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By D. M. Doyle, P. J. Hannon

"The Canada/Nova Scotia Mineral Development Agreement (M.D.A.) is a subsidiary of a five-year Economic and Regional Development Agreement, signed on June 11, 1984. The objectives of the M.D.A. are to strengthen and diversify the Nova Scotian mineral industry and to coordinate federal and provincial policies and programs. The agreement outlines the activities and responsibilities of each government, including the division of the cost (60:40, federal: provincial) between governments. The agreement established a framework for joint planning to avoid overlap or duplication of programs.The M.D.A. programs are broadly divided into five sectors (Table 1 and Fig. 1): Geoscience, Minerals Technology, Development, Firm Specific and Public Information. Each sector is administered by a subcommittee composed of one federal member and one provincial member. The subcommittees report to a Management Committee. The Canada/Nova Scotia Mineral Development Agreement (M.D.A.) Management Committee is made up of Dr. Jack Reeves from EMR and Dr. Richard Potter from the NSDME.The Geoscience subcommittee, co-chaired by Peter Giles (NSDME) and William Poole (EMR), manages the regional geologic, geochemical and geophysical studies in the province. Included in this sector are studies into mineral deposits and the basic geological relationships between mineral deposits.The Mineral Technology sector is co-chaired by the authors, Don Doyle (CANMET) and Pat Hannon (NSDME). This sector of the M.D.A. conducts studies into mineral processing, sampling, deposit byproduct commodities and project feasibility analysis."

Jan 1, 1987

By S C. Labram, A Stabile

"Modern train load out stations are becoming faster and have the ability to load a 100 wagon train in less than three hours. At 70 metric tonnes per wagon, that is an average fill rate of 2333 t/h and considering the train is continuously moving, the instantaneous feed rate into the wagon is in the region of 4200 t/h.The main feed conveyor that maintains the level in the surge hopper, above the final weigh hopper, operates in excess of 5000 t/h to ensure uninterrupted charging of the weigh hopper to not disrupt the train loading operation.To avoid demurrage costs (a charge by the Rail Operating Company for exceeding the contractually agreed loading time) the train has to be handled speedily and efficiently and spend the absolute minimum time at the loading facility. Everything else has a much lower priority, including sampling.Thermo Fisher Scientific’s brief was an engineering, procurement and construction management (EPCM) contract to provide a sampling scheme adjacent to the prepurchased and predesigned load-out facility, but which interfaced seamlessly with the loading system of the trains and provided the following: an instantaneous particle size distribution (PSD) result for each truck loaded, and a final PSD for the total train prior to the train leaving the loading facility a crushed (-3 mm) composite sample for chemical analysis for the whole train a composite physical analysis sample for the whole train.The design presented some interesting challenges to our team, but the end result is a system that not only ‘stands alone’ but interfaces seamlessly with the loading system. This paper discusses the design itself and potential improvements to future systems.CITATION:Labram, S C and Stabile, A, 2017. An automatic sampling methodology to determine contractual particle size distribution in rapid train load out stations, in Proceedings Eighth World Conference on Sampling and Blending, pp 229–234 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

