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By H. A. Buehler

The Ozark region occupies a large part of the southern half of Missouri, the northern portion of Arkansas and comparatively smalll areas in northeast Oklahoma, southwest Kansas, and southern Illinois. It is a roughly elliptical plateau which has been deeply dissected by stream erosion, and which slopes imperceptibly into the prairie regions to the north, east, and west, but which terminates rather abruptly along the Arkansas River on the south and Tertiary lowland to the southeast. The Boston Mountains extending for a distance of some 200 miles (321 km.) along the Arkansas River rise to an elevation of 2000 ft. (609 m.), forming the most prominent topographic feature of the region. The so-called St. Frangois Mountains in St. Frangois, Madison, and adjoining counties in southeast Missouri, form a prominent but more restricted topographic feature, the highest point rising to an elevation of 1800 ft. (548 m.). The remainder of the region is largely a plateau area varying from 800 to 1500 ft. (243 to 457 m.) in elevation. Stratigraphy The region is underlain chiefly by sedimentary formations, which outcrop in roughly concentric bands around a comparatively restricted area of igneous rocks in the St. Frangois Mountain district. Although located geographically near the eastern edge of the Osarks, the St. Frangois Mountains have apparently suffered the greatest orogenic movement and the lowest formations of the succession are exposed in their immediate vicinity. The various sedimentary formations of the region do not show any considerable metamorphism and have suffered little change since consolidation. With the exception of a few basic intrusions in Ste. Gene-vievg County and a pegmatite dike in Camden County, Missouri, there is no evidence of igneous action, since the first sedimentary beds were deposited.

Jan 1, 1918

By H. A. Buehler

LOCATION THE Ozark region occupies a large part of the southern half, of Missouri, the northern portion of Arkansas and comparatively small areas in northeast Oklahoma, southwest Kansas, and southern Illinois. . It is a roughly elliptical plateau which has been deeply dissected by stream erosion, and which slopes imperceptibly into the prairie regions to the north, east, and west, but which terminates rather abruptly along the Arkansas River on the south and Tertiary lowland to the southeast. The-Boston Mountains extending for a distance of some 200 miles (321 -km.) along the Arkansas River rise to an elevation of 2000 ft. (609 m.), forming the most prominent topographic feature of the region. The so-called St. Francois Mountains in St. Francois, Madison, and adjoining counties in southeast Missouri, form a prominent but more restricted topographic feature, the highest point rising to an elevation of 1800 ft. (548 m.).. The remainder of the region is largely a plateau area varying from 800 to 1500 ft. (243 to 457 m.) in elevation. STRATIGRAPHY The region is underlain chiefly by sedimentary formations, which outcrop in roughly concentric bands around a comparatively restricted area of igneous rocks in the St. Francois Mountain district. Although located geographically near the eastern edge of the Ozarks, the St. Francois Mountains have apparently suffered the greatest orogenic movement and the lowest formations of the succession are exposed in their immediate vicinity. The various sedimentary formations of the region do not show any considerable metamorphism and have suffered little change since consolidation. With the exception of a few basic intrusions in Ste. Genevieve County and a pegmatite dike in Camden County, Missouri, there is no evidence of igneous-action, since the first sedimentary beds were deposited.

Jan 10, 1917

By T. F. Witherbee

That the single charging-bell, properly proportioned, is a good all-around device for distributing material in a blast-furnace can scarcely be questioned; yet it is equally true that, in some cases, it has not given perfect satisfaction; as has been demonstrated by the substitution of double bells, at furnaces under the beat management—generally in cases where anthracite, either alone, or with a portion of coke, was used as fuel. The defect of the single bell seems to be that it does not permit, beyond very narrow limits, such changes in charging as may be required by the temporary conditioil of the furnace. Modern furnaces using Connellsville or other good coke and Lake Superior ores seem to be well served with a single bell. At least, up to a certain size, somewhere about 18 by 85 ft. to 19 by 90 feet, these furnaces make low-silicon pig-iron with about 1,800 lb. of fuel, or in some cases very much less. The following instances occur to me, in which other forms of charging-apparatus have been employed with advantage: 1. A charcoal-furnace, 61 ft. high, 11 ft. in diameter at bosh, and 8 ft. 2 in. at stock-line, with a rectangular crucible, 42 by

