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By CHARLES HOPPER

In preparing the Steep Rock Lake iron ore body for mining, it was necessary to drain Steep Rock Lake. Using diamond drills, a cut 1800 ft long, 100 ft wide, and maximum depth of 95 ft amounting to 300,000 cu yd of material, was excavated in Greenstone to provide a drainage channel. This was done under the general contractor C. A. Pitts, of Toronto, who in turn employed a diamond drill contractor, Boyles Bros., to do the drilling. The method of excavation was selected be- cause a clean square cut and grade could be maintained in the bottom of the channel. Bootlegged holes would have left blocks of rock in the corners which would have created a situation which would have become progressively worse as the open cut advanced from one side of the ridge to the other.

Jan 1, 1949

By A. W. Schlechten, R. F. Doelling

Parkes' process crusts were vacuum distilled using a shortened Pidgeon retort. Zinc was effectively removed below 800°C and recovered as a zinc sheet easily stripped from the furnace liner. Lead dasdistillationazinc required about 950°C and the resulting lead condensate tillationtended to stick to the thin metal liner. Final purity of residue is limited only by nonvolatile metals such as copper. THE successful commercial application of vacuum distillation for the removal of zinc from lead, as described by W. T. Isbell,' raises the question of whether or not a similar method can be used for other metal separations. Preliminary laboratory experiments on the vacuum distillation of Parkes' crusts by one of the authors' gave such encouraging results that it was decided to make a series of tests with large scale equipment to determine the results that might be expected in plant operation. The pilot-plant test results are reported in this paper. The great advantages of vacuum dezincing are its simplicity, the relatively low temperature of operation, and the fact that the zinc is recovered in the metallic state ready for use again in the desilveriza-tion of lead bullion. These same advantages and others would be realized from the vacuum distillation of the zinc-lead-silver crusts made in the Parkes' process. Zinc readily will distill over at temperatures of 800" to 900 °C as contrasted with the 1200°C or more used in the Faber du Faur furnace. Furthermore, the recovery of zinc as metal is practically perfect since there is no loss by oxidation or fuming. If the vacuum distillation is continued and hastened by raising the temperature to around 1000°C, the lead will distill over at an appreciable rate and deposit in a warmer zone of the condenser. This lead will contain a small amount of silver, but it can be returned to the desilverizing kettles and there is not the chance of loss that there is in disposing of the rich litharge from cupelling. The lower temperatures required will mean a saving of fuel and less wear on the equipment. One further point is that the vacuum distillation could be run on a small scale practically continuously, so as to avoid the accumulation of large amounts of silver-rich material and the need of assembling a large crew for special cupellation runs. It should be pointed out here that the final purity of the silver residue will depend on the copper or iron content of the crusts, as these and other metals with low vapor pressure will not be eliminated. The idea of treating the silver-zinc-lead crusts from the desilverization of lead bullion by vacuum distillation is not a new one, but such experiments have not been reported, as far as can be determined. In 1912, T. Turner presented a paper3 before the Institute of Metals on the vacuum distillation of brass, and a member of the audience suggested that he try his method on Parkes' crusts. In 1935, W. J. Kroll, noted for his discovery of debismuthiz-ing lead by calcium additions, published a paper' on vacuum distillation of metals and pointed out possible applications in the lead industry. Principles of Separation The successful separation of two or more metals by distillation depends upon an appreciable difference in the effective vapor pressures of the metals in the alloy to be treated. The vapor pressures of the common metals have been determined with considerable accuracy; for example, at 800°C the vapor pressures of the individual metals which are of particular interest in this paper are about .as follows: Zn, 241 mm; Pb, 0.0553 mm; and Ag, 0.0000276 mm. Unfortunately,, data are not available on the activities of the components in the ternary system Zn-Pb-Ag, but it can be estimated that on the basis of the vapor pressures of the individual metals a good separation can be made by distillation. The preliminary laboratory tests previously mentioned' supported this assumption. The pilot-plant size vacuum distilling furnace is shown in Fig. 1. It consisted of an abbreviated Pidgeon retort, 5 ft long with an ID of 8 in., heated by six silicon carbide elements placed symmetrically