May 9, 2017

By Creelman C. A

The removal of a horizontal piIlar between Nos. 9 and 11 levels (1,116 feet from the collar of main shaft) was successfully carried out by blast hole drilling, and, to the best of the authors' knowledge, this mass blast of 40,000 tons of ore (requiring the use of 7,150 Ibs. of explosive and 5,385 feet of primacord) was the first occasion on which a task of this magnitude has been carried out underground in an Australian mine.Blast hole drilling was introduced by this company approximately fifteen months ago at New Cobar Mine in sub-level stoping.This experimental drilling and blasting gave very valuable information as to the correct technique to be adopted when the final plans were being prepared for the removal of the horizontal pillar at New Occidental Mine.The dimensions of this pillar were approximately 15 feet long, and from 40 to 60 feet wide, and 65 feet in depth.The block of ore extended to 35 feet above the level and 30 feet below, the upper section being partially excavated to form the bell-hole floor and grizzly chambers used for drawing off the broken ore in the shrink stope above the No. 10 level.Description of Ore-bodyThe main ore-body is situated almost vertically in a well-defined shear zone at the junction of sandstone and slate, the sandstone forming the eastern and the slate the western wall.The ore-body consists of highly silicified slate (chert) sparsely impregnated with sulphides, containing a small percentage of copper and bismuth.The total length of the ore-body is approximately 450 feet, with a maximum width of 60 feet in the main section, which is about 200 feet long.This is the section shown in longitudinal and cross-section (Fig. 1).The ore-body is divided by a horse of mullock in the northern end into two legs, the eastern and western, which are approximately 250 feet long and averaging 25 feet in width.This ore-body is strong, hard and abrasive with extremely strong walls, an ideal proposition for open stoping and mass blasting.MiningIn all levels above No. 9 the cut and fill method of ore extraction was adopted. On Nos. 10 and 11 levels the shrinkage method was introduced with economical results, and no difficulty had ever been experienced with shabby or bad walls.Shrink stoping had been carried upwards from No. 11 level to within 10 feet of No. 10 level, when mining hazards developed, and instead of...

Jan 1, 1944

By W. S. Steinberg

Keywords: Pyrometallurgy, furnace, open slag bath, Evraz, Highveld In 1963, the full-scale engineering study for the development of an Iron, Steel, and Vanadium plant was initiated. This was the birth of Evraz Highveld and its process flow as it is known today. Highveld started out as a steel works that produced steel and vanadium-bearing slag by using a titaniferous magnetite iron ore mined from its own mine site near Steelpoort. The process utilized co-current rotary kilns for pre-reduction, and submerged-arc furnaces for the production of pig iron with approximately 1.15% vanadium. Because of various technical challenges that the submerged-arc furnaces experienced over the years, a decision was taken in 2004 to convert the furnaces to Open Slag Bath (OSB) operation, in order to enable the maximum production of iron and vanadium while improving the control over operational parameters. The conversion proved to be highly successful and, since 2005, three furnaces have been converted to OSB operation. The fourth furnace will be converted in the first half of 2011. The conversion to OSB furnaces not only yields higher vanadium recovery to the metal at the iron plant, but also improves the overall throughput of the steel plant. This is the result of a higher carbon content of the liquid iron from the OSB furnaces, eliminating the need for anthracite addition to the shaking ladle process. This reduces the blowing time by approximately 60%. The conversion of the furnaces to OSB configuration is seen as an enabling technology change and a prerequisite to the implementation of other plant improvements that are envisaged. The long-term vision of EHSV is to convert five furnaces to OSB mode, and permanently decommission the two remaining submerged-arc furnaces. This will allow EHSV to maintain current steel and vanadium production but at a significantly improved overall efficiency.

Jan 1, 2011

By J. W. Paul

The United States Bureau of Mines has made a study of roof control in 68 mines operating in the Pittsburgh coal bed in Pennsylvania, West Virginia, and Ohio. In making these studies, much information was also obtained on various exact methods of development and pillar extraction as practiced by this large group of mines. It is proposed in this paper to describe these methods, by which this group of mines in 1931, a year of reduced industrial activity, produced 28,123,000 tons of bituminous coal - 7.4 percent of the national production for the same year and an amount equal to the combined production of the coal mired in the States of Alabama, Colorado, New Mexico, Utah, and Wyoming. The results of these studies, insofar as they pertain to the prevention of injuries from falls of roof, have been published by the Bureau of Mines. In this paper only a few mines typical of each system will be described; however, should the reader be interested in other details with respect to these mines, he may refer to the following publications: T.P. 522, 1932, 43 pp. Falls of roof and coal in mines operating in the Pittsburgh coal bed in Marion and Monongalia Counties, W. Va., by J. W. Paul and J. N. Geyer.- A study of nine mines in these two counties of northern West Virginia.

