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By R. G. B. Pickering

The purpose of this report is to chronicle the interaction between Anglo Platinum and Sandvik Tamrock during the development and application of Long Hole drilling as a viable, practical method of narrow reef mining. Appendix I identifies the people from Anglo Platinum and Sandvik Tamrock who participated in the project. Experiments in long hole drilling and blasting were conducted in a number of mines during the late 1950s and early 1960s. Some of the advantages envisaged were: ? The face would require no daily examination ? No barring down or immediate face support required ? No sockets to clean ? Only a few well drilled holes rather than large numbers of short blast holes ? Avoidance of poorly-marked holes, inaccurate drilling and improper charging ? Better supervision and better face advance. Holes 50 mm in diameter were drilled parallel to the stope face by pneumatic rock drills from advance strike gullies. Holes were charged with amon gelignite cartridges. All holes were traced with detonating cord to ensure propagation of the explosive along the entire hole length. Rock was blasted onto a rolling scatter pile. A number of reasons have been cited for the failure of this mining method. The two most important were: ? problems in controlling hole deviation which was exacerbated by the fractured ground ? difficulty of charging up holes in fractured ground, due to relative movement of rocks along fracture surfaces over the hole length. The idea of drilling a few long holes parallel to the face rather than a large number of short blast holes perpendicular to the face was an attractive option to mechanize narrow reef mining. This particular project started in 1998. Initial trial The Union Section of Anglo Platinum purchased a Solo H 606 RA long hole drilling rig from Sandvik Tamrock in 1997. In order to evaluate long hole drilling as a basis for face mechanization. The Solo series of drill rig was selected because it combines high drilling performance and good drilling accuracy with high versatility with a compact size. The mining method proposed by a consultant to Anglo Platinum was based on the current trackless mining method practised by Union Section. The trackless mining method is better described as a trackless development and transport method. Reef strike drives were installed approximately 40 metres apart on dip using trackless equipment. Reef was mined using hand-held pneumatic rock drills and scraper winches and the excavation was supported by sticks and crush pillars at the top of the panel. Rock scraped from the panel was loaded by Lad Haul Dumps (LHDs), trammed to an inclined haulage, and loaded into trucks for transport to the surface. It was intended that the Solo would drill from the bottom advance strike drive to the top advance strike drive. Blasted rock would still have to be scraped from the panel. It is also difficult to understand how well the support would have survived. In the initial motivation it was stated that drilling accuracy was of the utmost importance. Previous work by Sandvik Tamrock had shown that, in other parts of the world, 20 metre long holes could be drilled with a deviation not exceeding 100 mm provided that: ? holes were collared with low thrust pressure ? operating percussion pressure was kept low -often not ? exceeding 10 MPa ? drilling was to be with tubes, not rods. In early 1998 initial trials with the Solo H 606 RA were inconclusive; attempts were made to drill 40 metre long holes from one advance strike drive to another. Difficulty was experienced in measuring hole straightness and accuracy, there was no survey of the holes, and no blasting was done to expose hole barrels. A long hole mining system was then defined, with a corresponding face mining system, which led to the design of a test programme for the Solo drilling rig. Earlier experience at Telfer The proposed face mining system was based on a mining method that is practised at Telfer Gold Mine in Western Australia. Telfer has a series of domed reefs that dip at angles of less than 20° to more than 40°. The ore bodies are all thin, with widths from 0.6 metres to 0.8 metres. Figure 1 shows the face layout. Experience at Telfer indicated the following: ? The spacing and size of the dip pillars should be based on the rock mechanics of Union Section ? The length of the face is determined by the ability to correctly position the blast holes. Positioning of blast holes is a function of hole straightness and hole direction. Blast holes in Telfer were drilled 64 mm diameter with retrac bits from Sandvik. Extension steel was T38 by 1.8 metre lengths. A guide rod was fitted immediately behind the bit. To achieve greater accuracy' operators drill at relatively low percussion and thrust pressures. Results show that very good accuracy is achieved in 20 metre stopes and only fair accuracy in longer stopes. In 34 metre long stopes 10-15% of the holes are not sufficiently accurately positioned and have to be re-drilled. All blast holes are