Jan 1, 1952

The Institute is at present in sound condition professionally and financially. During the past year 546 new members were elected, the total membership now numbering 5,221. The Treasurer's Report shows that the expenses of the Institute are met by its income. Your Directors have thought it wise to issue a referendum to the members asking for a vote upon the suggestion that the annual dues be raised from $10 to $12. This moderate increase is proposed not because the Institute needs it at present but to insure a continuation and increase in its activities and usefulness, and to provide a reserve fund for future emergencies. In well-conducted organizations, such as clubs, it is a wise and safe course to so adjust the clues of members that the total receipts from this source will about equal the annual expenditures, the initiation fees being used for reserve purposes. Our dues alone do not at present meet the expenses. This, I think, is the whole situation, so far as this question is concerned. Expenses might be reduced in some measure, but they cannot be materially reduced without affecting the integrity and usefulness of your Institute. Nothing of consequence in our internal affairs has occurred during the past year except, perhaps, the establishment of the Arizona Section and the prospect now assured of a Nevada Section. During recent months there has been a marked change in the relations of the Institute to its fellow societies and to the Government of the United States. That change is one of cooperation. The Secretary of the Navy has appointed two of your members to the Naval Consulting Board to cooperate with members of 10 other scientific organizations. This Board is now organized, and through its work your Institute is brought in close touch with the activities of the other societies and with the United States Navy. The President of the United States has requested us to recommend to the Government one of our representatives in each State in the Union to act in collaboration with a member in each State from the. American Society of Civil Engineers, the American Society of Mechanical Engineers, the American Institute of Electrical Engineers and the American Chemical Society, thus organizing a directorate of engineers for each State, that will conduct, through the members of each of the five societies referred to, living in each State, a campaign of industrial preparedness. No more important step than this, it seems to me,

Jan 3, 1916

By Louis D. Huntoon

(New Haven Meeting, February, 1909.) THE Hammond Mining and Metallurgical Laboratory is the gift of Prof. John Hays Hammond to the Sheffield Scientific School of Yale University. Professor Hammond was graduated from this school in the class of 1876, and has always shown great interest in its welfare and progress. In 1903, he offered to build and equip for it a mining and metallurgical laboratory. This offer was gratefully accepted by the trustees; and in October of the same year the architect, W. Gedney Beatty, and the writer were employed to draw up plans for the building, which is the largest of its kind in this country. The Treasurer's report shows the total gift of Professor Hammond, including land, building, and equipment, to have been $128,741.53. This does not include $5,000 given by Professor Hammond to defray incidental expenses. The detailed cost of the machinery and installation, included in the above sum, is as follows: Machinery, $9,846.00 Pipe and fittings, 925.05 Hangers, shafting, pulleys, belts, etc., . 357.83 Steel-work, tanks, launders, etc., . 272.19 Screens for sizing ore, 411.37 Lumber, 1,896.36 General Supplies Hardware, . $135.96 Tools, 263.66 Paints, . 04.63 Electric and incidental, . 439.57 - 1,043.82 Machine-shop material, 259.44 Labor, 722.29 - 981.73

Mar 1, 1909

JOSEPH W. RICHARDS, Chairman. J. E. JOHNSON, JR., Vice-Chairman. ARTHUR S. CALLEN, Secretary, 453 Chestnut St., So. Bethlehem, Pa. SUB-COMMITTEES IRON ORE DWIGHT E. WOODBRIDGE, Chairman. ERNEST F. BURCHARD, WILLIAM KELLY, CHARLES F. RAND, FELIX A. VOGEL. BLAST FURNACES II. A. BRASSERT, Chairman. WILLIAM H. BLAUVELT, AMBROSE N. DIEHL, JAMES GAYLEY RICHARD V. McKAY. STEEL WORKS HENRY D. HIBBARD, Chairman. GUILLIAEM AERTSEN, C. F. W. RYS, .1. S. UNGER. FRANK D. CARNEY, A. A. STEVENSON, WILLIAM It. WALKER. MECHANICAL TREATMENT CHARLES A. BUCK, Chairman. ROBERT W. HUNT, GEORGE MESTA, FREDERICK W. WOOD.

