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By Andrew T. Zador

The growing demand for energy combined with the need to protect and preserve the environment have become prime considerations in the design of efficient coal handling ports and their associated stockpiles. The transportation system is analogous to a connected chain which originates at the mines and terminates at steel mills or power plants, in which the port and the stockpile reserves are a vital link. Railroad transportation costs can be minimized by rapid discharge of the cars which, in turn, requires that an adequate stockpile system be available. Ocean transport costs can be minimized by having an efficient high capacity ship loading system and sufficient quantities of coal on hand -all of which also requires an efficient stockpile system to meet shipping demands. The connector between the two modes of transportation, therefore, is the port stockyard. It is here at the interface that optimization of land utilization and materials handling equipment results in a total transportation chain that can operate as a smooth efficient system. Multiple origins of coal require the simultaneous handling of large quantities of coal of many grades. Railroad transportation logistics often limit and define the incoming arrival patterns and limited rolling stock often require priority treatment for railroad operations. The physical limitations of prime ocean frontage require that coal stockpiles be built close to residential or industrial areas which entail strict environmental controls. Large capacity, enclosed coal stockpiles for ocean shipping have not yet been considered feasible but may be forthcoming in the near future. Several have been analyzed.

Jan 1, 1978

By E. B. Thorn

The production of portland cement requires that four basic elements -- calcium, silica, alumina and iron -- be ground to a fine powder, homogenized, and fed to the pyroprocessing unit, where it is burned to a cement clinker. The raw material must either be ground wet in a slurry form or dried and ground to a dry powder, with a fineness of approximately 85% passing the 200 M screen. According to the Portland Association's Plant Information Summary, approximately 38% of the U.S. capacity is ground wet and 62% ground dry. The following data is taken from the Portland Cement Association's 1981 Energy Survey: [Million BTU/Ton kW h/Ton Wet Process Plants 5.86 138 Long Dry Process Plants 4.97 123 Dry Preheater Plants 4.03 157 ]

Jan 1, 1984

By Bo-Hyun Kim, Mark K. Larson

"Researchers at the National Institute for Occupational Safety and Health (NIOSH) have advocated the use of instrumentation in underground mine ground control studies to aid in understanding failure mechanisms and to calibrate numerical models. The purpose of these efforts is to reduce accidents and fatalities in underground mines resulting from inadequate design of mining layout or supports. A recent examination of some results of stress change measurement raised some questions about the validity of elastic solutions in ground, such as coal. Particularly under deep cover as is prevalent in the western United States, the coal around a borehole would likely have yielded so that post-yield behavior would govern the response of the instrument.A numerical experiment was conducted by NIOSH researchers using the FLAC software (Itasca Consulting Group, 2016) to investigate the response of the borehole pressure cell (BPC) and the biaxial stressmeter to varying degrees of plastic behavior. The pre-encapsulated BPC cell pressure exhibits increasing departure from the elastic case with decreasing post-peak-strength slope of the stress-strain curve. On the other hand, because the BSM usually provides much more confinement to the surrounding rock than does the BPC, the measurements obtained from the BSM do not depart significantly from the changes in secondary principal stresses and directions as calculated by the elastic solution.INTRODUCTIONSeveral years ago, Larson et al. (2000) described several instruments that NIOSH had used in underground field studies and how to monitor them with dataloggers. This information is still relevant because of the need to understand how excavation affects the stability of mines and thus, the safety of mining personnel. One class of instruments described in that paper includes those instruments that measure change in rock stress. Although several such instruments might be used to measure stress change, the focus of this paper is on two instruments with which NIOSH has had more recent experience—the borehole pressure cell (BPC) (Geokon, 2018a; RST Instruments, 2018) and the biaxial stressmeter (BSM) (Geokon, 2018b). Other instruments, such as the CSIRO hollow inclusion stress cell (HICell) (Earth Sciences, 2019; Worotnicki and Walton, 1976) may be included in future studies.Early studies in the development of the BPC were those of Panek and Stock (USBM, 1964), Sellers (1970), Lu (1984), and Babcock (1986). Babcock’s work is particularly useful, but his setting and testing range was far below what we experienced in the NIOSH study, and it was Babcock’s data reduction scheme from his laboratory experiments that was used to calculate stress change in the NIOSH study (NIOSH, 2019a)."