Jan 1, 1935

By Robert L. Lappi

Volume I of this report described the demonstration of cross-ridge mountaintop coal mining. This document, Volume II, is intended for use as a supporting document to Volume I. This report describes the MATHMINE Mine Simulation System, a modularized software system used to support the demonstration of cross-ridge Mountaintop coal mining. The MATHMINE Mine Simulation System was used to generate mine plans for coal extraction at lowest cost. The functional details, data requirements and output of the software system are detailed here. The computer program was written in FORTRAN. One module which generates resource location data required the use of SAS?.

Jan 1, 1984

By G. W. Jones

In cooperation with the Edison Electric Illuminating Co. of Boston, Mass., and the Boston Consolidated Gas Co., a survey has been conducted continuously since 1929 by the Bureau of Mines to determine the general hazards of combustible gases and vapors in manholes of the Edison Electric Illuminating Co. The New England Telephone & Telegraph Co. also has cooperated in the investigation, and tests have been made in its manholes since May 1, 1933. Yearly reports6/ on the results obtained since the beginning of the survey have been published. The method and procedure of conducting the survey in 1936 were the same as heretofore.

Jan 1, 1938

By G. W. Jones

In cooperation with the Edison Electric Illuminating Co. of Boston, Mass., and the Boston Consolidated Gas Co., a survey has been conducted continuously since 1929 by the Bureau of Mines to determine the general hazards of combustible gases and vapors in manholes of the Edison Electric Illuminating Co. The New England Telephone & Telegraph Co. also has cooperated in the investigation, and tests have been made in its manholes since May 1, 1933. Yearly reports6/ on the results obtained since the beginning of the survey have been published. The method and procedure of conducting the survey in 1936 were the same as heretofore.

Jan 1, 1938

By R A. Stevenson

The purpose of this paper is to outline a recent development in ground fault protection. The paper deals with the A problems of both grounded and ungrounded systems, with regard to hazards to personnel and equipment, and shows that the increased sensitivity and speed of operation of this new detector results in increased safety to personnel and reduced stresses on equipment, and permits the detection of leakage currents before they become serious faults. The detector can be applied to all distribution systems, either A.C. or D.C., up to 15 kv., and the detector itself can be tested at regular intervals without interrupting line circuits. T HERE is always controversy among engineers as to whether distribution systems should be grounded or ungrounded. There are many experi-enced engineers on both sides of this controversy, and each have valid reasons for their choice

Jan 1, 1964

By G. Prabhulingaiah, D. M. R. Sekhar

"A phosphate deposit of estimated 6 billion tons of P2O5 is located 180 to 300 meters below the sea level, 185 kilometers north of port Luderitz, Namibia. The ore bed is 6 kilometers wide and 220 kilometers long. The ore deposit analyzing 16% P2O5 is loose sediment of 1.5 meters thick without over burden. Sizing of the ore to - 1.4 mm + 0.150 mm improves the grade to 23.54% P2O5 with a P2O5 recovery at 83.75% containing phosphate pellets and flakes of sea shells. Clay, bacterial debris as organic matter, flakes of broken sea shells and fine phosphate containing particles report to - 0.150 mm fraction and + 1.4 mm fraction contains sea shells and few phosphate containing pellets. The mineralogical composition of phosphate containing pellets is estimated (Droste, T.C., 2004) 80% frankolite, 6% pyrite, 4% illite, 3% quartz, 3% amorphous SiO2, 1% carbonates, 1% organic carbon and = 1% ilmenite, magnetite, zircon etc. The phosphate content in the pellets is around 27.7% P2O5. Trials to separate by gravity or flotation processes were not successful. Leaching of carbonates (sea shell flakes) under mild acidic condition (= 2% acetic acid or HCl solutions) or calcinations followed by leaching however yielded residues (concentrates) of grades 27.85 - 32.74% P2O5 at P2O5 recoveries ranging from 78.45 - 96.8% P2O5."

