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By Joseph F. Shirley

A froth flotation plant to recover byproduct molybdenite must be designed specifically for the selected separation process and the type of ore being treated. However, some general rules concerning byproduct molybdenite plant design can be suggested that would be valid for most processes. The term byproduct molybdenite in this paper is limited to the recovery of molybdenite from copper porphyry ores, which essentially amounts to selectively separating molybdenite from copper and iron sulfide concentrates. In addition, most of the discussions would apply only to the almost universally used approach of floating the molybdenite and depressing the copper and iron sulfide minerals. Thickener It is considered essential that the feed to a byproduct molybdenite plant be thickened prior to being treated for molybdenite recovery for the following reasons: 1) Thickening allows the rougher and any preconditioning stages to be operated consistently at the optimum percent solids. 2) Generally, the size of equipment used for pretreatment and the rougher flotation is considerably smaller and less expensive than if the copper flotation concentrates were treated directly. 3) The removal of as much water as practical by thickening removes flotation reagents added in the copper flotation circuit, especially frother which can be quite troublesome in a byproduct molybdenite plant. Although thickening of the molybdenite plant feed has long been recognized as necessary to optimum metallurigcal results, very often this area has been designed poorly. The thickener should be sized so that it can produce a consistent thickened product containing 60 to 65% solids without the use of flocculants. Some plants are only able to thicken to 45 to 50% solids consistently while other plants let the percent solids vary from 45 to 60% over a 24-hr period. While the former is more desirable than the latter, in many cases better metallurgical results could be obtained if the percent solids could be held consistently close to 60%. Probably the simplest and most satisfactory system of handling the thickener underflow is to spigot it into a sump, using a sphincter type valve for the orifice, controlled by a density meter. Another satisfactory approach is to use an ODS type pump to meter the underflow. There are other methods of handling thickener underflow which work satisfactorily as long as the amount of copper concentrates entering the thickener is relatively uniform. However, they often fail to operate satisfactorily when there are large fluctuations in the thickener feed rate due to fluctuations in ore tonnage and/or copper grade. If the percent solids in the plant feed could be held close to 65%. There is the possibility the molybdenite plant tailing would not have to be thickened prior to the copper filter plant. This means that the molybdenite rougher float would have to be conducted at 45 to 50% solids with all the cleaner tailings returned to the rougher for dilution. Admittedly, in some situations this approach would be impossible, but if capital or plant area is limited, it should be considered.

Jan 1, 1981

By W. H. Wamsley

Used in combination with conveyor belt haulage, bucket wheel excavators offer unusual possibilities for low excavation and haulage costs. Originating in Germany, these machines are now in use or on order for Australia, Northern Rhodesia, the Belgian Congo, Italy, Yugoslavia, France, Belgium, Indonesia, and the Union of South Africa. A diamond mine in South Africa that has been using German bucket wheel excavators since 1928 purchased American cable-way excavators and American scraping equipment after the war. Operating cost for this equipment was double that for the bucket wheel machines operating in the same mines with crawler-mounted stackers. Machines built and operated in the U. S. are, as far as the writer knows, confined to the northern Illinois coal field, where they are used in conjunction with large shovels to strip overburden too high to be reached by the 30 to 40-yd shovels. Their long stacker belts can deposit the spoil far enough behind the coal face to avoid slides, in such a manner as to make reclaiming the land easier. These large, specialized machines are now being made by Bucyrus-Erie Co. but are not adaptable to many open pit mining situations.

Dec 1, 1956

By Robert H. Merrill

During the period from 1960 through 1966, the U.S. Bureau of Mines and Kennecott Copper Corp. performed an experiment to determine the changes in stress, strain, and displacement created by changes in slope angles. The experiment was performed at Kennecott's Kimbley pit near Ely, Nev. The supporting laboratory studies were made in the Bureau's Denver Mining Research Center, where final reduction of the data and summary reports are now in progress. The experiment was divided into two phases. The first consisted of the laboratory and field investigations required to select a new slope angle and configuration; the second phase consisted of a full-scale instrumented study in the west wall of the pit. The west slope was changed from an average of 45º to a slope angle that started at 45º on the south wall and gradually in- creased to 57 ½ º at the center of the west wall, and to 61 º on the north wall. The pit was deepened from 500 to about 550 ft, and approximately 2.55 million tons of waste and ore were removed from the wall.