Jan 1, 1905

By J. A. Herbst, J. L. Sepulveda, J. D. Miller

Effective bismuth removal from flue dust residues has been accomplished for flue dusts from both zinc and copper smelters. Selective dissolution of bismuth, with respect to silver and lead, can be achieved by the proper manipulation of temperature, sulfuric acid addition and sodium chloride addition. Optimum levels of these variables were determined using a response surface technique to maximize the objective function: Y = %Bi/(1 + %Ag). In the case of the flue dust residue from a zinc smelter, which contains bismuth as an impurity (0.05%), 86% bismuth removal can be realized with less than 1% combined silver and lead dissolution. Similarly, in the case of flue dust from a copper smelter, which contains bismuth as a primary constituent (3.3%), 93% bismuth removal was obtained with little dissolution of silver and lead values.

Jan 1, 1977

By R. L. Perrine

The rise of drain holes drilled from a well to improve productivity is becoming increasingly common. This paper presents data on the productivity increase that may he expected. The data were obtained by use of an electrolytic tank mode1 of a reservoir. Parameter varied in the study to show their effect on productivity include the length, number, and arrangement of drain holes, formation thickness, and damage near the well-hore. The most important results are that very large increases in productivity can be expected only ill damaged wells. The distance, to which drain holes penetrate is of primary importance. Additional length is more effective than increasing the number of drain holes. INTRODUCTI O N In planning the use of drain holes to improve productivity, it is important to he able to estimate the increase in production that can be obtained for a given expenditure. It is also important to know what combination of parameters related to the use of drain holes can he chosen to give the largest production increase. This paper presents data on the productivity increase that may be expected, as obtained by use of an electrolytic tank model of the reservoir. In this study, :! number of parameters were varied to show their effect on productivity. These included the length, number, and arrangement of the drain holes. The presence or absence of a damaged zone of low permeability around the well-bore and the thickness of the producing formation were also considered. The productivity increase that might be expected due to drain holes is presented as the ratio: productivity of a well after drain holes are drilled and completed to the productivity of the same well in an undamaged state without drain holes. The symbol Q/Qo is used to represent this productivity ratio. Thus, the comparison in each case is with a formation of uniform and undamaged permeability. Additional notation includes I;, for the radial extent of the damaged zone, which is assumed to be cylindrical and concentric with the wellbore. Thus the effect of an actual block is approximated by a discontinuous varia-tion in permeability. For purposes of this paper, the permeability within the damaged zone, denoted by kb, is assumed to be 1/25 that in the remainder of the formation, or k. A necessary assumption in making comparative tests using an electrolytic model is that reservoir flow behavior can be approximated by single-phase, steady-\tate flow according to Darey's Law. In addition, the formation must be assumed to be homogeneous with equal horizontal and vertical permeability. The well-hore diameter in this study is assumed to be 11¾ in. and the drainage radius of the well equal to 500 ft. The drain holes are considered to have been completed without damage. THE ELECTROLYTlC MODEL 'She producing formation was modeled by forming a sheet of phosphor-bronze into a cylinder 40% in, in diameter to represent the boundary of the drainage area. This was enclosed in a waterproofed plywood box, and the box filled to an appropriate level with a dilute sodium chloride solution. Solution concentration was adjusted to give a total resistance from

Jan 1, 1956

By George T. Bator

If confronted with the problem of pumping a mixture of slime-free sand, gravel and boulders up to six in. in diameter, at the rate of 175 tons per hr in one single-stage pump against a static head of 90 ft and a total head of 140 ft, of which 20 is on the suction side, an engineer might be forgiven for backing away from the job, especially if he were also told that the pump wearing parts must have a minimum hardness of 550 Brinell for the sake of abrasion resistance.

