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By William Hustrulid

INTRODUCTION The Idarado mine lies high in the San Juan Moun¬tains on the divide between the Uncompahgre and San Miguel Rivers and consists of a consolidation of a number of old and prominent mining properties through which course some famous and very produc¬tive veins. Among the better known are the Smuggler, Tomboy, Montana-Argentine, Black Bear, Liberty Bell, Virginius, Flat, Barstow, and Japan, as well as many others. Most of these veins have been extensively mined over the past century and in the last three decades the Montana-Argentine and Black Bear veins have been the backbone of the Idarado mine; therefore, the descrip¬tion of the veins is limited to them. Access to the mine is through either the Treasury tunnel, whose portal is below Red Mountain Pass on US Highway 550 at an elevation of 3244 m (10,643 ft), or the Mill Level tunnel entrance 3.2 km (2 miles) east of Telluride, CO, at an elevation of 2761 m (9060 ft). The Treasury tunnel intersects the Black Bear vein 2643 m (8670 ft) from the portal and the Mill Level tunnel intersects the Argentine vein 2179 m (7150 ft) from the portal. Mining is by shrinkage stoping from slusher sublevels. The size of the scope blocks varies somewhat, but the standard size is 67 to 76 m (220 to 250 ft) long and 61 to 76 m (200 to 250 ft) high. The mine ranks either first or second in Colorado in yearly production of gold, silver, copper, lead, and zinc. It is 9.6 km (6 miles) from the Red Mountain plant to the Pandora plant, via interconnecting drifts and raises. There are engineering offices at both plants as a matter of convenience. The Red Mountain plant in¬cludes the company general offices, warehouse, carpen¬ter and machine shops, and mine change room. The Pandora plant consists of the mill and assay office, ma¬chine shops, and mine change room. The flotation mill has a capacity of 1632 t/d (1800 stpd), making a bul¬lion product and separate concentrates of lead, copper, and zinc. HISTORY OF THE MINE The Montana-Argentine vein was first extensively worked by the Tomboy Gold Mines Co., Ltd., a British concern. This company mined the stoped areas above the Ophir level between 1910 and the late 1920s and most of the stoped areas above the 2100 level between 1900 and the late 1920s. Gold was the principal ore metal mined. The area between the Revenue and Ophir levels was mined chiefly by the Revenue Mines Co. between 1900 and 1910. The ore was worked from the Revenue tun¬nel, which portals in Canyon Creek. Gold and silver were the chief metals recovered. The stopes between the 1700 and Revenue levels, as well as some higher stopes, were mined by Telluride Mines Inc. during the 1940s. The Mill Level tunnel was driven by that company from 1945 to 1948. Lead and zinc then became economically more important than the precious metals. In 1953, Idarado purchased Telluride Mines, which merged with the parent company in 1956. The Black Bear vein was first extensively worked by the Black Bear Mining Co. in the 1900s and by the Colorado Superior Mining Co. from about 1914 until snowslides at the mine camp [altitude 3750 m (12,300 ft)] terminated the company's operations in 1924. Leasers operated at intervals until 1934. The Treasury tunnel, formerly the Hammond tunnel, had been started before 1900 and reached the 1646-m (5400-ft) mark early in the 1900s, at which time activity lagged until the late 1930s. In the early 1940s, Idarado extended the Treasury tunnel from its heading at 1646 m (5400 ft) to the Black Bear vein and established a raise connec¬tion with the 600 level, the lowest level in the old mine. Since completion of initial work in the mid-1940s, systematic development of the mine, both in the driv¬ing of new headings and the utilization of older open¬ings, has resulted in the present extensive network of workings. GEOLOGY Introduction The oldest rocks of the area are the Precambrian metamorphic rocks, the Uncompahgre formation which are massive quartzites, some phyllites and slates, and dolomitic or limestone beds. West-dipping Paleozoic and Mesozoic strata lie on the Uncompahgre formation. These units are separated by a major angular uncomformity from overlying, essentially horizontal, Tertiary formations that include the basal Telluride conglomerate and several thousand feet of overlying volcanic rock. Intrusive rocks, mostly Tertiary in age, are common in the area, and are in the form of dikes, stocks, and sills. Dikes are very closely associated with some of the ore veins (Mayer). Description of Veins The Black Bear and Argentine veins range from 0.6 to 7.6 m (2 to 25 ft) in width, but in most places are 1.5 to 2 m (5 to 7 ft) wide. They vary in character from a well-defined tabular structure between sharp "frozen" walls, or gouge seams, to an irregular zone of quartz and quartz-sulfide stringers. Many gouge seams within the veins make the veins blocky and loose. Common gangue minerals in the veins include quartz, pyrite, rhodonite, chlorite, sericite, clay minerals, epidote, calcite, adularia, rhodocrosite, fluorite, and specularite. Quartz constitutes 60 to 70% of the veins and varies widely in character, ranging in color from clear through white, gray, and green, to amethyst. Chlorite, sericite, the clay minerals, and fine-grained quartz are common alteration products of the vein walls and of wall rock fragments in the veins. Epidote is an altera¬tion of the dike or, less commonly, of tuff-breccia. Calcite, adularia, fluorite, and rhodocrosite, which are wide¬