Jan 9, 1969

Solution mining got back in the news recently when Kalium Chemicals, Ltd., a joint subsidiary of Pittsburgh Plate Glass and Armour & Co. announced its plans to go full speed ahead in mining potash by the solution method in its vast holdings 30 miles west of Regina, Saskatchewan. The company is keeping the technical details of this mining technique under wraps for competitive reasons, but company officials insist that the major difficulties in solution mining of potash have been overcome and that the final product will have definite competitive advantages on the open market. Kalium's holdings in the Regina area total about 500 sq mi. The ore-primarily sylvite mixed with appreciable quantities of halite-occurs in the Prairie Evaporite formation at a depth of 5300 ft. The ore zone averages about 100 ft in thickness. Too deep to be economically mined by conventional underground methods, such as is being done at International Minerals & Chemical Corp.'s Esterhazy mine (depth about 3100 ft), Kalium gambled that it could develop a successful method of solution mining even though earlier attempts in Germany, the U.S., and Canada had failed.

Jan 6, 1964

By John F. Shaw

The Denver Mining Research Center headquartered at Denver, Colo., is typical of the new research units recently activated as a result of increased emphasis by the U. S. Bureau of Mines on research concerned with metal and nonmetal (except mineral fuels) ore extraction. The Denver Center, together with the recently established Minneapolis Research Center and the Offices of Mining Research at Reno and Spokane, and the older, well-known Applied Physics Research Laboratory at College Park, Md., share the responsibility for the comprehensive mining research program. The Denver Center specializes in two areas of research: (1) The development of engineering principles and mathematical techniques applicable to exploration, development and operational problems in mining, and (2) the study of rock mechanics as applied to ground control. The first area presently includes three categories: research on methods of ore sampling, mine-systems analysis, and applications of statistics to problems of mineral exploration. The second area emphasizes research in two main categories: laboratory investigations designed to develop a fuller understanding of how rock be- haves under the loads imposed by mining, and field investigations to evaluate stress-strength relation- ships in full-scale, operating mines.

Jan 3, 1962

By A. B. Young

T HE smelting industry in Utah is represented by four plants: The Midvale of the United States Smelting, Refining & Mini.ng Co., the Murray of the American Smelting and Refining Co., the Garfield of the same company and the Tooele of the Inter- na.tiona1 Smelting Co. The first two are lead smelters, the Garfield plant is a copper smelter while the last named embraces both a lead and copper smelter. A total daily smelting capacity of over 10,000 tons is represented, of which 4000 tons is lead smelting capacity. With the exception of Tooele, forty miles distant, all of these smelters are within twenty miles of Salt Lake City. Pyrometallurgical operations in this district are con- fined entirely to the production of unrefined bullion, which is shipped east to the various refineries of the respective companies. The smelting industry in this state had its beginning away back in the latter days of the Civil War, many years before the completion of the trans-continental railroad. Ox teams were still bringing their companies of better fixed Mormon emigrants across the plains to the valley of the Great Salt Lake and those in lesser circumstances were still pushing their hand- carts over the same roads to the temporary abode