Jan 1, 2019

By M. C. Manderson

Phosphate rock is a naturally occurring mineral which, in upgraded form, is important to both agriculture and industry. Phosphate rock normally must be upgraded before it can be assimilated and used. Most of this upgrading is now carried out by chemical means, largely through the reaction of a strong inorganic acid (primarily sulfuric acid) with the phosphate rock to remove major impurities and to produce soluble phosphate. A less important but yet still commercially significant conversion method involves the thermal reduction of phosphate rock at high temperature in an electric furnace to produce elemental phosphorus. The elemental phosphorus is then converted to usable end products. This conversion route is a more costly method than the chemical route, but it has the advantage of being able to produce chemically pure phosphorus derivatives. On the other hand, the use of phosphate rock for agricultural application-primarily for use as a plant nutrient ingredient for fertilizer -does not require the same level of purity. As a consequence, the chemical method of phosphate rock upgrading is the predominant technique for producing phosphatic fertilizers.

Jan 1, 1977

By Yang Zhao, Wanping Wang, Siqing Liu, Shuming Wen

"In China, most reserves of rich iron ores have now been depleted and sustainable development of low-grade iron ores has become a critical discussion topic in the mining industry today. The aim of this study was to evaluate the applicability of beneficiating a low-grade, hematite-magnetite ore (assaying 18.64% Fe) for subsequent utilization. In this study, process mineralogy and beneficiation of the ore were carried out in the laboratory. First, mineralogical investigation shows that the main metal minerals are hematite, magnetite and a little octahedrite that cannot be used at present, and the main gangue minerals include albite, quartz and ankerite. Based on the above, a combined process of magnetic separation and gravity concentration has been developed by using laboratory low- and high- intensity magnetic separators and a slimes shaking table, respectively. Factors affecting these separation methods are discussed, including process mineralogy on the tailings of the high-intensity separator, where the iron minerals were found to be finely disseminated in the gangue as inclusions. By applying a two-stage grinding step and a magnetic separation stage together with gravity concentration, a marketable iron concentrate assaying 60.79% Fe at a recovery of 60.23% can be obtained, when the feed assays 18.75%Fe, which provides some useful information for the sustainable use of low-grade iron ore.IntroductionIron is the world’s most commonly used metal. Steel, of which iron ore is the key ingredient, represents almost 95% of all metal used per year, primarily in structural engineering applications and for maritime purposes, automobiles and general industrial applications. By the end of 2009, the iron and steel production in China ranked the first in the world, and reached 0.567 and 0.537 billion tons, respectively. The domestic supply of iron ore was 0.88 billion tons and about 0.628 billion tons of iron ore were imported in 2009."

Jan 1, 2015

By Jorgen Elbrond

The calculation of open pit operations' capacities has been the subject of many studies using computer simulations. The common problems of simulation are that if one desires a realistic picture of an operation the programming work becomes long and expensive and that the program requires a large computer and long time to run. The method presented in this paper overcomes these problems by using a straightforward calculation procedure. It is based on the queuing theory's simple formula for waiting time in a closed circuit and uses a system of correction factors, which reflect the loading, travel, and dumping times' relative variabilities and which have been found by 1) mathematical analysis and 2) simulations of basic circuits. The method includes the main parameters of the operation: numbers, sizes and operational availabilities of trucks, shovels and dumping point, distribution of the production over shovels, trucks' loading, travel (loaded and empty) and dumping times as well as their variabilities, shift's composition. The procedure models: the average duration of a complete truck cycle during stationary and non-stationary periods during the shift as a function of the number of trucks in the circuit, the composition of the average 100% operated shift, the simultaneous probabilities of operationally available shovels and trucks and the distribution of shovels and trucks according to production and cycle times. The procedure can be programmed for a small computer and requires only short computing time, which makes extensive studies of various parameter combinations practical to carry out. The development of the method is an extension of previous work by Morgan and Peterson {1}, who by simulation established the relation between the match factor (which is the ratio between the sum of one loading time pr truck in the circuit and the complete net cycle time) and the shovel's efficiency, and by Hayes, Splaine and Whitaker {2}, who established this relation from a large number of observations in the Nchanga pit in Zambia. The present method establishes this relation by a straightforward calculation procedure. Theoretical solutions to some situations are found in the litterature on the queuing theory. The papers {3},{4} and {5} give solutions to situations, which fall outside those encountered in the shovel-trucks circuit. These solutions together with simulations of in-termediary situations have been used to establish an interpolation procedure for the correction factors to be applied to the waiting time formula in order to cover the situations encountered in practice.

