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By D. G. Hulbert

"Mineral processes are complex operations that have been modelled, controlled and simulated by diverse technologies. A review is given of some models of mineral processes in the literature. A new variation of a dynamic process model of flotation is presented. Process models, describing real and relevant physical variables, give a good understanding and good qualitative simulations, but can sometimes require prohibitively difficult and extensive parameter estimations. Process control is usually best done on operational plants by the fitting of system models, which can home in on the plant?s observability, controllability and optimisation. System models can be of a form amenable for use in appropriate techniques for controller design and implementation. The control of milling circuits is discussed in some detail. While simulators use models to predict dynamic or steady-state information, part of their strength lies in their ease of use and their ability to test different types of specifications. Some results are given for a flotation circuit simulator that can be made to solve virtually any permutation of specifications.INTRODUCTIONThe development and application of appropriate technology can lead to better operation of mineral processing plants in industry, so that substantial benefits can be obtained from improved throughputs, efficiencies and product qualities. This paper addresses some of the principles and methods used in the quest for the development and exploitation of such technology. Although the design of processes is often the motivation for modelling and simulation, this paper also targets the use of modelling and simulation for the control and optimum operation of mineral processes. The two processes given the most attention here are flotation and milling.Technologies used in mineral processing range from simple to very complex. Simple technical principles, such as mass and component balances, can yield some simple and effective strategies for modelling and control. However, minerals in particulate form have such complex physical and chemical properties that there appears no end to the complexity that emerges, as more and more factors are taken into consideration in the pursuit of better results and understanding. The situation is aggravated by the complexity of mineral processing plants, which have spatially variable conditions within process units, and complicated interactions between these units."

Jan 1, 2003

By Yushuai Xian, Youjun Tao, Fangyuan Ma

This study compared particle size distribution and density distribution of crushed products for waste printed circuit boards (WPCBs) before and after heat treatment. The crushing process for heat-treated WPCBs is a selective crushing process; more metal is concentrated in the coarse particle range of 1.0–0.5 mm and high density range of 2.8–2.96 g/cm3. WPCBs were separated using the Falcon separator to recycle useful metals. The Box-Behnken module of Design-Expert was used to obtain optimal separation parameters, namely, rotational frequency of 25 Hz, recoil water pressure of 0.025 Mpa and feed speed of 34 ml/s. On this basis, the separation indexes for WPCBs with and without heat treatment were compared. Experimental results show that the recovery effect is better for heat-treatedWPCBs than for thosewithout heat treatment. Comparedwith the non-heattreatedWPCBs, the product grade of heat-treated WPCBs is 1.95% higher, and its recovery is improved by about 14%. Finally, characteristics ofWPCBs before and after heat treatment were analyzed on the basis of stress experiments and scanning electron microscopy (SEM), respectively. Results of mechanical characterization indicate that the stress experiment for non-heat-treated WPCBs has a maximumpressure load of 212 N and displacement of 20mm, but heat-treatedWPCBs only need 140 N stress and 14 mm displacement. SEM shows that copper foil and fiberglass cloth are deformed in different directions and sizes after heat treatment to form gaps, which promote liberation of copper foil and fiberglass cloth.

Sep 10, 2020

By J. B. Zadra, H. J. Heinen

Electrowinning molybdenum from oxide feed materials dissolved in a variety of molten baths was investigated . A NaCl - NaF - Na₂ B₁ O -Mo0, system operating at 1,000 ° C in a cell without a protective atmosphere gave the best current efficiency , metal - to - cathode adherence , and purity of electrowon Technical - grade molybdic oxide , containing associated impurities of aluminum , iron , copper , silicon , and tungsten in combined quantities of 10 percent , was used as feed material to produce dendritic crystals of molybdenum metal containing less than 0.5 percent total base metal , carbon , and oxygen impurities . Arc - melting molybdenum electrowon from either chemically pure or commercial - grade Mo0, produced ingots that contained less than 0.1 percent total impurities and had hardness readings of 168 to 200 Vhn .

Jan 1, 1964

By McMillan W. D.