Jan 1, 2014

By Whitman E. Brown

"As part of a Nation-wide program of the United States Government to develop natural resources, funds were made available to the Bureau of Mines for the investigation of the ""Reduction of Iron Ores by Solid and Gaseous Fuels."" These investigations are concerned with the direct production of iron from its ores without the use of a blast furnace. Many processes have been proposed for the low-temperature reduction of iron oxides, but few have been developed to commercial use. High-quality sponge iron is produced com-mercially only in Sweden.The cost of iron production depends in part on the grade of the ore and increases with decrease in grade. The substitution of cheap natural gas in a sponge-iron process for expensive coke in smelting in the blast fry pace would partly compensate for the use of low-grade ores. If a process can be developed in which inexpensive natural gas can be used as fuel and reducing agent, the large deposits of low-grade iron ore and the natural-gas resources of east Texas may be utilized in the commercial production of sponge iron.Sponge iron may possibly be used to supplement scrap iron, pig iron, or, in special instances, wrought iron. Sponge iron will vary but slightly in composition and will contain little or no nickel, cobalt, manganese, copper, and other allowying metals. It is becoming increasingly difficult to obtain choice scrap iron of nonalloy composition. Pig iron contains 1.5 to 3.0 percent silicon and 3.0 to 4.5 percent carbon. Sponge iron generally contains less carbon than pig iron.The Madaras Steel Corp. in Longview, Tex., built a set commercial-scale sponge-iron plant which was intended to be operated in connection with a steel foundry. The process is based upon the reduction of preheated iron ore, with hot reducing gases under pulsating pressure. Although the Madaras Corp. had previously experimented with the production of sponge iron by reducing the ore with pure hydrogen, several attempts to use a re-formed natural gas as the reducing agent in this pilot plant were unsuccessful. The Bureau of Mines leased this plant and conducted an extensive investigation of the Madaras process on a semicommercial scale.The experiments were not entirely satisfactory, partly because of mechanical difficulties and partly because the reducing gas was not of the best quality. However, much information on the production of reducing gas by re-forming natural gas and on the reduction of iron ore with re-formed natural gas was obtained.The sponge iron produced contained approximately 0.1 percent phosphorus; and, after melting in an electric furnace, the iron contained approximately 0.15 percent phosphorus. The results of attempts to improve the ore by ore-dressing methods were not encouraging."

Aug 1, 1946

By F. T. Wang, Y. Yuan, S. H. Tu, X. G. Zhang

Hydraulic support instability (HSI) is one of the most common causes of disasters in underground coal mining, posing a threat to the safety of mine workers and normal operation of the equipment. It is prone to occur in fully-mechanized mining faces with a large mining height (FMMLMH), especially when the dip angle of the coal seam is large. The key to controlling HSI is to deduce its mechanism and employ effective control techniques. This paper focuses on the analysis of HSI types, the key parameters and techniques to control HSI in FMMLMH, the establishment of a model of HSI in FMMLMH, and a multi-parameter sensitivity mechanical model of different HSI forms in the no. 7219 longwall face in Xutuan Coal Mine, Huaibei Mining Group, by using sensitivity analysis . The results show that HSI mainly presents in three forms: hydraulic support gliding (HSG), hydraulic support tilting (HST), and hydraulic support tail twisting (HSTT). The occurrence of the above three forms depends mainly on support anti-instability capability. In the no. 7219 longwall face, HSG and HST are the main two forms of HSI. The dip angle of the working face and the friction coefficient between floor and hydraulic support are the sensitive parameters for HSG, while HST is strongly dependent on the dip angle of the working face and the friction coefficient between roof and hydraulic support. By the applications of measures such as the oblique layout of the working face, cutting the floor into a step pattern, moving the support under pressure, and raising the setting load, the support stability was controlled effectively.