Jan 1, 1977

By Joseph W. Murray

GENERAL DESCRIPTION The undercut-and-fill method of mining was devel¬oped by Inco Ltd. in the Sudbury district of Ontario, Canada, to deal with abnormal ground conditions en¬countered in pillar recovery. Although it is now some¬times utilized as a primary stoping method, the main application continues to be in pillar recovery. In any mining method in which fill is employed, pillar systems tend to fail gradually as stoping pro¬gresses. As a result, most pillars are completely destressed and broken by the time pillar recovery com¬mences. Control of the back in such ground becomes a major problem and is contingent on local conditions as they exist in the workplace. In most mining methods, operations such as drilling, blasting, slushing, and tim¬bering can be readily assessed and the mining cycle pre¬dicted. However, in the mining of pillars with overhand methods, the time and effort required to maintain a safe back varies to a quite unpredictable extent. For years the square-set method was the most effective means available for pillar recovery. However, in areas where bad ground was encountered, square setting was a costly and sometimes hazardous method. As the tonnage mined from pillars increased, it was apparent that a safer and less expensive approach was required. Past experience indicated that the best way to deal with broken ground was to mine from the top down. Top slicing systems were examined but were rejected due to the major sub¬sidence which resulted. However, ideas generated from studying top slicing finally developed into the present method of undercut-and-fill. This method, which has almost completely replaced square-set mining, exhibits reduced costs, has improved safety and efficiency, elimi¬nates the factor of bad ground overhead, and allows a closer prediction of the actual mining cycle. The result is more accurate scheduling of pillar production. This factor is of increasing importance as the tonnage from pillar recovery increases in relation to total mine pro¬duction. Undercut-and-fill is a method of extracting a block of ore by mining successive cuts, working from the top down. After a cut of ore is completely mined out, lami¬nated timber stringers are constructed along the sides for the full length of the cut. Round logs covered by a wire screen are laid across the stringers to form a mat. The cut is then tightly filled with hydraulically placed cemented sand fill. Mining is then resumed on the cut immediately below the mat. As the cut advances, the timber stringers are supported by round timber posts seated on the solid bottom of the cut. Drilling, blasting, muck removal, and timbering operations are repeated until all the ore in the cut has been removed. At this time, another log mat is laid and the cut is filled. This sequence is continued until the block has been mined from level to level. The basic principles of undercut-and-fill can be adapted to a variety of mining situations. It can be applied transversely or longitudinally in ore bodies of varying width, that dip from vertical to flat lying. The method can be applied using conventional equipment or it can be mechanized through the introduction of small diesel or air-powered units for muck removal. Undercut¬and-fill is a selective method, ore recovery is high, and dilution can be controlled to an acceptable level. How¬ever, it is a high-cost, labor-intensive method and as such its use is restricted to areas where the less costly meth¬ods of ground support are not successful or where bulk mining methods are not adaptable. One basic require¬ment for undercut-and-fill is an adequate source of hydraulically placed cemented sand fill and a dependable sand plant and distribution system. The success of the undercut method is to a large degree dependent on the competence of the fill placed in each of the successive cuts. In theory, the weight of the fill on each cut is supported in part by the cemented sand itself as it binds to the rough sides of the mined¬out areas next to the pillar walls, and in part by the tim¬ber support. As mining progresses beneath the mat, a downward pressure is exerted within the cemented fill and a voussoir arch develops above the mat, across the full width of the cut. This supports the major portion of the load. The log mat and posts have a double func¬tion in that they carry the remaining weight not sup¬ported by the voussoir arch and in addition prevent any pronounced settling of the fill mass. Excessive settling breaks the bond between the cemented fill and the pillar walls, eliminating the arch effect and transferring the entire load to the mat and posts. The theory of a voussoir arch forming within the cemented sand fill mass is substantiated in practice by the fact that pillars have been mined from level to level with no more significant weight on the bottom cuts than was experienced on the top cuts. In effect, as mining progresses downward, there is no additional transfer of weight from previous cuts to the cut being mined. APPLICATION OF UNDERCUT-AND-FILL Because of their size, the Sudbury ore bodies are, for the most part, mined transversely and, to a considerable extent, by filling methods. Transverse stoping blocks are designed with 9-m (30-ft) crown pillars established be¬low each level and with stopes separated by vertical rib pillars. In past years, floor pillars were left above the level as well, but present practice favors the silling of primary stopes at the base of the rail. Rib pillars are designed to meet the ground control requirements of the primary stoping methods. Local conditions and past experience generally dictate their size. Historically, since the square-set method was to be used for recovery, rib pillars were designed to accommo¬date standard 1.7-m (5'h-ft) sets. A typical rib pillar is four sets wide [6.7 m (22 ft)] when located between two square-set or cut-and-fill stopes having a width of 8.4 m (27.5 ft). When the undercut-and-fill method was de¬veloped, a mat width of two sets was adopted. This di¬vided each cut of the common four-set pillar into two

