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By G. Pahomova

"It developed an alternative to traditional flotation enrichment refractory of complex lead-zinc ores to obtain conditional lead and zinc products. The scheme involves the use of the processing method of gravitational enrichment of the primary cycle and the combined gravity-flotation-hydrometallurgical in finishing. The method has a high level of comprehensive utilization of mineral raw materials in relation to the elements that are extracted in the process of enrichment in lead concentrate, and the increase in recoverable value of processed ores. INTRODUCTION The primary source of lead are complex sulphide lead-zinc ore, lead ore actually rarer. The value of lead and zinc ores (except lead and zinc) is determined by the presence in them of a wide range of associated components: silver, contained mainly in galena; gold; bismuth in the form of its own minerals, but is associated with galena; cadmium contained mainly in the sphalerite; indium, thallium and gallium contained in sphalerite and galena; selenium and tellurium are present as impurities in the sulfide, tellurium is also found in the form of its own minerals. Of note harmful impurities antimony, which is often associated with lead sulphosalts and arsenic, having a substantially own minerals. By quality (content of lead, or the total content of lead and zinc) lead-zinc ores are divided into rich, ordinary and poor; by oxidation - to sulfide (with a lead oxides content less than 15% and zinc oxide content less than 10%), mixed and oxidized (50% oxide); the latter have a subordinate importance as a source of zinc and lead. Traditionally, the most contrasting technological properties of sulfide ores of nonferrous metals considered flotation properties, which determines the priority in the selection of flotation technology of their enrichment. Other methods of enrichment - gravity (jigging and enrichment in heavy suspensions), radiometric sorting (separation) are used as preliminary operations whose main purpose is to increase the content of lead and zinc in the ore sent to flotation, as well as the removing of part of empty rock in dumping tails prior to flotation. Effective implementation of these methods is possible at large size emphasizing in the ore of lead and zinc minerals, which is typical of the rich ores and which so far mainly worked. At the same time, non-ferrous metal ores (polymetallic, lead-zinc, copper-zinc, etc.) have a sufficiently high contrast to gravitational properties since ore minerals have a higher density compared with a rock minerals. But the gravity of the practice of non-ferrous ores enrichment is not widely used, despite its high degree of environmental safety and lower cost expensive."

Jan 1, 2018

By Marcus Vinicius Laudares Mendonça

SMS calculates and improves tooth and flank profiles individually to realise a homogeneous pressure distribution over the complete load flank. With a tooth profile and flank profile according to AGMA, ISO or DIN standard the contact pressure varies along the load flank. To realize a homogeneous distribution of the contact pressure and to avoid a premature mesh, SMS improves the tooth profile and flank profile in three steps. First SMS is doing a preliminary calculation and design according ISO, DIN or AGMA standard, with safety factors based on more than 125 years of experience. Second SMS calculates and checks the local contact pressure at the load flank under consideration of shaft, gearing and housing deformation, as well as the bearing behaviour. Finally SMS calculates and designs an improved tooth profile and flank profile to receive a homogeneous contact pressure at the load flank. The result is an increase of the torque transmitting capacity of up to 30% of the SMS advanced gears with the same outer dimensions than traditional gears.

Aug 17, 2017

By Shosei Serata

The Stress Control Method improves productivity and safety in underground coal mining. The method stabilizes the roofs and floors of mine openings in both shallow and deep coal beds, regardless of whether or not there are ground stability problems. This paper describes the means by which a tapered pillar experiment is used to collect in situ data, which is then used as input for a computer model. The resulting computer model is then used to optimize Stress Control design for the mine under study. The behavior of any underground coal mine is not elastic, as is generally conceived in the conventional view of rock mechanics, but rather is highly time-dependent and site-specific. Unfortunately, conventional rock mechanics methods based on the elastic principle are not able to capture the highly non-elastic nature of ground behavior which actually occurs. To capture actual behavior, a new method of study is required. In this regard, geomechanical study of an underground coal mine may be compared to studying the behavior of a living fish. Its true behavior zannot be understood by examining it "dead in the laboratory. It Is the "live" nature of the coal mine that is required for design optimization of the Stress Control Method in the given mine. The Stress Control Method is based on a time-dependent theory of the global behavior of the mine ground, which generally contradicts the conventional elastic theory of ground behavior. Therefore, site-specific adaptation of the Stress Control Method required a comprehensive understanding of the global behavior of the mine as a function of time and site-specific ground conditions.

