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By John C. Ferm

This report summarizes the results of an investigation of geologic factors that contribute to roof quality in underground mines in the Pocahontas #3 coal bed in southern West Virginia and southwestern Virginia. The investigation consisted of three phases, the first of which was selecting for detailed study five mines which represented the range of roof conditions, mining practices and geologic phenomena. This was accomplished by a statistical analysis of questionnaire data from 35 mines. The second phase consisted of detailed mapping in a "good" and a "bad" top area in each of the five selected mines and resulted in a qualitative assessment of geologic factors contributing to roof £alls. In the third phase, five mines were selected in which most non-geologic factors believed to contribute to roof falls were relatively constant but displayed a wide variety of geologic characteristics. Data from these mines were analyzed to test the qualitative assessments from Phase II. The primary result of this investigation is that slickensides arising from different geologic causes are the major geologic features leading to roof falls. These slicken-sides are found in ancient slump deposits, fire clays and sandy fire clays and channel contacts. Falls also occur where thick sandstone bodies directly over-lying the coal contain abundant shale or ironstone pebbles and coaly strips. The best top material is a sequence which consists of shale grading upward to sandy shale and sandstone within 30 feet of the top of the coal. Thick sand-stone units (301) without pebbles also provide good top.

Jan 1, 1978

By S. Banisi, M. Pourkani, A. Kargar, D. Hamidi, M. Sarvi

"Flotation circuit of the Sarcheshmeh copper mine, which processes 40,000 t/d of ore at 0.9% Cu, consists of rougher, cleaning and recleaning stages. To study the stage recovery of copper, one of the eight identical flotation banks was selected for sampling. It was observed that 17% of copper goes to the tailings and the lowest copper recovery was for particles coarser than 105 microns. The third five-cell unit of this bank not only did not recover appreciable amount of copper but also decreased the concentrate grade. The grade of the last five cells was lower than the feed grade suggesting that mixing of the concentrates of the three units was not beneficial. To alleviate the problem, the concentrate of the first four-cell unit was directed to the cleaners feed box (i.e., by- passing the regrind mill). This increased the overall copper recovery by 1.7 %. Also rather than the current practice of adding the reagents only at the feed tank (i.e., 100-0-0%), various combinations of stage addition were examined. It was found that 75-25-0% addition regime could increase the total copper recovery by 1.3%. At the recleaning stage in order to increase the final copper concentrate grade, a froth washing system was used. To ease the maintenance of the system, a shower type washing system was installed. The optimum wash water superficial flowrate was found to be 0.06 cm/s. Continuos sampling of the recleaners with and without wash water over a period of two months showed an increase of 1.5% in the final copper concentrate grade at constant overall copper recovery.INTRODUCTIONThe Sarcheshmeh copper ore body that may rank as third or fourth largest in the world contains 1 billion tonnes averaging 0.9% copper and 0.035% molybdenum. It is located in southeastern Iran and has been processing 40 000 t/d of ore since 1982. The mine produces 100 000 tonnes of copper and 2200 tonnes of molybednite concentrate per year.After three stages of crushing, the ore with a F80 (80% passing size) of 12.7 mm is fed to eight parallel ball mills in closed circuit with cyclones to produce a product 70% finer than 75 microns. The concentrate of rougher flotation circuit is ground; rougher tails constitute the main component of the plant final tailings. A concentrate with an average grade of 31% Cu (chalcocite and chalcopyrite) is obtained after cleaning and re-cleaning stages. Total recovery of the plant varies between 83% and 87%, depending on operating conditions and ore type."

Jan 1, 2001

By R. G. B. Pickering

Resue mining was practiced extensively in the East Rand in the first half of this century. The obvious benefits were that a higher grade of ore was trammed to the surface. The disadvantages were slow rates of face advance and an extremely labour-intensive method of mining. A collaborative effort by personnel from Avgold?s Target Gold Mine and Sandvik Tamrock has resulted in the design of a method of resue mining which can be totally mechanized. The purpose of this paper is to describe the methodology used and the proposed new mechanized resue mining method.