Jan 1, 1982

By George D. Brunton

The past and present argillaceous rock studies for the U.S. National Waste Terminal Storage Program consist of: (1) evaluation of the geological characteristics of several wide- spread argillaceous formations in the United States; (2) laboratory studies of the physical and chemical properties of selected argillaceous rock samples; and (3) two full-scale in situ surface heater experiments that simulate the emplacement of heat-generating radioactive waste in argillaceous rock.

Jan 1, 1981

By Terrence R. Lejcher

Strip mining has played a key role in the production of coal used to produce energy in this country. About 56 percent of the 1.3 x 106 hectares (3.2 x 106 acres) that had been disturbed by surface mining as of 1965 can be attributed to stripping (Udall, 1967). The runoff from these stripped areas is usually heavily polluted with high concentrations of total dissolved solids, iron, aluminum, manganese, sulfates, and has low pH values with high acidity. Another problem of a most serious nature in our large cities is the volume of treated solids to be disposed of after sewage treatment. Chicago is presently producing about 664 metric tons/day of dry weight solids (Horn and Earles, 1968). Disposal methods most commonly used include lagooning, incineration, and landfilling. The high cost of land in an urban community is prohibitive of lagoon storage. Incineration and landfill disposal create air and water pollution problems. Many countries including Australia, England and Israel have long realized the value of sewage solids if they are recycled back to the land. In the United States, the value of such solids in reclaiming strip mined land has been recognized by relatively few people. This paper is concerned with the use of anaerobically digested sludge to reclaim strip mined land.

Jan 1, 1972

By Shu Fan Chau, Innes Duncan, Harry Asche, Bill Newns

The North Auckland Line (NAL) in Northland, New Zealand, is being upgraded to pass Hi-Cube containers. The line includes many tunnels and bridges which need to be refurbished to improve structural clearances and line capacity. Tunnel 2 along the line, the Makarau Tunnel, has posed special challenges. Opened in 1898, the tunnel suffered multiple collapses during construction and has shown significant distortion during operation. The tunnel has been repaired many times since opening. Successfully enlarging this delicate and distressed structure has involved historical research, advanced analysis and design, and careful construction. The results provide lessons for dealing with brick lined tunnels and poor ground conditions