Jan 1, 2006

By Robert M. Funk, James E. Lawver

The development of a laboratory and pilot-scale high-tension electrostatic flowsheet for the production of iron-ore super concentrates having silica contents in the range of 0.1% is presented, A variety of hematite and magnetite ores were tested with this flowsheet, and the pertinent metallurgical data are given. The applications and limitations of this process are brought out in terms of both practical and theoretical considerations. There have been many new developments in the past few years that have required mineral beneficiation engineers to consider new techniques for iron-ore concentration. Developments in iron powder metallurgy, and particularly the making of sheet metal directly from high-purity iron-ore concentrates, and the concept of using metallized pellets as a substitute for scrap in electric-furnace steelmaking have created demands for what is commonly referred to as iron-ore superconcen-trates. In most cases, specifications for superconcentrates are that they contain less than 0.1% total gangue, with some applications requiring less than 0.05% gangue. To produce these concentrates, new and modified processes are being developed, and the three methods receiving attention as the primary concentrating methods are: 1) Magnetic separation in which dry rotor-type equipment is used for magnetite and the high-intensity magnets are used for hematite. 2) Amine flotation. 3) "High-tension" electrostatic separation (or, elec-trodynamic concentration). This paper examines the high-tension electrostatic process in terms of its limitations and abilities for the production of iron-ore superconcentrates. A technically effective and simple flowsheet was developed through laboratory and 5 and 100-ton pilot-plant studies for the production of concentrates containing from 0.02 to 0.20% SiO, with iron recoveries between 85 and 90%. The process is also shown to be economically feasible. Laboratory Test Program Separation Characteristics: The suitability of the electrostatic high-tension method for iron-ore supercon-centration stems from the fundamental mechanism governing the process, i.e., particles are separated on the basis of differences in electrical conductivities. This conductivity is a characteristic property of minerals, and if the difference between the conductivities of two min- erals is great enough, then their separation can be performed by a high-tension machine. As shown in Figs. 1 and 2, the critical factor is the time necessary for the particle charge density, which is established by a corona discharge field, to decay to a level such that the centrifugal force acting on the particle exceeds its electrostatic, or image, force. At this point, the particle can leave the rotor at some trajectory and fall into an appropriate product receptacle. This decay time is directly proportional to electrical conductivity, and the process variables, such as feed rate, rotor speed, electrode configuration, and rotor composition, are selected to optimize the ability of the separator to resolve different conductivities. Table 1 lists the measured conductivities for some of the iron and gangue minerals used in developing the flowsheet. These results were obtained by using the conductivity cell described by Lawver and Wright.' The discharge curve of most minerals that lend themselves to super concentration by high-tension is so fast that the

Jan 1, 1971

Development of New Comminution Testing Methodologies for Geometallurgical Mapping of Ore Hardness and Throughput The emerging discipline of 'geometallurgy' is becoming increasingly recognised as a discrete and high-value activity that reflects an ongoing trend towards more effective mine site integration and optimisation. Constrained sampling that reflects and defines inherent ore body variability is a key geometallurgical requirement. This requires use of larger numbers of low-cost physical testing which can be applied to small sample volumes suitable for defining natural variability. The AMIRA P843 'GeMIII' project (Geometallurgical Mapping and Mine Modelling) is a major industry-supported research initiative designed to develop new tools, methods and protocols to support geometallurgical integration. As part of this integrated research a new more rapid low-cost comminution test (GeM Comminution index) has been developed which can be employed as a front line tool for geometallurgical mapping purposes and predictive throughput modelling. The test has been designed to be inserted into routine assay sample preparation and is based on constrained jaw crushing protocols linked to analysis of resultant size distributions. Extensive validation and modelling has shown the GeM Comminution Index (Ci) is correlated with the Drop Weight index A*b and Bond Mill Work Index (BMWi). Large scale trials have been conducted within a commercial assay laboratory to demonstrate and optimise incorporation of the Ci test into routine sample preparation protocols. While Ci based estimates of A*b and BMWi are not as precise compared to larger volume more expensive test-work, the ability to undertake large numbers of tests typically in a systematic down-hole manner, provides a high level of data support through adjacency. A Ci based approach is highly suitable for variability mapping, domaining and selecting large composite samples for more precise testing. Within the context of the AMIRA P843 project this forms part of an integrated work flow designed to support geometallurgical integration. The development of a comminution test linked to routine assay sample preparation represents significant value adding to a process that in most cases is already going to be carried out. This paper introduces the Ci test concept and application to geometallurgical testing for throughput modelling