Jan 1, 2000

The objective in control blasting is to reduce overbreak to control the final pit wall slope, shaft, drift, ditch, bench, etc. to the final planned excavation limit.

Jan 1, 1991

By K. A. Lewandowski

Leach heaps are plagued with problems due to fine material migrating downwards, clogging the spaces between the coarser ore particles, resulting in decreased metal recovery rates. The use of a binder in agglomeration would allow the fine material to be immobilized, increasing the availability for the solution to come in contact with all the ore. As much as 1.64*1012 BTU per year would be able to be saved if a 25% increase in copper recovery was experienced, assuming that 70% of heap leaching units converted to using agglomeration technology. This is equivalent to approximately 18% of the energy currently being used in leaching heaps.

Jan 1, 2008

By P. Sharma

Longwall panels facing wind blasts have been investigated at both Newstan and Moonee Collieries in the Newcastle Coalfield under the Wind Blast Project of the School of Mining Engineering, The University of New South Wales. The fluid mechanics involved in the compression and expulsion of air during wind blasts have been defined and the overpressure and air velocity time histories utilized to determine the wind blast parameters and define the relationship between them. The characteristics of wind blasts pressure pulses in mine roadways show some similarity with transient phenomenon like air blasts from explosives and shock waves from the failure of pressurized vessels. The peak wind velocities in roadways, however, exhibit marked deviations from theoretical predictions based on compressible fluid flow and Rankine-Hugoniot relationships applied to goaf air displacement and panel geometry. Peak overpressures and wind velocities have shown a positive correlation with the areas of goaf falls, with the values peaking off for larger areas. The field investigations have helped the mine management in risk assessment and in evaluating the success of induced caving by hydrofracturing to mitigate the hazard.

Jan 1, 2004

By L D. Queen

"An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation. The Tifalmin porphyry copper district is located in the Star Mountains roughly 35 km north-east of the Ok Tedi mine. The district is a cluster of temporally and geochemically related porphyry copper deposits and their associated skarns. It covers an area of roughly 10 × 6 km and comprises at least three intrusions (Olgal, Futik and Kum Kom) where drilling has demonstrated the presence of porphyry copper mineralisation, and another three intrusions (Rattatat, Unfin and Pad 48) where porphyry copper mineralisation is strongly suggested by surface investigations and geochemical or geophysical signatures.Copper and gold mineralisation was first discovered in the area in the 1960s as part of the regional exploration program by Kennecott that located Ok Tedi. The mineralisation is associated with poly-phase porphyritic diorites, quartz diorites and tonalites that intrude the Upper Eocene to Middle Miocene Darai limestone and underlying Cretaceous to Eocene Feing Group. Radiometric K-Ar and U-Pb zircon ages cluster between 4 and 2 Ma for both the mineralised and post-mineral intrusions. In addition to porphyry-type mineralisation, variable amounts of calc-silicate and/or magnetite skarn occur near the contact of most of the intrusions. Mineralisation in the Olgal and Futik porphyries occurs as chalcopyrite replacement of mafic minerals and in a stockwork of micro-veinlets. Mineralisation mostly occurs in the intrusions but is also observed in hornfelsed siltstones and sandstones of the Feing Group on the southern margin of the Olgal intrusive body. Quartz veining is uncommon. Secondary magnetite is commonly associated with the highest copper and gold grades. Sericitic alteration is most prominent in the mineralised hornfels. Structurally, the area is dominated by syn- to post-mineralisation thrusting from the north. At least one of the deposits (Olgal) appears to be truncated at depth by a thrust fault.CITATION:Queen, L D, 2015. The Tifalmin porphyry copper gold district, Star Mountains, Western Papua New Guinea , in Proceedings PACRIM 2015 Congress, pp 357–360 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Mar 18, 2015