Jan 1, 2015

By R. L. F. Kay, T. A. Fletcher, A. J. Boyce, C. M. Rice

Paper presented at Mineralisation in the Caledonides, the Mike Gallagher memorial meeting held in Edinburgh, 27-28 June 1996. The synorogenic, mafic Ordovician intrusion comprises orthopyroxene-rich xenoliths and mafic-ultramafic cumulates with low-grade pyrrhotite, chalcopyrite and pentlandite mineralisation. Despite abundant textural, mineralogical and O isotope evidence for significant assimilation of country rocks, S isotopes indicate only a modest uptake and suggest that mineralisation is dominated by magmatic sulphur. The preferred explanantion is that the magma was close to S saturation on emplacement. Uptake of elements such as Si and O led to early crystallisation of orthopyroxene, resorption of olivine and early segregation of a sulphide liquid. Post-magmatic interaction with a high-temperature, surface-derived fluid, whose movement may have been controlled by regional shear zones, reset biotite and whole-rock O isotope values but not those of orthopyroxene. This model identifies the presence of olivine-rich rocks and evidence of extensive assimilation as guides to prospectivity

Jun 19, 1905

By R Thompson, D Tulloch

"Wet-weather or 24×7 wet-trafficability is a commendable design objective for unpaved mine haul roads, but practically often difficult to achieve, due both to the road sheeting material selected, and also its breakdown under the action of traffic and environment. Accidents allied to wet-trafficability issues, either due to rain or often more commonly, over-watering of mine roads, is increasing and especially the loss of control or uncontrolled vehicle movements categorisation. The dominant system failure is that once vehicle control is lost, control cannot be regained.The friction supply, or ‘skid resistance’ of a mine haul road sheeting is critical to the safe operation of the road and data has to be collated concerning the variation of friction supply with sheeting material engineering or indicator properties. In the light of the paucity of data existing in this field, this paper presents the contributory research that informs the development of improved and validated unpaved mine road friction supply estimates, based on a number of mine site evaluations across a range of sheeting material types and characteristic parameters, leading to improved predictive models of friction supply.Results from this preliminary review will enable operators to assess their road sheeting materials, improve and validate unpaved mine road friction supply estimates. This will contribute to improved road design, management and 24×7 operability decisions and contribute to a reduction in heavy vehicle loss of control incidents on gravel-surfaced mine haulage roads.Citation:Thompson, R and Tulloch, D, 2016. What's Stopping You? The Role of Sheeting Material Selection in Safe Haulage Operations, in Proceedings Ninth AusIMM Open Pit Operators' Conference 2016, pp 253–264 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Nov 15, 2016

Jan 1, 2003

By M Hourn, D Turner, P Voigt, D Mallah

The GPM Gold Project is located in Armenia, and consists of an open cut mine at Zod, near the Azerbaijan border, and a CIL processing plant at Ararat near the Turkish border. Mining at Zod commenced in 1976, and focused on near surface oxide ores, which overlay deeper refractory sulphides. Historical mining has now almost depleted the oxide ores, and the sulphide content of ore delivered to the processing plant at Ararat is increasing. Gold and silver recoveries through the Ararat plant were declining steadily.GeoProMining, the owners of the project, have expanded the Ararat facility to deal with the increasing sulphide content in the ore. In 2014, GeoProMining refurbished an existing concentrator on site to recover a sulphide concentrate from the ore, and have constructed an Albion ProcessTM plant to oxidize the refractory concentrate. Glencore Technology (GT) provided the Albion ProcessTM plant as a technology package.In July 2014, the progress of the GPM Gold Project Albion ProcessTM Plant was reported (Hourn, Voigt and Turner, 2014). At the time of writing, the construction of the plant was nearly complete. This paper presents an update of project progress, covering commissioning and ramp up of the GPM Gold Project.The commissioning phase occurred over June to August 2014, which included an M3,000 IsaMillTM fine grinding plant, a 6 tph limestone milling plant, a 60 tpd vacuum pressure swing adsorption oxygen plant, a 10 m residue thickener and a 12 tph Albion ProcessTM oxidative leach plant. Since commissioning was completed, ramp up occurred over the following three months with downstream gold recoveries from cyanide leaching reaching over 98%.CITATION:Voigt, P, Hourn, M, Mallah, D and Turner, D, 2015. Commissioning and ramp-up of the Albion Process at the GPM Gold Project, in Proceedings MetPlant 2015, pp 207–219 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 7, 2015