Jan 1, 1986

By Hargraves A. J

With increasing depth, stresses and stress phenomena are becoming more important considerations in Australian coal mining.A large number of compressive and tensile strength tests was made on coal and rock cubes cut from samples taken in various Australian collieries. In addition, an empirical impact strength test was carriedout on fragments of the same samples, to obtain a correlation with compressive strength.In one colliery, flat-jack measurements were taken to assess stresses in coal ribs and in the roof.INTRODUCTIONBy world standards, some of the New South Wales collieries are approaching depths of cover considered a maximum for bord and pillar the normal Australian coal mining method. Several South Coast collieries are winning coal at depths of about 1500 ft., and future operations are being planned at deeper cover. Some areas of these deep workings produce considerable quantities of seam gas.Phenomena associated with high stress and copious gas under pressure in the seam violent phenoma presenting a danger and a nuisance in mine working - have occasionally occurred, and the phenomena are being investigated as a whole (Hargraves, 1958), and separately as a gas problem (Bolt and Innes, 1959; Hargraves, 1962) and as a stress problem. This paper relates to the latter - to investigations into the strengths of coals and coal measure rocks, and into the stresses in and about coal seams...

Jan 1, 1962

The original leases making up today's North Kalgurli Mines were developed by independent companies on an ad hoc basis. The major disadvantage which now arises is the varying levels of mining horizons. From Kalgurli Shaft in the south, through main Shaft and to Croesus Shaft in the north, there are three separate sets of levels. work has commenced to link the three areas with either common levels or new development which will facilitate common access and haulage via Main Shaft.Generally, the workings at North Kalgurli were established at the turn of the century. The introduction of techniques and equipment to bring the mine into the 20th Century has been difficult and slow. The general size of openings, a widespread workforce and complex ore distribution often make innovation and change more than a challenge.

Jan 1, 1985

By J. S. Martin

The redesign and reinforcement of a highwall located adjacent to a $10,000,000 shopping center, destabilized by active subsidence, was conducted after an extensive geotechnical investigation. The highwall was excavated in a hillside where a coal seam had been retreat mined and augered. Engineering geological mapping, determination of the properties and behavior of the intact rock, fractured rock, and urethane grout material, and soil were required to characterize the highwall and conduct a stability analysis. Remedial measures included re-excavating portions of the highwall to reduce the slope angle, and pressure grouting of the immediate roof.

Jan 1, 1988

By L. C. IlsLey, E. J. Gleim

It is the firm belief of the Bureau's engineers that the introduction of electrical mining machines has done much to increase safety in mining as a whole . There can be no doubt but that the substitution of machine mining together with permissible powder for the method of " shooting from the solid " has been the means of saving the lives of many shot- firers . Machine mining also leaves the roof in a better condition as lighter charges of powder are required to bring down the coal . Thus the danger from falls of roof are lessened . It is recognized that every effort should be made to prevent all accidents due to the use of electrical machinery but it is felt that any action tending to limit the use of such machinery would be a backward rather than a forward step in the safety measures laid out to decrease the hazards in coal mining .

May 1, 1920

By Magnus Troilius

The importance of having methods for the rapid and accurate determination of manganese in modern steel manufacture cannot be overestimated. The method adopted in the laboratory of the Midvale Steel Works is very satisfactory, and I think it will be of interest to some of the members to give a description of it. Before giving the more detailed account of the manipulations, a review of some well-known methods and chemical facts may not be out of place. The " bromine and ammonia" process, as fully described in the tenth volume of the Transactions of this Institute, was certainly an improvement on the methods in which fixed alkaline salts were used, as these gave invariably too high resnlts, as generally carried out. But the separation of all the iron before the manganese was, in both methods, a source of great inconvenience and delay.