Jan 1, 1982

By Kalle Hakalehto

Automation and mechanization in mining has no value of its own. It has to be measured in economic terms or in Increased safety or in the utilization of deposits which otherwise were unavailable. The mechanization and automation of workphases in the mining and rock construction process has led to the decrease of manual work. This has had a very essential importance in the countries where the labor is expensive or where there has been difficulties in the availability of labor. The work has become much more efficient. We would like to argue that the need of mechanization has been initiated mostly from the above mentioned reasons. The origin has seldom been the "process" approach or the real need or will to change the way mining is done. Through mechanization bigger unit sizes of machines and operations have become possible and better efficiency is achieved. This has led to new problems as well, but in most cases they have been solved so that better and more economic outcome than before has been the end result. The changes have thus been justified. It is perhaps feasible to state already at this stage that we see that the mechanization andautomation steps in the difficult conditions of mining have been remarkable. There is not a slightest underestimation of the achievements and importance of this development in our mind. Where we are to- day is the result of the experience, know-how and technological excellency of all the parties involved. The natural approach to automation has been through mechanization and automation of individual unit operations and machines. This has been the way as well in manufacturing industry as in mining. There may not have been any other way. The other presentations in this session tell us very clearly of the approach .which the machine manufacturers have taken. They have not done it alone but 'most of the initiatives have come from them. A good use of their present achievements gives many possibilities to increase the, efficiency in mining. We would, however, like to extend the view one step further to the future:

Jan 1, 1989

By Jack Parker

Cycles of mine convergence coincide with annual cycles of temperature and humidity in the mine. Laboratory measurements confirm that the presence of moisture reduces rock strength and induces internal rock stress. Findings are used in the design of mine openings.

Jan 1, 1969

By Michel David

Once a mineralized body has been delineated, drilled and estimated from surface samples, it often happens that before the final go-no-go decision is taken, an underground sampling program is decided, very often via a ramp which hopefully will be later used to extract the ore. The cost of such a program can be close to a million dollars. On other occasions, when a property changes hands and is reexamined maybe 20 years after the original drilling program, a "check sampling" program is carefully, designed to make "sure" that the old assays are "correct". One may fear poor sampling practice in old days, or deliberate highgrading by promoters. Again one may be talking of resampling cost which may be in the six-figures range. Other similar situations include the drilling of a few large diameter holes to ascertain the validity of a majority of small diameter drill holes, or auger drilling, or back-rimming. In uranium checking radiometric information is also a problem related to the above ones.

Jan 1, 1979

By P. Magnani

Quarry environmental reclamation can be achieved with favorable economical advantages and high energy efficiency, using stabilized materials obtained by the treatment of organic fraction of municipal solid waste (MSW).