Jan 6, 1915

The following is an incomplete list of members and guests who called at Institute headquarters `during the period Aug. 10, 1918 to Sept. 10, 1918. Amil A. Anderson, Hill City, S. D. A. A. Holland. H. G. S: Anderson, Hurley, N. Al. Paul S. King, Wilmington, Del. Louis A. Barton, Capt., Engineers, U. S. A. Herbert G. Officer, Washington, D. C. K. Baumgarten, lst Lieut., U. S. R. A. B. Parsons. Charles W. Boise, Belgian Congo. Robert R. Pollock, San Francisco, Cal. Henry C. Carlisle, The Pas, Manitoba. Frank A. Smith, Edgewood, Md. J. Ross Corbin, Boyertown, Pa. Edward C. Thompson, Kennett, Cal. S. Le Fevre, Forest Glen, N. V. Malcolm M. Thompson, Morsemere, N. J. Harold P. Ford, Cleveland, 0. Lester C. Uren, Berkeley, Cal. William L. Hogg, Waynesboro, Va. Harold W. Aldrich has been appointed assistant general superintendent of the Inspiration Consolidated Copper Co. at Inspiration, Ariz. Capt. Percy E. Barbour, formerly with the New York State Troopers, has been granted a leave of absence from the Department for the period of the war, by the Governor, and has been appointed Captain in the Engineer Corps, U. S. A. He served on the Mexican border in the New York State National Guard. A. C. Barke, formerly with the United States Smelting, Refining and Mining Exploration Co., has accepted a position with the United States Fuel Co. at' Salt Lake City, Utah.

Jan 10, 1918

Jan 1, 1923

By F. O., Garrabrant

ELECTRIC power requirements for Climax are similar to those of most metal mines, except that large blocks of power are used underground and there are a number of other unusual applications. Power is furnished to the Company by the Public Service Co. of Colorado over two 100,000-volt lines to a bank of three 3333-kva. transformers 100/13.8 kv. These transformers are designed so that with the addition of forced cooling their capacity may be increased a third at 12,000-ft. altitude. On the secondary side of this bank is a 15-kv., 1200-amp. oil circuit breaker provided with automatic reclosing relays allowing three successive short circuits before lockout. From this breaker 13.8-kv. power is transmitted to a central switching station where all circuits radiate to the different operations of the property. Each circuit is protected by a 23-kv., 600-amp. oil circuit breaker with by-pass and isolating disconnecting switches. (A more complete description of the electrical distribution system at Climax, with detailed drawings, is contained in T.P. 1734, A.I.M.E., Mining Technology, July, 1944.)

Jan 1, 1946

By S. S. Clarke

THEN sufficient ore has been blocked out on a lease to warrant a shaft, the proposed. Location of the shaft should be carefully considered, as the old promoter's method of sinking wherever his hat landed is a sad mistake as some: of us have learned by trying to operate such mines. In some instances certain conditions exist, such as mill location, railroad right of way, etc., that necessitate a change of location from the. ideal, but under average conditions and with the knowledge of the formation gained from the drill logs and with the general trend of the orebody known, the shaft can be located from. the standpoint of its future relation to the mine with re-gard to hauling and' hoisting a maximum tonnage: at a minimum cost per ton, also taking into account the minimum amount of expense and trouble- in sinking. Quite a few shafts are located in the wall of a run or just in the thin edge of the orebody: As a general rule the formation here is solid and permits rapid and safe sinking, and the shaft pillars are left in practically valueless material. This permits the withdrawal of pillars and still keeps the shaft protected. By opening a shaft in thin ground, you are not tempted to cut it high, wide, and handsome, but to shape it up-with ample clearance for lay-bys and with more of an idea of permanency, as a few hundred extra dollars spent around the shafts for proper track installation will re-turn, a good rate of interest in ease of handling "cans," quick get-away; and fewer wrecks. Nearly all the shafts in this district are sunk on a contract basis, the con-tractors furnishing all labor and explosives, and the company building the sinking-derrick, dump-track, and furnishing, all equipment, material, and power. The following costs are the averages, for three shafts sunk through good solid formation, with no water to, contend with, and approximately 260 -ft., deep, having about-100 ft. of shale. The contract price was $9 per foot with the shale and $13 per foot in the solid, making a total of $2,980 paid to the contractor. The wages paid by the contractor were, $5 per shift for the hoist-man, who also sharpened steel, $6 for the lead shaft man, and $5 for the helper;, the cribbing was sawed on contract for 40c. per shaft: foot; about. 94 board feet. of cribbing was used per foot of staggered cribbing and 168 board feet per foot of solid cribbing.