Jan 1, 1925

By Charles M. Means

A thoroughly modern coal-handling system has been installed at the Pittsburgh Terminal H. H. &. Coal Co.'s new No. 8 shaft, or Cover-dale mine, about 11 mi. (17.7 km.) from Pittsburgh on a spur of the Montour R. R. The Pittsburgh scam, which is the one worked, is reached at a depth of 342 ft. (104 m.). Power is delivered to the mine at 22,000 volts over a transmission line of the bow-arrow type of construction. A private telephone line connecting the various mines and the main office in Pittsburgh is carried on the same poles. The high-tension line is brought to an outdoor, step-down, transformer substation, where three 833-kv.-a., single-phase, 60-cycle, oil-insulated, self-cooled transformers reduce the potential to 2300 volts. The transformers are protected by electrolytic, aluminum-cell lightning arresters, horn gaps, and choke coils. High and low-tension busses, together with fuses, choke coils and air-break horn-gap switches with thcir remote-control operating mechanism are mounted on a substantial structural-steel tower 25 ft. (7.6 m.) in height. Mine Substation The transformer substation is adjacent to a. pressed-brick building that houses direct-current substation equipment together with switchboards, etc., for the control and distribution of power. This building also contains the hoist for the men and supplies. Main Switchboard The 2300-volt circuit is carried underground in a conduit from the transformer substation to an oil switch in the substation and supply hoist building. This is the main switch controlling the 2300-volt busbars mounted back to a twelve-panel black-slate switchboard. From the switchboard, 2300-volt circuits are carried to transformers feeding the street and house lighting systems and to other transformers furnishing power for the elevator and lighting circuits in the store building. Other 2300-volt circuits run from this switchboard to the main hoist building, fan house and auxiliary or man-and-supply hoist. High-tension wiring,

Jan 1, 1922

R. W. Raymond, New York City: In connection with the subject of multiple tuyeres, my attention has been drawn to the practicability of gaining, without the multiplication of tuyeres, the advantages which that measure is intended to secure ; and learning that this had been attempted in the invention of the Gaines radial-discharge or fan-shaped tuyere, I wrote to Messrs. Gustafson Brothers, of Sequachee, Tenn., who manufacture that tugere, for information concerning the invention itself and its performance in practice. To their courtesy I am indebted for most of the facts given below. I present these statements without expressing any personal opinion on the subject (having no personal knowledge of it), for the purpose of bringing this device to the notice of members, and calling forth both criticism and testimony concerning it. Admitting that definite advantages are to be secured by the multiplication of tuyeres, we must still confess that this change involves considerable expense in the increased number of blastpipes, tuyeres, coolers, connections, and pipes for water-circulation. There is, moreover, a serious question involved in the weakening of the furnace-wall by extra perforations. This consideration applies with special force to furnaces already built without reference to such an increase of tuyere-openings. Ledebur, in his Handbuch der Eisenhüttenkunde, says on this point (I translate from the edition of 1893, page 365): "A limit to increase in the number of tuyeres is given, in any case, by the circumstance that the perforation of the wall weakens it; in other words, that there must be left between the tuyeres a mass of masonry sufficient to carry the parts above. In most cases an appropriate number may be secured by giving to the furnace as many openings as there are meters in its circumference at the tuyerelevel. For instance, for a hearth-diameter of 2.5 meters, i.e., a circumference of 7.7 meters, seven or eight openings might be made, of which, however, one would

Jan 1, 1899

By J. L. Sepulveda, R. F. Yap, R. Jauregui

INTRODUCTION The Third Theory of Comminution, oftentimes called the Bond Theory, was first published by Fred C. Bond in 1952. Since then, it has been widely used in the milling industry to size crushing and grinding equipment as well as to predict and evaluate their performance. Although two other comminution theories were advanced almost a century earlier, one by Rittinger, 1867, and the other by Kick, 1885, their use was limited because they did not agree with actual commercial comminution systems (Bond, 1961). According to Bond's Third Theory, the work input is proportional to the new crack tip length produced in particle breakage, and equals the work represented by the product minus that represented by the feed. For particles of similar shape, the crack tip length is equivalent to the square root of one-half the surface area, and the new crack length is proportional to [(1/vT - vF)]; where P and F are particles diameter of product and feed, respectively. For practical calculations, the particle diameter can be expressed in terms of the 80% passing sizes, P and F, respectively. Therefore, the work input is given by, [10 _ 10', W = Wi [7P_7F(1)] Where W = the energy input to the mill, (kwh/t) Wi = Work Index, (kwh/t) P = square sieve opening which 80% of the product passes, (microns) F = square sieve opening which 80% of the feed passes, (microns) Equation (1) is the mathematical statement of the Third Theory, or the Bond equation. Numerically, the work index is the energy required in kwh/t to reduce a given material from theoretically infinite feed size to 80% passing 100 microns. The work index of a particular material is determined in the laboratory by conducting the standard Bond grindability test. Bond developed separate tests for the rod mill and the ball mill. The simplified test method described in this paper, which is used at Anaconda Research Center to approximate the work index, is an alternative procedure to the ball mill

