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By G. G. Gunther, C. L. Boeke

DISCUSSION I. M. Symonds, (Cia. Minera de Penoles, Monterrey, N. L., Mexico)—Messrs. Boeke and Gunther have written most interesting papers describing their zinc de-leading operations. Mr. Gunther was kind enough to show me his operation last fall. Aside from the excellent results, I was most impressed by the ease of control. I had expected a difficult operation requiring careful control and supervision, but found one requiring little attention. The only disappointing fact of the separation is the poor copper recovery in the concentrate of the delead-ing float. The final zinc concentrate still has about 3 pct chalcopyrite, assuming the copper is chalcopyrite, and adversely affects the zinc grade. If the copper mineral is free, the poor recovery is probably due to the high cyanide concentration. Fortunately, about 75 pct of the copper is recovered before the pulp goes to the zinc circuit. Mr. Gunther's remarks about proper cyanide to zinc sulphate ratios cannot be overemphasized. Although our experience has been confined mostly to the conventional lead-zinc separation, we have found these ratios most important, and more important in the mill than in the laboratory. The laboratory appears to give optimum results over a wider range of reagent control than the mill, especially in regard to the effect of cyanide. In handling relatively high grade ores in the mill we usually find that ratios of cyanide to zinc sulphate of 1 to 10 or 20 or higher work best. Where there is a substantial amount of chalcopyrite present it has been noted that cyanide in the range of 0.02 to 0.10 lb per ton of ore is often beneficial in zinc rejection and possibly in copper recovery in the lead concentrate, while larger amounts, say 0.20 to 0.25 lb per ton, will float more zinc and cause a serious rejection of chalcopyrite. On an ore with 15 to 20 pct zinc content we have had to use 0.8 to 1.2 pounds of zinc sulphate per ton of ore with the above cyanide to obtain optimum results. Some time ago we made a laboratory study following Parral and Santa Barbara practice on an ore which had never a very satisfactory rejection of zinc in the lead concentrate. The results were not encouraging and a mill test was not run. Since then the study has been reopened using in the laboratory fresh zinc concentrate from the mill instead of one made in the laboratory. A reasonably clean lead float from the zinc concentrate was obtained. It has been checked several times and each time the zinc concentrate made in the mill responds and the one made in the labora-

Jan 1, 1953

By G. G. Gunther, C. L. Boeke

DISCUSSION I. M. Symonds, (Cia. Minera de Penoles, Monterrey, N. L., Mexico)—Messrs. Boeke and Gunther have written most interesting papers describing their zinc de-leading operations. Mr. Gunther was kind enough to show me his operation last fall. Aside from the excellent results, I was most impressed by the ease of control. I had expected a difficult operation requiring careful control and supervision, but found one requiring little attention. The only disappointing fact of the separation is the poor copper recovery in the concentrate of the delead-ing float. The final zinc concentrate still has about 3 pct chalcopyrite, assuming the copper is chalcopyrite, and adversely affects the zinc grade. If the copper mineral is free, the poor recovery is probably due to the high cyanide concentration. Fortunately, about 75 pct of the copper is recovered before the pulp goes to the zinc circuit. Mr. Gunther's remarks about proper cyanide to zinc sulphate ratios cannot be overemphasized. Although our experience has been confined mostly to the conventional lead-zinc separation, we have found these ratios most important, and more important in the mill than in the laboratory. The laboratory appears to give optimum results over a wider range of reagent control than the mill, especially in regard to the effect of cyanide. In handling relatively high grade ores in the mill we usually find that ratios of cyanide to zinc sulphate of 1 to 10 or 20 or higher work best. Where there is a substantial amount of chalcopyrite present it has been noted that cyanide in the range of 0.02 to 0.10 lb per ton of ore is often beneficial in zinc rejection and possibly in copper recovery in the lead concentrate, while larger amounts, say 0.20 to 0.25 lb per ton, will float more zinc and cause a serious rejection of chalcopyrite. On an ore with 15 to 20 pct zinc content we have had to use 0.8 to 1.2 pounds of zinc sulphate per ton of ore with the above cyanide to obtain optimum results. Some time ago we made a laboratory study following Parral and Santa Barbara practice on an ore which had never a very satisfactory rejection of zinc in the lead concentrate. The results were not encouraging and a mill test was not run. Since then the study has been reopened using in the laboratory fresh zinc concentrate from the mill instead of one made in the laboratory. A reasonably clean lead float from the zinc concentrate was obtained. It has been checked several times and each time the zinc concentrate made in the mill responds and the one made in the labora-