Jan 1, 1884

By R. C. Canby

The difficulty of exact neutralization by ammonia of the acid solution obtained in the determination of arsenic by the method of fusion with carbonate of soda and nitrate of potassium, led me to try the effect of neutralization by oxide of zinc. I obtained most satisfactory results, and found a great saving of time. The oxide of zinc being added in excess, there is no delicate testing with lit? mils paper and tedious alternate adding of dilute ammonia arid acid until the proper point is reached. The excess of oxide of zinc, too, seems to hold, from the filter, the gelatinous silica and alumina, which are usually precipitated upon neutralization, and thus to render the washing of the precipitated arseniate of silver almost invariably much more rapid. Except the substitution of the neutralization by oxide of zinc for the more tedious method by ammonia, the details of the process are practically the same as those published by Mr. Richard Pearce, of Argo, Colorado (Proceedings of the Colorado Scientific Society, vol. i.), who shows what very accurate results can be obtained by the fusion process. I have made numerous tests of comparison between the two methods of neutralization, and find that those by the oxide of zinc agree almost exactly with the old method. Increasing the quantity of nitric acid, even to considerable excess, to be neutralized by the oxide of zinc did not affect the precipitation of arseniate of silver. The process may be outlined as follows : Take 4 to 1 gramme finely pulverized ore, and mix in a porcelain crucible with eight to ten times its weight of a mixture of equal parts of carbonate of soda and nitrate of potassium (Kerl. Metallurgische Probirkunst gives 4 to 5 nitre to 14 soda). Gradually bring to fusion and keep so, three to five minutes. Cool. Disintegrate in hot water. Filter and wash with hot water. Acidify the filtrate with nitric acid, and boil to expel carbonic acid and nitrous fumes. Cool. Add an excess of an emulsion of oxide of zinc. The bottom of the beaker should have quite a layer of oxide of zinc, after stirring.

Jan 1, 1889

By D Feng, G C. Lukey, J S. J, R R. Dagastine

An ongoing challenge in the coal industry is pyrite removal using the flotation method for fine particles. The use of a æseparation in frothÆ (SIF) technique to recover coarse particles of up to a few millimetres in size has presented a potential solution. However the commercial success of SIF is still limited due to the poor understanding of the separation mechanism. The particle floatability in SIF is determined by the constructive or destructive stresses resulting from the forces exerted by the particle on the bubble films. Hence, the surfactant properties and the mineral characteristics (eg hydrophobicity, shape, etc) are the key factors that control the interaction between the particle and froth. In the present work, the interaction forces between different particles (ie crushed glass, pyrite and coal) and a single air bubble are measured in a non-ionic surfactant solution (pentaethylene glycol mono n-dodecyl ether, C12E5) using an atomic force microscope (AFM). Experimental results show that the interaction is primarily dependent on the contact angle (ie the degree of hydrophobicity of the minerals). For crushed glass, the transition from repulsion to attraction was observed as the contact angle increased. As the concentration of C12E5 increased, the adhesion force was suppressed due to wrapping effect. For coal and pyrite with much higher hydrophobicity, long-range attraction was observed. The effect of shape factor was only visible at high surfactant coverage where sign of destabilisation became noticeable on the retracting force curve. These results provide a fundamental understanding of the direct interaction forces of a complex real system between an irregular particle of either crushed glass, pyrite or coal and a single air bubble. Theories developed for interpretation and modelling are thus far based upon the study of ideal systems of glass or polystyrene spheres and an air bubble and were incapable of explaining the interactions in complex real systems. The presented measurements were a first approach to the problem of real mineral-bubble interaction in flotation systems. Thus, the results can only best be interpreted semi-quantitatively.