Sep 13, 2010

By D Brooks, S Gunn, T Perkins, J Siliezar

Ernest Henry Mining (EHM) is a copper, gold and magnetite mining and processing operation located 38km north east of Cloncurry. The EHM ore body has an average grade of 0.9% copper (chalcopyrite), 0.5g/t gold and 22% magnetite.Open pit mining commenced in March 1998, but EHM transitioned to underground mining in December 2012, extending the life of operations beyond 2024. The underground mine currently produces up to 3Mt/year of ore, hauled up a decline by truck. This figure shall double in 2014 with the completion of a 1000m shaft. The EHM Concentrator has a nameplate capacity of 10Mt/year. In June 2011, Xstrata Copper acquired the Mount Margaret resource to supplement underground production and utilise the full capacity of the Concentrator. The Mount Margaret Mining (MMM) operation was established to develop the five deposits making up the resource by way of open pit mining.Copper concentrate is produced by the EHM Concentrator via grinding, flotation and dewatering circuits. The grinding circuit comprises a SAG and ball mill in a closed AB configuration. Flotation consists of a rougher train, regrind and three stages of cleaning, with the tail from the first stage reporting to the tailings thickener.The mineralogy of the MMM ores is complex, comprising fine textures of active pyrite, copper oxides, native copper, clay and secondary copper sulphides. This paper provides an overview of the metallurgical test work completed to forecast the performance of the MMM ores through the EHM Concentrator and identify flow sheet modifications for maximising copper recovery and producing a saleable concentrate. A discussion of the challenges experienced in the plant to date follows.CITATION:Gunn, S, Siliezar, J, Brooks, D and Perkins, T, 2013. Adapting the Ernest Henry Mining concentrator to process the Mount Margaret ores, in Proceedings MetPlant 2013 , pp 347-370 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 15, 2013

By Dunkin H. H

Flotation tests in a laboratory flotation cell at pH 8-8.5 and other sui}plementary tests were carried out on sphalerite-quartz and sphalerite-pyrite-'quartz mixtures. Depression of sphalerite resulted vvhen the addition of copper sulphate exceeded an optimum value. Evidence that this depression was due to collector deficiency because of reaction between copper sulphate and collector is presented. It was not found possible to devise a direct analytical method suitable for .plant control of reagent addition. A possible means of control using potassium cyanide and a higher pH is discussed.INTRODUCTIONAt the Rosebery mill of Electrolytic Zinc Co. of Australasia Ltd., recovery 'of sphalerite in the zinc concentrate occasionally drops unexpectedly. Experiment in the mill and mill laboratory has suggested that one of the factors causing this may be the addition of more than an optimum amount of copper sulphate. On such occasions, an increase in the collector addition improves zinc recovery in the zinc concentrate, but only at the expense of sphalerite-pyrite selectivity.To examine the phenomenon more thoroughly, and if possible to find some means of determining the optimum addition of cepper sulphate and controlling the actual addition at this optimum, one of the authors (W. J. T.) was employed by the Electrolytic Zinc Co. of Australasia' to carry out investigations in the Mining Department, University of Melbourne, under the joint direction of R. B...

Jan 1, 1954

By O. Celep, T. Uslu, Savas

This study investigates the enrichment of a low-grade colemanite concentrate (-3 mm) using a Knelson centrifugal gravity concentrator. Due to its low boron content, the concentrate is unsaleable and has to be stored under appropriate conditions to avoid potential environmental problems. The low-grade colemanite concentrate was comminuted to size fractions of-1 mm, -0.5 mm, and -0.15 mm before treatment in the Knelson Concentrator. The effects of particle size, fluidizing water velocity, and bowl speed on the enrichment process were examined. The B2O3content of the concentrate was increased from 33.96% to a maximum of 45.52%. B2O3recovery increased with increasing bowl speed and particle size, and decreased with increasing fluidizing water velocity. The enrichment process also rejected arsenic and iron to some extent, with a maximum reduction of arsenic from 1360 g/t to 765 g/t and iron from 0.88% to0.33%.

Jan 1, 2015

Size Separation Pretreatment and Alumina Recovery from Bayer Red Mud

Sep 13, 2010

Geophysical methods are useful non-invasive means of collecting subsurface information on the environmental impacts of resource developments. Techniques such as seismic refraction, ground probing radar and magnetics are well established tools for environmental studies of waste disposal sites. Repeat resistivity surveys, gravity surveys, and infrared or multispectral remote sensing hold great promise for monitoring environmental changes. Examples from New Zealand include geothermal field monitoring, which may involve the use of geophysics for .detecting changes of surface thermal activity, hydrothermal eruptions, induced earthquakes, contamination or depletion of shallow groundwater, and subsurface fluid recharge. Other examples include resistivity monitoring of contained effluent disposal ponds, tailings dams and landfills. Experience has shown that care is necessary to identify the natural variability of the parameter measured before drawing conclusions regarding an observed change.