8053. Andreev, B. A. Zakashanakii, M. S., Samsonov, N. N., and Fotiadi, E. E. A Textbook of Gravitational -Surveying, (in Russian). Gosgeolizdat, Moscow-Leningrad, 1941, 432 pp. This textbook was written at the suggestion of-the faculty of geophysical exploration of the Leningrad Mining Institute. It contains two parts: (1) Theory and technique, and (2) Applications to geology and exploration. The features bringing; the text up-to-date include: (a) Description of the principal modern gravimeters; particular attention to the method of simultaneous oscillation of 4 pendulums; (b) discussion of the aperiodic oscillation of torsion balances; outline of the construction of variometers Z -20 and Z-40; (c) material for aid in interpretation, such as reference-charts, formulas with their derivation, atlas of theoretical curves, and examples of interpretative calculations; (d) considerations in the planning of surveys and the densities of various rocks; and (e) detailed analysis of significant examples. The text is provided with a bibliography, V. S.

Jan 1, 1946

By Stewart W. Towle

In 1989, the u.s. Bureau of Mines initiated a contract to gather engineering, operating, and environmental cost data for 1988 for 30 major foreign primary copper smelters in market economy countries. Data were collected for 29 of the designated smelters together with information on applicable environmental regulations. Materials balance data obtained were used with available data for the eight U.S. smelters to determine the approximate extent of copper smelter sulfur emission control in 1988. A broad characterization of the status of sulfur emission control regulation was made. The 37 U.S. and foreign smelters represented roughly 73.2% of world and 89.3% of market economy primary copper production in 1988. The 29 non-U.S. smelters attained 55.3% control of their input sulfur in 1988. Combined with the 90.4% control of U.S. smelters, an aggregate 63.4% sulfur control existed. Roughly 1,951,100 mt of sulfur was emitted from the 37 market economy smelters in 1988. Identifiable SO2 control regulations covered 72.4% of the 29 foreign smelters, representing 66.5% of smelting capacity. Including U.S. smelters, 78.4% of the major market economy smelters were regulated, representing 73.1 % of smelting capacity. Significant changes since 1988 that may increase sulfur emission control are noted.

Jan 1, 1993

By John P. Capp

Information is summarized on the major uses of fly ash and prospects for utilization in agriculture and in brick and mineral-wool manufacture. Technological aspects of utilization are discussed and references are given to the work of many of the leaders in the fields of fly ash production, processing, and utilization.

Jan 1, 1970

By Seong-Taek Yun, Chil-Sup So, Seung-Jun Youm

The greisen-bordered quartz veins are spatially, temporally and genetically associated with a late Cretaceous granite intruding an early Palaeozoic, limestone-rich formation. Paragenetic data indicate three stages of hydrothermal mineralisation, the first typically in or within 500 m of the granite stock and showing vertical zoning between deeper Mo and higher W, the other two more abundant in the upper and outer parts of the mineralised area. Fluid inclusion and stable isotope data indicate that hydrothermal veins formed over a temperature range of 120-395 degrees C from low-salinity fluids. Early wolframite and molybdenite deposition in stage I occurred at temperatures of 300-395 degrees C from magma-derived, low-salinity, CO2-rich fluid, mainly as a result of fluid boiling and associated CO2 effervescence. Progressively larger volumes of local meteoric water entering the system led to the successive deposition of scheelite in stage I vugs, stage II base-metal sulphides and stage III fluorite-carbonates. Remobilisation of W and its later deposition of scheelite in the outer and shallower portions of the system occurred at 215-275 degrees C and stage II and III mineralisation from lower-temperature fluids that became progressively lighter in stable isotopes. It is concluded that the influx of meteoric water increased with distance from the cooling pluton

Jun 18, 1905