Jan 1, 1950

By D. J. Brown

Froth flotation is now used to recover millions of tons of fine coal,(-1/2 mm), every year. For example, in Great Britain, 4-million tons of flotation concentrate was produced in 1960, about 3% of the mechanically cleaned out- put and 2% of the total salable output. Coal flotation is not so widely practiced in the U. S. as it is in Europe since much American coal occurs in thick seams of high grade and does not require cleaning. However, in the U. S., as elsewhere, the increasing mechanization of coal mining machinery has lead to a lowering in size and grade of product from the mine and to an increasing need for cleaning of coal to improve the grade before sale. With an increasing proportion of fines in the product, froth flotation has assumed a greater importance as a cleaning process for coal. In the last three years, over 40 flotation plants have been put into operation in the Appalachian area of the U. S.80 Fine coal recovered by means of froth flotation may be used as a powdered fuel alone or mixed with coarser grades. The removal of this fine coal from washery water makes possible the recirculation and re-use of the water and also leads to improved and more consistent operation of jigs and other separating machines which function less effectively when there is a build-up of fine solids in the washery water. Finally, froth flotation is of assistance in the production of a clarified effluent which may be discharged from the washery without polluting local rivers, streams, or wells. The 1961 cost of installing a new coal flotation plant in Great Britain would be about $19,000 per ton of dry solids input per hour, including the cost of new buildings and coal and shale filtration and water clarification equipment. The figure for the gravity concentration of the large coal, the main plant, would be about $4,800 per tph. The operating cost for the new coal flotation plant would be about $1.10 per ton. Even at this level of cost the cleaning operation is economic in Great Britain for coking coals. In the U. S., a froth flotation cost of 14$per ton of cleanfilter cake was quoted early in 1958 for one operation,l though admittedly this was a special case. Assuming a cake moisture con- tent of 30%, it is estimated that this cost is equivalent to a figure of 256 per ton of dry solids input, for the incorporation of coal flotation in an existing washery. A cost oh 416 per ton of concentrate was quoted76 in a recent analysis of flotation economics, this cost being brought up to $1.11 per ton after payment of royalty, welfare and association dues and taxes. Capital cost was $120,000 for a plant producing 133 tpd of concentrate. Before considering the treatment of coal by flotation it is appropriate to survey some of the properties of coal which are of importance in the process.

Jan 1, 1962

By I. W. Alcorn

THE cost of drilling in the past few years of proration and ever decreasing allowables has received increasing thought and study. It seems to parallel the strides made with respect to pumping problems in 1932 and 1933. Low profit margins stimulate the search for economies. Any discussion of slim-hole drilling involves a multitude of abstract values, and the problem can be attacked only by a study of records and rational thinking. In a measure, traditional methods must be studied with a questioning attitude. The words "slim hole" may be somewhat misleading, as the term is entirely relative. For that reason I wish to define it as used in this discussion. On the Gulf Coast, I would define a normal hole as one in which 10 3/4-in. surface casing has been set and below which a 9 7/8-in. hole is dug. With this as a datum, holes of larger diameter below surface would be called "large holes," with 12 1/4 in. as the "extra large" class and holes of smaller diameter classed in the slim-hole range, 6 1/4 in. representing about the minimum size. The optimum size has not been definitely determined, but the preference is for a program that would call for setting 8 5/8-in. surface casing and digging a 7 3/4-in. hole. Different sections of the country might very properly deviate from this practice because of differences in depth, formations, regulations, or other factors. This paper will be confined to a discussion of the drilling of exploratory wells, as the drilling of this class of wells is quite a different problem from that of drilling a proven field, although many of the same considerations apply in both cases, as will be apparent. Although for many years shallow slim-hole and core-test holes of small diameter had been dug, it was not, I believe, until the fall of 1937 that serious consideration was given to its application in drilling wells 7000 ft., 8000 ft. and 9000 ft. deep on the Gulf Coast. This does not mean that nearly all such wells drilled on the Gulf Coast now are of the slim-hole variety, but only that efforts have been aimed in that direction with very encouraging results and there has been a very definite trend to reduce the size of holes, a trend that probably will continue. As to exploration policy, there appear to be two distinct and opposite views. The one usually followed in the past was more or less designed to "shoot the works" on one well. Generally on the Gulf Coast, the depth at which production might be obtained is limited by the drilling equipment and economics. With this thought in mind, wildcat wells are often planned with the idea of carrying the depth to 10,000 ft., or even 15,000 ft. All of the possible hazards are anticipated and the general drilling and casing programs are so predicated. Such wells contemplate the drilling of large-diameter holes; the setting of at least one protective string of casing and an extensive coring policy, all of which finally adds up to an expensive well. As opposed to this general policy is the slim hole. This, on the Gulf Coast, recognizes that new discoveries represent gam-