Jan 8, 1928

By J. F. Shirley, H. K. Burke

In January, 1964, three years of laboratory and pilot plant test work reached its culmination. The San Manuel Div. of the Magma Copper Co. switched its molybdenite extraction circuit from the sodium hypochlorite-ferrocyanide process to the new process described in this paper. The San Manuel Div. Concentrator of the Magma Copper Co. is located near San Manuel, Ariz., approximately 45 miles northeast of Tucson. The concentrator treats approximately 35,000 tons of ore per day containing about 0.85% copper and 0.025% molybdenite. The San Manuel ore body is a disseminated copper deposit in quartz monzonite and quartz monzonite porphyry. The primary copper mineral is chalcopyrite; minor amounts of chalcocite, bornite and covellite are also present in varying degrees. Pyrite makes up somewhat more than half the sulfide content of the rock. The molybdenite is nearly always found on fracture faces or associated with quartz veinlets.' On January 9, 1964, the sodium hypochlorite-ferrocyanide process that had been used for molybdenite recovery was replaced by the following process: 1) A conditioning step with hydrogen peroxide, sulfuric acid, sodium cyanide and zinc sulfate. 2) A two stage rougher flotation step with stove oil and sodium ferrocyanide. 3) One cleaning step with sodium ferrocyanide. 4) One cleaning step with sodium hypochlorite and potassium ferricyanide. 5) Four cleaning steps with an anti-foam Exfoam 636 and potassium ferricyanide. COPPER CONCENTRATOR Since molybdenite recovery is closely allied with previous steps in the operation, the copper concentrator will be briefly described first. The ore is ground to approximately 5% plus 65-mesh and 65% minus 200-mesh in a two-step grinding circuit consisting of rod mills in open circuit and ball mills in closed circuit with rake classifiers. A straight forward rougher circuit producing only one concentrate is used. The rougher concentrate is combined with the discharge from the regrind ball mills and is then cycloned, with the overflow going directly to the cleaner circuit and the underflow going to the regrind mills. The feed to the cleaner circuit is approximately 90% to 94% minus 325 mesh. The cleaner circuit consists of two stages in which the tailing from the final cleaner stage is returned to the primary cleaner and the primary cleaner tailing is returned to the rod mill discharge at the head of the circuit. Approximately 0.013 lb of the allyl ester of amyl xanthate (S-3302) per ton of ore is added to the rod mills and approximately 0.006 lb of sodium isopropyl xanthate (2-11) per ton of ore is stage added down the rougher banks. The frother used is methyl iso-butyl carbinol (MIBC) and it is added at various points in the grinding and the rougher flotation circuits. Lime is added to the ball mills and/or to the rod mills to maintain the desired pH in the rougher flotation. The rougher flotation has been operated at various pH levels between 10.5 and 11.5. Fuel oil is added in the regrind mills at a loading of approximately 0.01 lb per ton of ore. Factors Which Directly Affect Molybdenite Recovery: It has been found at San Manuel that, to obtain maximum molybdenite recovery, an oily type flotation collector must be added in the grinding circuit. When molybdenite is crystallized in the characteristic platy form, which appears to be the case almost all of the time at San Manuel, its surface has a greater tendency to adhere to oils than water. 2 Of the several oily type collectors available on the market, the allyl ester of amyl xanthate seems to be the best molybdenite promoter. The floatability of San Manuel's molybdenite does not appear to be affected, at least up to a pH of 11.5, by lime. It has been reported that at several plant operations high lime alkaline circuits have adversely affected the molybdenite recovery.3,4 It has been found through microscopic studies that the bulk of the chalcopyrite will float before the molybdenite in both the rougher and cleaner circuits. It is necessary to add fuel oil to the regrind mills to enhance the

Jan 1, 1965

Jan 1, 1924

Lead-zinc production in Canada accounted for 19% of the total value of metals and minerals produced in 1976, says Keith C. Hendrick, president of Noranda Sides Corp. Mine production of recoverable zinc w as 1. 158 million toils of 27.4% of free world production in 1976. Slab zinc pro¬duction was placed at 472,000 tons that year. Canadian mine production of recoverable lead was 247,000 tons or 10.6% of the free world production. Refined lead that year was at 177,000 tons.