Vermiculite is the name of a group of micaceous minerals characterized by their property of exfoliating when heated and thereby increasing many times their original volume. The resulting product has excellent heat-and sound-insulating properties, is used as a lightweight aggregate, and has numerous other industrial applications. Deposits of vermiculite occur in areas of pre-Cambrian rocks of the Llano uplift of central Texas, particularly in Burnet and Llano Counties. Because of the growing economic importance of vermiculite products, the Bureau of Mines undertook a detailed investigation of several deposits in Llano County. This investigation was completed in February 1947. The work comprised topographic surveys of five deposits, drilling with a 4-1/4 inch Parmanco power auger drill, and laboratory evaluation of samples. 342 drill holes, aggregating 4,905.5 feet, were completed, and 377 samples were taken. Laboratory evaluation of the vermiculite comprised expansion tests on both crude and beneficiated samples. The products obtained were compared with commercial-grade vermiculite from Montana.

Jan 1, 1949

By G Hill, D Grant, M Tyler

Located approximately 570km NNW of Adelaide in South Australia, the Olympic Dam mine (OD) is the world’s fourth largest remaining copper and gold resource and the largest uranium resource. Discovered in 1975 and going into production in 1988, the operation consists of a fully vertically integrated underground mine, mineral processing plant, copper smelter and refinery, as well as associated infrastructure. BHP achieved production of 203,000 tonnes of copper cathode plus associated by-products in financial year 2016. The immense size and high-grades of the Olympic Dam resource offer a range of growth pathways, making closure timing estimation and planning a challenge. Current estimates put the OD resource at approximately 79 million tonnes of copper. Major factors influencing the closure timing estimation, as a result of the large resource and grade tonnage profile, include the mining method that may be used and the downstream processing method, ultimately influencing such things as the total disturbance area and the extent of infrastructure to be demolished at the end of mine life. In addition, factors such as the mining and processing method and production rate have a direct impact on larger rehabilitation items such as the total number of tailings cells required.CITATION:Grant, D, Tyler, M and Hill, G, 2018. Olympic Dam mine closure planning process, in Proceedings Life-of-Mine 2018, pp 43–45 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 25, 2018

By Robert G. Clarke

The production of natural abrasives varied in quantity and value compared with those of 1974. Output of tripoli-type materials and garnet both decreased in quantity and value. Special silica stone products decreased in quantity but in-creased in value. The production of emery increased 38% in quantity. The production of manufactured abrasive materials decreased for all types and the decreases were more in quantity than in value. [ ] FOREIGN TRADE Imports of abrasive materials were 15% less in value than in 1974, and exports plus reexports decreased in value 10%. Net exports, the excess of exports and reexports over imports, were $9.3 million, about four times that of 1974. The volume of all abrasive materials exported decreased and the value of nearly all decreased.