Jan 1, 1953

By Orville A. Derby

In view of the recent important discussions before this Institute on the genesis of metalliferous deposits, certain features in the occurrence of gold which seem to be more clearly developed in Brazil than elsewhere can hardly fail to be of interest. As the observations here recorded have necessarily been of a somewhat cursory character, they will be presented in the form of brief notes, without any attempt to draw conclusions as to their genetic significance. Gold in Gneiss.—The statement made by me many years ago† that in the districts of Campanha and S?o Gongalo, in southern Minas Geraes, gold occurs in gneiss, can now be reaffirmed in the most positive manner, on the strength of a recent examination of fresh material, confirmed by the testimony of an experienced mining captain from the Morro Velho mine, who, being skeptical on this point, made a prospectingtour in the district. The specimens examined, though so far decomposed as to be readily disaggregated, were clearly of

Jan 1, 1903

By Grafton M. Thrasher

Discussion of the paper of GRAFTON M. THRASHER, presented at the New York meeting, February, 1916, and printed in Bulletin No. 106, October, 1915, pp. 2129 to 2138. RICHARD MOLDENKE, Watchung, N. J.-Mr. Thrasher is to be commended for developing a very interesting comparison between high- and low-carbon and silicon malleable cast irons for an equal degree of chill. His deductions are unquestionably correct, though the application to daily practice is open to serious criticism, from the standpoint of the man who receives the castings, as will be shown later on. Running over some of the points which affect the tests in question: Mr. Thrasher admits that the temperature of pouring has a serious influence on the degree of chill shown in the test plugs he made. This is a well-known fact, and I would like to add that another serious trouble comes from oxidized iron when the heats have been unduly prolonged. Such metal always shows more chilling action for a given silicon and carbon content than normally melted iron. Indeed, the test piece quoted giving the extreme of the chill line (1.07 per cent. Si and 2.47 per cent. total carbon) was such a poor heat, a ferro-silicon addition never "doctoring" up the metal sufficiently to show normal good strength. In spite of these irregularities, which would simply change the curve slightly, I take it that Mr. Thrasher's results can be accepted as reliable enough to establish the relationship of silicon and total carbon as shown. When, however, the claim is made that silicon and carbon can be substituted for each other in mixture-making (a claim which may be perfectly correct metallurgically), a very important factor is forgotten, causing the purchaser of the castings much annoyance. This is the contraction (erroneously called shrinkage) of the castings. They will no longer be true to pattern. The lower the total carbon is allowed to drop, the greater the contraction, and increasing the silicon correspondingly will not correct this trouble. No wonder that recently, in revising the Standard Specifications for Malleable Castings it was necessary to increase the variation of castings from requirements considerably over what was formerly found sufficient. The habit of running up the proportion of silicon and escaping danger from grayness of the metal by copious steel additions was evidently too tempting to manufacturers, as their molding losses were thus kept down and a very soft malleable iron resulted. The customer, however, got serious variations in his work, which was not the case when the total carbon was kept between 3.25 and 2.75, and the silicon held as low as possible without undue cracking of the hard castings.

Jan 5, 1916

By Bradley Stoughton

A LL over our great country I have been privileged to see, during the last six weeks, the manifestation of a new spirit among engineers. Partly under the inspiration of leaders whose influence has been felt not only nationally, but internationally, - partly through the recognition of the service that engineers have been able to render to the nation during the recent crisis, a new self-consciousness has been acquired by the individuals, and now this self-consciousness seeks expression in opportunities for service and participation in public affairs. The engineer . no longer sees himself a mere reservoir of information to be tapped when some one is in need of cold facts, no. more an administrator of mere material things, but a potential administrator of- public affairs, a citizen and a human being, with the responsibilities, as well as the privileges of citizenship.