Jan 1, 2005

By Graeme J. Jameson

A form of column flotation cell is described in which the contact between the feed and the air stream is made in a mixing device at the top of a vertical downcomer. The air-liquid mixture flows downward to discharge into a shallow pool of pulp in the bottom of a short cylindrical column. The bubbles disengage and rise to the top of the column to overflow into a concentrate launder, while the tails are discharged from the bottom of the vessel. The main advantages of the device described are that the overall height of the column is reduced to about one metre (3 ft), and the column can be self-inducing with respect to the air supply. Experimental results are given from tests on a feed stream to a conventional zinc cleaner circuit. Using wash-water to reduce gangue entrainment, the column gave high concentrate grades and high recoveries.

Jan 1, 1988

By L. P. Meade

The need to feed an established talc mill from a crude ore source located closer to the mill prompted the 1986 drilling of a previously mapped, talc-carbonate deposit. This talc-bearing rock had never been considered a potential source of crude talc ore because of the high carbonate content and other deleterious minerals observed in outcrops. Selective mining, in-pit grade control and creative processing have allowed this deposit to be developed into an ore body. By definition, an ore body has to be a source of positive cash flow. With industrial minerals, this is determined as much by market acceptance as by ore tenor.

Jan 1, 1994

A form of column flotation cell is described in which the contact between the feed and the air stream is made in a mixing device at the top of a vertical downcomer. The air-liquid mixture flows downward to discharge into a shallow pool of pulp in the bottom of a short cylindrical column. The bubbles disengage and rise to the top of the column to overflow into a concentrate launder, while the tails are discharged from the bottom of the vessel. The main advantages of the device described are that the overall height of the column is reduced to about 1 m (3 ft), and the column can be self-inducing with respect to the air supply. Experimental results are given from tests on a feed stream to a conventional zinc cleaner circuit. Using wash water to reduce gangue entrainment, the column gave high concentrate grades and high recoveries.

Jan 1, 1987

By James R. Miller

This paper will describe the balling disc operation of the Marcona Mining Company at San Nicolas, Peru, with special reference to green ball quality. The report covers experience during the period of September and October 1965 and will be, more a state-of-the-art at Marcona report rather than -a technical treatise.

Jan 1, 1966

By C. R. Rinschler, G. R. S. Smith

A high-strength, industrial-grade membrane filter- cloth capable of high flow rates and sub-micrometric particle retention has been developed. A filtration system using this filter- combined with special filtration techniques and using tubular-filters makes possible the filtration of many difficult-to-jilter slurries. Small, discrete, slowly-settling particles are formed into filter cakes that when pulsed of the membrane filter cloth do not redisperse but settle rapidly to form a dense filter under- flow. A positive separation of solids from liquids is obtained.

Jan 1, 1991

By L. F. Moran

Rod Mill Grinding is an important and fundamental step of the concentrating process in Marcona Mining Company's plants at San Nicolas Bay, Peru. This Paper summarizes experience in regard to consumption and specifications of grinding rods based on twelve years operation of the grinding circuits. Marcona operates a Plant complex producing 10 million tons per year of beneficiated iron ore products including pellets, sinter feed concentrates, high grade concentrates as slurry and filter cake. Processing capacity is 15 million tons per crude ore per year of which 13 million are around in rod mills. A continuous testing program was carried out over the years. Initial studies of daily production operations gave indications of poor grinding rates in relation to mill size and power consumption, with a high rate of broken liners and rods. The broken nieces of rod were not automatically expelled and further interfered with grinding efficiency, while replacement of liners caused unwarranted do n tire. Objectives were to improve all phases of the Process by studying the effects of changes in rod diameter, length, alloy constituents, and volume and weight of rod charge.

Jan 1, 1975

By J. B. Pautler, A. E. Gross, M. G. Strominger

The U.S. Bureau of Mines research on the use of cement as an agglomeration aid allowed precious metals heap leaching to become an economically viable method of precious metal recovery during the 1980's. However, many agglomerated ore bodies still exhibited poor heap/agglomerate strength. Precious metal recovery was often too slow or required prohibitively high levels of cement to be economically sound. Problems with heap ponding, blinding, and solution channeling and "wash-out" indicated that there was a need within the industry to improve on existing agglomeration technology. An in-depth research effort was undertaken to satisfy this need. Three years of research led to a patent on the use of vinyl addition polymers as effective agglomeration aids. A close working relationship between Brewer Gold and Nalco Chemical Company aided in the development of this program. This paper details the pilot test work, the full-scale evaluation at Brewer Gold Company, and conclusions.