Jan 1, 1982

By Carl W. Volz

NORWAY'S metallurgical development, which has extended over many centuries, is intimately associated with that country's unique topography and climatic conditions. It is a rugged mountainous country, with a most irregular shore line, bordered by thousands of islands forming a natural protection for shipping. There are an endless number of lakes, and innumerable rivers and waterfalls, many fed by glaciers and melting snow. Rainfall is abundant. Though half of the country is north of the Arctic Circle, it is warmed by the Gulf Stream, so that the average temperature is about 45°, or 20° higher than other parts of the world in the same latitude. A sub-

Jan 1, 1935

By Walter C. Smith

THE Cerro de Pasco Copper Corp. began to produce bismuth at the Oroya smelter in 1929, at which time the only important consumption of that metal was in the manufacture, of pharmaceutical compounds, and the normal supply of metal exceeded the demand. With a new and major producer entering the field, additional uses for bismuth had to be developed. A survey indicated that the most promising opportunity lay in fusible alloys containing bismuth. A large number of bismuth alloys were known and a few of them were being used in the manufacture of automatic sprinkler heads and safety plugs. The literature cover in,- the- alloys was found to be sketchy and contradictory. Research, development, and educational work carrried on by the Corporation during the past fifteen years has resulted in greatly enlarged uses for the metal. In general the low-melting bismuth alloys are of high specific gravity, relatively soft, of low tensile strength, brittle to shock, yet subject to plastic

Jan 1, 1945

By JOSEPH MACDONALD

STRIPPED of its romantic possibilities, mining is a commercial business, carried on for the profit there is in it, and the business of the manager, in its ultimate analysis, is to make the profit as large as possible. The general manager of a mining company should keep himself personally familiar with all the details of its operations. If be fails to do so, chaos results, labor becomes inefficient, costs go up, and profits go down. Nowhere is this emphasized more than in Mexico. If the mining operations are extensive and complex, involving, besides the work of mining the ore, its transportation and treatment by various milling operations, and the marketing of the products, the business, to be carried on properly, must be divided into a number of main departments, which are subdivided into minor departments. Each main department should be under the immediate charge of a superintendent, or a foreman, and each minor department under a shift-boss. Organized in this way, and with all the principal centers of activity connected by telephone, the manager can keep in touch at all times with the details of every branch of the business, and be in a position to perform effectively the full duty of manager. The ?total cost of mining operations may be broadly divided under the heads of Labor, Supplies, and Marketing. Of these, it will be found, on averaging a number of cases, that the cost of labor will be about two-thirds of the total cost. The importance, therefore, of an efficient labor-department will be apparent. This department must be sub-divided and organized differently, according to the peculiar conditions at each mine, and no one system, however well suited for one particu-

Jan 1, 1909

W. N. Triplett (Geology of the Silver-Lead-Zinc Deposits of the Avalos-Providencia District of Mexico, P. 583) was born in Tecumseh, Mich., and attended the University of Michigan and Massachusetts Institute of Technology. From 1912 to 1917 he worked for Moctezuma Copper Co., a Phelps Dodge subsidiary at Nacozari, Sonora, Mex. He served in Army as Second Lieutenant Field Artillery from 1917 to 1918 and from 1919 to 1922 returned to Moctezuma Copper Co. From 1922 to 1952 he has been with Compania Minera de Penoles, S. A., a subsidiary of American Metal Co., Monterrey, N. L., Mexico in Unit Dept. as chief geologist and at present is consulting geologist. Mining geology and operation is of particular interest to Mr. Triplett. He presented a previous paper before with M. W. Hayward entitled, "Occurrence of Lead-Zinc Ores in Dolomitic Limestones in Northern Mexico." Besides being an AIME member, he holds membership in Society of Economic Geologists. Fishing, hunting and photography are his favorite hobbies.