By J. R. Hart, V. C. Smith, V. P. Babarovich-Hansen, J. H. Schubert

"An introduction and case study of Mintek's latest generation of Milling Circuit controller, the MillStarTM with IES (Interpreting Expert System) is presented. This paper also describes as a case study its installation on a Southern African base metals concentrator plant which has demonstrated improved operation, greatly reduced variation in product quality, as well as increased capacity and availability of the circuit.In this case study, the MillStarTMw as installed along with several other enhancements, including hydrocyclone underflow monitoring, product particle size estimation and flotation circuit stabilisation. MillStarTM utilised the particle size measurement to control grind to within 2% of setpoint, all the while ensuring sump level control, density control as well as mill power optimisation. Mintek' s expertise in the modelling of tumbling mills allowed the development of a specialised optimising controller that ensured the mill was kept in the most efficient operating region.MillStarTM augments the automatic control with useful user interface features, such as on-line trends of any logged variables, historic data warehousing and retrieval and production summary reports.A novel feature that distinguishes the MillStarTMi s the IES (Interpreting Expert System). This is a feature of the software that works with the automatic control code to generate explanations for control actions made by the controller. Often complex control systems are mysterious in nature, reducing trust and eventually resulting in them being switched off. The IES method of exposing the heart of the controller has proved popular and is seen to have potential as a training tool. 314"

Jan 1, 2000

Some crystal chemical properties of several apatites from Brazilian deposits and occurrences were determined; the samples had different origins (igneous, metamorphic and sedimentary). The aim was to establish relationships to their froth flotabilities behaviour. Detailed X-ray diffraction studies were carried out, to determine precisely the unit cell parameters and the degrees of crystallinity of the apatite samples. A thorough chemical analysis of these minerals was also conducted. Microflotation studies were carried out, using sodium oleate and dodecylammonium chloride as collectors. The crystallinity levels were determined for all samples studied, with the establishment of a crystallinity index, which varied from 10.0 to 4.5, on a normalised scale. A marked dependence was observed between this crystallinity sequence and flotability; for both collectors, flotability dropped drastically for crystallinity indices lower than 8.0. Also, the low crystallinity apatites showed a decrease in floatability, for longer conditioning times.

Jan 1, 1993

By William R. Archibald

This report presents the results of a program demonstrating the technical and economic feasibility of hydraulic borehole mining of uraniferous sandstone. This mining technique involves the drilling of a borehole from the surface down into the ore body. A borehole mining tool consisting of a high-pressure cutting jet and a slurry jet pump is lowered down the borehole. Water from the high-pressure cutting jet cuts into the sandstone formation generating a water-sand slurry which flows to the slurry jet pump and is pumped to the surface. The cutting jet is rotated and moved up and down from the surface, thereby generating a cavity in the formation. When the cavity size reaches a point beyond the effective radius of the cutting jet, the mining tool is moved to an adjacent borehole and the process is started again. This program was carried out in four phases. Phase I included modifications to the borehole mining system (built and tested under a previous contract, No. H0252007, for the mining of coal) and factory tests of the jetting system. Phase II covered the field tests and a public demonstration. Phase III involved nozzle and equipment modifications and factory tests. Field testing continued under Phase IV along with an evaluation of this method. The field tests were conducted successfully on Rocky Mountain Energy Company's Nine Mile Lake property near Casper, Wyoming. Results indicate that; hydraulic borehole mining of uranium is technically feasible (300 tons of uraniferous sandstone were mined from a depth of 90 ft) and the hydraulic borehole method is economically attractive. However, additional testing is required to clearly define the economic feasibility.

