit be added to the existing gravity mill to handle the mill slimes, using sodium sulfite as the zinc depressant Owing to the poor lead and zinc markets and the unfavorable royalty schedules offered by the Minerals Separation Co., the project fell through at this time. In 1921, a number of laboratory tests were made at Midvale. on other ores, among them ore from the Sunnyside mine at Eureka, Chlo. Both sodium and calcium sulfites were used with succcss on this ore. Early in 1923, further tests were made at Eureka, shortly after the Sunnyside company had resumed operations after a two-year shutdown. AS a consequence sodium sulfite was given a trial in the mill. The results showcd sufficient improvement over previous work as to warrant continuous use of this reagent in the lead circuit Exeept for a week or so in 1926 and for several weeks in April, 1927, when sodium cyanide and zinc sulfate were given trial runs in this mill. sodium sulfite has been used ever sinee.. Sodium sulfite was chosen rather than the calcim sulfite brcauscl of its gre1ater sollubility, making it possitdr. to introduee the reagent as a solutiun. This Solution is introdueed into the Marry ball mills along with the original feed of ore,. The consumption of sodium sulfite in the Sunnyside mill varies from 1/4 to 12 lb per ton of ore, the smaller amount probably being nearer the average. In February, 1924, laboratory work at Midvale demonstratcd conclusivcly that good results could be obtaincd on the company's Bingham ore and on most of the custom ores tested by using sodiuin sulfite as the zinc depressant, xanthate as the lead enhancer in the lcad circuit, followed by copper sulfate and additional xanthate to float the zinc Sodium sulfide was used to float the iron pyrite after the lead and

Jan 1, 1928

Jan 1, 1915

August 10 to 14, 1914, inclusive Headquarters will be at the Hotel Utah. The following program of the meeting has been arranged y the Local Committees: Monday, August 10.-Members and guests will register at the Secretary's desk in Room C-38, Hotel Utah, and receive the official badges, booklets and programs of the meeting. At 2:30 P. M., there will be an organ recital at the Mormon Tabernacle, after which members and guests will be taken for a trip around Salt Lake City in sight-seeing motor cars. At 8:30 P. M., the ladies will assemble in the lobby of the Hotel Utah, where they will be met by the members of the Ladies' Entertainment Committee, and will be taken for an automobile drive about the city. At 8:30 P. M., a technical session will be held in the ballroom of the Hotel Utah. The following papers will be presented: F. W. BACORN: An Amendment to Sales's Theory of Ore Deposition. C. W. GOODALE: Drumlummon Mine, Marysville, Mont. I. B. JORALEMON: The Ajo Copper-Mining District. STUART CROASDALE: Leaching Experiments on the Ajo Ore. FRANK R. VAN HORN: The Occurrence of Bournonite, Jamesonite, and Calamine at Park City. HOWLAND BANCROFT: Dip Chart. M. N. ALLING: Ancient Auriferous Gravel Channels of Sierra County, California. B. S. BUTLER AND H. D. MCCASKEY: Copper Ores of the New London Mine. GEORGE A. PACKARD: Rope Idlers in the Raven Shaft. ARTHUR CROWFOOT: Development of the Round Table at Great Falls. FREDERICK LAIST AND A. E. WIGGIN: The Slime-Concentrating Plant at Anaconda. W, L, AUSTIN: The Treatment of Copper Ore by Leaching Methods. W. L. AUSTIN: Leaching Copper Products at the Steptoe Works.

Jan 8, 1914

By W. W. Norton

The gradually increasing proportion of sulphide ores which lead smelters of to-day are called upon to handle has caused the roasting problem to become one of ever greater importance. Mines have increased in depth, methods of concentration have improved, competition in the purchase of ores has become more keen and it has been clear that new and modern methods of eliminating sulphur must receive a large share of attention. We may look back a score of years or so and recall the old hand roasters, some of which turned out a finely divided roasted product, and others fashioned with a "fuse-box" wherein the roasted ore was slagged. Then came mechanical furnaces of several types, the Bruckner cylinder, the Brown-O'Harra, the Ropp, and others, all designed to cut out the cumbersome hand labor of the old reverberatories. Roasting costs were thereupon reduced and tonnages stimulated to a gratifying extent. However, it seems to have been early recognized that this very feature of a greater amount of roasted sulphide ore as compared with oxide ore brought with it a train of difficulties at the blast furnaces. The mechanically roasted ore was fine physically, the blast furnaces were choked and the troubles of the smelterman were not at an end. Then the Huntington-Heberlein pot system of roasting appeared upon the scene, and later on the Dwight-Lloyd machines were invented, and inasmuch as both of these processes were designed not only to roast the ore but to agglomerate it as well, it was apparent that distinct steps in advance had been achieved. At the Murray plant, modern up-to-date roasting practice is fully exemplified and there are now in successful operation roasting furnaces or devices of several sorts: namely, Godfrey revolving-hearth furnaces, Wedge multiple-hearth mechanical roasters, Dwight-Lloyd sintering machines, and Huntington-Heberlein pots. Godfrey and Wedge furnaces will properly handle material high in sulphur, say ores with 25, 30, and 35 per cent. of that element; D. & L. machines and H. & H. pots will positively not treat efficiently ores or mixtures containing anywhere near the sulphur content mentioned, but are confined to charges containing in the neighborhood of 15 or 18 per cent. In passing, it may also be explained that, so far as the knowledge of the writer goes, Godfrey and Wedge furnaces do not economically eliminate sulphur to an extent

