Search Documents

The following brief paper is intended, to apply to the concentration of the silver-lead ores at Broken Hill, New South Wales, generally, but more particularly to the process as applied to the ores in the Block 10 mine. The attached plans are similar inmany respects to those prepared, recently for the Block 10 company.Most of the information here given is already well known to those members who are conversant with the concentration practices of this field, but it is the writer's object to give just a few of those figures which he found difficult to obtain when first engaging in practical work. At that time no text-books were at hand from which it was possible to get sufficient details for practical working purposes.It is first supposed that the mine has been thoroughly explored and has been found to contain a lbde sufficient in size ·and quality to warrant the erection of a concentrating plant. The points then to determine are:-(1.) The quantity and cost of water available for steam-making and ore-dressing purposes.(2.) The most convenient site for the mill.

Jan 1, 1904

By C. Sawmliana, P. Pal Roy

Near-field blast vibration response of restrained pipelines either buried or exposed was studied during rock excavations at sensitive locations for Stage-II (4x500 MW) commissioning work in a super thermal powerplant of the National Thermal Power Corporation Ltd. (NTPC) of India. The existing running plant of capacity 1000 MW has many important installations including pipelines of varying diameters. These pipelines were very close to the blasting locations (mostly within 12m). Blast optimization and consequent monitoring of vibrations were carried out simultaneously for more than 250 experimental blasts conducted by the authors. This paper emphasized, particularly, on the safety of pipelines and their responses due to repeated blasting impacts. A few guidelines supported by theoretical derivations are provided for futuristic research.

Jan 1, 2001

Interest in mining narrow, vein-type ones has increased world-wide during the last few years despite the fact that prices of metals have not risen in proportion as production costs of mining, such as labour and energy. In many mines, where narrow ones are mined, the mining technique and equipment used have been poorly mechanised. The mining methods have been cut-and-fill or shrinkage stoping. Pneumatic jack-legs have been used for drilling and slushers have been used for loading. The methods that have been used can be called the 'first-generation' methods. Work has been very hard for miners. It has been dangerous and exhausting. Production volumes of mines, and a s a result of that, also profits have been low.

Jan 1, 1990

Reduce the potential for injuries in underground mines by providing an alternative to manually lifting and holding heavy roof-support beams. The Problem I-beams, rail sections, and wooden cross-beams are often used in mine haulageways where extensive roof support is needed. Studies by the Bureau of Mines have shown that many mines must resort to manually lifting, moving and installing these beams. Such manual-handling activities have the potential for causing injury to mine workers, particularly back injuries. I-beams used for roof support can typically weigh from 400-500 lbs. Rail sections weigh about the same as the I-beams. Wooden beams can weigh from 200 to 300 lbs. To lift a single beam usually requires the efforts of 3 to 5 mine personnel.

Jan 1, 1988

By J. L. Fuentes-Cantillana, Á. Rodríguez, P. Morillo

Evacuation operations from work places in the case of an emergency are the key to minimising the effects of catastrophic type accidents such as fires, explosions, cave-ins, floods, etc. The correct planning of these operations, having the necessary means for sounding alarms and communicating and suitable training of personnel are the key factors for guaranteeing successful evacuation. The article gives an analysis of the problem in the case of subterranean mining and various types of work developed by AITEMIN in this field, including the development of new alarm and communication means specially adapted to the characteristics of mining operations, development of tools for help in the preparation of emergency plans and experiences in the education and training of personnel.

Jan 1, 2003

By David Harrison, Edward J. Walter, Meitra Ferek

Agreat tool in the prevention of blasting complaints and subsequent damage claims. In today’s highly competitive industry, we are always looking for ways to become more cost effective. This article investigates the importance of pre-blast surveys-not just for regulatory compliance, but as an overall long term cost cutting and public relations tool.

Jan 1, 1995

By P J. Minehan

The investigation of the potential of the Webb/Baynes block of the Buller coalfield for anode carbon manufacture has generated much information on the physical beneficiation characteristics of the Buller coals in general. A further 36 drill holes of the block in question and several other samples have been analysed by sink/float analysis, either at a single density or over a range of densities. In addition, many samples have been beneficiated on a laboratory shaking table or by flotation. The ash content can be reduced by a variable factor. The reduction factor of between two and four by sink/float analysts is understandably less for tabling where separation is not perfect. All Buller coals behave similarly, there being greater ash reduction factors for higher ash coals.