Jan 1, 1990

By H. E. M. N. Moursy, Hassan El-Shall, M. N. Ahmad, Ponisseril Somasundaran, C. R. Holland, D. Chandler, G. Walker, Regis Stana, Patrick Zhang

The reaction kinetics of the acidulation of Moroccan phosphate rock (32% P?05) using 45% nitric acid were investigated using the Shrinking Core Model (SCM) , a newly developed Diffusion Model (DM) and a newly developed Combined Reaction and Diffusion Model (CRDM). This was achieved buy conducting a number of experiments in a controlled batch reactor system at different temperatures. The conversion and the reaction rate were calculated at different time intervals. The SCM based on reaction time and the CRDM characterize the acidulation process of phosphate rock. Diffusion of nitric acid and the products of acidulation at the early stages of reaction through the phosphate rock were found to be the rate-controlling step, resulting in a dif- fusion coefficient, which controls the mass transfer process. The reaction kinetics were found to fit first-order kinetics especially during early stages of reaction time. The deviation of the kinetics from first-order at later stages was due to the fact that the overall reaction was mass transfer controlled with respect to nitric acid and products diffusion.

Jan 1, 2002

By K. I. Cunningham, C. S. Spirakis

Two types of native sulfur deposits have been found in Lechuguilla Cave: thin late-stage crusts and massive early- formed deposits. Most, if not all, of the crusts appear to have formed by remobilization of native sulfur contained within gypsum and sulfur masses. The process is occurring today and is slowly converting the sulfur to gypsum. The absence of primary gypsum or reductants in the cave indicates that the early-formed massive sulfur deposits did not form by in-place reduction of sulfate, which is the process that formed the commercial sulfur deposits in the Delaware basin. Instead, it appears that H2S dissolved in basinal brines was oxidized by oxygen in near-surface ground waters in the cave. Fluid- inclusion data, which suggest the presence of both fresh water and saline water as the sulfur formed, support this interpretation. Although individual sulfur deposits in Lechuguilla Cave are no more than several tens of metric tons, the abundance of secondary gypsum (as much as 10 meters thick) indicates that native sulfur may once have been much more extensive.

Jan 1, 1992

By G. S. Austin, L. A. Brandvold, J. Renault

A series of tests were run on soil samples collected near the site of a Socorro, NM lead smelter that was abandoned more than 98 years ago. Three 30.5- x 30.5-m (100-X 100-ft) squares provided 13 samples each. As described in Part I (this issue), each sample was divided into seven splits according to particle size. A mineralogical examination of the fine-size fraction of Pb-contaminated soils failed to reveal crystalline Pb compounds. These authors conclude that Pb is present as amorphous phases or is adsorbed on the surfaces of soil particles. Site vegetation contained elevated Pb levels that increased with increasing Pb amounts in the soil. However, the biological absorption coefficient (BAC) indicates that some of the site's Pb is unavailable, or the plants are able to limit their uptake. Of the vegetation tested, rabbit bush showed the greatest increase of Pb levels in its roots, whereas levels in creosote were highest in the foliage. In cactus, the Pb levels varied. The amount of Pb due to dust incorporated in plant tissue is almost negligible compared to the amount absorbed through the plant vascular system.