Jan 1, 1994

While ring-blasting techniques offer substantial benefits, there are intrinsic problems that are exacerbated when large blocks of ore are mined in attempts to improve the production economics. These problems relate to both the convergence of holes around the ring drive and the charging of up-holes. The size and shape of the ore block concerned should be carefully considered so that good drilling, patterns can be achieved. The hole diameters chosen, both affect and are affected by the size of the ore block. The explosive selected depends on the hole diameter and together with the diameter, determines the nominal drilling pattern. Ring-drilling patterns are best derived through the use of computer software together with engineering judgement. The loading of explosive is necessarily high, and charging patterns require careful thought if overcharging is to he minimised and adequate energy is to be available at all parts of the ore block. Initiation systems are critical to the blasting results, but sub-optimal systems must be adopted if collar priming is practised. In view of all these factors, idealised guidelines can be given for ring design, but it should be recognised that developments in technology may in time change some of these parameters.

Jan 1, 2000

By K. Donahue

Clay minerals play an important role in the gravitational stability of man-made rock piles by affecting geotechnical and hydrological characteristics of the material. The Goathill North rock pile at the Questa Mine in northern New Mexico has been the focus of a multi-disciplinary study to determine the effect of weathering on rock pile stability and in particular to determine if clay minerals are forming in the rock pile due to surface weathering. The results of x-ray diffraction (XRD) and scanning electron microscopy (SEM) analyses indicate the clay minerals found within the fine-grained soil matrix of the rock pile are of hydrothermal origins and did not detect any neoformation of clay minerals. The foremost mechanism for an increase in clay minerals within the rock pile is the physical break down of hydrothermally altered rock fragments to release hydrothermal clay minerals. The hydrothermal clay minerals within the rock pile show evidence of weathering by the dehydration of smectite from 2-water to 1-water interlayers within the clay mineral structure. This dehydration is most likely due to higher temperatures and low relative humidities in the Goathill north rock pile during surface exposure.

Jan 1, 2009

By H. L. Paterson, R. R. Keays, M. J. Donaldson, D. J. Boyack, M. F. Smith, J. J. Gresham

"Albite-rich sedimentary rocks that contain abundant Fe sulphides occur in the lower portion of the ultramafic sequence at Kambalda and their broadly antipathetic relationship to most of the nickel orebodies is well documented. These sediments generally contain less than 0.1% Ni, but nickeliferous equivalents with 0.3 to >6% Ni occur locally in close spatial association with more typical volcanic peridotite-associated massive, matrix and disseminated ores. Nickeliferous sediments have been mined at several locations at Kambalda but, collectively, they constitute less than 2% of the total Kambalda nickel reserve. The nickeliferous sediments occur both on the ultramafic-footwall basalt contact and within the lower part of the ultramafic sequence, and are usually less than 5 m thick. They are commonly finely laminated rocks with a cherty appearance, composed essentially of layers of albite+quartz±amphibole, alternating with pyrrhotite+pentlandite±pyrite layers. Some contain a minor amount of carbon. The sediments are only of ore grade (> 1% Ni) immediately adjacent to massive Fe-Ni-Cu sulphides, which are interpreted as normal magmatic ores on the basis of geochemistry and the presence of distinctive ferrochromites. The nickel content of the sediments decreases to background levels (<0.1%) within about 20 m of the massive sulphides. In most locations the massive sulphides clearly have been emplaced tectonically within or adjacent to the sediments, which are commonly highly deformed. Typical nickeliferous sediments contain about 1% Ni, 0.1% Cu, 0.1% Zn, 400 ppb Pd, <0.2 ppb Ir and 10-200 ppb Au. They have lower Ni/Cu (average 10), higher S/Ni (average 6) and higher S/Se (average 11 200) than associated massive sulphides (average 40, 4.8 and 7400, respectively). The Pd/Ir ratio of the nickeliferous sediments (typically > 1000) contrasts markedly with that of associated massive sulphides (average ca 16).The high Pd/Ir values are similar to those which have been reported for hydrothermal Fe-Ni sulphides and barren sulphidic sediments of probable volcanic exhalative origin. They preclude a direct magmatic origin for the sulphides in the nickeliferous sediments, though S/Se and S isotope data indicate a magmatic source for the sulphur. The restriction of nickeliferous sediments to within about 20 m of normal magmatic sulphides and the lack of a widespread Ni dispersion halo suggest local redistribution of originally magmatic Ni. This probably occurred by diffusion and hydrothermal mobilization during deformation and amphibolite-facies metamorphism. Although not totally ruled out, a volcanic exhalative origin for the Ni seems unlikely."