Jan 1, 1941

By F. S. McNicholas

In the first big blast at Climax, a "loop back" (three-wire system) was used (Fig, I), with the idea of securing a wiring system that would give to all series the same amount of current. Single-phase, 60-cycle, 13,800-volt power was furnished to the blasting transformers, each of which has a rated capacity of 600 amp. at 480 volts. Two blasting leads were taken from these transformers, giving a blasting current of single phase, 60 cycle, 480 volts, with an instantaneous capacity of more than 600 amp. The blasting switch was on the secondary leads from the transformers, to avoid lag in energizing them.

Jan 1, 1940

By F. S. McNicholas

In the first big blast at Climax, a "loop back" (three-wire system) was used (Fig, I), with the idea of securing a wiring system that would give to all series the same amount of current. Single-phase, 60-cycle, 13,800-volt power was furnished to the blasting transformers, each of which has a rated capacity of 600 amp. at 480 volts. Two blasting leads were taken from these transformers, giving a blasting current of single phase, 60 cycle, 480 volts, with an instantaneous capacity of more than 600 amp. The blasting switch was on the secondary leads from the transformers, to avoid lag in energizing them.

Jan 1, 1940

By Arturo M. Morales

INTRODUCTION Manganese carbonates are nodulized by Cia. Mineral Autlan, located in Ongongo in the state of Hidalgo in Mexico. This kiln is fed minus 15 mm material at the rate of 1870 tpd. About 920 tpd is nodulized into a commercial product having a size range of 15 mm to 12 mm. The kiln also produces about 320 tpd of minus 12 mm fines and 40 tpd of chimney dust. Currently, the minus 12 mm fines are being recirculated to the kiln and the chimney dust is not being utilized. It has been found that there exists a market for manganese to be used as feed to the ferromanganese and silicomanganese furnaces. However, a coarse, lumpy feed can only be used in these furnaces. Accordingly, it has been proposed to screen the fines at 6 mm and agglomerate the minus 6 mm fraction along with the chimney dust and suitable binders. This paper presents the results of a feasibility study and test work aimed at producing an acceptable agglomerated product from the kiln fines for use in the ferromanganese and silicomanganese furnaces. Since the agglomerated product will have to be handled and loaded into trucks in the kiln area and transported over long distances, a minimum strength of 200 N is expected to be necessary. They should also have enough initial strength to be stockpiled and result in a minimum of breakage. A third important requirement is that the agglomerates should be resistant to disintegration during combustion. The recirculating fines were to be agglomerated in their original state, if possible, that is, without grinding. This eliminated the possibility of pelletizing as an agglomeration method. Based on a few laboratory trials, the briquetting method seemed most attractive, and a test program was initiated to explore the briquetting process.