Jan 1, 1983

By Frederick F. Frick

SPONGE iron as produced at Anaconda is a fine, -35 mesh, impure product, about 50 pct metallic iron, obtained from the reduction of iron calcine at a temperature of 1850°F by use of coke resulting from slack coal. The metallic iron particles are bulky and spongey and precipitate copper readily and rapidly from a copper sulphate solution. Investigation of the treatment of Greater Butte Project, Kelley, ore at Anaconda early showed the desirability of using sponge iron as a precipitant for the copper in solution resulting from desliming of the ore in a dilute sulphuric acid solution. Anaconda had done considerable work on the production of sponge iron in 1914 for use as a precipitant of copper from leach solutions. Some success and considerable experilence were attained at the time. indicating that, sponge iron might be successfully made by a modification of the process used in 1914, a batch process in which an iron calcine was reduced by means of soft coke, resulting from noncoking coal, in a Bruckner-type revolving horizontal cylindrical furnace widely used 50 years ago. The coke and calcide formed the bed in the Bruckner furnace, which was rotated at about 1 rpm. The bed was brought to a temperature of about 1800°F by means of an oil flame over the surface. Although results were reasonably satisfactory, they did not warrant full development of the process at that time. A good deal of work has been done in the last 50 years on the production of sponge iron. The objective in some cases has been the production of a precipitant for copper from solution, but the bulk of the work has been done for the production of open-hearth steel furnace stock. The production of an open-hearth stock presents two problems rather than one: first, producticon of the sponge iron, and second, what is perhaps of equal difficulty and importance, conversion of the sponge iron into a form suitable for use in the open-hearth furnace. So far as is known to the writer, none of the sponge iron processes tried in the past have proved to be economically feasible. However, Anaconda had a combination of conditions appearing to justify an attempt to produce sponge iron which would serve for the leach-precipitation-float process. It was thought that the process used in 1914, if changed to a continuous one, might work out satisfactorily. The following favorable conditions at Anaconda justified the investigation: 1—A sufficient tonnage of good grade iron calcine resulting from the roasting of a pyrite concentrate in one of the acid plants, at substantially no cost. 2—Reasonably cheap natural gas. 3-—The fact that there was no need for production of a high grade product. 4— The fact that there was no need for obtaining a consistently high reduction of' the iron in calcine. A small revolving Bruckner-type furnace about 2 ft ID by 4 ft long was set up for early pilot work at the research building. This pilot furnace showed that a satisfactory product could be obtained at reasonable cost. It also indicated a marked advantage in preceding the reduction furnace with a furnace of similar size and capacity for preheating and roasting out any residual sulphur from the feed. The small furnace was operated for several months, various details of the process were worked out. and sponge iron was produced to supply a pilot LPF plant which treated 300 lb of Kelley ore pel- hr. Later a second pilot furnace 5 ft in diam and 12 ft long inside was set up at our reverberatory furnace building. This furnace confirmed the data of the small furnace and gave a basis for design of the final plant. At Anaconda a pyrite concentrate, running about 48 pct S, is recovered from copper concentrator tailings by flotation. This concentrate is roasted to sulphur of 3 pct or less at the Chamber acid plant. The iron calcine contains about 57 pct Fe and 18 pct insoluble. The iron calcine feed, as mentioned before, is re-roasted and preheated in a reroast furnace preceding the reduction furnace. Both are of the Bruckner type. The reroasted calcine is fed into the reduction furnace at 800" to 1000°F along with 30 pct slack coal. In the feed end of the furnace the volatile is burned from the slack, giving a soft coke which readily serves for reduction of the iron. Hard metallurgical coke will not serve the purpose. since it does not reduce CO readily at a temperature of 1850°F. All indications are that the actual reduction of the iron is accomplished by carbon monoxide below the surface of the bed, which is 30 in. deep at its center. Apparently there is a constant interchange: Fe²O³ + 3CO = 2Fe - 3CO², CO² + C = 2CO Actually iron oxide is reduced by CO at somewhat lower temperature than the 1850 °F used in the process. but this temperature is necessary to obtain a satisfactory rate of furnace production. The furnace atmosphere is generally reducing, and typical blue carbon monoxide flames satisfactorily cover the bed. Gas flames from four 3-in. Denver Fire Clay Inspirator burners are played directly on the bed, which is slowly cascaded by the 1 rpm of the furnace. An excess of coke is necessary to assure maintenance of good reducing conditions in the furnace bed. Part of this coke is recovered for re-use.

Jan 1, 1954

By William P. Jones

THE consumption of sulphuric acid, one of the most important commodities in our modern industrial world, is often used as a barometer for industrial activity. The economics of acid manufacture are largely dependent upon the location of the place of consumption and the availability of raw materials in that locality. Sulphuric acid is made from SO,, oxygen from the air and water. Therefore the sulphur dioxide is the only raw material to be considered in an economic study. SO, can be obtained from almost any material containing inorganic sulphur, such as elemental sulphur, pyrites, coal, sour gas and oil, metallurgical gases, waste gases, or gypsum and anhydrite. Many tons of acid can also be reclaimed by the recovery and concentration of spent acids. The aim of this paper is to present a guide to the economic aspects to be considered when the installation of an acid plant is contemplated. It must be remembered that 1 ton of elemental sulphur produces 3 tons of sulphuric acid and that the shipping of sulphuric acid by tank car is very costly. The size of the plant must also be given careful consideration. For instance, operation of a plant producing 5 tons of acid per day might be warranted in Brazil or Pakistan, whereas economics usually favor buying quantities up to 50 tons per day for use within the United States. Elemental sulphur, when available at the low price of 1M4 per lb delivered at an acid plant, has always been the raw material most frequently used for sulphuric acid. All conditions favor its use at this price. The so-called sulphur shortage has been the subject of so many technical papers, magazine articles, and newspaper items during the past y6ar that it hardly seems necessary to mention it again, but a very brief review of the matter will serve as a foundation for the discussion that follows. There is no shortage of sulphur. Only a shortage of low-cost Frasch-mined brimstone exists today. Other more expensive sulphur-bearing materials are plentiful, both in the United States and abroad. The low cost of Frasch-mined brimstone has discouraged the development of higher cost sources. However, the approaching depletion of Gulf Coast dome deposits and the greatly increased demand for sulphur here and abroad have made it necessary for industry to prepare for conversion to utilize sulphur in other forms. For future planning this situation must be considered permanent and not temporary. This conclusion is based on the fact that although sulphur demand will continue to rise, the production of Frasch-mined sulphur probably will not increase greatly beyond its present level of about 5,000,000 long tons per year. The International Materials Conference in Washington estimates 1952 requirements of the free world at nearly 7 million long tons; and the Defense Production Administration has recently set a new goal for 8,400,000 long tons annual domestic production by 1955. The total sulphur equivalent produced in this country in 1950 was 6 million tons. What, then, are the alternatives for the manufacture of the vital chemical, sulphuric acid? Today about 85 pct of this country's sulphur, and nearly 50 pct of the world supply, comes from our Gulf Coast salt domes and is extracted from the earth by Frasch's hot water process. The Gulf Coast salt dome deposits have been the most important known natural deposits in the world, producing 90 million tons of sulphur during the past 50 years. However, at the present rate of extraction these deposits cannot be expected to last indefinitely. Pyrites Pyrites are, and have been for many years, the source of more than 50 pct of the world's sulphur requirements. The principal use, of course, is in the manufacture of sulphuric acid. The use of pyrites in the United States has diminished greatly because of the availability of low cost native sulphur, but pyrites have continued a major source of sulphur in many other countries. The most available pyrites for use in this country are in the form of lump pyritic ore and in mill tailings from flotation of other minerals such as lead, zinc, copper, gold, and silver. An important factor, when the use of pyrites for acid manufacture is being considered, is the disposal of calcine. A ton of sulphuric acid requires approximately ton of high-grade pyrite and results in 1/2 ton of calcine. If the calcine is a fairly pure oxide, free of harmful impurities, it can be used, after sintering, in an iron blast furnace burden. Its value might be as high as 15d per unit of Fe at the blast furnace; or possibly $10.00 per ton of sinter, if it assays 65 pct Fe. This might result in a credit of $4.00 per ton of acid if the sintering plant and blast furnace are both located adjacent to the acid plant. On the other hand, several factors must be considered before this credit can be realized, i.e., freight to blast furnace, availability of sintering facilities, methods of eliminating impurities, and the removal of valuable metal values. In some locations it would be most economical to dump the calcines.