Jan 1, 1978

By Gerasimenko MP

The authors consider the features and advantages of a new promising trend in mineral processing - column flotation. Design aerohydrodynamic and process characteristics of pilot-scale columns are studied. Optimum process parameters for various ores and operations are studied. Flow patterns in cells of various cross-sections are investigated and the relationship between Peclet criterion and specific capacity of cell is determined. On this basis options were developed for optimisation of hydrodynamic characteristics of large capacity cells, including longitudinal sectionalising of the collection zone. It is shown that advantages of high column cells (10 - 13 m) produced in the West can be implemented in a sectionalised machine of smaller height (5 - 6 m). Investigations on the influence of slurry and air bubbles flow patterns indicated that to increase flotation rate of particles of various flotabilities it is expedient to combine co- and counter-current flows of phases in a single column. The proposed designs allow tailings to repeatedly return tailings to the collection zone and alternate co- and counter-current slurry and air flows. For the first time columns with alternating slurry flows (combination of co- and counter-current flows) are developed and commercialised. Unique multi-sectional cells having a number of essential advantages compared to monocells are developed. Design features of the cell allow: implementation of different flotation process flowsheets in a single unit, improved beneficiation process results, reduced specific power requirement, and elimination of sludging of the aerator and cell bottom part. Operational problems restricting introduction of column cells are connected with the aeration unit. The new cells are supplied with pneumohydraulic and stepped airlift spargers, allowing control of bubble size distribution with a service life of not less than three years. Successful operation of the new generation of columns with cell volumes of 15, 25, 40, and 80 cu m and the height of 5 - 6 m in copper, molybdenum, antimony, tungsten and nickel ores and coal slurry processing at different plants in this country illustrated their considerable process and operational advantages and verified the proposed options and relationships. A simulation technique of the flotation process and models for the solution of various problems are developed. A modular model of column flotation is constructed taking into account the kinetic relationship of the process and the hydrodynamic regime in the cell. Based on this, a technique of calculating the column parameters based on laboratory tests is developed (scaling-up), and applied for calculating the design of commercial-scale columns is developed.

Jan 1, 1993

By J. Jerome MacDonald, Stephen B. Flewelling

"In the base metals industry of today, operations required to continuously improve productivity and reduce unit production costs. Consistent with these goals is the improvement of the muck haulage cycle from open stopes with the No. 12 Mine.Existing haulage from open stopes occurs with LHD units experience at Brunswick indicates that LHD units have a higher unit cost per tonne mucked than loader-truck combinations when the haulage exceeds approximately 150 m.The reintroduction of open stoping at Brunswick made existing loader-truck combinations unsuitable. It was recognized that application of recent continuous loading technology could potentially result in significant unit cost savings and productivity increases.A review of the technology resulted in the testing of the Voest Alpine AL60 loader in the open stopes located on the 1000m Le vel. Testing began in January 1988 and finished in late November 1988. The test will provide experience as well as quantified production and maintenenace parameters f or the purpose of quantifying production unit costs and system productivity, ultimately resulting in a decision to further utilize or not utilize the loader as a primarymucking unit at Brunswick .INTRODUCTIONBrunswick Mining and Smelting Corporation Ltd. produces approximately 10500 tonnes per day from the No. 12Mine located 35 km southwest of Bathurst, New Brunswick Mining is the largest (single mine) producer of zinc in the world having mined over 52 million tonnes since 1964. Many years of production still remain with proven and probable reserves in excess of 90 million tonnes. Reserves grade at an average of 3.65% Pb. 8.99070 Zn, 99g/mt Ag and 0.32% Cu . The operation employs 1446 people with approximately 800underground including operations and maintenance staff.In the No. 12 Mine two major mining methods are presently utilized . The primary method is Mechanized Cut and Fill (MCF) accounting for approximately 6500 tonnes per day from four major areas. The remaining 4000 tonnes per day is produced from open stopes with the majority of 2800 tonnes per day being produced by the 1000 level.The 1000Level contains 30 million tonnes segmented into open stope blocks 30 m high, 15m and 30 m on strike , with a transverse width of 10m to 50 m (Fig. I). Stopes are organized on a primary secondary stope arrangement along strike. Primary stopes are undercut and overcut full size with the back fully cablebolted. All stope and lateral development is driven 4.0 m high. All stopes to date are flat bottom (Fig. 2). Primary stopes are filled with cemented rock- fill , secondaries with conventional rockfill, Haulage from the stopes is primarily with 8-yard LHD units ; mucking is approximately 50% conventional and 50% remote. Haul distance to date has averaged approximately 150m with current stope haul distances of approximately 300m. Haulage is presently to a small jaw crusher situated on the 1000 Level. From August 1989on, haulage will be to a set of ore passes feeding to a large gyratory crusher situated on the 1200 Level. Start up of the gyratory crusher corresponds to an increase in tonnage from 1000 Level to 4100 tonnes per day."