Jan 1, 1952

By John K. Berry

The handling and disposal of mine water is a much larger problem than is apparent at first glance. Many more tons of water are removed from underground coal mines in the United States each year than tons of coal. In the old anthracite region of Pennsylvania the average was around 36 tons of water for each ton of coal produced, while bituminous mines average about 5 or 6 tons. Water encountered in the mines may come from several different sources. In mines with light cover and without a firm solid roof, rain may seep directly into the workings. This is also true of deeper mines where pillar falls have broken the roof to the surface. The top under creek beds is generally faulty and broken and water comes through the roof at this point. Deeper mines are frequently overlain by water-bearing strata and pillar breaks and faulty top can admit water, while places running to the dip often accumulate water which comes up through the bottom. In all cases, except where passing under a year-round stream, the amount of water made is usually pretty closely connected with the weather-more water during the rainy season, less in the dry, although the deeper a mine is, the longer it is before changes in rainfall are felt. Another source of water, not connected with seasonal changes, is old water, accumulated over a period of time in abandoned workings of the mine, or from flooded adjacent mines. Here the water filters through the coal, the parting, or comes up through cracks in the bottom. Similarly, core holes, drilled to prospect the property, can be a source of unexpected water and modern safety requirements for dust-allaying water at the working faces can turn a pleasant working area into a soggy mess over the weekend if the system is not properly maintained and operated. Why water must be disposed is obvious. If allowed to accumulate, sooner or later the mine will fill up, or at least it will become "roofed" in low places, blocking ventilation and traffic. But it must also be controlled long before it reaches this stage. Water standing in the working faces is unpleasant for the men who have to work there and also cuts down on the amount of work they can do, as they must watch where they step all the time. It is especially unpleasant in low coal.

Jan 1, 1973

By William P. Blake

Postscript to the paper read by Prof. William P. Blake at the Lake Superior meeting, September, 1904. POSTSCRIPT.*-Since the publication of my paper upon the blackening of the surface of rocks in desert regions, I have noted that similar views of the origin of superficial discoloration in the case of certain rocks from the Salt Lake basin of Utah have been expressed by Prof. George P. Merrill. In the pamphlet, he described the brown-coated weathered boulders from Tooele county, reaching the conclusion that the discoloration was due to the solution of the manganese and iron compounds 'in the interior of the boulder, while they were in the water, and the gradual bringing of this material to the surface through capillarity and its oxidation when exposed.

Mar 1, 1905

By J. Gordon Parr, A. J. Goldak

OgdEN et al.1 and Bumps et al.2 suggested that the solubility of aluminum in a titanium extended to 30 pct.* Sagcl,3 Clark and Terry,4 Anderko et al.,5 Ence and Margolin6 and Saulnier and croutzelles,7 subsequently proposed the existence of an intermetallic compound in the region of 25 pct Al. Although Anderko et al.5 suggested that the structure of this phase, Ti3Al, was of the DO,, (Ni3Sn) type, their conclusion was based only upon the results of a powder photograph that included reflections from ? = 0 to ?= 27° (CuK),some of which were reported to be indistinct. Ence and Margolin,6 referring to a 25 pct alloy (but without indicating whether the figure referred to weight or atomic pct) stated: "The structure of Ti2Al is isomorphous with Ti3Sn DO19, as pointed out by Pietrokowsky". But the point was not proven, because Ence and Margolin did not report values of calculated intensities of Ti2Al with the DO,, structure. In fact, the observed intensities they reported are in poor agreement with the values we calculate for DO19 (see Table I). Saulnier and croutzelle7 concluded, on the basis of electron micrography and microdiffraction studies, that the structure of the 25 pct A1 alloy was ordered hexagonal. Our purpose has been to unequivocally determine the Ti3Al structure. Alloys containing 25 pct Al were prepared from iodide titanium and spectrographic standard aluminum by levitation melting. The levitation chamber was evacuated to 10-6 mm Hg before filling with 99.999 + pct A. After casting, the alloy was chilled to liquid air temperature and crushed to -200 mesh in a percussion mortar. The powder, wrapped in molybdenum foil, was annealed (by induction heating) at 1000 "C for 2 hr in the argon atmosphere of the levitation chamber. No sublimate was seen on the foil or elsewhere, and therefore the assumption is