Jun 13, 2021

By Thomas B. Borak, Keith J. Schiager, Janet A. Johnson

OBJECTIVES OF RADIATION MONITORING Monitoring is a tool used in the practice of radiation protection. The primary reasons for monitoring are to reduce radiation exposures to levels that are as low as reasonably achievable and to assure that no individual receives a dose exceeding the maximum individual dose limit. The documentation of radiation doses for legal, medical, or epidemiological reasons is a subordinate function of any monitoring program. The investment in radiation monitoring programs should be guided by four criteria: (1) the detection and avoidance of unnecessary exposures, (2) the magnitude of potential health risks, (3) the determination of combined doses and risks with adequate confidence, and (4) the verification of compliance with established limits. FIRST CRITERION: DOSE REDUCTION - DETECT AND CORRECT UNNECESSARY EXPOSURES The system of dose limitation advocated by the International Commission on Radiological Protection (ICRP, 1977), and subscribed to in a broad sense by various regulatory agencies, is comprised of three essential ingredients: (1) [justification] of any practice that produces radiation doses by some commensurate net benefit, (2) [optimization] of radiation control measures by reducing doses to levels that are as low as reasonably achievable, and (3) [limitation] of individual doses to preclude inequities and [moldistribution] of risks. All too often, only the third criterion - the limitation of individual doses to prescribed regulatory limits - is explicitly addressed in everyday radiation protection programs. The emphasis of most exposure control and monitoring efforts appears to be directed toward limiting and documenting individual doses that might approach the legal limit. The first two criteria, i.e. justification and optimization, should contribute to a rationale for allocating monitoring efforts. When applied to individuals, these criteria mean the detection and elimination of unnecessary exposures. This should be a high priority of any monitoring program. Measurements should be directed toward detecting inoperative or ineffective control measures, whether or not there is a risk of exceeding the individual dose limits. The ICRP recommends a procedure that can be used effectively to reduce unnecessary exposures. A n investigation level should be established at an exposure rate substantially lower than the regulatory limit, e.g. 30% of the limit (ICRP 26, 1977, p.33). Measurements obtained during routine monitoring that exceed the investigation level are evaluated with respect to cause and potential reduction. To be effective, the evaluation results should be formally recorded and conveyed both to management and to the workers involved. Although the investigation level recommended by the ICRP is based on a fixed exposure rate or derived air concentration, an equally effective evaluation program may be based on the investigation of a percentage of all measurements. For example, one might investigate the highest 5% or a random selection of all measurements. In any case, the objective is to detect and correct situations that are producing unnecessary exposures. SECOND CRITERION: MONITOR IN PROPORTION TO THE MAGNITUDE OF RISK The ICRP criteria apply generally to all radiation exposures. However, a second priority for monitoring programs should be established on the basis of the nature of the exposures and the magnitude of the health risks involved. Current practices in radiation protection are based on dose limits to specific organs and assessment of individual exposure pathways, with little consideration of the combined doses from various pathways. Although the intent of the ICRP recommendations was to limit the total dose to each critical organ, combined doses from external and internal sources are rarely determined. The recent recommendations of the ICRP (Publ. 26, 1977) are based on the limitation of total health risk from occupational radiation exposures. Implementation of this concept would necessarily require the measurement, calculation and summation of doses and concomitant risks to all organs of the body from all exposure pathways. The U.S. Environmental Protection Agency has taken the first step toward translating the ICRP recommendations into regulations. The proposed recommendations for Federal Radiation Protection Guidance for Occupational Exposures (USEPA, 1981) include provisions for summing the risk-weighted organ doses to determine compliance with an effective whole-body dose equivalent limit. Whether or not the proposed guidance is modified before final adoption, it seems clear that some version of dose summation and combined risk limitation will be included in future regulations. With this

Jan 1, 1981

By John H. Bassarear

The never ending search for lower unit costs led Pima Mining Company's management to study various alternatives shortly after Expansion III was completed. It soon became apparent that a considerable amount of ore assaying 0.35% copper had to be moved in order to maintain mill feed at the desired 0.54% copper content, which is the overall grade of the ore reserves. Detailed investigations revealed that for a period of 8 - 10 years, about 14,000 tons per day of 0.30 - 0.40% copper ore would pass by the crusher en-route to the low grade stockpile. This ore is very hard, very clean and has to be mined at a relatively uniform rate. A six month pilot plant run testing autogenous and semi-autogenous grinding gave very encouraging results. Since no mining costs need to be charged to the extra tonnage (it had to be mined anyway) the only major costs would be against milling. The decision to proceed was made in August 1970 and Fluor-Utah was retained as the prime design construction contractor. Ore was first put through the new mill on December 29, 1971. By March of 1972 design tonnage of 14,000 TPD had been attained. The first three expansions at Pima Mining Company, 20 miles south of Tucson, increased mill capacity from 3,000 TPD to 36,000 TPD. These expansions included a separate, conventional, three stage crushing plant plus additions to the existing concentrator that primarily duplicated the then current flowsheet. Expansion IV, however, consisted of an entirely new and separate primary crusher and a completely new mill building located about 2,000 feet southeast of the 01 d concentrator. The new mill, using semi-autogenous grinding, produces scavenger and rougher flotation concentrates that are pumped to the old mill for regrinding, cleaning, thickening and filtration.

