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By J. Moore Samuel

Before the year 1909, no measurements of dust losses and flue gases had been made at the Copper Queen Reduction Works, at Douglas, Ariz. At that time the "unaccounted" loss of the smelter had reached a figure that drew attention to the possibility of large stack losses. The measurement of dust losses had not been studied much at that time, and estimates of dust losses, with a few exceptions, were the results of tests by what has been termed '(chemical methods," that is, the aspiration and filtration through cotton, wool or some similar medium of a few liters of gas, the dust being weighed on a chemical balance, and the volume of gas computed from the volume of aspirator water. The possible error of this method is enormous, and the probable error much too large for reliable results. Since the object of these tests was to discover not only the extent of the dust losses, but also the possibility of profitably reducing them, it was decided that the method adopted must have the following characteristics: 1. The sample of gas taken for filtration must be representative of the whole volume of gas, that is to say, it must have the same dust content. This condition will be fulfilled in all cases if the gas sample is drawn off without change of rate of flow. Then it will contain the same proportion of entrained or suspended particles of dust as the original gas. In the case of flue gases which contain only fume, at a temperature above or near the sublimation point, this condition need not be adhered to since the fume will not behave as entrained particles, but more nearly as a true gas. Consequently the gas sample in this case only may be drawn off at any rate of flow. 2. The gas sample taken must be large enough (a) to yield sufficient dust to permit of accurate chemical and physical examination; (b) so that the ratio of the total volume of gas passing through the flue to the volume of gas sample taken shall not be too large.

Jan 1, 1917

By Herbert Haas

ChEap power is essential to large-scale mining and metallurgical operations, particularly where fine grinding of large tonnages has to be resorted to, as is the case with an increasing number of mines treating disseminated copper ores, or low-grade gold ores, in the grinding of which considerable power is consumed. A comparison of the cost of generating power with Diesel engines and steam turbines and a brief discussion of the different factors governing the profitable use of either type of prime mover may be timely in view of the importance of a cheap supply of power to the mining industry, and the prominence that has lately been given to Diesel engines in America. As many of our most important mines and largest producers of metals are located in sections where neither cheap coal nor water power is available, and the price of oil fuel is increased by long hauls, the Diesel engine will become a prime mover of increasing importance. This applies particularly to the Southwestern States, where an ample supply of cold water for condensing purposes is unusual and local climatological conditions interfere in securing the high vacua essential to high turbine efficiency. About 12 times as much water will be required for condensing purposes with steam turbines under general conditions as is needed for jacket-cooling of two-stroke cycle Diesel engines; four-stroke cycle Diesel engines require only one-twentieth of the cooling water. Another condition favoring the use of Diesel engines is the high load factor at which most mine power plants operate. The influence of load factor on the cost of generating power is shown in Tables 1. and 2 and in Fig. 1. The higher the cost of fuel and the load factor, the greater will be the proportion of the fuel to the other costs making up the total cost of generating power. Capital charges (interest and redemption), maintenance, operating labor, and lubrication are practically constant amounts regardless of the load-output of the station; the fuel consumption and its cost are practically the only variables with the load. The costs here given are based on conditions as generally found in the Southwestern States, and on Diesel engine operating experience in European countries, where this type of prime mover has reached a high state of perfection, units upward of 4,000 hp. having proved entirely successful in operation.