Jan 1, 1977

By Joel K. Waterland

GENERAL DESCRIPTION Open cut-and-fill stoping for many years was prob¬ably the most widely used mining method in under¬ground metal mines. Then for a time this method was largely supplanted by the blasthole stope. It again be¬came popular as many mines reached depths or grades where methods requiring large open voids to remain open for extended periods of time became unsuccessful, often as a result of excessive dilution. The open cut-and-fill method is very flexible and is readily adaptable to almost any ore body. The standard application requires that a slice of ore usually 2.4 to 3 m (8 to 10 ft) thick be removed from the back of the stope, and as the ore is taken down, the back is dressed and rockbolted. After the back is secured, the broken rock is removed through rock passes to the level below. When the rock has been removed, the rock passes are extended upward the height of the ore removed, the stope is backfilled, and another cycle is mined. This method is best employed in plunging ore bod¬ies with considerable vertical extent, ore areas that re¬quire selective mining, ore areas where weak wall con¬ditions exist, and ore bodies that have an ore value that will carry this relatively expensive mining method. Blast¬hole stoping, shrinkage stoping, and other mining meth¬ods that do not employ rock passes in a stope are not efficient in plunging or flatly dipping ore bodies because the footwall makes ore removal quite difficult. Since mining is accomplished by taking down slices of the back, only small areas of the wall rock are exposed at any one time, and these only for short periods. Due to limited back height, areas of uneconomic rock may be left in place, or they may be mined and the material gobbed in the stope. Because the miners in the stope must work under freshly blasted areas, the amount of ground control is usually great. The volume of rock that is broken during one section of mining is relatively small and the amount of nonproductive work required is high. This results in limited productivity for the scope and, be¬cause of the nonproductive work that must be done on a regular basis, the production from the stope may be quite cyclical. SUITABLE ORE BODIES The open cut-and-fill method may be adapted to al¬most any type of ore body with a relatively high vertical extent. The ore body must be accessible at both top and bottom as well as at regular intervals throughout its vertical extent. Although adaptable to most ore bodies, the method is probably best employed where the ore has poor con¬tinuity and where most types of bulk mining would pro¬duce excessive dilution. In areas of poor ore continuity, the capability of continuous and extensive sampling dur¬ing the mining of each cycle makes this method very desirable. This capability also minimizes the amount of evaluation sampling that must be done before mining is started. Because of the extractive system used, the size and shape of the stope may be as readily changed as the sampling mandates. Probably the only ore characteristic demanded is that the ore has strength enough to be sup¬ported through the use of rockbolts or cable bolts dur¬ing the mining and backfilling cycles. Good planning, systematic sampling, and careful supervision will pro¬duce a product with less dilution than any other open stoping method. PLANNING Evaluation Planning Once it has been decided that the open cut-and-fill method would be the most efficient for mining a par¬ticular ore body, the next considerations would probably be the availability of an economical backfill material and the selection of an efficient transport system for this material. Although hydraulically transported mill tail¬ings are the most widely used product, this is not always practical due to mill location or the quality of the tailings. In such cases, backfilling may be used. The type of backfill and the type of equipment used will determine if a floor or capping on the backfill is required to minimize dilution during ore removal. The early selection of rock removal equipment is im¬portant since equipment usually determines the amount of development work required to bring a stope into pro¬duction and the size of the openings needed. The size and continuity of the ore body will usually determine the type of loading equipment. The use of slushers or load¬haul-dump (LHD) equipment captive in the stopes will minimize the amount of development. If the ore con¬tinuity is such that a ramp system for extraction can be used, the cost of development will be increased but the flexibility of continuous mining will minimize the cycli¬cal nature of the production. The height of the mining section usually is deter¬mined by the strength of the wall rock and the amount of back bolting required. Once this has been decided, the appropriate drilling equipment can be chosen. The number and sizes of the rock passes employed depends upon the type and size of the extractive equipment and the type of backfill that is to be used. Since the miners must enter and leave the stope each shift, the level inter¬val is usually maintained at approximately 45 m (150 ft). Access from the level above into the stope must be main¬tained at all times. The employees perform all the work in the stope and adequate ventilation must be provided. Stope Planning Due to the flexibility of the method and the vari¬ability of the ore zones, layout is usually done on a stope basis. In areas where continuity is a problem, the size of the stope is usually determined by the boundary of the ore (with all of the ore within that boundary being removed). In areas of good continuity where ramps are to be used, the length of the stope may be determined by the length of time each of the cycles (preparation, back¬fill, mining, and ore extraction) requires. The ramp work is then laid out so that access to the various parts

Jan 1, 1982

By R. R. Graham

Echo Bay Minerals operates a 650 tpd gravity float gold circuit at their Manhattan facility. The objective of this paper was to document the gold recovery by size distribution. This was accomplished by studying the material balance of each size fraction as it progressed through each gravity Circuit separator. It proved to be an impossible task. Figure I depicts the gravity circuit. The gravity cyclone underflow goes to two duplex Pan American type jigs. The jig overflow reports to the OK Flash Float Cell. The cyclone hutch product is reground and goes to a full size wilfly table. The rougher table concentrate reports to a pilot size Wi1fly cleaner table. The cleaner and rougher table tails circulate back to the jig feed. The gravity circuit was composite sampled for two to four hours per day for four days. Each sample was taken every 20 minutes. Each composite sample was filtered, dried, repu1ped, and screened through a wet/dry screening procedure. Each size fraction was checked for gold value by fire assay fussion technique. A material balance was run on each size fraction. The tables listing the results are enclosed.

Jan 1, 1989

Mark R. Lamb, Torres, Sonora, Mexico (communication to the Secretary*): I have read Mr. Merrill's paper with much interest, and cannot but accept his challenge as regards the cost of the clean-up at the Homestake mill. If I understand him correctly, the two men take the precipitates from the presses in about six hours, and do nothing towards refining them. If that is so, it would correspond to the work four men do here in about eight hours—that is, to clean the box and to place the precipitates in the driers. In order to make a fair comparison, the bullion in each case should be stated in ounces, since it is as diflicult to precipitate silver as gold, and our bullion is mostly silver. Bullion to the value of $50,000 is equivalent to a weight of about 2,500 oz. At our last clean-up eight standard bars, each 1,100 oz. in weight, were made, which equals a total weight of 8,800 oz'. Four men had the precipitates drying by 8 p.m. This plant, the Minas Prietas works of Charles Butters & Co., at Minas Prietas, has square-bottomed electrolytic vats, an arrangement which necessitates the handling of all the precipitates in buckets. The Butters plant at Virginia City, Nev., is equipped with conical vats, and probably could make a still better showing under similar conditions. Another point is that common labor only is required. (Ordinary laborer's wages at Minas Prietas are $0.70 gold.) I may say also that, except for the convenience of the refinery, the box could be cleaned monthly, instead of bi-monthly, at only a slight additional expense. The comparison, if made a few months ago, would have been much more in favor of the electrolytic process, as the tailings handled at that time contained a large quantity of copper, and the amount of zinc-dust required would have been enormous.