Jan 1, 1989

By W. I. Nissen

This Bureau of Mines report presents the design for a commercial-sized flue gas desulfurization (FGD) demonstration plant that uses the citrate process. The goal of the Bureau's citrate process is to minimize the undesirable environmental impacts of industrial plants emitting SO2-bearing gas. The FGD plant is located at the George F. Weaton powerplant, Monaca, Pa. Construction was completed in April 1979 and will be followed by preliminary testing and a 1-year testing and evaluation program. Design capacity of the FGD plant is 156,000 scfm of 0.2-volume-percent-SO2 flue gas yielding about 16 tons of sulfur per day. The plant is intended to (1) clean fly ash, SO2, and Cl- from the gas while cooling the gas in a venturi scrubber, (2) absorb SO2 from the gas using 1,200 gpm of a countercurrent-flowing citric acid-sodium citrate-sodium thiosulfate solution, (3) react the absorbed SO2 in two 13,000-gal stirred closed vessels with added H2S, thus precipitating elemental sulfur and regenerating the citrate solution for recycle, and (4) recover the sulfur from the slurry by air flotation, followed by melting in a heat exchanger and separation from the occluded citrate solution in a sulfur decanter at 35 psi and 135° C.

Jan 1, 1979

Myths and misconceptions regarding cyanide in the mining industry are common. This is not surprising, given the other uses to which cyanide is, and has been put. Unfortunately, the economic benefits, largely due to the introduction of cyanide in towns such as Waihi are often unrecognized. In the current "green" environment, the New Zealand mining industry faces constant scrutiny and challenge. The use of cyanide in hard rock mining is one area which can stir the emotions of the public.   Cyanide can be transported, stored, used and disposed of in a safe manner. The modem mining industry has an excellent record regarding the safety and environmental issues relating to cyanide. Many of the fears relating to cyanide in the mining industry have proven to be unfounded. It is important that this reputation is maintained. This is possible given proactive and ongoing management of cyanide issues.

Jan 1, 1997

By Robert Merritt

Approximately 2,500 tons of low-grade uranium ore were used to conduct a field-scale heap leaching test as well as throe column leach tests. The ore used to construct the field-scale heap leaching test had a grade 0.023% U308 while the ore used fur column leach tests had a grade of 0.027% U308. The ore was crushed to minus 4 in. The field-scale heap was constructed on a clay base with parallel effluent lines running under the heap spaced 7.5 ft apart. The heap was leached for 45 days, drained, washed, and dismantled. Based on residue analysis, approximately 205 tons of ore were leached with a final U308 extraction of 96%. The heap was 17.9 ft in height and leached in a near-vertical pattern down to approximately 1 ft, at which point the leached zone become bell-shaped. Two 8-ft column charges and one 18-ft column charge were leached for 36 and 45 days, respectively. Based on residue analyses, all U308 extractions were greater than 96%. The conclusions generated by the studies showed that (1) flow rates were similar for column leaching and field-scale leaching, (2) final U308 extractions were the same for both types of leaching; however, the deeper ore beds tended to require longer leaching times, and (3) the reagent requirements were reduced with increased bed depth.

Jan 1, 1980

By William H. Shockley

THERE was an, extensive mining and industrial exploitation of Russia, about 20 years ago, by Belgian, French and British capitalists; but the results were discouraging. It is said that the Belgian and French public together lost approximately $400,000,000, chiefly in the coal and iron industry of southern Russia. This loss is largely accounted for by the fact that most of the companies were formed, as usual in " boom " times, to sell stock; and that thorough investigations were not made. In 1906, renewed interest was taken in Russian mines, especially in those of gold and copper, and a considerable literature is growing on this subject; but as yet I have seen no account of one of the large Russian mining estates. The following notes on such property, where the conditions are so different from those in the mining regions of the United States, should be of interest This inadequate and probably somewhat inaccurate account is made up, partly from my own notes made on the ground, in the summers of 1904 and 1905, and partly from a volume on the geology and resources of the estate issued several years ago, containing surface-and mine-maps, and from a pamphlet published in 1904, giving a general account of the condition and prospects of the property. This pamphlet includes a report made by Arthur L. Pearse, a member of this Institute, in 1899. Both these books are in Russian, and were prepared by the Bogoslovsk Mining Co., which owns and operates the property. Articles in many other languages have appeared dealing with various matters connected with the estate; but I know of no other general account of it. In numerous books and articles on the Urals and Siberia, however, localities are described in which somewhat similar conditions prevail. Recent papers by