Jan 1, 1920

By W. D. BRENNAN

FOR a considerable number of years past, there has been a shortage of experienced coal miners at the coal mine of the Phelps-Dodge Corp. This is probably due, to some extent, to the mines being so far south of the main western coal producing regions, so that experienced men secure employment before reaching this point. These mines are at Dawson, New Mexico, which is in Colfax County, near the Colorado line. The district is known as the Raton field, and the seam being worked is known as the Raton or Blossburg seam. The mines operated are Nos. 1 to 6, 8 and 9. The daily out- put at the present time, with restricted operations, is about 4000 tons, but the daily capacity is about 5500 tons. The mines are so grouped that the coal is dumped over four tipples. The mines are about twenty-five years old and extend as much as three miles under ground; they are all drift openings, with trolley, storage battery, and mule haulage. At the same time that the shortage of miners existed there was a surplus of common labor seeking employment. We learned that the Copper Queen. Branch of Phelps Dodge Corp. conducted a training school in one of its mines which had worked out very satisfactorily, and thus gained the idea of attempting such a school in the coal mines at Dawson.

Jan 1, 1929

By R. W. Raymond

The immense commercial value of the Gayley dry-blast process has been established beyond controversy. The testimony of practical blast-furnace managers, on both sides of the Atlantic, agrees that it reduces the cost of pig-iron about $1 per ton; but from the same testimony it appears that this direct economy does not by any means cover its commercial advantages—in fact, that under conditions frequently encountered in practice, it presents other benefits of even greater immediate financial importance. To this point I shall recur in a later part of the present paper. The history of the reception given to this invention, especially on the other side of the Atlantic, has repeated an experience familiar to American engineers. First, the reports of our practice were rejected as theoretically impossible, according to generally accepted notions and formulas; then the figures were reluctantly accepted; and finally the attempt has been made to explain how they could have been obtained. As to the Gayley dry blast, me Americans have been for several years in the last stage of the discussion, and the rest of the professional world is rapidly coming to it. Meanwhile, the remark has been repeatedly made that, whatever Mr. Gayley's process may turn out to be worth, it was not an invention on his part, but simply the energetic and persevering execution of a procedure already proposed by others. This question I wish to discuss first in the present paper. I. What Did Mr. Gayley Invent? Realizing that this is not the place for detailed argument on an issue of patent law, I shall not undertake to quote and criticise the publications prior to Mr. Gayley'a first patents. It may be admitted, without such critical analysis, that the possible advantages of removing moisture from the air forced into an iron blast-furnace were recognized and talked about a hun-

Jan 1, 1909

A. O. GATES, Salt Lake City, Utah (communication to the Secretary*).-The writer is delighted by the results shown in Mr. Bell's paper, which prove in an experimental way different from that followed by the writer' that "the work clone in crushing is proportional to the surface exposed by the operation," per Rittinger's Law. Mr. Bell's work was all clone in commercial types of crushing machines, while my work was clone in a testing machine; his work represents fully one hundred times the expenditure and time that my work represented; a great number of tests were made by his organization to determine single points, while I was obliged to depend on the general trend of a curve for results; we agree absolutely that "Kick's law," for which Mr. Stadler was sponsor, does not apply to rock crushing. While my work indicated the reliability of Rittinger's theory, Mr. Bell's work clearly establishes the reliability and makes it Rittinger's law.

Jan 5, 1917

BUMPS in No. 2 Mine, Springhill, N. S., furnished the main feature for discussion at the morning meeting* on Ground Movement and Subsidence on Feb. 18. Walter Herd, the author of the paper by which the discussion was introduced, not being well, was unable to be present. The paper was presented by T. L. McCall, who laid stress on the presence of sand- stone as the cause of bumps, on the value of retreating longwall in eliminating such phenomena at the coal face and on the advantage of wide headings with packs on either side, leaving space behind them for any coal that bumps might dislodge. George S. Rice, who presided, referred to his visits to the Nova Scotia mines and said that bumps usually occurred only where the coal was thick and strong and the roof and floor both more than usually hard. If any of the three were weak, bumps were not likely to occur. In the Nova Scotia mine, to which Mr. Herd referred, there was no material for packwalls and for this reason as well as for safety he had been convinced that long- wall retreating with a small percentage of pack walling was the most feasible plan.

Jan 1, 1929

By Sidney W. Poole

THE control of deoxidation, together with other factors involved in electric-furnace melting practice, to consistently produce heats of high-quality alloy steel to stringent specifications as to nonmetallic inclusion content is a problem of considerable magnitude. Steel melting as an industrial process is subject to many variables that can influence the quality of the product made. In order to modify the size, amount and distribution of inclusions and to promote their elimination, deoxidation, among other practices, plays a major part. The primary purpose of deoxidation in the production of electric-furnace alloy steels is the removal of dissolved oxygen present as FeO in the bath, which is produced as a result of prior oxidizing conditions. The elements usually added to effect this removal of soluble oxygen are those that will, upon addition, cause the formation of an oxide compound insoluble in the molten steel bath as silica, alumina or silicate combinations.