Jan 11, 1977

Dr. F. W. C. Schniewind, New York, N. Y.:—Mr. Vogel speaks of briquetting the flue-dust by means of lime. I learned recently in Europe of a process employed with considerable success at one of the blast-furnaces, where blast-furnace cinder is used with lime. The lime and cinder are mixed with the flue-dust, and exposed to superheated steam, which converts the blast-furnace cinder into cement. Briquettes are then formed in the usual way. Mr. Vogel :—The process referred to is the " Scoria process," which has not been in use for some time past. For that reason, I did not mention it. Prof. Arthur L. Walker, New York, N. Y.:—I would ask the commercial name for magnesium chloride, and where it can be obtained ? Mr. VoGel :—I mentioned magnesium chloride, because it has been used abroad to a large extent. In Germany it is obtained as a by-product in the manufacture of potash-salts; in the United States we have been using chloride of lime, which is cheaper. Otto Sussman, New York, N. Y.:—What is the cost of this process ? Mr. Vogel :—The cost is exceedingly low. I can give you the figures from abroad, but not for the United States, because the system has not -yet been used here commercially to

Jan 1, 1913

(Members are urged to send in for this column any notes of interest concerning themselves or their fellow-members.) Members who registered at Institute headquarters during July and August : R. H. Sweetser, Columbus, Ohio. C. R. Hill, New York, N. Y. Raymond Brooks, Goldfield, Nev. H. A. Morrison, Candor, N. C. Edward M. Shepard, Springfield, Mo. Frederick J. Patchell, Washington, D. C. Rudolf Gahl, Clifton, Ariz. Walter F. Pyne, New York, N. Y. Edmund von Maltitz, Pittsburg, Pa. Harvey B. Small, Mangangue, Colombia, S. A. M. K. Rodgers, Seattle, Wash. M. B. Cohen, Miami, Ariz. J. F. Erdlets, Jr., New York, N. Y. A. P. Watt, Newark, N. J. D. H. Ladd, Houghton, Mich. H. H. Stoek, Urbana, Ill. Roy F. Cornell, Santa Cruz, Cal. D. W. Brunton, Denver, Colo. W. B. Foote, Geneva, N. Y. C. S. Robinson, Youngstown, O. H. Sjogren, Stockholm, Sweden.

Jan 9, 1913

By S. J. CRAIGHEAD

IN 1912 the United States Smelting Refining and Mining Company made a large investment in a number of coal properties in Utah and in 1915 a subsidiary, the United States Fuel Co., was organized to take over these properties. Operations started in 1916 and the Company soon became, and remains, one of the major coal producers of the State. It now produces approximately one million tons of coal annually, which is about 22 per cent of Utah's output. The U. S. Fuel Co. owns large reserves of good coal and is in a position to expand operations with the industrial growth now taking place in the West. Traveling through the Western States, one is struck by numerous advertising signs portraying the picture of jovial Old King Cole of nursery rhyme fame. His rotund figure and beaming countenance are the symbols of Utah King Coal, the product of the King mine of U. S. Fuel Co. King Coal is a by-word of quality coal throughout all the Western States.

Jan 1, 1948

By C. F. Weinaug, P. M. Blair

Many difference equations used to approximate reservoir flow problems treat the phase pressures implicitly but not the mobility-density coefficients. Such difference equations are neither wholly explicit nor implicit, but might be described as mixed. Mixed equations are relatively easy to apply. But the associated time truncation error is relatively large, and when used to solve problems characterized by high flow rates, these equations may be unstable for practical size time steps. This paper outlines the development of a completely implicit difference analogue for reservoir simulation, along with a Newtonian iterative method for solving the resulting nonlinear set of algebraic equations that arise at each time step. While this implicit equation requires two to three times more work than does the mixed equation, it is shown to markedly decrease the time truncation error and to yield a stable solution for much larger time steps than does the mixed equation. INTRODUCTION Calculation of multiphase, multidimensional flow in porous media is generally accomplished by numerical methods that involve approximating systems of partial differential equations by systems of partial difference equations. The early development of difference equation techniques was directed toward the solution of linear differential systems. However, the equations of multiphase fluid flow through porous media are highly nonlinear in that mobility and density often are strong functions of pressure and saturation. Thus, solution of the flow problems by difference equations involves solving sets of algebraic equations whose coefficients change from step to step of the calculations. Current practice is to evaluate most coefficients at the beginning of a time step and then to apply difference methods well suited for solving linear problems. At least two major difficulties arise from the use of this practice. First, evaluation of coefficients at the old time level and evaluation of pressures at the new time level results in larger time truncation errors near displacement fronts than if all quantities in the distance derivative were evaluated at the same time level. Second, evaluation of mobility coefficients at old time levels results in an unstable difference equation in regions of high flow rate. The first of these difficulties often results in optimistic recovery behind the displacement front, and the second requires the use of small time steps or special techniques in solving coning problems1 or problems of gas percolation.2 Evaluation of all quantities in the distance differences at the new time level results in a completely implicit difference system. Such a difference system has a lower time truncation error than equations in which mobilities are evaluated at the old level and pressures evaluated at the new level. The algebraic equations resulting from the fully implicit difference equation are, in this case, nonlinear and require some iterative method for their solution. This paper develops fully implicit difference equations for two-phase flow in porous media, describes their solution by Newtonian iteration, and gives examples of problems more easily solved by the new method than by previous methods. EVOLUTION OF DIFFERENCE EQUATIONS FOR NONLINEAR PROCESSES The completely implicit difference equation used in these calculations results from evolution beginning with explicit difference equations, proceeding through what we will call "mixed" equations, and arriving now at a completely implicit equation. Let us illustrate the evolution by considering different analogs of the nonlinear equation