Jan 1, 1953

By William P. Jones

THE consumption of sulphuric acid, one of the most important commodities in our modern industrial world, is often used as a barometer for industrial activity. The economics of acid manufacture are largely dependent upon the location of the place of consumption and the availability of raw materials in that locality. Sulphuric acid is made from SO2 oxygen from the air and water. Therefore the sulphur dioxide is the only raw material to be considered in an economic study. SO2 can be obtained from almost any material containing inorganic sulphur, such as elemental sulphur, pyrites, coal, sour gas and oil, metallurgical gases, waste gases, or gypsum and anhydrite. Many tons of acid can also be reclaimed by the recovery and concentration of spent acids. The aim of this paper is to present a guide to the economic aspects to be considered when the installation of an acid plant is contemplated. It must be remembered that 1 ton of elemental sulphur produces 3 tons of sulphuric acid and that the shipping of sulphuric acid by tank car is very costly. The size of the plant must also be given careful consideration. For instance, operation of a plant producing 5 tons of acid per day might be warranted in Brazil or Pakistan, whereas economics usually favor buying quantities up to 50 tons per day for use within the United States. Elemental sulphur, when available at the low price of 1 ½ ¢ per lb delivered at an acid plant, has always been the raw material most frequently used for sulphuric acid. All conditions favor its use at this price. The so-called sulphur shortage has been the subject of so many technical papers, magazine articles, and newspaper items during the past year that it hardly seems necessary to mention it again, but a very brief review of the matter will serve as a foundation for the discussion that follows. There is no shortage of sulphur. Only a shortage of low-cost Frasch-mined brimstone exists today. Other more expensive sulphur-bearing materials are plentiful, both in the United States and abroad. The low cost of Frasch-mined brimstone has discouraged the development of higher cost sources. However, the approaching depletion of Gulf Coast dome deposits and the greatly increased demand for sulphur here and abroad have made it necessary for industry to prepare for conversion to utilize sulphur in other forms. For future planning this situation must be considered permanent and not temporary. This conclusion is based on the fact that although sulphur demand will continue to rise, the production of Frasch-mined sulphur probably will not increase greatly beyond its present level of about 5,000,000 long tons per year. The International Materials Conference in Washington estimates 1952 requirements of the free world at nearly 7 ½ million long tons; and the Defense Production Administration has recently set a new goal for 8,400,000 long tons annual domestic production by 1955. The total sulphur equivalent produced in this country in 1950 was 6 million tons. What, then, are the alternatives for the manufacture of the vital chemical, sulphuric acid? Today about 85 pct of this country's sulphur, and nearly 50 pct of the world supply, comes from our Gulf Coast salt domes and is extracted from the earth by Frasch's hot water process. The Gulf Coast salt dome deposits have been the most important known natural deposits in the world, producing 90 million tons of sulphur during the past 50 years. However, at the present rate of extraction these deposits cannot be expected to last indefinitely. Pyrites Pyrites are, and have been for many years, the source of more than 50 pct of the world's sulphur requirements. The principal use, of course, is in the manufacture of sulphuric acid. The use of pyrites in the United States has diminished greatly because of the availability of low cost native sulphur, but pyrites have continued a major source of sulphur in many other countries. The most available pyrites for use in this country are in the form of lump pyritic ore and in mill tailings from flotation of other minerals such as lead, zinc, copper, gold, and silver. An important factor, when the use of pyrites for acid manufacture is being considered, is the disposal of calcine. A ton of sulphuric acid requires approximately ¾ ton of high-grade pyrite and results in ½ ton of calcine. If the calcine is a fairly pure oxide, free of harmful impurities, it can be used, after sintering, in an iron blast furnace burden. Its value might be as high as 15¢ per unit of Fe at the blast furnace; or possibly $10.00 per ton of sinter, if it assays 65 pct Fe. This might result in a credit of $4.00 per ton of acid if the sintering plant and blast furnace are both located adjacent to the acid plant. On the other hand, several factors must be considered before this credit can be realized, i.e., freight to blast furnace, availability of sintering facilities, methods of eliminating impurities, and the removal of valuable metal values. In some locations it would be most economical to dump the calcines.