Jan 1, 1904

By Yi Yang, CHENGLONG XU, HAORAN LI, YALI FENG, JINRONG JU

Steel slag is an industrial byproduct of steelmaking that is mainly composed of calcium, iron, silicon and magnesium. In addition, depending on the different raw materials entering the furnace, steel slag contains a certain amount of vanadium that is the main source of vanadium production, accounting for 69 percent of the species of the total vanadium raw material [1]. From experience, the production of one ton of steel will generate 1.0 to 1.5 tons of steel slag [2]. However, the utilization efficiency of steel slag is less than 30 percent, and a large amount of steel slag is identified as waste and stacked in tailings. The random stacking of steel slag not only wastes large amounts of limited steel and land resources but also exerts pressure on the environment [3], making it imperative to research the effective utilization of steel slag and recycling of metal resources.

By Almon F. Robertson

"INTRODUCTION At various times, North Carolina has supplied small quantities of brown-iron ore to the furnace o the South. Between 1917 and 1920, the quantity reported to the Federal Geological Survey was 126,000 tons. Most of this came from Cherokee County. 3/In furtherance of the program initiated by the Bureau of Mines to develop additional reserve of iron ore during the emergency, the Cherokee County deposits were examined by an engineer 4/ of the Bureau of Mines, and work was recommended on,12 properties.The project was begun in April 1944 and continued intermittently until April 1945, when work was stopped. During this period, eight properties were explored, and an additional property was mapped.ACKNOWLEDGMENTSIn its program of exploration of mineral deposits, the Bureau of Mines has as its primary objective the more effective utilization of our mineral resources to the end that they make the greatest possible contribution to national security and economy. It is the policy of the Bureau to publish the facts developed by each exploratory project as soon as practicable after its conclusion. The Mining Branch; Lowell B. Moon, chief, conducts preliminary examinations, performs the actual exploratory work, and prepares the final report. The Metallurgical Branch, R. G. Knickerbocker chief, analyzes samples and performs beneficiation tests. Both these branches are under the supervision of Dr. R. S. Dean, assistant director."

Nov 1, 1946

By S. R. Zimmerley, C. H. Schack

"INTRODUCTION The pilot-mill concentrator at Boulder City, which has a nominal capacity of 40 tons per day, was planned and built to investigate further the beneficiation of domestic manganese ores that have been concentrated successfully in the ore-dressing laboratory. As milling of manganese ores in general is a new field, it is necessary to prove the success of laboratory processes by means of large-scale operations employing standard milling equipment, and thus to obtain engineering and metallurgical data essential to designing and operation of full-scale concentrating mills.The first ore tested in the pilot plant was the Three Kids Annex manganese ore from the Las Vegas Wash area 15 miles southeast of Las Vegas, Nev. The deposit was worked for its high-grade ores during the First World War From 1930 to 1939, 4,500 feet of diamond drilling was completed by a private company. Diamond drilling started by the Bureau of Mines in November 1940 and completed in July 1941 aggregated 6,600 feet. Exploration by the private company and the Bureau outlined reserves of about 1,000,000 tons of ore assaying plus 28 percent manganese. A much larger tonnage of lower-grade ore also was outlined by the drilling.Because of its physical nature, the Three Kids Annex ore had long proved unamenable to concentration by ore-dressing methods. The ore is a mixture of manganese oxides (wad) filling the pores and to some extent replacing a silt composed of very fine grains of volcanic tuff, quartz, feldspars, other silicates, and clay. Some gypsum occurs as fillings along fracture planes. The ore is light and very porous. Upon wetting, it disintegrates into granular but friable particles and a large amount of slime. It is these characteristics that make the ore so difficult to concentrate."