Jan 1, 1994

By Paul M. Jr. Musgrove, Donald C. Moore

Conventional Smelting Practice Conventional copper smelting practice varies from smelter to smel¬ter, but generally consists of some or all of the following unit processes: roasting, smelting, converting, and fire refining. Roasting. Copper sulfide concentrates can be smelted directly or after an initial roasting step. Roasting is used in some smelters because roasting prior to smelting increases smelting capacity, less energy is required to melt hot roaster calcines than wet sulfide concen¬trates, roaster off gases are high in Sot concentration, 5-15% SO2, and some volatile impurities are removed from the concentrate prior to smelting. However, many smelters do not use roasters, because the problems associated with handling hot dry calcines outweigh the advantages mentioned. Concentrate roasting is performed in multiple hearth or fluid bed roasters. If the moisture is low, roasting can be performed autogenously, usually at 500-600°C. High roasting temperatures are avoided because excess oxidation of the iron compounds may lead to magnetite formation. Magnetite is detrimental to reverb operation because mag¬netite can combine with refractory minerals to form a highly viscous slag. This slag prohibits efficient matte-slag separation and leads to excessive copper losses. Also, magnetite can settle through the matte layer, deposit on the furnace bottom, and consequently reduce furnace capacity. Roasting is carried out only on sulfide concentrates prior to smelt¬ing in reverb or electric furnaces. For smelting processes, such as the flash and continuous that rely on the exothermic heat of oxidation of the sulfur minerals, roasting is not practiced. Reverberatory Smelting. The predominate copper smelting fur¬nace for the past 50 years has been the reverb. These furnaces are typically 100-120 ft long, 30-35 ft wide, and 12-15 ft high. A typical furnace layout is shown in Fig. 2. Refractory brick linings cover all internal surfaces of the furnace. Originally the flame was directed to reverberate or reflect off the furnace ceiling and melt the feed material. Current practice is to direct the flame down the furnace length to melt the concentrate. A method of charging the concentrates or calcines, generally along the side walls to minimize refractory erosion, is incorporated in the furnace design. The copper concentrates, calcines, and fluxes charged into the reverb undergo a series of complicated reactions as the temperature of the mixture increases. The reaction of the iron and copper sulfides with the oxygen in the furnace produces a molten Cu25-FeS mixture called matte. Copper smelting metallurgy is based on the fact that sulfur has a greater affinity for copper than for iron and most other common metals. Therefore, in a system containing copper, the copper will preferentially remain as a sulfide compound until all of the other metals have been oxidized. The oxidized metals combine with silica to form a silicate slag that floats on the matte and is removed from the system. Reverberatory furnace smelting chemistry can be approximated by the following chemical equations: FeS2 + O2 - FeS+ SO2 (1) The formation of FeS ensures that any copper present other than as sulfides will be reduced by the relationship: CuO2 + 2FeS + O2 - CuS + 2FeO+SO2 (2) or 2Cu +FeS - Cu2S + Fe (3) As the molten charge travels down the furnace, continued oxida¬tion of the iron minerals and sulfurization of the copper minerals occurs. When all of the copper has been converted to sulfides, the iron sulfides can then be further oxidized as: FeS + (3)2 O2 FeO + SiO2 (4) The FeO reacts with the silica added as flux in the furnace charge. A simplified equation is: FeO + SiO2 -FeO SiO2 (5) The iron silicate slag formed is skimmed from the surface at the end opposite the burners. The copper content of reverb slag is usually less than 0.6% Cu and is discarded. Matte is removed along the side wall and is taken to the converter for oxidation of the remaining sulfur and iron. The main objectives in reverberatory smelting are to produce a molten Cu2S-FeS matte containing 30-60% Cu and a throwaway slag. Production of matte permits complete conversion of all copper minerals into copper sulfides, which can migrate because of specific gravity differences, through the lighter slag layer. Also, the molten matte droplets collect the noble metals, gold and silver, as the matte settles in the furnace. The large settling area of the furnace provides enough separation time to produce a low grade slag, which can be discarded without further processing. High heat losses are associated with reverberatory smelting be¬cause of the large volume of gases sweeping through the furnace. Therefore, an outside source of heat is required to keep the smelting reaction going. Natural gas, fuel oil, or pulverized coal are used as this heat source.