Jan 1, 1962

By O. S. True, C. H. Herty

It is well known that manganese will desulfurize molten iron through the formation of manganese sulfide, which, being only slightly soluble in the metal, rises to and enters the slag where it remains as such (mixer slag) or is converted partly or wholly to calcium sulfide (basic open-hearth slag). The reaction by which manganese sulfide is formed is Mn + FeS = MnS + Fe. This reaction is generally considered to be reversible, although McCancel states that such is not the case. The experimental work was carried out to confirm the results of Roh12 as to the reversibility of the reaction. Six trials are herein considered. In each 200 gm. of electrolytic iron free from manganese and sulfur were melted in a pure zirconia crucible in a carbon resistance electric furnace. To the melted iron in runs 2, 3 and 4 metallic manganese and manganese sulfide were added to give 1, 2 and 4 per cent. manganese respectively, with 0.50 per cent, of sulfur in each. In runs 5, 6 and 7, metallic manganese and iron sulfide to give the same proportions of. manganese and sulphur as in runs 2, 3 and 4, were added. The melts were held molten in the furnace for an hour and then cast into small ingots. The results of these melts are as follows: Ingredients of Charge Analysis or InRot ManOaneBe, SulfUR, ManOaneBe. SULFUR, Run Material Charged PeR Cent. PeR Cent. Per Cent. PeR Cent. 2 MnS + Mn............... 1.0 0.452 0.636 0.367 3 MnS + Mn............... 2.0 0.50 1.57 0.310 4 MnS + Mn............... 4.0 0.50 2.81 0.230 5 FeS + Mn................ 1.0 0.50 0.647 0.397 6 FeS + Mn................ 2.0 0.50 1.45 0.374 7 FeS + Mn................ 4.0 0.50 2.73 0.195

Jan 1, 1925

US 4,132,758-Leaching of copper sulfide ore using nitrogen dioxide as the oxidant A slurry of ore in sulfuric acid is contacted with a nitrogen dioxide-containing gas at a temperature below 11 5" C and a pressure of from atmospheric to 35 PSI, the slurry agitated, and the resulting nitric oxide from the slurry regenerated to nitrogen dioxide for reuse About 90% of the copper In the ore IS dissolved, and the leach solution is separated for further processing. This procedure is also useful for sulfide ores of silver, molybdenum, nickel, cobalt, and zinc T C Franklewicz and R E Lueders, assigned to Kennecott Copper Corp Jan 2, 1979 7 pp (423-27) Filed 1 1-30-77 (855,983) (Drawings below ) US 4,138,231-Wet-cleanlng of gases from the roasting of copper sulfide ore, zinc sulfide ore, pyrite, or arsenopyrite The gases are washed in dilute sulfuric acid circulating in a closed system, the wash water treated to remove arsenic, the washed gas condensed by cooling; and the condensate separated, purified of residual arsenic, and neutralized B G V Hedenas, J E Wiklund, T R H Johansson, and KA Melkersson, assigned to Boliden AB Feb 6, 1979 7 pp (55-71). Filed 3-30-77 (782,908) Priority Sweden, 4-9-76 (76/04,219) (Flow diagram at right ) US 4,148,862-In the acid-leaching of silicate-bearing calcined zinc sulfide ore, finely divided calcine is leached at elevated temperature with sulfuric acid, and additional calcine is added at an optimum rate to precipitate SI~ICIC acid in the same stage In an easily settling, readily filterable form Reactions take place at a silica concentration so low that the solution is not greatly oversaturated An anionic flocculant- may be added to the slurry immediately before the filtration step S. P Fugleberg and J T. I Poijarvi, assigned to Outokumpu Oy Apr. 10, 1979 4 pp (423-106). Filed 10-6-77 (840,041) Priority Finland, 10-8-76 (76/2880)