Sep 1, 1947

By M. R. Geer

The Bureau of Mines surveyed the solid wastes that have accumulated from coal mining in Washington, Oregon, and Montana. Location, amount, character, and condition were determined, and both economic and environmental factors were considered. Areas covered by waste comprise 800 acres in Washington, 5 acres in Oregon, and 1,200 acres in Montana. In a few cases waste is being utilized, particularly cinder from burned refuse dumps, as fill in road construction and as a ceramic raw material. It is estimated that wastes deriving from coal mined for electric power generation by 1974 may amount to 3 million tons of washery refuse, 2.7 million tons of fly ash, and 1,000 acres of spoil banks.

Jan 1, 1969

By P. Tong-jiang

We studied mineralogical characteristics of the vermiculite samples from seven mines in China. We found three types of interstratified structures in the investigated samples: 1:1 regular interstratified structure, segregated interstratified structure, and random interstratified structure. The vermiculite samples from six mines are mainly composed of phlogopite?vermiculite interstratified minerals, and the samples from the other mine are composed of chlorite?vermiculite interstratified mineral. The samples with phlogopite?vermiculite structure have high contents of K2O, TiO2 and Na2O. The samples with chlorite?vermiculite structure have the highest contents of MgO and Al2O3. The cation exchange capacity of the samples is 17.3?99.0 mmol/100g.The exchangeable cations of phlogopite?vermiculite samples from Weili Mine mainly include Na+ and Ca2+, while those of phlogopite?vermiculite samples from Tongguan Mine mainly include Mg2+ and Ca2+. The thermal expansion ratio of the samples is between 9 and 32. The exfoliated vermiculite products from Weili Mine have the best mechanical strength.

Feb 27, 2013

By L. E. Djingheuzian

"IntroductionIN 1940, in a series of articles, M. W. von Bernewitz(l) reviewed in detail roasting of gold and silver sulphide ores and concentrates. Since that time some significant developments in roasting have taken place, and a number of papers have appeared in the Canadian technical literature, dealing severally with the roasting of a particular ore or concentrate.The purpose of this paper is not only to review some of these Canadian papers and bring the information on Canadian roasting practices to .date, but also to analyse the mechanism of roasting. Hence, it is proposed to start with a study of physico-cl1emical factors fundamental in roasting, the knowledge of which, by creating a sound basis for a proper roaster design, makes good roasting possible.Physico-Chemical FactorsRoasting can be defined as heating sulphide ores in presence of air below fusion points of the constituent minerals, at such temperatures as to convert metallic sulphides to oxides. Hence, roasting is a chemical process.The object of roasting is, by changing, chemically, mineral constituents of sulphide ores, to make the economic recovery of the desired •Constituent metal by a subsequent standard treatment possible. For instance, pyritic ores, containing gold in sub-microscopic form which no amount of grinding would release, are roasted in order, by oxidizing the sulphides, to open up their structures to such an extent as to enable cyanide solutions to come into contact with the gold and dissolve it. The production of porous calcine is, therefore, the object of roasting of refractory auriferous sulphide ores or concentrates. Such a porous calcine may be defined as a desirable calcine, and the roasts which produce desirable calcines can in their turn be termed desirable roasts."

Jan 1, 1952

By John C. Taylor

A resource document, prepared for the Royal Commission on the Health and Safety of Workers in Mines, which outlines aspects of occupational health and safety in the Canadian metallurgical industry has been expanded to provide examples of some practical solutions. General process descriptions are provided in conjunction with some of the major potential hazards associated with each step. Emissions to the working environment, particularly around converters, are discussed. Newer metallurgical processes, the role of automation and other advances in the metallurgical industry are described with respect to their impact on the work place. The role of staff and unit employees in providing safe working conditions in the complex metallurgical plant of today is outlined and the need for close cooperation within the human sector emphasized. The technology is available and is being applied to integrate the complex process requirements with realistic regulations to control emissions and provide a safe, healthy environment in metallurgical plants.