Jan 1, 1985

By Ed Ramsay

The Hill is a cyanidation plant with flotation and roasting. The grinding and treatment have at the present time a capacity of 1400 tons of ore per day and the precipitation has a capacity of 6000 tons per day. The ore is delivered by a 36" conveyor belt and tripper to 8 mill bins having a capacity of 6000 tons. The ore from these bins is discharged onto 20" gathering conveyor belts which are regulated with p.i.v. speed changers to control the belt weight. A weightometer determines the tonnage milled. The feed enters the 8' x 12' rod mill through a rubber lined chute and the product is discharged through a launder to a 7' rake classifier in closed circuit with 10' by 6' hardinge ball mill. Lime slurry is added at the rake classifier to give a strength of .30#/ton. The classifier sand is fed into the ball mill with a scoop feeder and the mill discharge flows in a launder back to the classifier. The classifier overflow is pumped to a 24' by 12' bowl classifier in closed circuit with a 8' by 15' regrind mill. The bowl classifier sands are fed to this mill and the mill discharge is pumped back to the bowl. The bowl overflow at 8% - 200 mesh is the final product from the grinding circuit.

Jan 1, 1975

By Rudolf Kvapil

INTRODUCTION Sublevel mining is a mass mining method based upon the utilization of gravity flow of the blasted ore and the caved overlying waste rock mass. As with any other mining method, sublevel caving has advantages and dis¬advantages which must be carefully considered and evaluated. The major advantages of sublevel caving are dis¬cussed as follows: Because all of the mining activities are executed in or from relatively small openings, sublevel caving is one of the safest mining methods. Drifts, which are the pri¬mary working places, are distributed in a uniform pat¬tern on all levels. Normally the maximum dimensions of the sublevel drifts are about 5 m wide and 3.7 m high. The transportation drifts can have the same section, or the height may be increased to about 4.5 m when trucks are loaded in the transport drifts. The stability and safety of such drifts in competent rock can be easily controlled by smooth blasting or by a combination of smooth blasting with shotcreting. In less competent rock masses, stability can be achieved by combined reinforc¬ing, for example, by a combination of smooth blasting, shotcreting, and rockbolting. The major mining activities can be broken down into three groups: drifting and reinforcing; ore fragmenta¬tion, i.e., production drilling and blasting; and ore draw¬ing, loading, and transportation, and all are relatively simple. Because of the repetitive nature of the mining system, one can standardize almost completely all min¬ing activities. This means that a high degree of work efficiency can be achieved. Because the components of mining production in sublevel caving can be standardized, a high degree of mechanization is possible. In modern sublevel caving the sections of drifts and tunnels are sufficiently large to allow the introduction of large trackless mining equip¬ment. The advantages of a trackless system can be then broadly utilized not only for direct mining but also for all services, including the transportation of mining per¬sonnel to the working place. The flexibility of mining is very good. Standardiza¬tion and specialization of mining activities and equip¬ment on separate levels (lower level or levels in de¬velopment, upper level or levels in production mining) together with the trackless system yield a high degree of flexibility. This allows a rapid start-up of mining and good flexibility in making production rate changes. The method lends itself to good work concentration, organization, and working conditions. Normally, on the lower levels, various phases of development are under¬way. Upper levels are in various stages of extraction. Therefore the work can be easily organized into a sys¬tem which excludes interference between mining activi¬ties. Safety of mining (in small dimension openings), good work organization, high mechanization using large modern mining equipment, etc., comprise very good working conditions. Naturally such a system enables a high work concentration and rationalization of separate specialized mining activities and therefore mining by sublevel caving can be effective and relatively in¬expensive. The major disadvantages of sublevel caving, on the other hand, are: There is a relatively high dilution of the ore by caved waste. Various types of ore loss can occur. When the ex¬traction limit (that point yielding the maximum accept¬able amount of dilution) is reached, the remaining highly diluted ore represents an ore loss. Some ore is lost in passive zones located on the level of extraction between the active zones of the gravity flow. Part of the ore from these passive zones can be recovered together with ore extraction on the lower sublevel, but some un¬diluted and often not fragmented ore located in passive zones above the plane of the footwall is lost. In gen¬eral, these losses are larger as the inclination of the ore body and the footwall is reduced. A relatively large amount of development is re¬quired. This includes transport drifts, usually located in the footwall waste rock on each sublevel, and sub¬level drifts, which connect the active mining areas to the transport drifts and as a result are partially in ore and partially in the waste rock of the footwall. The waste rock length increases as the inclination of the ore body and footwall decreases. It also includes orepasses, used for transport of the ore or waste from the separate sublevels downward to the main haulage level, and normally driven in waste; and inclined drifts or tunnels, which provide a connection for the trackless equipment between the main haulage level and the separate sublevels and are driven in waste. Finally there is the de¬struction of the surface through subsidence. To maximize the ore recovery, minimize the dilu¬tion, and achieve a high efficiency of mining by sub¬level caving, good data regarding the gravity flow pa¬rameters for the blasted ore and the caved waste are of utmost importance. The exact type and amount of data required depend upon the purpose and needs of the study. For the first feasibility study, it may be sufficient to utilize the data from other sublevel caving operations with similar conditions and circumstances. For any higher level of mine planning it is clear that more exact data, including analytical and experimental analyses up to full-scale in-situ testing, are necessary. Basic gravity flow principles and design guidelines for the application of the sublevel caving mining method are presented in the following sections. Although some¬what simplified, they should provide a basis for mine planning and operation. The gravity flow principles described can be effectively applied to other mining situations, with some modification. Also, steep dipping coal seams can be effectively mined by modified sub¬level caving.