Jan 1, 2005

By J. A. Marshall

"INTRODUCTION The purpose of this paper is to investigate the potential for use of UAVs in underground mines and present a prototype design for a novel autorotating UAV platform for underground 3D data collection. The mining industry has recently shown increased interest in the use of unmanned aerial vehicles (UAVs) to assist in everyday operations [20, 11, 4]. Above ground, small UAVs are in some cases a more efficient, inexpensive, and safer alternative to manned aircrafts currently used for photography, inspection and security [6, 10]. For example, attaching camera, infrared and LiDAR payloads, UAVs can provide a low-cost method of obtaining highly accurate 3D photogrammetric data and aerial photography. UAVs are now commonly used in open pit mining operations for applications that include stockpile surveying, 3D pit modelling, facilities management, accident reporting, progress monitoring, and environmental assessment [11]. There are numerous companies offering UAV services for mining applications; e.g., UAV Geomatics, Leica Geosystems, Sky Futures, SenseFly and FlyTerra, to name just a few. Common services include stockpile and open pit volume computations, environmental assessment, aerial mapping and photogrammetry. Although the aforementioned services are thosemost commonly provided by UAV surveying companies, they may also add value to other daily mining operations. For instance, SenseFly offers pre and post-blast monitoring in order to identify the presence of misfires and wall damage. This data can also be used to reconcile the blast results with expected results. GEM Systems offers a UAV equipped with a magnetometer for mineral exploration surveys. Companies such as Barrick have even used UAVs for solids modelling at tailings dams and stability monitoring [11]. UAVs in mining have so far been mostly limited to surface applications. Harsh underground environments pose many obstacles for flying UAVs. The confined space, dampness, reduced visibility, air movement, and lack of control signal propagation hinders most operators from being able to fly a drone underground. It may be that truly practical uses for UAVs underground will require either autonomous or semi-autonomous flight capabilities. Although there are many difficulties with flying underground, the potential benefits from a working system could greatly improve mining operations. The potential benefits of deploying UAV platforms underground include access to unreachable and dangerous locations and removing personnel from unsafe situations. These benefits have the potential to greatly improve mine monitoring and mine safety. Research has shown that current UAV technologies exist that allow for autonomous indoor flight. Extensive research has been done to develop UAV systems that are capable of performing on-board simultaneous localization and mapping (SLAM), which can allow them to navigate and map a foreign environment autonomously [5, 7, 1, 13]. Grzonka [7] successfully used an open hardware quadrotor to autonomously navigate and map an office building. The research outlines the localization, mapping, path planning, height estimation and control of the autonomous quadrotor. Other research has been done that exploits autonomous UAVs for search and rescue. Kassecker [9] proposed a software and hardware framework for a quadrotor capable of indoor and outdoor urban search and rescue and Rudol [19] developed a system for human body detection and geolocalization using an autonomous UAV.The use of autonomous UAVs in search and rescue has the potential to improve situational awareness and surveillance for a rescue team. Similar research has been done that uses autonomous UAVs for indoor exploration. Pravitra [15] outlined a strategy for autonomous exploration of miniature air vehicles (MAVs) within indoor environments and Rudol [19] developed a method for using an unmanned ground vehicle (UGV) and a UAV for cooperative indoor exploration. These studies show that aut"

Jan 1, 2017

By Donna L. Boreck

Introduction Coal has been mined in the western United States since the mid-1800s. The resource is abundant in the West. Despite recent economic conditions, it remains an important part of the nation's energy warehouse. Many of the seams that comprise the western resource base are thick, some exceeding 61 m (200 ft). Multiple, closely-spaced seams or thick coals that split into two or more minable seams are also common. Historically, subsurface development recovered from less than 15% to 50% of these deposits (Matson and White, 1975). This resulted in a substantial loss of inplace reserve due to the limited height that could be mined using the available technology. In an effort to increase future recovery from such deposits, the US Bureau of Mines analyzed the potential for incorporating three thick-seam mining methods into western coal mines. These methods are high-face longwall, multislice longwall, and longwall caving. They were developed in Europe and are being used to mine thick coal seams in Europe and Asia. One aspect that will determine their success in the US is the geology of the individual deposit; namely, are the methods adaptable to the geologic conditions characteristic of western coal? This paper analyzes some of the geologic factors that would affect development of thick or closely-spaced seams using these mining methods. Occurrence of thick and closely-spaced seams Since mining conditions and methods differ around the world, a thick seam is often defined as a seam whose total thickness cannot be extracted using the available equipment. In the western US, the maximum extractable mining height is about 4.3 m (14 ft). For this research, a thick seam was defined as a minable seam 4.6 m (15 ft) thick or more. Since one of the seams is usually lost using standard mining methods, closely-spaced or split seams were also addressed. These were defined as two minable seams 1.5 m (5 ft) thick or more separated by 9 m (30 ft) of interburden. Thick or closely-spaced coal seams are common in the West. Their importance has not been documented due, in part, to the methods used to measure resources. In the past, all reserves greater than 1.1 m (42 in.) thick for bituminous seams and greater than 3 m (10 ft) thick for subbituminous seams were grouped together and were differentiated only by rank. Further differentiation by seam thickness or minability was not done. For this reason, reserve figures for thick or split seams are difficult to determine. By analyzing available published resource data, Ferm and Muthig (1982) addressed thick seam resources. They estimated a resource of 90 Gt (99 billion st) of coal (16% of the Western resource base) in seams 4.6 (15 ft) or greater at less than 610 m (2000 ft) overburden. Yet, a later study of Pierce and others (1982) on one thick seam deposit in the Powder River Basin estimated a resource of 103 Gt (113 billion st). Pierce (1986) reported that the overburden reached a maximum of 731 m (2400 ft). The resources in this single deposit exceeded the 90 Gt (99 billion st) total projected for the entire western US by 12.7 Gt (14 billion st). Due to such differences, this author chose to look at reported occurrences of thick or closely-spaced minable seams in place of resource figures, to verify the need to research thick seam mining in the West. An occurrence was defined as an isolated site where a thick seam or two or more closely-spaced seams were reported. The occurrences were derived from published material - mostly mine descriptions, measured sections, coal exploration logs, and oil and gas logs. Due to time constraints, the reserves associated with each occurrence were not determined. All data points were random. The analysis was limited to occurrences in Colorado, Wyoming, and Utah. Although source material available to the author was limited to general coal summa-