Jan 1, 1977

By E. L. Larison

DURING the early years of the present century a notable interest appeared in American industry in the matter of recovering and render-ing profitable byproducts of manufacturing operations. Among the byproducts of reduction of sulfide ores in many plants large amounts of sulfur dioxide gas were available. In certain instances the concentration of S02 in the gases evolved was, or could be made, sufficient to make them suitable for manufacturing sulfuric acid. At the Anaconda Reduction works such a condition existed. Large tonnages of copper concentrate were being roasted and the roaster gases were available and suitable as a source of sulfuric acid. Earlier work on a large scale at Martinez, Calif., and in the Ducktown region, Tennessee, had shown the feasibility of manufacturing sulfuric acid from the S02 contained in gases arising from the reduction of copper sulfide ores. In California and Tennessee the acid found a ready market in manufacture of fertilizers, while the situation regarding disposal of the sulfuric acid was quite different at Anaconda. No local use for acid existed, and a commodity of such low value could not stand freight charges to consuming centers. There was a very large accumulation of sand tailings from the gravity concentration mill below the plant at Anaconda. This material, made by the comparatively crude methods in use up to about 1915, contained about 13 lb. of copper per ton, largely as sulfide minerals. In 1912 serious research work toward recovery of this copper was instituted and after about two years of laboratory and pilot-plant development, a process for accomplishing it was perfected. This process involved roast-ing the sand tailings, leaching the roasted tailings with sulfuric acid, and throwing out the copper by passing the solutions over scrap iron. Under this plan each ton of sand required the introduction of about 70 lb. of 60° Be. sulfuric acid.

Jan 1, 1934

By R. P. Carreker

ONE of a series, this report is concerned with the experimental documentation of the deformation behavior of pure metals over a wide range of temperature. Previous reports in the series describe the creep behavior of platinum1 and aluminum2 and the tensile behavior of copper,&apos; silver,& and molybdenum.5 Subsequent reports will describe the creep behavior of copper and silver and the tensile behavior of platinum, thus providing extensive creep and tensile data on four face-centered-cubic metals tested under comparable conditions. Material Two lots of high-purity aluminum were used in this investigation. One lot, designated AC, was obtained direct from the Aluminum Co. of America as high-purity aluminum in notch-bar form. A typical analysis of this type of aluminum is 0.006 pct Si, 0.015 pct Cu, 0.006 pct Fe, and 99.975 pct Al. Lot BS was obtained from J. I. Hoffman of the Bureau of Standards as a portion of the Aluminum A described in his report of the redetermination of the atomic weight of aluminum.6 Lot BS originated with the Aluminum Co. of America also, and was received in the standard notch-bar form. A detailed chemical and spectroscopic analysis of lot BS has been published.6 Principa1 impurities were 0.006 pct Si, 0.003 pct Fe, 0.002 pct Cu, and >99.987 pct Al. Samples were processed identically by room-temperature swaging and drawing to 0.030 in. diam wires from approximately ½ in. diam rods that were machined from the notch-bar pigs. Samples were cut from the cold-drawn wire, placed in a grooved graphite block in groups of thirteen, and annealed in air for 1 hr at each of several temperatures, with the results shown in Table I. The two lots of aluminum responded quite differently to annealing, the behavior of aluminum AC suggesting that its grain growth was controlled by impurities. The recrystallization texture of aluminum of similar purity comparably treated has been reported as predominantly <111> fiber with some <100> fiber present.7 Testing Procedure The testing procedure and method of analyzing the data have been described previously.2,3 Annealed specimens of 5 in. gage length were tested in an Instron tensile testing machine at a constant head motion corresponding to a strain rate of 0.04 min-&apos; at a number of temperatures. Instantaneous tenfold rate changes were employed in some tests to deter-