Jan 1, 1970

By Walter Bradley

MINING by hydraulic process of the important gold-bearing gravels of the Sacramento Valley in the basins of the Yuba, Bear and American rivers began in 1853, and continued at an ever-increasing rate for 30 years. In its most active days it reached an annual production of $15,000,000, and averaged $10,000,000 per year over the 30 years. During that period no provisions were made to prevent the debris from being dis-charged into the water courses, and ultimately large quantities of silt, sand and gravel were washed down to the lower valleys; filling channels, causing overflow of agricultural lands and difficulty in maintaining navigable depths in the Sacramento River. Agitation and suits filed because of these damages finally resulted in the injunction decision of Judge L. B. Sawyer,1 in the U.S. Circuit Court, Jan. 23, 1884, which closed down practically all hydraulic mining operations not having means to restrain their debris. In 1893 (approved March 1, 1893), Congress passed the act introduced by Representative Anthony Caminetti, of California, creating the California Debris Com-mission, composed of three officers assigned from the Engineer Corps of the U.S. Army, and since that time it has been legal and possible within the drainage systems of Sacramento and San Joaquin rivers to conduct hydraulic mining of gold-bearing gravels behind dams built under this commission's permit and supervision. In the watersheds of the Trinity and Klamath rivers, which flow directly to the Pacific Ocean and do not affect navigation or agriculture, operating permits from the California Debris Commission are not required. There is a measure of restriction there, however,. through what is known as the Trinity and Klamath River Fish and Game Dis-trict. Chap. 215, Stat. 1931, in order to secure relatively clear water for the fishing season, provides that between July 15 and October 15 each year there must not be discharged into these rivers debris that will cause a turbidity greater than a content of 50 parts per million in the

Jan 1, 1936

By W. J. O'CONNOR

THE production of lead in the United States for the period from 1720 to 1912 was 10,432,668 tons valued at $924,600,000. The average price during this period was 4.4c. a pound, although lead sold at times below 3 cents. We know that in the early days of lead smelting in Montana lead was deliberately lost in the slag and only enough reduced as bullion to collect the silver. This bullion was freighted by ox team to Corinne, Utah, the terminus of the railroad at that time, and shipped to Swansea or Omaha, Neb., for refining. The slag from some of the above mentioned operations contained as much as 15 per cent lead and has since been resmelted. Since 1921 the production of lead has increased each year and the price and tonnage have been as follows :

Jan 1, 1926

By Reaves. M. J.