Jan 3, 1988

By C. H. Smith

The present interest in beryllium seems to be the result of recent developments in the uses for the metal discovery of significant deposits of disseminated beryllium ores, and the development of chemical and physical processing techniques, for recovering the beryllium values contained in these ores. At the present time most of the new and potential uses for beryllium are in the research stage of development and are not consuming large quantities of the metal. However, they do point to an expanding market. Interesting new discoveries or extensions of previously known deposits have been reported during the past year from the Boomer Mine in Colorado, the Mesquite area of Nevada,, the Chemalloy pegmatites in Canada, the Brazilian and African pegmatite districts and from other areas, However, this paper will discuss only four of the beryllium deposits, which appear to be of special significance due to their; large reserves of ore and promise economic development. The new chemical and physical processing techniques are not discussed in detail, but it should be stressed that the development of an economical process to handle the pegmatitic and low-grade non-pegmatitic beryllium ores is of great importance to the development of the beryllium industry. Scientists in both government and industry are co-operating in this effort and it is expected that commercial processes will be developed in the near future.

Jan 1, 1962

Zinc sulphide concentrate from Mount lsa, Australia, was leached in acidic sulphate-chloride solutions using ferric ions as oxidants under atmospheric pressure. The parameters of temperature, particle size and concentrations of ferric and chloride ions were evaluated.The shrinking core model can be used to describe the leaching which is limited by the electrochemical processes in the initial leaching stage, and by both surface reaction and pore diffusion processes subsequently. The apparent activation energy measured in the initial leaching stage is 49 kJ/mole over the temperature range 60 to 90°C.About 99 per cent of the contained zinc in the -45+38 µm particle sample dissolved at 90°C in a 0.5 M H2SO4/0.1 M Fe2(SO4)3/0.8 M NaCI solution within two hours. Some nine per cent of sulphide sulphur was oxidised through to sulphate; the remainder was convened to elemental sulphur.The mixed sulphate-chloride system has advantages over separate chloride and sulphate systems due to the low reagent cost, high leaching rate and possible selective leaching of zinc sulphide concentrate with high lead contents.

Jan 1, 1994

By D Millette, F Gay

Surface tailings are finely ground waste materials produced in all mines after the mineral or metal has been removed from the metallurgical process. Current methods for disposal of tailings involve constructing dams, dumping into existing waterways, and depositing as backfill underground. Each method has its drawbacks in terms of cost, level of complexity, safety and environmental concerns. Backfilling is the process whereby waste material (generally metallurgical tailings, prepared aggregates or a combination) is placed from the surface back underground into open mine workings via conveyor belts, hydraulic or pneumatic pipes or gravity. In the backfill process common difficulties are expensive placement costs, equipment wear and the ability to meet early and long-term strength requirements. Traditional tailings transportation in pipes involves the use of water as the transport medium. Foam technology involves the introduction of micro-air bubbles that replace the water as the main transport medium. This air can be removed after placement if necessary, by the addition of a defoaming agent. This could produce an under-saturated stable cake. Foam technology has the potential to revolutionise surface tailings disposal and backfilling by providing a less expensive and environmentally safe alternative to those just described.