Jan 1, 1973

By Daniel E. Collins

Massive banded Fe-Zn-Cu-Pb sulfide deposits in the Bagdad are associated with metamorphosed eugeosynclinal mafic volcanic and intrusives rocks Precambrian age. The geologic environment is very similar to that of the Jerome-Prescott area which lies 70 km. east of Bagdad. Mining of massive sulfides deposits began in 1917 on the Copper King mine in spite of its early discovery and the discovery of the Old Dick mine in 1882. The first shipments of zinc ore from the Copper King were hand sorted and averaged 46.89% Zn, 5.7 oz. Ag, 3.58% Pb and 1.68% Cu. Oxidized copper was first produced from the Old Dick mine in 1943~ In all, five deposits of massive sulfide ores have been mined in the Bagdad for a total production of nearly 2 million tons with an average grade of 12.0% Zn and 3.6% Cu. The largest deposit, the Bruce mine is currently closed because reduced metal prices and has produced a total of 823,000 tons of Zn-Cu ore to date. Mining of massive sulfide ore in the Bagdad area is presently idle. However, the potential for locating yet undiscovered ore bodies is good. Exploration for new ore bodies is difficult at best due to the complex geologic history of the area, the small size of individual ore bodies and lack of understanding as to the formation of massive sulfide deposits. The purpose of this investigation was to obtain a better understanding of the host rock formations through detailed geologic mapping, petrography and chemical analysis.

Jan 1, 1977

By Charles E. Michener

This paper deals with the occurrence of basic and ultrabasic rocks and their relationship in space and time to major structures of the earth. In order to introduce the paper properly, I must go back a little and try to put the subject in perspective. In the brief time that geology has been a scientific occupation, ore deposits have been grouped and classified in various ways depending on the theories that happened to be popular at the time. Some of the early scientists said that ore came from fire, other schools argued that ore originated from water. Then came the lateral secretionists and in later years the hydrothermalists led by the Lindgren school and the latter has largely been accepted as the modern concept of ore deposits. It now seems apparent that the hydrothermal school is being seriously challenged by other schools of thought. It also seems that we have reached the point where we should be willing to admit that we do not really know in most cases how the ore got to its present position. However, for practical purposes, we can classify ore deposits as to the environment in which they do occur. This is not difficult and gives the field geologist and the mining engineer something to work with and to believe in, and it gives them a basis for carrying out their field work. On the screen, Figure 1, there is a classification of ore deposits based on the environment. Under each classification is listed the more prominent metals which are characteristic of that

Jan 1, 1971

By Jack E. Bailey

From a start in 1957, milling 3,000 TPD of copper ore, the Cyprus Pima Mining Company, which is located 20 miles south of Tucson, has expanded its operations four times, and is currently treating over 55,000 TPD of ore.1 The most recent expansion added a 14,000 TPD semi-autogenous concentrator to the operation. Shortly after startup in January, 1972, the semi -autogenous plant exceeded the designed tonnage rate.2 Ore from the mine is delivered to a 54 inch gyratory crusher and is reduced to a minus 8 inch product. From the crusher the ore is stacked on a surge pile and then reclaimed and fed to the two parallel grinding circuits. A 28 ft. diameter by 12 ft. long semi-autogenous mill and a 16.5 ft. diameter by 19 ft. long ball mill make up each grinding circuit. The semi-autogenous mills are in closed circuit with screens and the ball mill circuits are closed with hydro- cyclones. Rougher flotation follows the grinding operation. Concentrate regrinding and cleaner flotation is performed in Mill I, the older mill. The generalized flowsheet is shown in Figure 1.