Jan 1, 1952

By William P. Jones

THE consumption of sulphuric acid, one of the most important commodities in our modern industrial world, is often used as a barometer for industrial activity. The economics of acid manufacture are largely dependent upon the location of the place of consumption and the availability of raw materials in that locality. Sulphuric acid is made from SO,, oxygen from the air and water. Therefore the sulphur dioxide is the only raw material to be considered in an economic study. SO, can be obtained from almost any material containing inorganic sulphur, such as elemental sulphur, pyrites, coal, sour gas and oil, metallurgical gases, waste gases, or gypsum and anhydrite. Many tons of acid can also be reclaimed by the recovery and concentration of spent acids. The aim of this paper is to present a guide to the economic aspects to be considered when the installation of an acid plant is contemplated. It must be remembered that 1 ton of elemental sulphur produces 3 tons of sulphuric acid and that the shipping of sulphuric acid by tank car is very costly. The size of the plant must also be given careful consideration. For instance, operation of a plant producing 5 tons of acid per day might be warranted in Brazil or Pakistan, whereas economics usually favor buying quantities up to 50 tons per day for use within the United States. Elemental sulphur, when available at the low price of 1M4 per lb delivered at an acid plant, has always been the raw material most frequently used for sulphuric acid. All conditions favor its use at this price. The so-called sulphur shortage has been the subject of so many technical papers, magazine articles, and newspaper items during the past y6ar that it hardly seems necessary to mention it again, but a very brief review of the matter will serve as a foundation for the discussion that follows. There is no shortage of sulphur. Only a shortage of low-cost Frasch-mined brimstone exists today. Other more expensive sulphur-bearing materials are plentiful, both in the United States and abroad. The low cost of Frasch-mined brimstone has discouraged the development of higher cost sources. However, the approaching depletion of Gulf Coast dome deposits and the greatly increased demand for sulphur here and abroad have made it necessary for industry to prepare for conversion to utilize sulphur in other forms. For future planning this situation must be considered permanent and not temporary. This conclusion is based on the fact that although sulphur demand will continue to rise, the production of Frasch-mined sulphur probably will not increase greatly beyond its present level of about 5,000,000 long tons per year. The International Materials Conference in Washington estimates 1952 requirements of the free world at nearly 7 million long tons; and the Defense Production Administration has recently set a new goal for 8,400,000 long tons annual domestic production by 1955. The total sulphur equivalent produced in this country in 1950 was 6 million tons. What, then, are the alternatives for the manufacture of the vital chemical, sulphuric acid? Today about 85 pct of this country's sulphur, and nearly 50 pct of the world supply, comes from our Gulf Coast salt domes and is extracted from the earth by Frasch's hot water process. The Gulf Coast salt dome deposits have been the most important known natural deposits in the world, producing 90 million tons of sulphur during the past 50 years. However, at the present rate of extraction these deposits cannot be expected to last indefinitely. Pyrites Pyrites are, and have been for many years, the source of more than 50 pct of the world's sulphur requirements. The principal use, of course, is in the manufacture of sulphuric acid. The use of pyrites in the United States has diminished greatly because of the availability of low cost native sulphur, but pyrites have continued a major source of sulphur in many other countries. The most available pyrites for use in this country are in the form of lump pyritic ore and in mill tailings from flotation of other minerals such as lead, zinc, copper, gold, and silver. An important factor, when the use of pyrites for acid manufacture is being considered, is the disposal of calcine. A ton of sulphuric acid requires approximately ton of high-grade pyrite and results in 1/2 ton of calcine. If the calcine is a fairly pure oxide, free of harmful impurities, it can be used, after sintering, in an iron blast furnace burden. Its value might be as high as 15d per unit of Fe at the blast furnace; or possibly $10.00 per ton of sinter, if it assays 65 pct Fe. This might result in a credit of $4.00 per ton of acid if the sintering plant and blast furnace are both located adjacent to the acid plant. On the other hand, several factors must be considered before this credit can be realized, i.e., freight to blast furnace, availability of sintering facilities, methods of eliminating impurities, and the removal of valuable metal values. In some locations it would be most economical to dump the calcines.