Jan 1, 2000

By George H. K. Schenck

INTRODUCTION Boom-type miners mechanize the cutting and loading of a solid face of rock having a compressive strength as high as 124 MPa (18,000 psi). These machines differ from continuous miners in that they can cut much harder rock for any given cutter-motor power. In addi¬tion, a boom miner is more flexible, selectively cutting any cross-sectional configuration from circular to square arched faces, 1.57-rad (90°) turns, and mixed faces. The basic operational concept of these machines, utilizing a rotary cutterhead mounted on a slewing boom, gained importance in Europe during the 1960s, where, due to their specialized use in roadway development in coal mines, they were called roadheaders. Along with continuous miners, boom miners share the ability to fully mechanize the cutting and loading of rock, while eliminating the need for explosives and the shattering and overbreaking of roof and rib rock that is normal with conventional mining methods. The machines are crawler-mounted and most have a gathering-arm load¬ing system that, moves the cut rock onto an internal chain conveyor, subsequently discharging the rock into a transport device such as a shuttle car or a bridge conveyor. Light- to medium-weight boom miners are cutting softer rocks and minerals successfully, including coal, shale, rock salt, potash, phosphate, weak sandstones (e.g., those associated with uranium ore), and other softer ores such as iron. Heavy third-generation ma¬chines have extended rock cutting to much harder ores such as borate, oil shale, and almost all rock associated with coal seams, except for the harder sandstones and limestones. The mention herein of specific manufacturers, brand names, or models is made by way of example only and does not constitute endorsement by the author or The Pennsylvania State University. All costs cited herein are in 1977 US dollars. TYPES OF BOOM MINERS All boom miners utilize drag bits fixed on a rotary cutterhead that is powered by an electric motor mounted in the slewing and elevating cutter boom. The cutter¬head is either a ripping or a milling type, and the machines have evolved through three generations (Boldt, 1977). As shown in Fig. 1, only one boom miner cur¬rently produced can use both ripping and milling cutter¬heads on one machine. Table 1 lists boom-miner models built in large num¬bers and/or successfully used or marketed in North America. Table I excludes machines of which only a few have been produced and none sold in North America, such as those manufactured by Demag, Eick¬hoff, Paurat, and Thyssen-Rheinstahl. The only Ameri¬can boom miners are the F-16 and Super ROC¬MINER"', manufactured by AEC, Inc., State College, PA. Machine Generations Three generations of boom miners can be distin¬guished according to their times of introduction, their rock-cutting abilities, and their weight classifications. Fig. 2 illustrates typical best-practice production rates possible from boom miners for selected cutter-motor power sizes (other power figures can be interpolated). Productivity is related directly to the energy applied to the cutterhead and indirectly to machine weight; higher

Jan 1, 1982

By L. Ou

This paper reports on research on leaching pyrite in a germ-free system. The effects of mineral particle size, pulp potential and various oxidation agents on pyrite leaching were studied. The results indicated that a smaller mineral particle size leads to a higher leaching rate of pyrite. The optimum results were obtained with a mineral particle size of -0.037 mm. The pulp potential reflected the leaching process and an increasing pulp potential improved pyrite leaching. The leaching rate rapidly increased with the addition of oxidation agents such as Fe3+ and H2O2. However, the precipitate generated by ferric ion coated the surface of the pyrite and prevented further leaching.

Jan 1, 2007

By Edward D. Thimons

The Bureau of Mines has developed a damage-resistant brattice, intended for use as a stopping in locations where blast damage to stoppings is a problem, the roof is relatively flat, and the differential pressure across the stopping is not high. Damage-resistant brattices consist of a series of vertically hung brattice panels joined to each other by Velcro connections that improve the airtightness of the brattice. When the brattice is subjected to strong blast forces, the Velcro connections peel apart, allowing the panels to open without incurring damage. The Velcro connections can then be resealed in a matter of minutes. A preliminary test in the Bureau's Safety Research Mine and subsequent long-term testing at the White Pine copper mine in Michigan and the Retsof salt mine in New York show that damage-resistant brattices withstand blasting better than conventional stoppings, while still performing as efficient air barriers when the Velcro connections are sealed.