Jan 1, 1917

By A. G. McGregor

During the past 5 years, five new copper-smelting plants have been built and put into operation in the State of Arizona. The monthly copper output from these plants averages from 5,000,000 to 18,000,000 lb. Previously, there never was as much activity in copper-smelting plant construction in the same length of time in all the rest of the world. Naturally, in this amount of work, some new problems were met and new features in plant design and equipment developed. Some of these features are described in this paper. Concentrate Hauling, Unloading and Sampling System (a) Concentrate Car The International Smelting Co. has provided at its plant at Miami specially designed cars for hauling the concentrates from the mills of the Inspiration Copper Co. and the Miami Copper Co. The car has a bottom that is readily made tight to hold flotation concentrates and is quickly and cheaply unloaded. Figs. 1, 2 and 3 show views and details of the car. The main feature of this car is the slot in the bottom throughout its length, which is closed by means of short planks that overlap like ship-lap as shown by Fig. 1. At one end of the car an iron gate operated by a screw and hand wheel is provided, which clamps or presses the planks together, making the bottom tight. A removable tapered plug rests vertically over this gate when the loading of the concentrates at the mill commences. A concentrate unloading pocket (Fig. 4), 180 ft.long, is provided at the smelting plant, which has a conveyor belt running throughout its length underneath the track. To unload a car, the vertical plug at the end of the car is lifted by means of a chain hoist which is supported by a trolley overhead. The gate at the end of the car is opened approximately 18 in. The removal of the plug leaves a hole down through the concentrates. The concentrates are poked through the hole from the top of the car until it is convenient for a man to stand on the bottom and draw one of the slats in the bottom toward the gate end of the car. The concentrates are then raked down with hoes on their natural angle of repose,

Jan 1, 1917

By Albert E. Wiggin, Frederick Laist

Early ill 1914 it was decided to test, on a fairly large scale, the treatment by flotation of Anaconda slime and mill tailing. For this purpose a standard-type Minerals Separation machine was installed at the Washoe Reduction Works during May and June, 1914. This was followed by the installation of a full-size Callow pneumatic machine plant. Experiments were also made, on a smaller scale, with the Froment, the Towne, the Fields, and the Anaconda flotation machines. The last-named machine was developed at this plant. In addition to the tests made in the standard-type Minerals Separation machine some tests were made using a Minerals Separation machine of the sub-aeration type. During the series of experiments a large variety of oils was tested. Experiments were also conducted using both round-table feed and tailing to determine whether it would be better to displace the round tables by flotation for the treatment of the slime, or to supplement the round tables by flotation of the round-table tailing. A series of tests was also made on the treatment of the mill tailing by grinding followed by flotation to determine the relative merits of flotation and leaching for the treatment of this product. In addition, flotation tests were made on mixtures of mill tailing and slime. The round-table feed referred to above is the total slime from the mill. It contains about 35 per cent. colloidal solids and approximately 90 to 95 per cent. of the total solids will pass through 200 mesh (0.067 mm.). It assays from 2.3 to 2.6 per cent. Cu. The mill tailing referred to above is the total discard from the mill, exclusive of the slime. It is all finer than 2 mm. and about 90 to 95 per cent. will remain on 0.25 mm. It assays about 0.60 per cent. Cu. A brief summary of the experimental flotation results follows:

Jan 1, 1917

By Leon Dominian

Page I. Introduction......................237 11. Coal..................239 (a) Western Asia Minor.........239 I-. Lignites of the Marmora-Bgean zone.239 2. The Black Sea basin.............241 (b) Eastern Asia Minor..........244 (c) Syria................246 (d) Mesopotamia................ 246 (e) European Turkey.'.....'..... 247 111. Oil and Allied Hydrocarbons...............248 (a) Early Use of Oil and Bitumen in Turkey........248 (b) Syria........................ 250 (c) Mesopotamia.................... 250 (d) European Turkey.................. 256 I. INTRODUCTION Apart from local needs in the country's development, the interest attached to Turkish fuels grows chiefly out of the importance of the railway industry in Asiatic Turkey. Geographical conditions make of the tracks in this region the most essential link in a chain of world-girdling lines. By its position Asia Minor can be likened to a bridge in the through routes destined to connect European factories with Asiatic and African markets. An uninterrupted right-of-way from points in central Europe to populous Indian cities must necessarily pass through the valleys of Anatolia. Crossing thence the Taurus Mountains, its natural passage is indicated as the Mesopotamian lowland and the shores of the Indian Ocean. Similarly, land connection between Europe and Africa can be obtained only by passage through Asia Minor, Syria and Palestine. The occurrence of fuel within an area of such significance from the