Jan 1, 1958

By Frank Linforth

THE object of this paper is to present a brief outline of the methods of geologic mapping employed in the Geological Department of the Anaconda Copper Mining Co., at Butte, and to show by means of a few typical examples the practical nature of the results obtained. The extremely complicated geological conditions early encountered in Butte mining led mine operators to realize the necessity of accurate detailed geologic mapping as an aid to successful development of the ore bodies. The organization of the Anaconda Geological Department was undertaken in 1900 by H. V. Winchell, who, in conjunction with D. W. Brunton, worked out the essential methods of procedure which are, in the main, followed at the present time. Since the year 1906 the geologic work has been under the direction of Reno H. Sales, and in some respects the scope of the work has been slightly enlarged. Au excellent article describing the general equipment of the geological department was presented by D. W. Brunton at the British Columbia meeting of the Institute in 1905.1 His idea, that data collected underground and recorded should be put to practical use and not remain a mere " inventory of the company&apos;s underground possessions," is the keynote to the success of applied geology in Butte. The geologic notes are taken as soon as possible after the ground is broken so that any mistake in the mining may be corrected at once, or any particularly advantageous procedure may be suggested before any useless work is done. Taking the notes underground is a comparatively simple matter, but a few necessary precautions may be pointed out. The essential to success is that the notes shall show exactly what geological facts are disclosed. The observer must discriminate carefully between important and unimportant exposures in making the record, especially as regards the relation between veins and stringers, between faults and minor slips, or between the characteristics of veins of different ages. A simple color scheme has been adopted for making this record: Red pencils are used to indicate.

Jan 11, 1913

By L. M. Barker

THE New Cornelia Branch of the Phelps Dodge Corporation is at Ajo, Pima County, Ariz. It is at the southern terminus of the Tucson, Cornelia and Gila Bend Railroad, which connects with the Southern Pacific Railroad at Gila Bend, Ariz., approximately 43 miles to the north. HISTORY Early history of the Ajo district has been related by A. W. Allen1 and by T. A. Rickard.2 Briefly, steps looking to the large-scale development of the ore body were instituted as early as 1912. This work culminated in the equipment of the property for open-pit mining by steam shovels. Ores initially mined carried economic amounts of copper carbonates as the principal copper-bearing mineral. Treatment methods developed for the extraction of the copper content involved leaching and a 5000-ton leaching plant was constructed and placed in operation in early 1917. Description of methods and equipment employed in this plant has been given by A. A. Tobelmann and James A. Potter.3 As exhaustion of the carbonate ore body neared, and mining oper-ations exposed an adequate tonnage of sulphide ores, a 5000-ton con-centrator was constructed for their treatment. That plant has been described by H. Kenyon Burch.4 Construction was completed and the plant was placed in operation in early 1924. Changes made in the original plant to improve practices and increase production will be the subject of subsequent paragraphs of this paper.

Jan 1, 1939

By Franklin R. Carpenter

Pyritic smelting, so-called, as practiced in the Black Hills is pyritic smelting only in the sense that Dr. John Percy uses the expression in his " Metallurgy of Silver," where he describes a process for smelting dry silver-ores in connection with iron pyrites, to form a regulus. The ores treated by this process in the Black Hills are very siliceous, averaging about 74 per cent. of silica, from 10 to 20 of iron oxide, 4 or 5 of alumina, and from 2 to 4 of lime. In 1889, while Dean of the South Dakota School of Mines, I took up the question of their smelting. There were neither copper- nor lead-ores in the Hills; hence pyrite and pyrrhotite, of which there are large deposits, were employed to form a regulus. This mixture of pyrite and pyrrhotite, besides being absolutely barren of gold and silver, carried an average of 30 per cent. of silica. The only basic material available as a flux for the great excess of silica in the ores and in the pyrite was a dolomitic limestone. Moreover, the local coke gave 24 per cent. of ash, 60 per cent. of which was silica—as unpromising a set of conditions as can well be imagined. We had before us the examples of Mr. W. Lawrence Austin, of this Institute, who had experimented at Toston, Montana,"

Jan 1, 1901

By Robert A. Cummings, Francis S. Kendorski

The behavior of the rock mass during block caving affects the production of the mine and dictates the loads and deformations placed on required drifts. Traditionally, the drifts used are massively reinforced and buttressed against very high and erratic mining- induced abutment loads or swell muck loads, resulting from undercut- ting. Extensive field investigations in numerous block caving mines in the porphyry intrusive environment, coupled with models developed for tunneling and longwall coal mining, have allowed the development of a unified hypothesis for ore bodies of various degrees of competence. In general, the hypothesis incorporates several sequential steps in the driftlrock mass deformation history, each with characteristic magnitudes, duration, and areal extents depending on ore body competence. 1. Prior to initiating caving, but subsequent to drift development, the rock mass and drift are in a state of near-equilibrium. Zones of stress concentration have stabilized, as for a tunnel. 2. For a newly undercut panel, the rock mass and drifts directly beneath the undercut respond by dilating towards the void created, and undergo a release of compressive strain that manifests itself as tensile failures. 3. For adjacent areas, the rock mass is compressed as the load is shed from the undercut level.