The title of my talk, "What Happened to the Uranium Boom?" is old news. Certainly it is for this group. All of us that make our living in uranium know that the boom of the last half of the 1970's is over. U.S. production has been exceeding consumption by more than two to one. Mines and mills are closing and yellowcake prices have been dropping for over 20 months. The gloomy outlook for the industry in the near term has been well documented by soothsayers of various descriptions, your daily newspapers, and in the Nuexco Monthly Reports. I'd like to attempt to describe the next upturn in the market (speculate, really) based upon the clues we're seeing now. In order to do that, I'd first like to go over briefly, some of the market factors that contributed to the recent price drop and resultant production cutbacks, and then hypothesize on the way these factors are changing and will change. Market prices are greatly affected (maybe even entirely determined) by buyer perceptions. This is particularly true with uranium, because of the long lead times associated with nuclear plant construction and also with conventional mine/mill development. Before the price rise (say, 1975) utility uranium buyers believed that: 1) U.S. producers would have difficulty expanding to meet U.S. demand. 2) Australian and Canadian production was essential to avoid shortages in the early 1980's. 3) Uranlum prices would continue to rise as demand exceeded supply. 4) Enrichment capacity would become inadequate. It was thought necessary, therefore, to build enriched inventory in the early 1980's for use in the late 1980's. Artificially accelerated expansion of the uranium producer industry was necessary to accommodate anticipated enrichment demand. Current perceptions are largely the opposite. These are the beliefs that were held most of this year and late last year as prices dropped. 1) U.S. production is far in excess of domestic need. Contraction of the U.S. production lndustry is necessary. 2) Canadian and Australian supply is optional and not essential. Producers in those countries are expanding mainly by displacing higher cost production and not because they fill a void, 3) Prices may be essentially stable for some time. 4) Enriched uranium is in excess supply. That is 1981. 1982 is shaping up to look like this: 1) Prices will have bottomed out. (That is not Nuexco's opinion necessarily, by the way, but it is my opinion.) 2) There will still be substantial utility inventories, but fewer spot sales. 3) Canadian and perhaps Australian sellers will have made substantial sales in the U.S. and will be aggressively seeking more. 4) U.S. production will have been dramatically curtailed. U.S. utilities that wish to con- tract long term will have difficulty in finding domestic sellers. Concern will develop about the availability of U.S. production capability. Virtually all long term con- tracts signed will be with non-U.S. sellers. 5) An awareness will begin to develop among U.S. buyers that we are approaching a period of dependence upon foreign uranium (which will be true). The history of the uranium market has been one of dramatic changes and overreaction to those changes. The rapid price rise of a few years ago generated excess U.S. production capacity and the rapid price drop of the last two years will almost certainly result in too little capacity. It will soon be difficult for U.S. buyers to buy domestic material except on the spot market. The question is, "will they care?" The lack of demand, of course, is the underlying reason for the current poor health of the uranium industry. In 1972, 1973 and 1974 collectively, there were 105 nuclear reactors ordered in the U.S. That ordering rate was expected to continue and accelerate throughout this century. In 1975, 1976, 1977, 1978, 1979, and 1980 altogether, there were 56 more reactors cancelled than ordered. The net growth of our only customer since 1974 has been a negative 56. TO put this in perspective, if these 56 reactors were operating now it would more than double present U.S. uranium consumption. Underlying lack of demand is something that is simply not going to change in this decade. Time is going to be required. The NRC indicates that the maximum feasible number of new reactors that can be licensed each year is six. That would increase uranium consumption by only 10% per year. New reactors, if ordered tomorrow, would not generate new uranium demand until after 1990. Even so, United States' consumption of uranium will rise from the 1980 level of 18 million pounds per year, to

Jan 1, 1982

By Thomas M. Drown

Remarkable as have been the direct results of Danks's puddler, there are some indirect and incidental results, which are well worthy of study for their intrinsic value and suggestiveness. The success of Danks's machine is due mainly to the nature of the lining and the manner in which it is attached to the walls of the revolving chamber. Herein lies Danks's merit and good fortune. Given such a lining, and the success of the machine, as far as puddling iron is concerned, could easily be predicted. Other important results have, however, been obtained with this machine, which, though clearly explicable, were, nevertheless, unanticipated. These are the increased yield of bar-iron over the pig-iron charged, and the elimination of phosphorus, which, though not absolutely complete, is yet more decided than in the case of hand-puddling. As regards the increased yield, this is merely what we have a right to expect if we consider the puddling process to consist in the oxida tion of the carbon and silicon of the pig by the oxygen of the oxide of iron. That this result is attained, nearly to the extent that the theory requires, in Danks's puddler, and has never been more than partially attained by the most careful experiments in band-puddling, points clearly to the fact that the contact of the molten pig-iron with the oxide of iron is much more intimate and complete in the one instance than in the other. The action of puddling in Danks's machine may be considered to be twofold: first, the removal of the carbon, silicon, and phospho rus of the pig, which is the primary object, and, second, the produc tion of wrought iron direct from the ore, which is entirely an inci dental result. Intelligent metallurgists were not long in recognizing in this latter result a fact of deeper import and greater value than the attainment of a perfect puddling process. A new method