Jan 1, 1917

By Waldemar Lindgren

At the time of the discovery of America the old world had a scant supply of the precious metals. Both the northern and the southern part of the new continent proved wonderfully rich in gold and silver and its treasures were eagerly looted; though the looting has lasted four centuries, the mines of its mountain chains are far from being exhausted. Even the later discoveries in Australasia and eastern Siberia could not rob the Western Hemisphere of its position as the greatest gold and silver producing region of the world, though finally the developments in a narrow and circumscribed area in South Africa wrested from the Americas their supremacy in the production of gold. Nevertheless the history of the two parts of the great western continent has been strikingly different. At first the Spaniards extracted vast treasures of silver from Mexico, Peru, and Bolivia, while Colombia arid some placer deposits in Peru yielded a smaller quantity of gold. A couple of centuries later, a stream of gold began to flow from Brazil, the silver production from the countries mentioned above continuing strong in the meanwhile. Later on, the yield of South America diminished, but to offset this there began a wonderful series of discoveries in North America. The gold fields of California astonished the world; and when the cream of these had been skimmed off there began a no less amazing development of the Central Cordilleran gold and silver districts, which soon made the United States the greatest producer of the precious metals. Aided by ever improving technique, extensive exploration, and a system of railroads, the yield was maintained and increased. Still later followed the discoveries of the gold fields of the arctic region and silenced those who had maintained that the zenith in gold production had passed. Recently the province of Ontario in eastern Canada

Jan 1, 1917

By Rudolf Gahl

Page Tests Conducted in Small Test Mill................... 577 Demonstration Tests Conducted by Minerals Separation Co....... 577 Flotation Tcsts Conducted by Inspiration Co............. 577 Sampling of Orebody........................ 578 Other Results Obtained in Small Test Mill.............. 580 Tests Conducted in Large Test Mill................... 580 Flow Sheet............................. 580 Discussion of the Value of Preliminary Table Treatment......582 Callow Flotation Installation..................... 585 Flinn-Towne ~nstallation...................... 587 Cole-Bergman Installation...................... 588 Methods Followed in Competitive Tests................ 590 Advantage of Pneumatic Flotation Machines.............. 591 Pneumatic Machines of Minerals Separation Co............. 591 Float Skimming Device......................; 594 Experimental Inspiration Flotation Machine.............. 596 Arrangement of Callow Cells in Series ................ 598 Variations of Flow Sheet....................... 599 Coal Tar as Flotation Agent..................... 600 Experience with Primary Slime.................... 602 Influence of Iron on Flotation..................... 605 Influence of Chemicals on Primary Slime............... 607 Fineness of Crushing Desired.................... 607 Operation of Large Concentrator..................... 608 Flow Sheet of Callow Sections.................... 608 Flow Sheet of Inspiration Sections.................• 609 Inspiration Flotation Machine of Steel Construction.......... 610 Minerals Separation Section..................... 613 Filter Plant............................ 613 Results Obtained in Commercial Concentrator............... 616 Tonnage............................... 616 Air and Power Consumption...................... 616 Floor Space............................ 617 Water Consumption......................... 617 Recovery............................. 618 Oil Consumption.......................... 618 Operating Costs for Flotation..................... 620 Discussion of Results Obtained.................... 620 Prospects for Futurc Development of the Flotation Process......... 621 Porous-Bottom Experiments..................... 622 Raising the Grade of Concentrates.................. 624 Recovering Carbonates by Flotation................. 624 Recovery of Silicates by Flotation.................. 627 General Theory......................... 628 Conclusion............................... 630