Jan 1, 1947

By Erskine Ramsay

The following description of the Alabama coal-,fields is taken (with slight omissions) from the report " On the Warrior CoalField, by Henry McCalley, A.M., C. and M. E., Assistant State Geologist:" The coal lands of Alabama are of the great coal basin of the Ohio, or of the great Appalachian coal-field, and are estimated at 8660 square miles. They formed one connected field prior to the great Appalachian revolution, when there was pushed up through them long, narrow, anticlinal ridges which have been denuded into the present well-known anticlinal valleys, and thus this great field has been divided up into three more or less distinct parts. These parts were named by Prof. Tuomey, in 1849, the Warrior, the Cahaba and the Coosa coal-fields, respectively, from the names of the rivers which drain them. They are of very unequal size, the Warrior, with an area of 7810 square miles, being nearly ten times as large as the other two combined, and about two-thirds as large as the coal-area of Great Britain. This Warrior field is the most northwestern of the three coal-fields of Alabama, and comprises all of the coal-measnres in Alabama, northwest of the Alabama Great Southern railroad, or all of those drained by the Warrior and Tennessee rivers. The greater part of this field, almost all of it away from the present railroad lines, has never received more than a surface-examination. Still, we have in the Warrior coal-field one of the richest coal-basins of this or any other country. As a whole, the Warrior coal-field is a broad, shallow, tray-shaped depression, sloping towards the southwest, and with its southwest end covered by a newer formation and its southeast side, especially, complicated by folds and fractures. As is the case with all of our Coal Measures. it is composed of a series of sandstones. conglomerates, shales, slates and coal-seams, with some little limestone at several horizontal positions, but contains no thick bed of this last rock like the Measures of Pennsylvania and other States. The thickest conglomerates are near the bottom and top of the series, and hence these coals might be said to belong mainly to the

Jan 1, 1891

•Mineral Development Corp. of Salt Lake City, Utah purchased the Somerset and Oliver coal mines from the Kaiser Steel Corp. The mines are located at Somerset, Colo., about 75 miles east of Grand Junction, Colo. Somerset Town, with housing for 600 persons was included in the transaction. The mines employ 115 miners and have a combined capacity of 1750 tons per day. The Minerals Development Corp. is headed by Claude P. Heiner and is controlled by Godfrey L. Cabot, Inc. of New York and Boston. The Cabot Company is a major producer of carbon black. • The Dept. of Interior has announced that the name of the Defense Minerals Administration will be changed to Defense Minerals Exploration Administration. The new name of this defense agency more nearly describes its duties, since the mineral programming, procurement and claimant functions of the former DMA have been transferred to the Defense Minerals Procurement Agency. C. O. Mittendorf, formerly acting administrator of DMA, is serving as acting administrator of the new DMEA.

Jan 1, 1952

By H. J. Rahilly

Mine fires, in the Butte district, have been a source of trouble and expense for the past thirty years, for while the actual fire area in most of the mines has been comparatively small, the handling of the gas has generally been a much larger problem than the actual putting out of the fire. The chief causes of these fires have been: Defective electrical equipment, incendiarism or carelessness, and spontaneous combustion. While fires from the first two causes have been numerous, they have generally been discovered and extinguished before they became very serious. Spontaneous combustion has been the cause of most of the large fires. Sufficient heat has been generated by ground movement, high rock temperatures, and decomposition to set fire to inflammable materials (tarred ropes, canvas, dry timber, manure, and hay) left in the gobs some years ago. Also, the heat resulting from the oxidation of the fine broken sulfide ore is sometimes sufficient to set fire to the timbers. In 1906, a fire was discovered on the 1100-ft. level of the Minnie Healy mine about 400 ft. (122 m.) west of the Minnie Healy shaft, which presumably originated, in one of the upper levels, from spontaneous combustion. The fire was walled off by concrete bulkheads and a pillar of ground 100 ft. (30 m.) thick was left below and mining resumed underneath. During subsequent operations, this pillar became cracked and broken and the fire passed into the new workings. It travelled downward and spread rapidly until, in 1917, it had become necessary to wall off portions of the Tramway, Leonard, and West Colusa mines on the 1300, 1400, 1500, 1600, 1800, 1900, and 2000-ft. levels, and the tops of the 2200-ft. raises were sealed. The fire area now consists roughly of a block 1000 by 1200 ft. and from the 600-ft. to the 2200-ft. level. This block is composed largely of a series of parallel northwest-southeast nearly vertical veins and stringers that form a large orebody, with no well-defined walls but badly fractured and faulted, and known locally as "horsetail ore." These veins, or orebodies, are mined by the square-set or the rill method.