Jan 1, 1982

By George B. Rice

In a recent speech in Pittsburgh, Dr. George Keyworth, the President's Science Advisor, made a statement which I believe deserves our very careful consideration. Dr. Keyworth said that there is no energy crisis. The crisis, he explained, is simply that people refuse to accept the solution. The solution which Dr. Keyworth has in mind is increased utilization of our abundant supplies of solid fuels and, in particular, uranium. I share his view concerning the solution to our energy needs. The use of uranium fuel is a safe, clean, and dependable means to generate our electric power. It is time that we addressed the real energy crisis: the refusal to accept the nuclear solution. The reason for the refusal is not difficult to find. It is nihilistic thinking about risk. Under this thinking, we assume the worst possible case and act accordingly, simply because we cannot prove to a total certainty that nuclear energy is perfectly safe. If this absolutist approach were generally applied throughout our society, there is no doubt all of us would soon be sitting around our campfires fearfully holding the wild animals at bay with our trusty spears. Today I am here to enlist your support in reversing the regulatory trend that threatens the very extistence of the nuclear power industry. As distinguished scientists, engineers and businessmen, you can use your influence to help bring rational regulation to the industry. Our industry supports strong safety and environmental protection programs. We understand the need for and do not object to reasonable regulation. Many anti-pollution measures can be practical to implement, cost effective and highly successful in minimizing environmental impacts. However, it is a fact of life that in the field of health and safety regulation, the law of diminishing returns operates with a vengeance. Absolute or near-absolute safety is impossible and any attempt to achieve it is intolerably costly. Fixation on absolute safety is particularly acute in the regulation of the nuclear power industry. Government Agencies, overly anxious to allay the irrational fears of those opposed to nuclear power, are literally regulating the industry to death - exactly the result sought by the anti-nuclear groups. Dr. Robert L. DuPont wrote in a recent issue of [Business Week]: "The nuclear power industry has been virtually stopped in the U.S. [because of fear]. This is true despite the fact that for more than 20 years the commercial nuclear industry has operated under unprecedented public health scrutiny and that to date there have been no radiation-related injuries, let alone deaths, suffered by any member of the public."1 I believe a useful way to convey the nature of the problem faced by the nuclear industry is to review an example of [unreasonable] regulation. While the example relates to our domestic industry, I am certain there are similar situations in other countries. For the example I will use the Nuclear Regulatory Commission's recently issued regulations governing the stabilization of uranium mill tailings.2 These regulations, known as the Uranium Mill Licensing Requirements, specify, among other things, that radon emanation from uranium mill tailings be limited to no more than 2 pCi/m2-sec. First, one must understand that this standard will have virtually no impact on the total amount of radon to which the public is exposed. Radon emitted from even completely unstabilized tailings piles is a tiny fraction--much less than 1%--of the amount of radon released from natural soils in the United States.3 In fact, it is far outweighed by natural variations in the background flux. For example, changes in the level of the Great Salt Lake in recent years have had [eight times] as much effect on the amount of radon released into the Salt Lake City regional air than the annual release from the Vitro Mill tailings pile located in that city.4 Nevertheless, NRC claims that the standard is required to protect the public. The Commission admits, however, that there are no studies which establish that exposure to radon at the low levels associated with uranium mill tailings will result in any adverse health effects.5 In the absence of actual evidence, the Commission assumes that some such effects will occur on the basis of the linear, non-threshold model.6 Employing this model, NRC calculates that the maximum hypothetical risk for the average member of the population is only about 1 in 70,000,000 from the radon that would be emitted from [three times] the number of mills now in existence, even if the tailings produced through the year 2000 are left unstabilized.7 NRC has elsewhere explained that this level of risk would be equivalent to the risk posed by "a few puffs on a cigarette, a few sips of wine, driving the family car about 6 blocks, flying about 2 miles, canoeing for 3 seconds, or being a man age 60 for 11 seconds." This level of risk is [de minimis] in comparison to other risks commonly and readily incurred in our society.9 Moreover, even this remote risk is overstated. A group of prominent health physicists, including experts from the Department of Energy, The Environmental Protection Agency, Britain, Canada and Germany recently published a study indicating that the risk to the public per unit exposure to radon can be no greater than one-third that suggested by the Commission, and [may in fact be zero].l0 Regulators routinely rationalize the need for their regulations. For example, NRC attempts to justify the radon flux standard because it is necessary to reduce the risk to someone who builds a house on top of a tailings pile. This possibility, however, is totally unrealistic because the Mill Tailings Act requires that stabilized tailings be transferred to