Jan 1, 1983

By William J. Potter, Roger N. Pyle

Introduction The primary ingredient for survival of mining companies during a recessionary period is to be innovative and resourceful in the structuring and financing of operations. Over the past decade, many mining companies have had to retrench and restructure their operations and to redeploy their assets just to survive. During the same period, traditional sources of financing for mining companies -- commercial banks, subsidized export credit loans and long term straight debt financing in the public markets -- became less available and more expensive. Meanwhile, new markets and new financing instruments provided new alternatives for mining companies. Those companies which exploited these alternatives greatly enhanced their capacity for survival and even for profitability. This article considers the strategies adopted by some companies to adapt to changes in this recessionary environment; the inadequacy of traditional sources of financing; certain of the new financing instruments, structures and markets which have become available; and how several mining companies have utilized these new instruments, structures and financial markets to improve their competitive positions. THE IMPACT OF RECESSION ON MINING COMPANIES AND TRADITIONAL FINANCING SOURCES In its 1983 year-end article "The New Economic Order" Mining Journal characterized many of the large private sector mining companies with operations in the industrialized countries as being "gripped between the jaws of high working costs and stagnating metal commodity prices." The article went on to characterize the Canadian mining industry in particular as being "in the difficult position of having to seek to sell the bulk of its output in a world market characterized by over-capacity, where production costs were determined by internal political and economic factors while prices were set by external competitive forces of supply and demand." Miners of a wide range of commodities in the industrialized countries have come to realize that their domestic mineral resources are in many cases no longer competitive with global sources of supply. To compete and survive they would have to rely on their technical and financial expertise in mineral exploration and development and in mining and processing. Strategies For Survival Global competition has compelled management in the extractive industries to improve their worldwide cost position. This is the only means whereby a mining company can differentiate its product and attain a substantial competitive advantage. During the past decade, for example, ASARCO Inc. and Inco Ltd. chopped their payrolls, postponed expansion plans and otherwise retrenched. ASARCO reduced its capital expenditures and cash requirements for operations in 1984 to a level almost equal to its internally generated cash. The company&apos;s finances were also strengthened by increasing total lines of credit from outside sources, by selling common stock, and by reducing annual pension funding through a change in actuarial assumptions. Other companies returned to their traditional organizational structure where they had a demonstrated competitive strength. Consolidated Gold Fields, for instance, resumed its operation as a mining finance house. Its capital investment programs are now devoted solely to resource development. In mid-1984, at a time of firming zinc prices, Union Miniere S.A. used the opportunity to integrate its smelting capacity and to direct investment to certain subsidiaries at the expense of less viable operations. Smaller corporations realized that it was not enough just to be the low-cost producer. For example, Drummond Coal Company in Alabama, which dominates its regional market due to its low-costs, has also developed a world-wide trading capability and has placed incremental tonnage from its own mines as well as brokered coal produced by others, thereby minimizing capital risk. In Canada, Campbell Resources has exploited difficult times by acquiring assets cheaply. In the process it improved its capacity to withstand cyclical economic swings by diversifying from solely a copper and gold producer to a company based upon asbestos, coal, copper, gold and natural gas.