Jan 1, 1953

By K. S. Benson

DEPLETION is a subject of vital importance to the mining industry. Yet, in spite of its importance, its significance is not generally understood. The purpose of this discussion is to clarify the main aspects of the subject from the viewpoint of a metal mine taxpayer. To define the term depletion, it is necessary to distinguish among its various uses. In the economic or geological sense, depletion means the exhaustion of a natural resource. A mineral deposit is a wasting asset and once exhausted is nonrenewable. Millions of years were needed to produce an ore deposit, which may be consumed in a few years and which cannot be replaced except by the discovery of new sources of supply. The wasting asset feature of the mining industry has a vital bearing on the impact of the Federal Income Tax Law on this industry. This is recognized in the law by the various provisions dealing with the depletion allowance, and in this sense the term depletion has an income tax meaning. Depletion from the tax viewpoint means the statutory deduction from gross income designed to permit the return to the taxpayer of the capital consumed in the production and sale of a natural resource. The mining enterprise realizes income on the extraction and sale of minerals and a portion of the income realized represents capital consumed. As the resource is exhausted, the mining enterprise approaches the end of its existence unless new sources of supply can be acquired. Depletion from the tax viewpoint is a creature of statute with limited meaning and application and, in essence, is a method for amortizing the value of the primary asset of a mining enterprise. An example can best illustrate the significance of depletion from the tax viewpoint. Compare a manufacturing concern with a mining company. In computing taxable income of a manufacturing concern, consideraion is given to the cost of producing such income, the principal costs being capital investment for plant and equipment, labor, and raw materials going into the products produced. A mining enterprise, on the other hand, is faced with a different problem because its principal asset is the natural resource which it is producing. In computing its taxable income, consideration is given also to its capital investment for plant and equipment and the cost of labor; but in addition, recognition must be given to the fact that a portion of the proceeds realized on the sale of mineral represents capital. Without such recognition, the mining company would be taxed not on income but on capital and income, and Congress has never intended that capital be taxed as income. Thus, when depletion allowable is referred to in the mining industry, it means the statutory deduction allowable in computing taxable income of a mining enterprise. For guidance the appropriate provisions of the Internal Revenue Code, Income Tax Regulations, and the judicial decisions interpreting and construing them must be examined. It is important to identify and distinguish three methods of determining the allowance for depletion: 1—Cost depletion, 2—Discovery depletion, and 3—Percentage depletion. The basic method is cost depletion and in addition some taxpayers may be entitled to use discovery depletion and other taxpayers may be entitled to use percentage depletion. Discovery depletion and percentage depletion, however, are mutually exclusive and if a taxpayer is entitled to percentage depletion, he is not entitled to discovery depletion. By statute, a metal mine taxpayer is entitled to use cost depletion or percentage depletion, whichever produces the highest deduction. Thus, discovery depletion is merely of academic interest to such taxpayers and to most others. Briefly and broadly speaking, these methods of determining depletion may be described as follows: 1—Cost Depletion: Under this method, the allowable deduction for depletion is based upon the cost of the particular deposit to the taxpayer, unless the deposit was owned prior to Mar. 1, 1913, in which case the taxpayer may use the fair market value of the deposit on that date or actual cost, whichever is higher. This method is sometimes described as basis depletion or adjusted basis depletion, but in this discussion it will be referred to as cost depletion. 2—Discovery Depletion: Under this method, the allowable deduction for depletion is based on the fair market value of the deposit at the date of discovery or within 30 days thereafter and was originally designed to take into account deposits discovered subsequent to Feb. 28, 1913. 3—Percentage Depletion: Under this method, the allowable deduction for depletion is based on a specified percentage of the income realized during the taxable year from a particular property. As stated, the concept of depletion is based upon the exhaustion of a natural resource as distinguished, for example, from the concept of depreciation based on the exhaustion of property used in trade or business. From the tax viewpoint, depletion first became important in the administration of the Corporation Tax Act of 1909, which provided for an excise tax on net income. As soon as this act went into effect, mining taxpayers attempted to claim a deduction for depletion in computing net income although there was no specific mention of a deduction for depletion in the statute. The courts in these cases uniformly held that the statute did not permit an allowance for depletion in computing net income and also held that the provision permitting a reasonable allowance for depreciation did not include depletion. These early cases are quite significant because they establish the principle that the

Jan 1, 1952

By Bernard L. Cohen

I.INTRODUCTION The B E I R-III Report (NAS-1980) introduces large increases in the estimated health effects of radon as compared with previous work (NAS-1972). It is the purpose of this paper to point out some important failures of these new BEIR-III estimates, to offer a general critique of the procedures used in obtaining them, and to offer more rational estimates. In Sec. II we use the BEIR-III model to calculate the risk to non-smokers from environmental radon, and show that it predicts more than twice the total lung cancer rate actually experienced by nonsmokers. In Sec. III we review the histological evidence which shows that no more than about 10% of the lung cancers among non-smokers can be due to radiation. In Sec. IV, we discuss alternative causes of lung cancer, which further reduces the fraction that can be caused by radiation, and in Sec. V we summarize and conclude that the BEIR-III model over-estimates the lung cancer rate in nonsmokers due to environmental radon by at least a factor of 40. In Sec. VT we review the evidence on risk of radon exposure to smokers, and conclude that it is probably not more than four times the risk to non-smokers; this means that the BEIR-III model over-estimates the risk of low level radon exposure to smokers by at least a factor of 10. In Sec. VII, we consider the reasons for the large over-estimates in the BEIR-III. Report. II. BEIR-III LUNG CANCER RATES DUE TO ENVIRONMENTAL RADON AND COMPARISON WITH TOTAL LUNG CANCER RATES AMONG NON-SMOKERS The BEIR-III Report gives the following estimates of the lung cancer risk from low-level radon exposure in terms of working-level-months (WLM): age 35-49, risk = 10 x 10-6 /yr-WLM 50-64, 20 >65, 50 where ages refer to age at death. For latent periods between exposure and onset of these risks it gives age 0-14, latent period = 25 years 15-34, 15-20 years (we use 17 yr) >35, 10 years where ages refer to age at exposure. This is a clear and unambiguous model which is readily usable for deriving numerical estimates. We begin by using it to calculate lung cancer rates due to environmental radon. *This is an abridged version of a paper scheduled to appear shortly in Health Physics. The first step in this process is to estimate the environmental exposures; this was done in a recent paper (Cohen-1981) which concluded that these are about 0.22 WLM/year. In Table 1, this is used to calculate the BEIR-111 predictions for radoninduced lung cancer rates in the U.S. (Col. (5)), and by combining these with population statistics, it is shown (Col.(7)), that it predicts about 24,500 fatalities per year, almost one-third of all U.S. lung cancers. The comparison between the age-specific expected rates from Col. (5) of Table 1 and observed rates among non-smokers is shown in Table 2. The recent paper by Garfinkel (1980) presents the results of a 12 year follow-up on one million Americans in a study by the American Cancer Society. The paper by Hammond (1966) gave the results of the first four years of that study. The paper by Kahn (1966) is based on the so-called "Dorn Study" of 293,000 U.S. veterans of World War II who carry government health insurance. It represents 8 and 1/2 years of follow-up. A recent update on that study (Rogot-1980) does not give absolute lung cancer rates, but the age-standardized ratio between smokers and non-smokers has remained the same which indicates that there has probably not been an important change in the rates for either. The paper by Hammond and Horn (Ha-1958) was an early study by American Cancer Society. It is immediately evident from Table 2 that the BEIR-III estimates for lung cancer induced by environmental radon exceed the [total] lung cancer rates due to [all] causes among non-smokers by about a factor of two at every age. It is only fair to point out that this does not represent a direct discrepancy with the BEIR-III Report since the latter states that its estimates for non-smokers may be too high by a factor ranging from 1 to 6, favoring a factor intermediate between these. Comparisons can also be made with total lung cancer incidence for all ages. A paper by Hammond and Seidman (Hammond-1980) gives the rate for ages above 40 to be 177 x 10-6/year for men and 124 x 10-6/year for women, whereas the rate calculated in Table 1 from BEIR-III for ages above 40 is 309 x 10-6, a factor of two higher. For all ages, the rate among women was reported as 36 x 10-6/year (Hammond 1958) as compared with 114 x 10-6/year calculated from BEIR-III in Table 1, a discrepancy of well over a factor of two. All of the data we have presented are basically from three study groups, but in all three cases the BEIR-III estimates for lung cancer induced [by environmental radon alone] are a factor of two higher than actual [tota] lung cancer rates among non-smokers. Another approach to comparing the BEIR-III pre-