Jan 1, 1981

By Michael Robert Hodson

I. Introduction The free market cash price of aluminium in the third quarter of 1985 is approximately $1000/tonne, well below the short as well as the long run marginal costs of production. The current disparity between marginal costs and price, as well as the pattern of prices since their high level in 1980, has led many economists to question whether prices are governed by the 'fundamentals'. Prices, some would argue, are the result of speculation; and speculation is irrational, little influenced by the fundamentals of the industry. The object of this paper is first to examine the relation between the fundamentals and price. A number of simple models are formulated and some empirical evidence is given to show that prices are not very different from costs. There are good reasons why costs and prices have fallen since 1980. Second, the connection is drawn between the fundamentals and speculation. To do this the concept of 'rational expectations' is defined and applied to a model of the aluminium industry. Then speculation is introduced into the model and the characteristics of the solution are examined. The object is to discover the new factors that determine prices and the basic change in the nature of the solution that is caused by speculation. 2. Prices and the Simple Supply-Demand Model Consider a simple model of the aluminium industry. Demand, D, is determined by economic activity, A, given exogenously, and inversely by the price, P. Supply, Q, is determined by the inverse of the rnarginal cost curve, and so varies inversely with exogenous factor costs, C, and directly with price. Price is determined by the equality of supply and demand. Given the equilibrium level of demand, price in this model is equal to the rnarginal cost of producing the required level of demand, P = MC (Dl. Algebraically, [ ] How well does this simple model explain the change in prices during the last five years? Table I shows the LME cash price and the estimated marginal cost of producing consumption during the second quarters of 1980 through 1985. The figures in the table show that price and the marginal cost of supplying demand both fell between 1980 and 1982, rose in 1983 and by 1985 had fallen again. Only in 1984 is the direction of change different. And during the 1980 to 1985 period, marginal costs fell by approximately $560/tonne while the LME cash price fell by $678/tonne. There is a correspondence between the fundamentals of cost and price over most of the individual years and the period as a whole.

Jan 1, 1986