Jan 1, 1980

By Paul L. Russell

7.5-1. introduction. The possible engineering uses of nuclear explosives were recognized with the first nuclear detonation. Subsequent experiments have demonstrated the ability of nuclear explosives to move very large volumes of rock and earth efficiently and at costs below conventional earth-moving methods. The ability to package a quantity of energy equivalent to a few million tons of conventional explosive in a container approximately 20 ft long and 36 in. in diameter enables easy emplacement and thus provides the potential for excavation on scales heretofore impossible. Space does not permit detailed discussion of the many experiments with conventional high explosives and with nuclear explosives that were performed in developing the information presented here. A bibliography is included for those wishing detailed information. Material from those sources and from personal experience has been used freely. 7.5-2. Nuclear Explosives. An explosion, in general, results from the very rapid release of a large quantity of energy within a limited space. This is equally true for both conventional and nuclear explosions, although the energy is produced in quite different ways. In the case of a conventional explosive, this energy arises from chemical reactions. In a nuclear explosion, the energy is produced as a result of the formation of different atomic nuclei by the redistribution of the protons and neutrons within the interacting nuclei. Nuclear energy is of a much higher order of magnitude than chemical energy when equal masses are considered, and the release of this energy is a thousand times more rapid than the fastest chemical detonations. Explosions may be produced by fission or fusion reactions. FISSION. When a free (or unattached) neutron enters the nucleus of a fissionable atom, it can cause the nucleus to split into two smaller parts. This fission process is accompanied by the release of a large quantity of energy. The complete fission of 1 lb of uranium can produce as much energy as an explosion of 9,000 tons of TNT.

Jan 1, 1968

By Arthur Knapp

Oil-well depletion curves, to be of value, should show when a well or lease may no longer be operated at a profit. The difference, at any time, between the total expenditures and the total income of a lease or well may be called the lease status. Plotting this lease status against time will give a curve subject to more accurate and different interpretations than the barrels-time curves. According to the hypothetical barrels-time curves shown in Fig. 1, well A, at the end of 16 mo., is producing twice as much oil as well B and will continue to produce for another year whereas well B will cease producing in about 6 months. Data fob Lease Status-time or Dollars-timb Curve Time Lease Months Status Lease purchased for $3000.............................. 0 — $ 3,000 No expenses chargeable to lease for 3 mo................. 2 — 3,000 Well A is started; cost of drilling for month is $16,000..... 3 — 19,000 Well is completed at additional cost of $10,000............ 4 - 29,000 Well A is brought in and flows 10,000 bbl. first month (we Fig. 1). Necessary to invest in tanks, boilers, pipe lines, etc.; difference between expenditures and receipts gives profit for month of $5,000............................ 5 — 24,000 Investment is small, cost of operating well A is small, so profits are $15,000................................... 6 — 9,000 Well ceases to flow so there is additional investment for pumping equipment and additional operating expense; net earnings from 6,000 bbl. of oil produced is $4,000. 7 ... Lease operation nor becomes normal and curve becomes smooth........................................... 8 — 500 9 2,000 10 4,500 11 6,000 12 7,700 13 8,700 14 9,000 15 9,500 16 10,000 A hypothetical lease status-time, or dollars-time, curve of this lease is shown in Fig. 2. At zero time, the lease is purchased for $3000, which is plotted below the zero dollar line, All subsequent entries are plotted

Jan 1, 1921