Jan 1, 1917

By H. W. Morse

The experimental work on the oxidized copper ore at the New Cornelia mine at Ajo, Ariz., ended on Jan. 12, 1916. On that date final decision was made on the general nature of the process to be used in the 5,000-ton leaching plant, and on many of the details, as far as experience on a 40-ton scale could decide them. With the approval of the Board of Directors and the General Manager, John C. Greenway, we have compiled what seem to be the most interesting data on the results obtained during the experimental period. The most important part of the data resulted from the operation of a 1-ton and a 40-ton plant at Ajo. Experimental work on Cornelia oxidized ore dates back to April, 1912, and has been going on nearly continuously since that time. A good many variations from the original idea have been tried, but the process finally decided upon is in principle and in all of its details a simple one. The history of the leaching work on this ore has been brought down to about a year ago in papers presented to this Institute by Stuart Croasdalel and Dr. L. D. Ricketts.2 The present paper will, therefore, deal principally with the results obtained during the last year of the work. Preliminary Tests by Mb. Croasdale Preliminary tests were begun by Stuart Croasdale in July, 1912. A number of important points were definitely decided by his work. Among these were: 1. New Cornelia oxidized ore, not very finely crushed (to about 2-to 3-mesh) showed good extraction with 5 per cent. sulphuric acid. It appeared that better than 80 per cent. extraction could be expected. 2. The consumption of acid for this extraction was not prohibitive. It was apparently 3.5 to 4 1b. gross per pound of copper recovered.

Jan 1, 1917

By H. Kenyon Burch

PaQe Introduction.............................................................. 708 Plant Sites................................................................ 709 Type of Construction...................................................... 710 Mine Plant............................................................... 710 Dual Arrangement.................................................... 711 Underground Haulage............................................. 711 Tipples........................................................... 712 Underground Pockets............................................. 712 Automatic Measuring and Loading Devices......................... 712 Shafts ............................................................. 716 Ore and Spillage Skips............................................ 716 Automatic Hoists................................................. 716 Man Elevator..................................................... 717 Crushing Plant....................................................... 718 Storage Bin........................................................... 720 Sampling Plant....................................................... 720 Motors and Other Equipment.......................................... 721 Compressor Equipment................................................ 721 Concentrator............................................................. 722 600-Ton Test Mill..................................................... 722 The 15,OO0-Ton Concentrator...................................... 726 Grinding Section.................................................. 726 Floor Arrangement........................................... 726 Marcy Ball Mills............................................. 729 Performance of Crushing Machines............................ 730 Flotation Section.................................................. 780 . Flow-Sheet with Callow Flotation Cells....................... 731 Flow-Sheet with the Inspiration Flotation hfachines............. 731 Flotation Air Supply.......................................... 733 Concentrates Filter Plant.......................................... 733 Drag Classifiers................................................... 733 Tailing Settling System........................................... 735 200-Ft. Dorr Thickener ...................................... 736 General Equipment and Supplies.......................................... 737 Water Supply......................................................... 737 Power Supply......................................................... 737 Transformer Stations.................................................. 738 Shops and Railroad Facilities.......................................... 739 Conclusion................................................................ 740

Jan 1, 1917

By Robert E. Tally

Underground fires have been common in the mines of the United Verde Copper Co. for the past 22 years. The first fire started in the 300 Hampton stope in the fall of 1894, following a cave in that orebody. The soft and highly pyritic nature of the ores is responsible for most of these fires. Since ore characteristics, methods of mining and ventilation, have an important influence on mine fires, they will be described in this paper. Orebodies The ores are mostly chalcopyrite except on the upper levels where considerable chalcocite, black oxides and other secondary minerals occur. The orebodies are in the form of lenses and vary in size. Some are small, while others are several hundred feet in length by a hundred or more feet in width. These orebodies extend, in some cases, from the surface to the lowest levels of the mine and usually at a steep angle of inclination. There are three different classes of ores, classified in accordance with the gangue rock. The copper in all cases, except as above specified, is in the form of chalcopyrite. The largest and most important of these reserves are in a pyrite gangue. Next in order of importance are the schist ores and lastly the quartz porphyries. The pyrite, locally termed iron ores, average from 30 to 42 per cent. sulphur. The schist and quartz porphyries vary from 3 to 25 per cent. sulphur. Mining Methods All the orebodies were worked by the overhand square-set method of mining prior to 1908; since then the cut-and-fill, shrinkage and mill-hole systems have been largely employed. But 15 per cent. of the total tonnage now extracted is from timbered stopes. Most of this is from the upper levels where the ore is soft, heavy and more or less broken to the surface. Large reserves of these ores are being conserved for more efficient methods of mining.