Jan 1, 1978

By Ken Hester, Bruce Arthur

"Introduction""Pebble milling"" is a term used to describe secondary grinding done with rock pebbles rather than steel balls or other steel or metal shapes. This paper discusses pebble milling with particular attention being paid to uranium grinding circuits.Pebble milling is typically employed:a. where the cost of steel balls becomes exorbitant, as may be the case in an isolated area.b. where the extra steel added from steel ball consumption may be detrimental to further processing.c. where the over-all economics favour the use of pebbles rather than ball.Pebble milling differs from autogenous and semi-autogenous grinding in that a sized product has been removed from some part of the primary comminution circuit, and it is this sized product which is added as grinding media to the pebble milling stage. The pebbles do not necessarily have to be produced from the ore- ""foreign"" pebbles can be used. Also, pebble milling generally refers to grinding from a fine size, below about 6 mesh (3360 micron), with relatively small media, that is, secondary grinding. In South Africa, however, primary pebble mills are sometimes used to grind crusher house product, and thus function in place of rod mills. Pebbles for this service are large and heavy as they must weigh as much as balls used in similar service.The first large applications of pebble milling were in South Africa(n. ""Tube milling"" with pebbles started in 1904. The tube mill was simply a tumbling mill with a length to diameter ratio over two. A ""typical"" early tube mill was 5.5 ft. (1.7 m) in diameter and 16 ft. (4.9 m) in length. Diameters increased steadily. King(n mentions 8-ft. (2.4 m) units in his 1949 book. By 1961, the largest units discussed by Jackson('' were 12ft. (3. 7 m) in diameter."

Jan 1, 1989

By E. D. Fox

Introduction The financial statement of the hypothetical Gold Mining Company, which forms the background of this survey, was designed for the Institute with the express abject of having it assessed in respect to Profits by the several taxing bodies representing the Federal and Provincial Governments, which comprise the Dominion of Canada. On the completion of the figures, copies were sent to the several Ministers of Mines, accompanied by a request that they be submitted to their taxing authorities and assessed accordingly. The response, while it was not general, was very favourable, and at all times we were treated with the utmost courtesy and patience, and what has been set forth in the compilation bears the authentication of the various government officials who are immediately concerned with taxation. The data itself constitutes a fairly comprehensive digest of taxation as it directly applies to gold mining profits in Canada. The Hypothetical Balance-Sheet (See Exhibit A) In the balance-sheet schedules 1 and 2 (Exhibit A), income is divided between taxable and non-taxable classifications, and here it should be noted that the division contemplates only the assessment for Dominion income tax. The same applies to the factors of exchange and depletion. Therefore, in making comparisons, it is to be borne in mind that the set-up for assessment to the Provinces, that is, Manitoba and Alberta, will stand as if the gross non-operating income is taxable, with the reservations as shown in the notes below the schedules.

Jan 1, 1934

By D. Bienstock

Using a nitrided fused-iron catalyst, the Bureau of Mines investigated the Fischer-Tropsch reaction in the oil-circulation process. The nitrided iron was quite stable and permitted the longest run obtained thus far in the oil-circulation process without the need for reactivating the catalyst--3,700 hours of operation. The effects of several operating variables--temperature, pressure, gas-recycle ratio, space velocity, synthesis-gas composition, and synthesis with and without reflux of oil to the reactor--on the production of oxygenates were studied. The importance of removing reflux oil and of high gas recycle in increasing the oxygenate content was shown. A finished gasoline containing 28 percent by weight alcohols was produced. The research octane number (RON) was 92.9 clear, increasing to 98.5 with the addition of one milliliter of tetraethyl lead (TEL). INTRODUCTION The Federal Bureau of Mines has been engaged since 1944 in a research and development program on the synthesis of gasoline from coal. Gasification of coal with steam and oxygen to yield a "synthesis gas"--a mixture of carbon monoxide and hydrogen--is being investigated at Morgantown, W. Va., and the synthesis of gasoline by catalytically reacting carbon monoxide and hydrogen over iron by the Fischer-Tropsch reaction is being investigated at Bruceton, Pa.

Jan 1, 1960

By Richard D. Hagni

Microscopic examination of selected, representative, unweathered samples of Birmingham, Alabama, Silurian, sedimentary iron ores has fully revealed the nature of the phosphorous-bearing mineral grains. The grains are composed of carbonate-fluorapatite (collophane), range in size from 1 um to 2.25 mm, and exhibit varied types of intergrowth with hematite. Modal analyses conducted under cathodolumi-nescence have provided quantitative data on the a3undance of each of 15 types of phosphorous-bearing mineral grains and made it possible to assess the potential phosphorous liberation from four types of iron ore. The results indicate that 27% of the total phosphorous contained in the iron ores in the western part of the Birmingham district can be expected to be liberated by fine grinding; 60% for ores in the eastern part of the district.

Jan 1, 1982