Jan 1, 1923

By H. J. Rahilly

Mine fires, in the Butte district, have been a source of trouble and expense for the past thirty years, for while the actual fire area in most of the mines has been comparatively small, the handling of the gas has generally been a much larger problem than the actual putting out of the fire. The chief causes of these fires have been: Defective electrical equipment, incendiarism or carelessness, and spontaneous combustion. While fires from the first two causes have been numerous, they have generally been discovered and extinguished before they became very serious. Spontaneous combustion has been the cause of most of the large fires. Sufficient heat has been generated by ground movement, high rock temperatures, and decomposition to set fire to inflammable materials (tarred ropes, canvas, dry timber, manure, and hay) left in the gobs some years ago. Also, the heat resulting from the oxidation of the fine broken sulfide ore is sometimes sufficient to set fire to the timbers. In 1906, a fire was discovered on the 1100-ft. level of the Minnie Healy mine about 400 ft. (122 m.) west of the Minnie Healy shaft, which presumably originated, in one of the upper levels, from spontaneous combustion. The fire was walled off by concrete bulkheads and a pillar of ground 100 ft. (30 m.) thick was left below and mining resumed underneath. During subsequent operations, this pillar became cracked and broken and the fire passed into the new workings. It travelled downward and spread rapidly until, in 1917, it had become necessary to wall off portions of the Tramway, Leonard, and West Colusa mines on the 1300, 1400, 1500, 1600, 1800, 1900, and 2000-ft. levels, and the tops of the 2200-ft. raises were sealed. The fire area now consists roughly of a block 1000 by 1200 ft. and from the 600-ft. to the 2200-ft. level. This block is composed largely of a series of parallel northwest-southeast nearly vertical veins and stringers that form a large orebody, with no well-defined walls but badly fractured and faulted, and known locally as "horsetail ore." These veins, or orebodies, are mined by the square-set or the rill method.

Jan 1, 1923

By R. R. Knuckey

HAVING been associated with the operation of the de Vooys process for coal, which has treated 13,000,000 tons per annum, and recognizing the process as of value in ore sorting, Huntington, Heberlein and Co. Ltd. (Great Britain) in 1936 made an examination of all other existing methods. In none of the pioneer applications had recognition been given to the necessity for maintaining certain conditions that are essential for sorting certain ores if a truly efficient separation by gravity alone is to be assured. Consequently, in 1936-1937 Huntington, Heberlein and Co. Ltd., with Pearson, employing certain ideas connected with the de Vooys application, commenced the development of a separator that would allow conformance with optimum conditions in the zone where separation is effected. Later, Holt and Wade added to the value of the original application by Wuensch, Rakowsky and others, whose conceptions are being successfully operated today. Meantime, Trostler had enhanced the value of the original H-H. process by improvements that are now embodied in operating plants, and today the process incorporates an improved apparatus and a stable medium that produce the optimum conditions for separation. These conditions are: 1. Elimination of disturbing currents in the separating zone; 2. Reduction to a minimum of medium-density differences throughout. the separating zone; 3. Stable medium of minimum viscosity to permit the unretarded fall of mineralized particles; 4. Removal of products of separation with minimum disturbance to the separating zone. DESIGN OF SEPARATOR The H-H. sink-and-float separator consists of a pyramidal tank in which the separation actually takes place, connected at the bottom with an inclined bucket elevator. The tank and the elevator are filled with medium to within a few inches of the top of the tank. Medium is fed into the separator at a variety of points, according to the type of ore being treated, and it is withdrawn from the separator at two points: (1) together with the float product, by means of a variable-speed paddle, at the surface of the medium, and (2) by flowing over an adjustable weir cut in the side of the elevator casing. The prepared ore is fed onto the surface of the medium. The intermediate paddles, which revolve at a slow speed of 2 to 5 r.p.m., move the floating ore across the bath from feed end to float-discharge end. The ore, which is heavier than the medium, sinks below the sweep of the paddles and reaches the boot of the bucket elevator, which discharges it without disturbance of the separating zone. These arrangements for withdrawing medium from the separator allow the use of minimum flows, which connote minimum