Jan 1, 1917

By Wilbert G. McBride

Two Also 3 1/2-ton motor trucks were used by Young Bros. while operating at the Mammoth Collins mine at Shultz, Ariz. One was equipped with an oil tank holding 1,075 gal. and was used for the transportation of "tops." The other was fitted with a stake body and used to carry machinery, wood, rails, pipe and all classes of miscellaneous supplies. The bodies were made of oak with maple flooring and were attached to the frame of the chassis by U-bolts, to avoid drilling the main members of the frame. Most of the hauling was done from Tucson, a distance of 47 1/2 miles. During the first 3 months, part of the road was in very bad condition and the tire cost was excessive. After this part was repaired the road was in fair condition, but never good. There were no excessive grades or bad sand, but wagon ruts, too narrow for the truck wheels and of a different gage, caused heavy tire loss; while chuck holes, sharp curves and stones, both imbedded and loose, were objectionable features. During wet weather the trucks could not get sufficient traction to climb some of the hills and were likely to stick in the mud in certain places, so that no attempt was made to run them unless they were on the' road when the rain started. This lost time amounted to about 5 per cent. of the total, but, whenever possible, it was utilized in making minor repairs. The price of gasoline was from 17 to 21 c. per gal. Rubber tires were used throughout. Drivers were said $4.50 to $5 per shift, and a return trip to Tucson was counted as two shifts even when made in one day. Drivers were provided with a room in Tucson and were paid for a11 time lost due to causes beyond their control. Trucks were loaded one way only. Speedometers were placed on both trucks but the excessive vibration soon caused them to fail. For this reason, and because no account was taken of the distance covered in picking up a miscellaneous load or in other minor ways, the mileage given is under the actual distance traveled. Some of the weights had to be estimated, but care was taken to have the number of ton-miles low rather than high, to avoid underestimating the costs. The cost of hauling from Tucson to the mines was $12 per ton

Jan 1, 1917

By G. R. Lehman

The Inspiration Consolidated Copper Co. had an orebody at Miami, Ariz., of close to 100,000,000 tons of low-grade copper ore, and the method of mining this ore most profitably was of great importance. The selection of a method was practically limited to a "top-slicing," a "shrinkage-stope" or a "caving" method. The "top-slicing" method would give a high extraction of the developed ore, at a high mining cost, and the "shrinkage-stope" or "caving" methods a lower extraction of the developed ore, at a lower mining cost. Ore-Drawing Tests To decide on the method to be used it was necessary to know about what extraction of the ore could be obtained by the "shrinkage-stope" or " caving " methods. As no definite information regarding the,question was available at the time, C. E. Mills, the general manager, thought that some information could be obtained by experimenting in drawing ore, covered with capping, from an experimental box. Some work had already been done along these lines by W. C. Browning for the Inspiration Copper Co. and by myself for the Live Oak Development Co. The idea was to represent as nearly as possible, in the experimental box, the conditions within an area of caved stopes of caved ore all ready for drawing. That is, the area was supposed to represent a semibroken mass with the capping above it in the same condition and ready to follow the ore downward as it was drawn out of the chutes under the ore. To represent the above condition, in the experimental box, crushed ore from the mine was placed in it to a given height, and red barren capping from the mine was placed on top of the ore. The Experimental Box The experimental box shown in Fig. 1 was made of wood and glass. The glass was used for the sides so that the drawing action of the outside row of chutes could be seen and noted, the latter being cut in the center by the glass sides. The length of the box was 30 in., width 20 in. and the