Jan 1, 1981

By N. O. Gray, John Henderson

Hydrogen-reduction data for naturally occurring single crystals and Prepared polycrystals of dense hematite have been presented. Results cover the temperature range 400o to 1000oC, for particles from ten sources, ranging in size from 0.07 to 10 mm and in shape from spheres to cylinders, cubes, and thin slabs. A consistent pattern of behavior has been demonstrated for single crystals and the reduction mechanism shown to be temperature-dependent. Below 579oC reduction is a simple topochemical process but at higher temperatuves it is complex and occurs in two distinct stages. Prepared particles from these laboratories behave in a similar manner to the single crystals. Data from two investigators showing topochernical reduction of Prepared particles above 575°C are inconsistent with that for other dense hematites. It is concluded that topochenzical reaction should not be used as a model for generalized rate expressions for dense hematites. SINCE 1958, McKewan1-6 has brought to prominence a simple concept of dense hematite reduction. This model is that oxygen is lost from a hematite particle undergoing reduction only from an oxide-iron interface that recedes in such a way that the oxide remaining retains the original shape of the particle, i.e., reduction occurs topochemically. An adjunct to this concept is that any intermediate oxides in the transition from hematite to iron only form thin layers so that oxygen cannot be lost from the particle without movement of the oxide-iron interface. Further, the rate of oxygen loss from the particle is said to be proportional to the area of the receding interface so that the iron layer grows linearly with time and the over-all reduction process can be described by the equation where ro and do are the initial particle radius and density, respectively, R is the fraction of the original oxygen lost, i.e., the fractional reduction, at time t, and K is the rate constant. The idea of an underlying simplicity in hematite reduction is attractive because it gives a tractable basis from which general theories of hematite reduction can be developed and it has received wide support,7-14 based mainly on the large amount of data13 that can be fitted to Eq. [I]. However, despite the fact that this equation has been derived for dense materials, the bulk of the data that have been used to test it13 have been for materials of only about 90 pct of theoretical density (5.26 g cm-3) or less, so that its generality for dense hematites has not been demonstrated. In any case, as will be seen, adherence of reduction data to Eq. [1] does not necessarily imply that reaction occurs topochemically. In this work only data for hematites approaching theoretical density are considered and it will be shown that in only one study besides McKewan's is topochemical behavior observed over the whole range of temperature investigated. For the majority of materials a linear rate of interface advance is observed to complete reduction only when wustite is not a stable intermediate phase in the transition of hematite to iron, i.e., at temperatures below about 575°C. Above this temperature, reduction is an exceedingly complex series of reactions that takes place in two distinct stages and it is only in the first stage that reaction in any way resembles a topochemical process. This means then that, far from representing general behavior as has commonly been supposed, topochemical reaction for dense hematites is only a particular behavior that may be observed under some circumstances. EXPERIMENTAL Hydrogen-reduction data have been collected and cross-checked in these laboratories by three techniques, weight loss, collection of water evolved in the reduction reactions, and direct metallographic examination. Details of these techniques are discussed elsewhere," where it is shown that results obtained by the three methods are in close agreement. The weight-loss method, by which most of the results were obtained, consisted of hanging the sample in a platinum mesh basket from an Ainsworth Model AV-AU-1 vacuum recording balance inside a vertical l 1/4-in.-ID alumina tube furnace. Dry deoxidized hydrogen was flowed downwards through the tube at 2 to 4 liters min-1 (stp). Single crystals from two sources and artificial oxides from three sources have been examined. The single crystals were from hematite deposits at Yampi Sound, Western Australia, and Brazil, the latter being obtained from Gregory, Bottley and Co., London, U.K. The "Yampi Blue" crystals were hand-picked from washed, magnetically concentrated, sized fractions, -200 mesh BSS (mean diameter ca 0.07 mm) and —36 + 44 mesh BSS (mean diameter ca 0.4 mm) while 10-mm cubes were cut from the Brazilian crystals with a diamond saw. The starting materials for the prepared particles were hematites designated, respectively, "Specpure Iron Oxide", Laboratory No. S639 from Johnson Matthey & Co. Ltd., London, U.K., "Calcined Ferric Oxide" from B.D.H. Laboratory Chemicals Division, Poole, U.K., and "Pigment Grade Oxide", EPR-50, from C. K. Williams & Co., Easton, Pa., U.S.A. Approximately spherical particles were prepared from the artificial oxides by rolling the material, moistened if necessary, in a glass jar, and cylindrical compacts approximately 10 mm in diam and of approximately equal height were pressed in a steel die at 200 psi. These spheres and cylinders were subsequently fired in oxygen for 20 to 100 hr at 1370°C. The