Jan 1, 1943

By L. C. Glenn

In the preparation of this paper the writer has drawn freely upon the writings of Orton, Munn, Shaw, Mather, Miller, Hoeing, St. Clair, Jillson, and others, as well as upon his own personal knowledge of the fields of both states. It is to be regretted that certain data gathered by him and his assistants last fall are not available for publication. Oil in Tennessee A few wells drilled for brine for salt making in Tennessee between 1820 and 1840 obtained oil, but no definite search was made for oil until just after the close of the Civil War. Active drilling was then begun in Overton and counties southwest of it on the eastern half of the Highland Rim. A number of strikes were made at shallow depths in the basal part of the Mississippian but the wells were soon exhausted and abandoned. Drilling was revived, about 1892, when the Spurrier district in Pickett County was developed and was followed by the Riverton district in the same county in 1896. A pipe line was laid from the Wayne County, Ky., fields and about 60,000 bbl. of oil were run before a very heavy slump in the production, a failure to find an extension of the field, and excessive local taxation caused the removal of the pipe line in 1906. There was then no production in Tennessee until the discovery, in 1915, of oil near Oneida, Scott Co., at about 950 ft. (289 m.) in fissures in the Newman or St. Louis limestone. This field, however, soon failed and was abandoned. In 1916, oil was found at Glen Mary, Scott Co., in the Newman limestone at a depth of 1232 ft. (375 m.) A number of wells have since been drilled there, some of which were dry while others, close by, were produccrs. The largest one yielded, at first, about 340 bbl. per day and produced for several months, when it suddenly went dry. Several of the first wells began at 6 or 8 bbl. per day and are still maintaining that output. Production is from a fissured part of the limestone and varies greatly in accordance with the size and extent of the ramification of the

Jan 1, 1921

By C. W. Hall

Owin to its isolation and comparatively small tonnage, the Yoquivo district is not widely known; though financially important andgeologically, quite interesting. San Francisco de Yoquivo, the center of the mining district, is in the western part of the state of Chihuahua, Mexico; about 35 miles west of San Juanito, on the Kansas City, Mexico & Orient R. R. The town can be reached from the railroad by a good trail over a rugged country. The elevation of Yoquivo is 6500 ft., and peaks surrounding the town are 2000 ft. higher. The mining towns of Concheño and Ocampo are 25 miles northwest, and Uruachic is 20 miles southwest. The climate, temperate and healthful, is almost ideal. The rainy season begins in July and ends the latter part of September. There is usually, though not always, some precipitation in all months except April, May, and June. The mountains are covered with pine and scrub oak. Water, wood and lumber are plentiful for all ordinary purposes, and hydroelectric power is available for about six months of the year, when wood for steam power is utilized as an expedient. The district has been worked recurrently for over 60 years. The Yoquivo province lies on the steep western slope of the Tara-humara division of the Sierra Madre range of mountains. The province is traversed by many deep canyons, most of which flow west. The town is situated on the San Francisco arroyo, which makes a confluence with the Huevachic stream 236 m. southwest; the Huevachic in turn joins the Candemeña river farther down and these two drain the western flank of the Tarahumara mountains and form the headwaters of the Mayo river. The claims of the Yocluivo Development Co., which is the only organization in the district with a record of any considerable production, lie within an area roughly l½ m. square. Thc area extends from the San Francisco arroyo north, and from the Dolores arroyo west, and is confined to the andesite formation. General Geology The terrane about Yoquivo is composed of a series of Tertiary eruptives that consist of numerous phases of andesite and rhyolite, and