Jan 1, 1917

By W. E. Gaby

SINce the time of their discovery, the genesis of the ores at Mount Morgan, and the nature of the changes which have affected the surrounding rocks, have been the subject of investigation and speculation by geologists and mining engineers. Developments at this mine, for a long period the greatest gold producer of the world, show that with depth the orebody grades into copper, of which metal it now furnishes a large output, exceeding in value the gold. In earlier times, the purity of the gold and the first extremely heterogeneous character of the ore made it an interesting problem. Now, the probable depth of the copper enrichment, genesis, and true relations of the neighboring rocks furnish equally profitable material for investigation, and the present somewhat detailed description of these ores and rocks, from microscopic study, may prove of interest. , The mine occupies a highly silicified portion of a belt of quartz-porphyry, which is believed to be the oldest rock of the region. This belt of quartz-porphyry and allied rocks runs north and south between two areas of a later intrusive granite, as shown on the geologic map,' Fig. 1. Near the orebody the quartz-porphyry has been intruded by a dense basaltic mass called the " Old Basic." Cutting all three of these igneous masses are two sets of basaltic dikes, one running northwest-southeast, and the other, cross dikes, northeast-southwest. These dikes may be contemporaneous with, instead of later than, the ore. Their number has caused peculiar difficulties in extracting it, as the clay selvage of decomposed material affords no adhesion between dike and ore, causing great blocks to loosen and come down. The investigation was based on specimens furnished by B. Magnus, former general manager of the mine, and notes on the areal geology by

Jan 1, 1917

By Samuel S. Wyer

The magnitude and economic importance of the problem of correctly valuing natural-gas leaseholds become evident when me consider that: (a) Natural gas is handled in 55 per cent. of the gas distributing plants in the United States. (b) Present known natural-gas acreage forms 47 per cent. of the total known mineral-land acreage in the United States. (c) Five acres of land are now required to protect and maintain continuous service to each of the 2,000,000 domestic natural-gas consumers in the United States. (d) The cost of acquiring and maintaining this acreage of an expendible resource represents a substantial part of the cost of the natural-gas service to the consumer. (e) "The right of a citizen by means of his ownership of or his mining leases on land to draw gas from beneath its surface is property and sometimes valuable property."' A large number of other court decisions have established the basic legal principle that the right to drill for gas within a given area constitutes an interest in the land itself, and is necessarily an exclusive property right. Definition of Vested Interests or Rights "Vested interests are economic interests which are legally recognized to be such that they cannot be impaired by public action, directly or indirectly, without indemnification. Vested interests are largely property interests. The recognition of an interest as a vested interest gives it some of the attributes of property, and by American courts it would be comprised in their very inclusive concept of property. Otherwise than through property, vested interests generally arise through contract."

Jan 1, 1917

By Edward A. Sayre

In shaft sinking for cod mines, the cost item greatly influences the method adopted. This holds true especially when soft material must be traversed. The average life of a coal mine is short. This is due either to the limited area of the coal basin or to the great expense of maintenance and haulage underground when the entries are extended a considerable distance from the hoisting shaft. Therefore, the engineer in opening a coal mine is confronted with the fact that the cost must be kept within certain bounds else the venture will prove unprofitable. For this reason he may be prohibited from adopting certain safe, but expensive, methods that are used in sinking caissons for foundations, driving transportation tunnels, or other work of permanent construction. A common method of sinking through difficult ground employs the aid of a steel shoe pushed ahead of the shaft timbers. Another is the drop-shaft method. These two methods were used in sinking the main shaft and the air shaft of the Eagle No. 3 mine, at Des Moines, Ia., and the following data show the relative success and cost of the two methods of sinking under the same conditions. A drill log of the material to be penetrated was as follows (see Plate 1, Fig. 1): Thickness, Total Depth, Ft. In. Ft. In. Drift................................... 38 38 Sand..................................... 2 40 Drift.................................... 2 6 42 6 Jointed clay.............................. 5 6 48 Drift, firm............................... 10 6 58 6 Clay, with sand streaks.................... 3 61 6 Sand.................................... 12 73 6 Shale, with two thin beds of coal and thin strata of rock......................... 86 6 160 Rock.................................... 5 165 Coal..................................... 3 9 168 9 Main Shaft Sunk by Steel-Shoe Method This record shows the first 73 ft. to bc of drift material, the lower 12 or 15 ft. of which was sand. The remaining 92 ft. to the coal was shale and rock, and offered no trouble in sinking.