Jan 1, 1967

By D. S. Flett, D. W. West

The solvent extraction of chromium 111 has been studied for the system Cr 111, H,SO., H,O/RNH/RNH., xylene, where the primary amine used was Primene JMT. Rate studies have shown that extremely long equilibrium times are required, ranging from 1 hr at 80°C to 20 days at room temperature. Heating the solution prior to extraction increases the rate of extraction. The variation in the amount of Cr 111 extracted is an inverse function of the acidity of the aqueous phase. Thus, the slow rates of extraction appear to be connected with the hydrolysis of the Cr I11 species. Extraction isotherms for the extraction of Cr 111 have been obtained for two sets of experimental conditions, namely at 60°C and for a heat-treated solution cooled to room temperature. The separation of Fe 111 from Cr 111 and Cr 111 from Cu 11 in sulfate solution by extraction with Primene JMT has been studied and shown to be feasible. A survey of the literature relating to the solvent extraction of chromium showed that, although many systems exist for extraction of Cr VI, only a very few reagents have been found to extract Cr 111. The extraction of Cr III by di-(2-ethyl hexyl) phosphoric acid has been reported by Kimura.' A straight-line dependence of slope —2 was observed between log D,, and the log mineral acid concentration at constant extractant concentration. Since the slope of this plot reflects the charge on the ion extracted, it must be concluded that a hydrolyzed species of Cr III is being extracted. Carboxylic acids generally do not form extractable complexes with Cr III but di-isopropyl salicylic acie does extract Cr 111. Simple acid backwashing of the organic phase, however, failed to remove the chromium. Similar difficulty in backwashing was found by Hellwege and Schweitzer8 in the extraction of Cr I11 with acetyl-acetone in chloroform. The extraction of Cr 111 from chloride solutions by alkyl amines has been reported4-' but the maximum amount of extraction achieved in these studies did not exceed 10%: From sulfate solutions, however, Ishimori" has shown that appreciable amounts of Cr I11 were extracted by amines. The amines used were tri-iso-octyl amine, Amberlite LA-1 (a secondary amine, Rohm & Haas) and Primene JMT (primary amine, Rohm & Haas). The efficiency of extraction with regard to amine type was primary>secondary> tertiary. Appreciable extraction of Cr I11 was recorded for Primene JMT as the aqueous phase acidity tended to zero. The major difficulty with Cr I11 in solvent extraction systems stems from the nonlabile nature of the ion in complex formation. This accounts for the slow rate of extraction generally experienced and the difficulty encountered in backwashing the Cr I11 from the organic phase in the case of liquid cation exchangers. Consequently, the possibility of extraction of Cr I11 as a complex anion is attractive since the backwashing problems should be minimized in this way. From published data, it appeared that the extraction of chromium from sulfate solutions of low acidity by primary amines afforded the best chance of success for a useful solvent extraction system for Cr iii This paper presents the results of a study of the extraction of Cr I11 from sulfate solution by Primene JMT and examines the application of such an extraction procedure for the recovery of chromium from liquors containing iron and copper. Experimental Chromium solutions were prepared from chrome alum in sulfuric acid and sodium sulfate so as to maintain a constant concentration of sulfate ion of 1.5 molar. Solutions of Primene JMT were prepared in xylene and the amine equilibrated with sulfuric acid/sodium sul-fate solutions, of the same acidity as the chromium solution, until there was no change in acidity between the initial and final aqueous phases. The solutions of Primene JMT conditioned in this way were then used for the equilibration experiments. Equilibrations at 25°C were carried out in stoppered conical flasks shaken in a thermostat; equilibrations at all other temperatures were carried out in stirred flasks in a thermostat. After equilibration, the phases were separated and analyzed for chromium. In the tests on the rate of extraction, small samples of equal volume of both phases were withdrawn from time to time and the chromium distribution determined. The chromium analyses were carried out either coloi-imetrically using diphenyl carbazide, or volu-metrically using addition of excess standard ferrous ammonium sulfate and back titration of the excess iron with potassium dichromate. The oxidation of Cr 111 to Cr VI in the case of the raffinate solution was effected by boiling with potassium persulfate in the presence of silver nitrate and, for the backwash solution, by boiling with sodium hydroxide and hydrogen peroxide. Results Preliminary experiments indicated that extraction results were effected by the age of the chromium solution, higher distribution coefficients being obtained with solutions which had been allowed to stand for some time. Consequently a stock solution of chrome alum, 10 m moles per 1 Cr I11 in 1.4 M Na,SO,/O.l M &SO,,

Jan 1, 1971

By Huw Williams

It is becoming increasingly common practice to re-use existing foundations for new projects, especially on urban sites where geotechnical conditions are congested. A knowledge of the foundation dimensions and capacity is essential for this to take place, however, as-built records are often non-existent. Non-Destructive evaluation (NDE) is now becoming an important tool in assessing old foundation properties and their potential for re-use. Pile dimensions are also critical in the power industry, when they are supporting transmission towers which need to be upgraded, and may carry additional wind loading. This paper describes the application of various NDE techniques that are available and how they can be used to determine pile length, shape and capacity in difficult site conditions, including foundations under structures.

Jan 1, 2006

By J. G. Donaldson

Optimum conditions are given for chloridizing pyrite, marcasite, chalcopyrite, arsenopyrite, stibnite, and covellite with chlorine gas. The formation of metal chlorides is affected by the time and temperature of chloridizing, and by particle size of the sample. Chloridizing data also are given for chemically precipitated antimony (III) sulfide and copper (II) sulfide.