Jan 1, 1927

By Howard Eavenson

THE eleven mines of the United States Coal and Coke Co. in the Pocahontas coal field are situated in McDowell County, W. Va., which is a mountainous region. The valleys rarely exceed 200 ft. (60 m.) is width, the hill sides slope from 25° to 33°, and the tops, from 600, to 900 ft. above the valleys, carry practically no level land. At least 80 per cent. of the area is covered with timber. All the workings are drift mines; and while the average grades of the seams worked are ample for drainage, local dips and swamps necessitate, the use of a large number of local pumps or of numerous drainage ditches to keep the mine free from water during heavy rains. The water handled in any mine is not great, compared with the coal output, and usually causes little trouble except during extremely wet periods. In hew workings, the water encountered, except in rare cases, is trifling; it is only after rib drawing has started and the surface has been broken that water troubles begin. The strata overlying the coal are sandstone, shales, and slates; as the soil is thin and sandy, it does not tend to puddle and stop the cracks, which remain open for a long time. As mining progresses and the robbed area increases, the broken area becomes larger, increasing the openings into the mined territory. The cover over the robbed portions of the mines varies from 30 to 750 ft. (9 to 228 m.). When the ground is saturated, considerable water finds its way into the mines the day after a heavy rainfall and for three or four days it tends to fill the low places, wash the tracks, etc. and. unless from three to four times as many pumps are installed as are normally needed, considerable', output is lost. In most cases the water will .flow from the mines, so the obvious solution is to provide drainage .ditches of ample size to carry the maximum flow expected, so that the inflow will have no opportunity to flood the working places.

Jan 1, 1921

Page McGREGOR, A. G.-Features of the New Copper Smelting Plants in Arizona. Discussed by L. D. Ricketts, E. P. Mathewson 2145 HOWARD, L. O.-The Basic-Lined Converter in the Southwest. Discussed by Walter Douglas, E. P. Mathewson, Kuno Doerr 2145 SAMUEL, J. MOORE-Determination of Dust Losses at the Copper Queen Reduction Works. Discussed by Walter Douglas, E. P. Mathewson, S. J. Jennings, C. E. Arnold, A. G. McGregor, L. D. Ricketts 2148 MORSE, H. W., AND TOBELMANN, H. A.-Leaching Tests at New Cornelia. Discussed by H. W. Morse, Lawrence Addicks, G. D. Van Arsdale, S. J. Jennings, F. S. Schimerka, C. G. Grabill, F. N. Flynn, Stuart Croasdale 2151 LAIST, FREDERICK, AND ALDRICH, HAROLD W.-The 2,000-Ton Leaching Plant at Anaconda. Discussed by F. N. Flynn, E. P. Mathewson 2159 ADDICKS, LAWRENCE-Possibilities in the Wet. Treatment of Copper Concentrates. Discussed by F. N. Flynn 2160

Jan 12, 1916

By C. O. Brawner

INTRODUCTION Tailings dams up to about 200 ft. high are proposed to store tailings for a major mining operation on Marinduque Island in the Philippines. The original design of the dam utilized a compacted laterite starter dam followed by hydraulic placement of cycloned sands to form slime retaining dykes along the outer face of the dam. The upstream method of construction was proposed. In view of the failure of numerous tailings dams in Chile during the EL Cobre earthquake (I), an investigation was performed to assess the earthquake stability of the proposed dam. This included a soil investigation on the first cycloned sand dyke, which had been built to a height of 45 ft. Results indicated that the density of the sand was much lower than existing empirical criteria consider to be necessary for resistance to liquefaction. As a result, the dam was redesigned. The downstream method of construction was selected and a yearly construction program was developed to make use of sand available by cycloning. To increase shear strength, improve resistance to liquefaction, and to utilize steeper slopes, compaction of all sand in the dam was specified. A special underdrainage layer and toe drains were incorporated in the design. Details of the decant system and surface erosion alternatives are described. SITE CONDITIONS The mining project is located in the Philippine Islands. At the present annual rate of production, about 4,230,000 cu.m. of tailings will be produced yearly. The volume requirements for tailings storage design for the first 10 years of operation are in the order of 45,000,000 cu.m. The topography near the mine is mountainous with many gullies and canyons carved out by high intensity rainfall and runoff. The bedrock, which is generally shallow and weathered for a considerable depth, is overlain by laterite soils. Vegetation is comprised largely of palms, banana trees, and tropical shrubs and grasses. Precipitation is heavy. The average yearly rainfall over 24 years is 65.2 in. and the maximum recorded 24 hour rainfall is 9.0 in. High intensity rainfall is associated with typhoon storms, which are a seasonal threat inthe Philippines. Fig. 1 shows a summary of monthly precipitation.

Jan 1, 1972