Jan 1, 1917

By Philip D. Wilson

The mines of the Calumet & Arizona Mining Co. are situated in the Warren Mining District, Cochise County, Arizona, between Bisbee and Warren and adjoin those of the Copper Queen Consolidated Mining Co. and the Shattuck Arizona Copper Co. Paleozoic 1imestones.of considerable thickness have been intruded by granite porphyry. A great stock of the intrusive outcrops boldly near the city of Bisbee and is known as Sacramento Hill. From this central core the porphyry has tongued out into the limestone for many thousands of feet as dikes and sills of irregular shape and variable dimensions. Churn drilling has recently developed a large tonnage of secondarily enriched copper ore in the main porphyry mass of Sacramento Hill. While some ore has been found in porphyry in other parts of the district, it is relatively unimportant compared with that which occurs in the Paleozoic limestone as irregular lenses replacing favorable beds near porphyry contacts or associated with well-marked fracture zones often many hundreds of feet away from the intrusive. In a zone extending from Sacramento Hill for a distance of from 1,500 to 3,000 ft., the limestone has been extensively metamorphosed and in this contact metamorphic zone the ore occurs through a known vertical thickness of upward of 900 ft. and individual orebodies often show extension in a vertical direction of over 200 ft. At a greater distance from the intrusive stock, the ore occurs as lenses of greater horizontal than vertical dimensions with an average thickness of approximately 30 ft. An exception to this rule is found in the occasional steeply dipping mineralized fractures, almost true tabular veins, which are found long distances a way from the zone of contact metamorphism. The copper was probably introduced during or slightly later than the porphyry intrusion in the form of chalcopyrite and bornite. These primary copper minerals are always associated with pyrite and often occur as rich lenses in a huge mass or shell of pyrite carrying an insignificant percentage of copper. In many instances the copper-bearing solutions were preceded by siliceous solutions and these latter, often persisting during the period of mineralization, have thoroughly indurated the ore-

Jan 1, 1917

Clarence M. Haight, Franklin Furnace, N. J. (communication to the Secretary*).—In that part of Mr. Wilson's paper describing the Gilman cut-and-fill system, a few features do not appear to be fully explained. In what way is the waste used for filling obtained? The text mentions that if waste is encountered when driving the filling raises, these raises are stopped until the fill is needed; that seems to suggest mining the capping for fill, which might make the backs of the stopes bad as the stopes approach the top of the ore. If, on the other hand, the fill is brought to the raises in cars, are the crosscuts that are necessary in ore or waste? Is the cost of this fill included in the stoping costs given? The second, or lower, series of stopes will present a different problem to that which the first series (under the capping) gave. Although no mention is made in the text, it would seem that the second series of stopes will have to be worked on center lines which are offset half the stope width, 20 ft., from the centers of the stopes above, in order to take advantage of the "rafter" effect of the stringers so named. How is the fill to be introduced into the lower stopes? There are no crosscuts mentioned to provide for this; or is it intended to tap the filled stopes above for the fill? In the latter case the method of control of this fill will be of interest. Ore that will admit of a stoping width of 40 ft. with the lengths mentioned in the description is stronger than any with which many mining men are familiar. Has the practice of starting to mine a second stope before the preceding one is filly completed been tried to any extent? If

Jan 1, 1917

By David Cole

At the time flotation appeared upon the metallurgical horizon in Arizona, the writer, under the direction of Dr. Ricketts, was engaged in remodeling and enlarging the No. 6 Concentration Plant of the Arizona Copper Co. at Morenci, and the work had been in progress nearly a year before the Inspiration experiments with flotation had disclosed the revolution in concentration that was at that time impending. The simplified flow sheet worked out for the remodeling of the Morenci plant had been based upon the removal of the freed metal in a minimum number of stages by treatment upon tables equipped with Butchart riffles, the latter being adapted to accomplish both classification of feed and removal of the metal at one operation, substantially as described in my previous paper.l The scheme of treatment for the slime was based upon the well-known fact that after copper sulphides, such as chalcocite and chalcopy-rite, etc., are reduced to a certain extremely fine state of comminution they are definitely beyond the reach of separation upon any of the concentrating devices then known. By using drag-belt classifiers (which served as conveyors as well as separators), the overflow ,would be of the usual "slimes" class. Experiment had shown that when these drag-belt overflows were properly diluted (that is, to about 5 to 7 per cent. solids in the feed under treatment) the very fine sand, and especially the very fine but still granular sulphide particles, would, if given a short distance to fall, quickly settle out in prime condition to yield an excellent recovery on vanners; a further valuable feature would be that the very fine non-separable final overflow tonnage might be discharged direct to tailing, thus conserving space requirements so as to permit within the old building, without embarrassment, the increase of capacity required. For reasons that are obvious, this kind of feed preparation for the slime could not be successfully accomplished in any form of pointed boxes or