Jan 1, 1978

By S. K. Kawatra, R. A. Seitz

Various mathematical models and expressions to describe the behavior of a hydrocyclone have been presented in the literature (Lynch and Rao, 1977; Plitt, 1976; Plitt and Kawatra, 1977). Two of these models, by Lynch (1977) and by Plitt (1976), have been extensively used for process control studies (Kawatra and Thayer, 1879; Seitz and Kawatra, 1984). However, neither of these models relates the fineness of grind in a hydrocyclone overflow to the corrected efficiency curve of a hydrocyclone. If such a relationship were available, then one could relate fineness of grind to the operating variables using equation (1) or (2). A plant metallurgist is interested in knowing the fineness of grind rather than the dso(c) of a hydro- cyclone. As a consequence, several empirical relation- ships between fineness of grind and hydrocyclone operating variables are available. These relationships (Fewings, Wickham, and Wiseman, 1979; Manning, 1978; Watson, Cropton, and Brookes, 1970; Watts, Thompson, and White, 1978; Wiegal, 1973) have only a limited relationship with each other and have met with varying degrees of success. The object of this paper is to develop a fundamental relationship between fineness of grind and the efficiency curve of a hydro- cyclone

Jan 1, 1986

By K. -A. Björkstedt

INTRODUCTION Strassa lies in the central part of Bergslagen, a tradi¬tional mining district, on the eastern side of the Stora Valley at an elevation of about 200 m above sea level. A railway siding runs between the mine and the Stora railway station from which there are railway connections to the shipping port and iron and steel works in Oxelosund, about 224 km away. The distance to the provin¬cial capital Orebro is about 60 km. The climate is typi¬cal for this part of central Sweden and is illustrated by the diagram of monthly precipitation and temperatures for the years 1968-1975 (Fig. 1). HISTORY There is no certain information as to when the Strassa mine was first worked, but it is known from sur¬viving accounts of mine inspectors that there were smelt¬ing works in operation in nearby villages in the 12th century. An example is the Gusselhytta ore smelting works, 10 km south of Strassa, which dates from this period. Around the year 1540 there were two smelting works in Strassa, the Upper Karberg and Lower Karberg works. Ore for these smelters was probably taken from Strassa and from the adjacent Blanka mine. In the year 1624 Strassa is mentioned by the painter Jons Nils Krook in an account of the iron mines in the Linde mining district (Linde Bergslags Jarngruvor). Several mines were listed in the area, the deepest being about 30 m. An impressive power installation is mentioned in 1639, including a piston system of lashed poles for transmit¬ting power from the Stora River to the Strassa fields. Its length was 2670 m. Common ground comprising about 20.2 km2 (5000 acres) of forest was allocated in 1689 for the furtherance of mining operations. Until the beginning of this century only the rich cen¬tral parts of the ore body were mined and these yielded, after handpicking, lump ore suitable for smelter feed. An example of the ore grades from these early times is an analysis of ore from the "Big Mine" (Storgruvan) from the year 1873: 48.5% Fe, 0.008% P, and 0.06% S. This same year a total of about 18 000 t was ex¬tracted from the Strassa mine. OWNERSHIP The mine was owned and run until 1874 by a min¬ing association made up of 119 so-called "bergsman," who were homesteaders often engaged in agriculture and timber-cutting as well. In that year the Strossa Grufvebolag (Mining Co.) was founded. In 1906 it was con¬verted into a joint stock company, the Strossa Gruveaktiebolag. This was acquired in 1907 by Metallurgiska AB for the implementation of Gustav Grondal's beneficiating and briquetting methods, for which the Strassa ore was well suited. The same year saw the completion of a new ore dressing plant with an annual production of 46 000 t of ore concentrate. In 1911 the mine passed to new hands, and in 1913 it was purchased by an Austrian company. Extensive new installations were made and in 1915 a new dressing and briquetting plant was completed with twice the capacity of the old one. In 1917 the Strassa mine was acquired by Granges. Be¬cause of unfavorable business trends and technical diffi¬culties, mining operations were brought to a close in 1923. Pumping kept the mine free of water until 1933 but it was completely filled ten years later. Up to 1950 the surface buildings and installations remained intact but the large dressing and briquetting plant burned to the ground in that year. Today only the machine shop re¬mains from this earlier period of operation, now housing parts of the Mineral Processing Laboratory. The decision to take up mining operations again was made in 1955 and construction work began the follow¬ing year. Of the old installation, only the "southern shaft" could be used for some development drifting after it had been completed with a new headframe. Other¬wise, all the buildings and installations required for the operations had to be rebuilt. New installations ready by 1960 were office and personnel facilities, a new shaft and headframe, a sorting and concentrating plant, a macadam plant, settling basins, pump stations, and a railway and yard with transport equipment. The instal¬lation was completed with two plants

Jan 1, 1982