Jan 1, 1985

By L. Wahl, M. Allard, B. Trentini

REMOVAL of sulfur in steelmaking is presently a problem of utmost importance as requirements on final sulfur contents in finished steels become increasingly strict. This is in spite of often increased sulfur contents in the fuels used during the different stages of steel processing. According to local conditions or current economic factors, sulfur can be removed during one or more of the processing operations; for instance, in a blast furnace operated under high temperatures and with a high slag basicity ratio and also by means of high inputs of manganese into the burden. These methods of proceeding are, however, expensive and in contradiction with the acid practice which is gradually being developed owing to the fact that it allows coke and lime savings and usually permits an increase of productivity. In the melting-shop, desulfurization can proceed in the open-hearth or electric furnace with one or several very basic slags, or with special prefused slags (Perrin process) mixed with steels poured from the open-hearth, electric furnace, or converter. Liquid pig iron has also been treated separately before being used in the melting shop, in order to reduce high sulfur contents. This external desul-furization has been developed in Europe for the basic Bessemer Thomas process, which is again in favor owing to the great flexibility of such installations and to the fact that the steel produced is not polluted by impurities coming from the scrap. Since the basic Bessemer process, as usually practiced, has a low desulfurizing power, external desulfurization is generally carried out in ladles by means of alkalies such as soda ash. This method, which has the advantage of being simple and costing little, has only a medium desulfurizing power, which implies the necessity for repeating the operation several times when a substantial amount of sulfur has to be removed or when very low sulfur contents are desired. Thermal losses are high and there is a considerable decrease in output. Moreover, the resulting fluid and corrosive slag, which is extremely difficult to eliminate, attacks the refractory linings of the mixers and steel furnaces. Therefore, it seems necessary to devise a simple method for external desulfurization which would not only remove substantial quantities of sulfur from the hot metal coming from an acid blast furnace, but which would also bring the sulfur contents down to the low level required in the making of high quality steel and iron. Principle of the Process In this paper a new method will be described for desulfurizing hot metal. It has been perfected by

Jan 1, 1958

By J. S. McNeil, J. U. Messenger

In the drilling of oil wells the control and prevention of lost circulation of the drilling- fluid is a oroblem which is frequently encountered; in many cases existing materials and methods for alleviating this condition have not been adequate. For severe cases of lost circulation in which the simpler methods, such as application of various bridging materials to the mud system, have proven unsatisfactory. a new material and method for applying to the loss zones have been developed which appear to he superior to existing techniques in many respects. The material, called a clay cement, is capable of being handled as a drilling fluid after initial mixing of the solitl ingredients with water and may be pumped down the drill pipe and squeezed into loss zones. After a short period, the material develops a very high gel strength which seals off the zone against further losses of the drilling fluid. In field tests, the process has been demonstrated to be a very effective method for combatting lost circulation. INTRODUCTION Methods of treating lost circulation in rotary-drilled wells by the addition of special materials to drilling fluids have long been in the process of development and are continuing to improve.&apos; However, the annual drilling costs which can be traced directly to lost circulation difficulties still run into the millions of dollars. The loss of mud materials into highly permeable zones may range up to $50,000 per well in some areas. In addition to mud costs, excessive requirements of casing, cement; rig time, and the attendant blowout hazards further emphasize the importance of the need for improved methods of combatting lost circulation. The shortcomings of presently-used materials and methods have led to the development of a new lost returns material in the form of a time-setting clay cement. In field tests this material has proven very successful in plugging off loss zones. which other more conventional materials had failed to do. REVIEW OF PRESENT CORRECTIVE MEASURES The proper control of such operating factors as pump pressure, clearance between pipe and hole, and speed of running pipe in a holez is of primary importance in preventing and minimizing lost circulation problems. These factors, however, concern matters of drilling technique and do not directly involve lost circulation materials. Aside from setting casing past loss zones, three principal methods are generally employed for the purpose of avoiding or combatting lost circulation: (1) adjustment in properties (principally density) of the drilling fluid, (2) addition of fibrous, flaky, granular or other types of bridging materials to the mud system, and (3) the application of cements to the loss zones. The latter method is usually employed only after the first two have proven unsuccessful. In Table I are listed some advantages and disadvantages of the common corrective measures. Frequently, a combination of bridging material and cementing material is used." This combination is based on the premise that a bridge will

Jan 1, 1952