Jan 1, 1917

By E. D. Gardner

Apparently the widespread agitation for the codification of our mining laws has had its effect, and it is quite possible that Congress will take up the question during this present session. The greatest objection to our present statutes pertaining to metalliferous deposits seems to be directed against the law of the apex. Those who are not in favor of changing this law are very far in the minority and they apparently realize that it is useless to urge its retention. Next in importance to the repeal of the law of the apex, seems to be a widespread advocacy for the repeal of the law requiring discovery of mineral before a valid location of a mining claim can be made. To a lesser extent there appears to be a desire to have the law changed so as to allow patent for claims without mineral discoveries. The purpose of this article is to review the arguments in favor of the proposed change of the discovery requirement, to show some of the evils that may be expected to result therefrom, and to consider whether a system may be devised to correct the undesirable features of the present law in this respect and at the same time not expose the mining fraternity and the general public to the evils which would result from the repeal of this fundamental requirement. Under the law as at present framed, the validity of a mining location is dependent upon the fact of discovery of mineral within the boundaries of the land claimed. " * * * but no location of a mining claim shall be made until the discovery of the vein or lode within the limits of the claim located" (Sec. 2,320, U. S. Rev. Stat.). All authorities agree that discovery is the source of the miners' title. Lindley, in his valuable work on Mines (3d Ed., Sec. 335), lays down the rule in part as follows: "Discovery is the initial fact. Without that no right can be acquired. * * * Such discovery must precede the location, or be in advance of intervening rights. The proof of recording and marking a claim will not authorize the court to presume a discovery." As is well stated by the U. S. Supreme Court in Erhard v. Boaro (113 U.S. 536):

Jan 1, 1917

By L. O. Howard

What was perhaps the first attempt at basic converting in the Southwest was made by the late Charles F. Shelby at Cananea early in 1907, when he removed the acid lining from one of the 8 by 12-ft. barrel-type converters then in use at the reduction works of the Cananea Consolidated Copper Co. and substituted a lining of one course of magnesite brick. In this shell, he blew 21 taps (about 180 tons) of matte to white metal before the brick gave out along the tuyhre line. In every case, the white metal so made was transferred and blown to copper in other converters. The cost of the lining was $700 and Mr. Shelby gave up the experiment as impracticable. No further serious effort to bessemerize copper mattes in a vessel having a more or less permanent lining of basic brick was made by any of the large copper smelters in the Southwest until some months after Messrs. Peirce and Smith had brought the question permanently before the public, and had proved it to be a metallurgical and financial success by their work at Baltimore and Garfield. Early in 1911, the Cananea Consolidated Copper Co. had in operation at its reduction works at Cananea, Mex., several 8 by 12-ft. barrel-type converters lined with magnesite brick which had previously been operating with acid linings. In the 18 months following, these converters produced 50,000 tons of copper. In June, 1912, this company installed the first of its new equipment of six stands of electrically operated 12-ft. Great Palls type converters, and shortly afterward discontinued the use of the smaller shells. In September, 1911, the Consolidated Kansas City Smelting & Refining Co. had installed at its El Paso plant two 10 by 26-ft. Peirce-Smith converters. Later, in March, 1913, one standard Great Falls type converter was added. About the time the Peirce-Smith converters were going in at El Paso, the Calumet & Arizona Mining Co. was experimenting with basic linings in 7-ft. by 10-ft. 6-in. barrel converters at its plant in Douglas, Ariz. These efforts were sufficiently successful to enable the plant to "handle the total output of copper in these small shells until the standard Great Falls converters were substituted in June, .1914.

Jan 1, 1917