Mar 1, 1908

By AIME AIME

AT the annual .meeting of the Minnesota Section in December, Dr. Leith characterized as a question still open the exact method of origin of Lake Superior iron ores and emphasized it as an important practical problem for the future. He welcomed the step taken last year by Dr. Gruner in presenting a new hypothesis, and asked Lake Superior geologists to meet the challenge of such multiple hypotheses by unbiased, study.

Jan 1, 1932

By Martin Plante, Denis Hamel, Marie-Pier Bédard, Jeff McLaughlin

"The results and methodology for the optimisation of the zinc flotation circuit at the Matagami Mill using the 6-Sigma methodology are shown in detail while only the base line and final results are presented for the copper circuit. As a result of the six sigma project, the zinc concentrate grade was improved by 2% and the zinc recovery by 3%, with reduced dosage of reagents while the copper concentrate was improved by 2.4%. The process variation in the zinc concentrate grade and recovery was also reduced by over 50%. The main factors contributing to the project success were identified to be improvement in the plant measurement systems, an internal benchmarking exercise and statistically designed laboratory and in plant tests to guide the process improvement. The annual value to Noranda resulting from this project has been estimated at approximately $5 million.INTRODUCTIONThe Matagami concentrator started ore processing from the new Bell- Allard deposit in July 1999. Concentrator performance (grade and recovery of zinc and copper concentrates) was highly variable and work was ongoing to fix the sources of variation. In the mean time, Noranda embarked on its 6-Sigma initiative across the company. The 6-Sigma DMAIC (Define, Measure, Analyse, Improve, Control) methodology was used in the two flotation circuits to reduce variation and improve performance. The statistical 6-Sigma methodology moves towards identifying and controlling critical process operating variables (the X' s) to achieve customerspecified performance (the Y's). The goal of the methodology is to describe a process in a numerical equation known as the transfer function, given as Y = f (XI, X2, X3.. .), where Y could be concentrate grade or recovery and examples of X's are reagent dosage, particle size, etc. With the transfer function, optimal operating conditions and control tolerances are established for the key X's to ensure the desired Y performance is obtained.Each phase of the 6-Sigma methodology is explained below with examples from the Matagami project."

Jan 1, 2001

By Teele Tuuna, Voglers Eesti OÜ

Blasting operations are a strong and important part in any kind of building and mining industry, a shot-firer profession is one of the oldest professions in the civilized world. It still serves an important and ongoing purpose in the modern economy all over the world. Europe, nowadays, values strongly the mobility of workers, the movement of know-how and labor force between different countries inside Europe. There are no borders for most of the everyday professions.

Jan 1, 2018

By Mills R, Shannon W. T, Finch J

As an extension to investigations into methods of upgrading the Iitania content of ilmenite concentrate from beach sands using 100 g/day, a small pilot plant treating 23 kg/ day of ilmenite was used.In the pilot plant the iron oxide in the ilmenite was reduced to iron at 1200°C using carbon (coal/char) without melting the material, using a tunnel kiln heating cycle, the charges contained in saggars, The iron was then separated from the titania fraction by ball milling, wet magnetic separation, and screening. The resulting magnetic product was high in iron (about 80 per cent Fe) and the non-magnetic product high intitania (about 70 per cent TiO2).From this trial sufficient data were obtained to enable the preparation of a flow sheet for a plant producing 62,300 tonnes/year total products, and an estimate of costs to be made. This has shown that for an estimatedcapital expenditure of nearly $6,000,000 and annual production costs of about $4,000,000, the cost per tonne of high titania product (76 per cent TiO2) would be nearly $70.

Jan 1, 1977