Jan 1, 1975

By W. C. Dailey

The Groveland Mine. Concentrator and Pellet Plant are located in Dickinson County in the Upper Peninsula of Michigan. The construction of the Groveland Mine in 1959 was the first United States operation for the Hanna Mining Company in mining and beneficiating low grade iron formations. (1) The plant produced concentrates at 675,000 lt/yr. (746,600 mt/yr.) which was used as sinter feed in the steel industry from 1959 to 1963. Upon completion of a pelletizing plant in April, 1963 production increased to 1.6 million lt/yr (1.77 million mt/yr) of pellets. After the plant expansion, mill capacity increased to 2.2 million 1t/yr (2.43 million mt/yr.), producing 2.0 million It/yr (2.23. million mt/yr.) of pellets assaying 63.0% iron and 7.0% silica. (1) Prior to April, 1971 the concentrator operated with a flowsheet consisting of spirals and a petroleum sultonate iron notation circuit. This was presented in a paper by R. R. Smith and E. E. Sougstad, "Flotation at the Groveland Mine" in the 50th Anniversary Volume entitled "Froth Flotation". The ore body at the Groveland Mine consists basically of two types of iron formation (34% total iron, 20% magnetic iron head), the most abundant of which is a bedded magnetite-chert-silicate rock. This is intermingled to varying degrees with non-bedded to poorly bedded hematite-clastic quartz-chert rock. (1) In developing the Groveland flowsheet, Humphrey spirals were used because of their ability to produce a coarse size concentrate at a low cost. Tests run in the lab and pilot plant on the Groveland ore showed a. spiral weight recovery in the 27 to 33% range. (2) However as the ore body was opened up, it became evident that the most abundant ore type did not liberate at the coarse size fraction needed for spirals. Plant recoveries for the first five years operating the spirals only averaged from 10 to 15% weight recovery. In 1971 the weight recovery on the spirals had dropped to a low of 5%. To increase iron recovery and at the same time increase plant capacity with the least effect on the present operations the mill flaw-sheet proceeded from a spiral-notation circuit to an all flotation circuit to a magnetic separator-flotation circuit. During this changeover the concentrate grade also had to be improved to meet market demands. To do this the following changes evolved

Jan 1, 1978

By Yu-ren Jiang

Novel compounds of hydroxycarbamoyl carboxylic acids, including 2-hydroxycarbamoyl -3-methlcyclohexane carboxylic acid (HCMC), 2-hydroxycarbamoylmethyl decylen-4-yl carboxylic acid (HCDC) and 2-hydroxycar-bamoylmethyl tridecylen-4-yl carboxylic acid (HCTC) were synthesized. The flotation behaviors of diaspore, kaolinite and illite using these compounds as collectors were investigated. The flotation results revealed that the floatability of the minerals was, in descending order, diaspore and kaolinite, followed by illite, and the ability to collect the minerals was, in ascending order, HCMC and HCDC, followed by HCTC. At a dosage of 2×10-4 mol/L, both HCDC and HCTC show a strong ability to collect diaspore and a weak ability to float kaolinite and illite. The flotation recovery of diaspore is over 90% and that of kaolinite and illite less than 30%. Moreover, Fourier transform infrared spectroscopy (FTIR) spectrum and zeta potential, as well as the adsorption amount, indicated that the adsorption interaction between HCTC and diaspore was primarily a kind of chemisorption in the form of chelate rings, while the interaction between the collector and aluminosilicate mineral was dominated by physisorption. It was concluded that the novel hydroxycarbamoyl carboxylic acid compounds are highly selective collectors for the flotation of diaspore vs. aluminosilicate minerals.

Jan 1, 2011

By N. A. Abdel-Khalek

The objective of this investigation was to compare the results of counter-current flotation column versus conventional flotation cell for the cleaning of a sulphur ore and its discrete size fractions. The tests were carried out using diesel oil as a collector and alcohol as a frother. Generally, flotation column produced concentrates of higher grade. For example, the grade of a rougher concentrate for a given feed obtained with the column increased by 33% higher than that produced with the flotation cell. Such higher grade could not be obtained using conventional cell even after 3 cleaning stages. Moreover, the column gave superior results for flotation of fine and ultrafine particles.

Jan 1, 1991

By J. E. Garlanger

Each year, Florida mines approximately 40 million tons of phosphorite, most of which is ultimately converted to various commercial fertilizers. The phosphorite is mined from within the bottom portion of the 40-to 60-ft. thick surficial aquifer, using open pit mining techniques. The ore zone or "matrix" is generally overlain by relatively permeable fine sand to silty and slightly clayey fine sand. The ore itself is highly variable in nature but, in Central Florida, typically consists of less permeable clayey phosphatic sands and sandy clays with stringers of permeable sand. The ore is much more sandy in the north Florida mining area.

Jan 1, 1990

By C. P. Boddy

One of the most difficult tasks a project manager will have in pre-mine planning is dealing with the mine site's hydrologic regime. Numerous federal statutes passed by Congress in the last decade have made the coal mining industry more responsive to the environmental impact during all phases of the mining process. Permitting these mines has dictated a cradle to grave approach in dealing with water related issues. This paper will attempt to explore the application of these technology- based laws and their direct application to the unique environment found in Alaska's interior regions.

Jan 1, 1985