Jan 1, 1962

By E. F. Perry

Simulated weathering tests are used to estimate leachate quality of rocks containing varying concentrations of pyrite and carbonates at mines. This study reports the inter-laboratory results from the third phase of a three part effort to develop a standardized, reproducible simulated weathering test for coal mine overburden. Initial development in phases I and II included evaluation of different test apparatus, effects of particle size, gas composition, leaching frequency and other variables. Phases I and II results are reported by Brady et al (2004) and Hornberger et al (2004). In Phase III, a standard column leaching protocol was used by eight commercial, government and university labs to test five rocks of different lithology, pyrite and carbonate content. Ideally, a simulated weathering test should have three characteristics to be useful in mine drainage studies. These include the capability to produce repeatable results, accuracy in simulating field conditions, and a concise test protocol of weeks to a few months. Simulated weathering, or kinetic testing, produces an effluent that simulates mine drainage composition. This can be beneficial for predicting mine drainage quality in comparison to whole rock analyses. The effluent may be tested for the same water quality parameters as the drainage produced during mining, such as pH, acidity, alkalinity, sulfate, iron, manganese, aluminum or other parameters. If the physical, chemical, and biological conditions of the kinetic tests are representative of those found in the mine environment, leachate composition may be used to estimate the water quality parameter concentrations produced during mining. Hornberger and Brady (1998), and Geidel et al (2000), compiled comprehensive chapters on kinetic tests for mine drainage prediction. These sources included: (1) a chronology and synopsis of scientific literature on these kinetic tests as they have developed over approximately 50 years, (2) discussion of the influence of physical, chemical, and biological processes, and (3) general guidelines for test procedures, data interpretation, and recommendations for further research to develop standard methods. Most of the kinetic test methods in use today were substantially developed and applied more than 40 or 50 years ago, including leaching columns (Braley, 1949), humidity cells (Hanna and Brant, 1962), Soxhlet reactors (Pedro, 1961), and field scale tests (Glover and Kenyon, 1962). . The most commonly used kinetic tests for mine drainage prediction are leaching columns and humidity cells. While these kinetic test methods and others have been used in hundreds of mine drainage studies as shown in Hornberger and Brady (1998) and other references (e.g. Sorini, 1997), they have rarely been used in coal mine permitting by either regulatory agencies or the mining industry. The major impediment to routine use of kinetic test methods is the variation in the design and operation of these kinetic tests, and lack of a standardized, accepted (e.g. by EPA or ASTM) test method. Interpretation and data comparison from different studies can be difficult. METHODS Five rocks of varying lithology, mineralogy, carbonate and pyrite content were collected in bulk from Pennsylvanian age coal measures in Pennsylvania, West Virginia and Indiana. Mineralogical composition of the samples was characterized in detail by Hammerstrom et al (in press) using optical, X-ray diffraction, scanning electron microscopy and electron microprobe methods. Acid Base Accounting (Sobek et al 1978; Skousen et al, 1997, Kania, 1998) was performed on bulk samples for % total sulfur (%S) as a measure of pyrite and potential acid generation, and Neutralization Potential (NP) as an index of carbonate content. A modified NP test, following the procedure suggested by Skousen et al, 1997 was also conducted for four of the five samples. The modified NP test accounts for initial acid consumption, Fe oxidation, and subsequent Fe hydrolysis by the iron carbonate mineral siderite. Siderite, even though it is a carbonate gives zero net alkalinity. Lithologic, Acid Base Accounting and general mineral composition data for the five rocks are shown in table 1 (see Appendix). Whole rock elemental analysis for major and trace elements was conducted on a sub-sample of each rock type. One laboratory was contracted to mix, homogenize, and size the bulk samples and prepare sub-samples for all the participating labs. Splits were sent to eight labs, and each organization was provided a detailed set of instructions for column assembly and leaching procedure. The protocol was developed based on results from phases 1and 2 of this study. The principal features of the column assembly are: a specific composition of particle size by weight (table 2), weekly leaching cycles, and incubation under humidified air containing 10% CO2. Figure 1 shows the column construction. [ ] The particle size distribution of the crushed rock sample is largely an artifact of the crushing process, rather than a natural systems process like the particle size distribution of a soil or an unconsolidated sedimentary deposit. The standardized particle size distribution, shown in Table 2, promotes operational consistency of the weathering test procedure, and provides better control over reaction kinetics. Large amounts of fine particles within specific zones of the leaching columns have been found to impede uniform fluid flow and/or gas flow (Hornberger and Brady, 1998). Regarding reaction kinetics, the importance of surface area to volume ratios has been described in a previous phase of this study (Brady et al., 2004), and significant differences in crushed particle size distributions and effective surface areas were found among the lithologic units. Standardizing the particle

Jan 1, 2009

By T. Lenarcic, R. Svrkota, J. M. McNeal, E. Prohic, S. Pirc

Paper presented at the 28th International Geological Congress held in Washington DC, July 1989 (International geochemical mapping sessions). Surface drainage samples, the usual media employed in mapping national territories, cannot be used for the geochemical mapping of carbonate terrains because there is no regular surface-drainage pattern on karst. Suitable sampling media have been studied in Yugoslavia, one-third of which consists of karst terrain. Sampling designs for the work were based on analysis of variance techniques. During initial investigations of various sampling materials in NW Yugoslavia sediment and aquatic moss from karst springs, along with soil, were established as the most appropriate media. Modified sampling procedures were then carefully applied along the entire Yugoslav karst belt, which includes terrains of contrasting topography, climate, pedology and vegetation. The studies indicate that climate and relief do not systematically influence the geochemical composition of soil. Regional and local variability was assessed for 21 chemical elements and the feasibility of producing stable regional geochemical maps was determined for a number of elements on the basis of sampling aquatic moss and sediment from karst springs and soil

Jul 1, 1991

The paper considers Aboriginal responses to large-scale mining, including Indigenous attitudes towards land access for exploration and project development. The core question, in which I have been interested for some time (Trigger 1997a; 2000; 2001), is: what are the significant factors affecting responses among Indigenous communities to mining exploration and development project proposals?

Jan 1, 2002