Jan 1, 1915

By Frederick W. Bacorn

The paper of Reno H. Sales on Ore Deposits at Butte, Mont.,' is a careful and painstaking work, an important contribution to the literature of the subject. As is almost inevitable in a work of such magnitude, a few misstatements occur; and some of these I shall endeavor to point out. Mr. Sales has also fallen into error, as I think, in the application of his theory of ore deposition; and while the modification which I venture to suggest is a minor one geologically, it is of great economic importance. I refer to Mr. Sales's idea of an outward migration of mineralizing solutions from a supposed central copper zone. In order that I may make myself plain, I must first epitomize the theory as a whole, which I understand to be this: All of the Butte ore deposits (save those along the Continental fault, not herein considered) are derived from solutions of magmatic origin, which, at the beginning of their ascent from a great depth, were acid in reaction, hot, and were carriers of copper, zinc, manganese and other metals. As they departed from their point of origin they lost heat, and, because of reactions with the wall rock, became more and more alkaline. The metal deposited from these solutions in their hot condition was principally copper; and, as they cooled, they deposited less and less copper and more and more manganese and zinc, until the copper practically disappeared. With this theory in general I have no fault to find, nor have I, it must be stated, the qualifications necessary to enable me to pass upon it critically; but it seems to me to be reasonable and probable, and is one which I am quite willing to accept. But Mr. Sales, for what seem to me to be insufficient reasons, has assumed that there is one deep central source of copper, relatively rich and small, from which all the solutions emanated. From this central source the solutions rose, through one channel or through a group of channels, to the surface, at and within a small area called the central copper zone,

Jan 1, 1915

By Mark N. Alling

A few years ago, gravel mining in the ancient river beds of Sierra county was the principal industry of the county. Operating drift mines were very numerous and many large hydraulic mines were in action. Eventually the State laws put an end to hydraulic mining, it being claimed that the silt carried down the streams to the valleys filled the rivers and caused winter floods, which were damaging to the surrounding farms. The ancient river deposits which could easily be found were, in the course of 40 years, worked out or practically so. The prohibition of hydraulic mining was a severe blow to the industry in general. The deposits were not developed and the gravel-mining industry practically stood still for many years. In Sierra county, nearly all the hydraulic mines were at the points where the ancient channels were exposed to view by reason of the modern stream erosion, or, in the miner's language, where the channels "enter" or "break out of" the ridges. The claims covering them also covered large areas in which were long segments of lava-capped channels, suitable for drift mines, and the points being hydraulicked were usually the controlling points of attack. Men who have followed hydraulic mining for any length of time seldom are content to work a drift mine. Knowing the great value of their deposits and feeling that they had been robbed by the State, these men have stubbornly held to their claims. The gold that is carried in these gravel channel deposits is more or less scattered throughout the depth and width of the deposit; the percentage that can be recovered by the process of drifting ranges from 50 to 75 per cent. of the total. In hydraulicking, practically all of the deposit is put through the sluices, and as the cost per cubic yard is only about 10 per cent. of the cost of drifting, the deposits, if classed as hydraulic mines, are much more valuable than if classed as drift mines; hence the old timers have held their claims (now worthless as hydraulic mines) at prices which the drift-mining industry could not stand; and such mines

Jan 1, 1915

By L. O. Howard

The practice of using acid-lined converters at the plant of the Old Dominion, Copper Mining & Smelting Co. was discontinued early in January, 1913, and replaced by basic-lined converters. The new equipment consists of one electrically operated 12-ft. upright converter stand and two basic-lined shells for same. These shells are provided with twenty-four 1 1/4-in. tuyeres. The mouths are 5 ft. in diameter. Owing to the fact that there is no bedding system or other large storage capacity between the mine and the smelter, the furnaces must handle as fast as produced ore and concentrates, which may vary considerably from day to day, both in tonnage and copper content. Consequently the converter is often crowded to its utmost capacity by the frequent fluctuations in the daily matte supply. From the foregoing, it will be seen that in no way can the 'converter be favored by running light charges or by transferring to it from other stands in operation. Neither can it regularly be allowed to cool between blows, and only occasionally can it be run at as low a temperature as might be wished. No copper precipitate or other foreign material high in copper is available to help out the tonnage. Repairs and alterations or the removal of defective tuyeres must wait until such time as a slight falling off in.the matte supply allows the converting operations to ease up. These variable conditions all combine to interfere with the systematic operating of the converter. The first 12-ft. shell was blown in on Jan. 6, 1913, and ran until June 27, 1913, when it was removed to patch the brick. A record of operations covering this period follows: Shell No. 1. First Campaign, Jan. 6,to June 27, 1913 Commenced blowing............................Jan. 6, 1913. Taken off stand for patching.....................June 27, 1913. Total blowing time..........................2,689 hr. 40 min. Total number of, blows made.............................. 664 Average time of blow............................ 4 hr. 3 min. Tons matte charged.................................. 18,157 Average copper content of matte................ 43.9 per cent. Average time blowing one ton. copper................ 22.5 min. Total copper made............................... 7,250 tons

Jan 1, 1915