Jan 1, 1984

By C. L. Kosmoski

The mining industry faces many workplace hazards that are a threat to human health and safety. The Bureau of Labor Statistics has found that employees in coal mining are more likely to be killed or sustain an injury or illness than workers in private industry. These injuries are also likely to be more severe (Bureau of Labor Statistics, 2007). In recent years, coal mining injuries have attained national attention due to highly publicized disasters. Despite these threats to worker safety and health, the U.S. relies on the mining of coal to meet its need for electrical power, with42% of electricity in the U.S. generated through the burning of coal (U.S. Energy Information Administration, 2012b; United Mine Workers of America, n.d.). Furthermore, the production of coal continues to increase and reach record levels every year(National Mining Association, 2012; U.S. Energy Information Administration, 2012a). For this reason, the coal mining industry must continue to find ways to protect its workers while maintaining productivity. One potential approach to protecting workers while maintaining productivity is through improving the safety culture at coal mines. In order to achieve this culture, operators, employees, the inspectorate, etc. must share a fundamental commitment to safety as a value. This type of culture is known in other industries as a ?safety culture? and can be defined as the characteristics of the work environment, such as the norms, rules, and common understandings that influence facility personnel?s perceptions of the importance that the organization places on safety. Fundamental to improving safety culture within an organization is understanding the existing safety culture; improvements cannot be made without first elucidating areas that require improvement. This manuscript details the process undertaken by researchers at the National Institute for Occupational Safety and Health (NIOSH) Office of Mine Safety and Health Research (OMSHR) to assess the safety culture at five underground coal mines across the United States. The purpose of this manuscript is to introduce the mining industry to a multi-methodological approach to safety culture assessments. Further, it is the goal of this manuscript to highlight the value of including both quantitative and qualitative methods into safety culture assessments in order to obtain the most comprehensive view of an organization?s safety culture. Because this manuscript focuses solely on the methodology implemented, data and results obtained during the five safety culture assessments will not be presented in this manuscript, but will appear in a future publication.

Feb 27, 2013

By Jack Eloranta

This paper documents the relationship between material handling and processing coete compared to blasting cost. The old adage, "The cheapest crushing is done in the pit", appears accurate in this c!ase study. Comparison of the accumulated cost of: powder, selected wear materials and electricity; indicate a strong. inverse correlation with powder factor (lbe powder/long ton of rock). In this case, the increased powder cost is more than offset by electrical savings alone.

Jan 1, 1995

By Daniel Harrington

On the night of June 8, 1917, the flame of a carbide lamp accidentally set fire to the uncovered and frayed insulation of an armored power cable near the 2,400-foot level of the North Butte Mining CO.'s Granite Mountain shaft. How this, insulation accidentally became exposed is explained on page 15. The highly flammable oiled fabric set fire to the shaft timbers, and as this is a downcast shaft the fire spread with great rapidity, soon filling the mine workings with smoke and gas. At the time 410 men were working underground, 247 of whom escaped by various means, but most of the 163 remaining were probably overcome soon after the fire started, and perished. Only two men were actually burned. The cause of the disaster was purely accidental. The immediate effect was the unexpected reversal of normal conditions; the ultimate effect was great loss of life and the destruction of the main hoisting shaft, putting it temporarily out of service. The work of rescue and fire fighting continued eight days. The following pages present an attempt to draw some practical conclusions and suggestions from this disaster. The property of the North Butte Mining Co. consists of about 220 acres in the northeastern part of the Butte district, Montana. Next to the Anaconda Copper Mining Co. it is the largest producer of the district. In 1916, the year before the disaster, there were 1,160 men employed. The output was 560,947 tons of ore, which yielded 24,498,181 pounds of copper, 1,047,063 ounces of silver, 1,712 ounces of gold, and 412,953 pounds of zinc. The profit was $2,479,595. The total development work for that year was 21,694 feet. During the first half of 1917, the year of the accident, the daily tonnage of ore averaged 2,000, an increase of 400 tons over that of the year 1916.

Jan 1, 1922

By M. W. Hildyard

A blast was engineered close to a tunnel in a deep-level mine with the purpose of studying the wave interaction with the tunnel. Numerical modelling of seismic wave propagation was used in both the forward planning and the back-analysis of this experiment. This paper presents the forward analysis. A likely model for a propagating blast was investigated and developed. This model predicts the influence of the velocity of detonation of the blast and the rise-time of the blast on the radiation pattern, and hence on the expected positions of maximum velocity at the tunnel surface. An attempt was made to quantify the parameters of the blast model, using waveform data from two other blast experiments. The blast model caused particle velocities which were an order of magnitude lower than these measured velocities. Large values for blast pressure were required to match the amplitudes of the recorded velocities. An alternative ?effective? source with a similar mechanism but increased diameter was proposed to account for the higher velocities. It was established that distant velocities scale with the square of the borehole diameter. A calibration blast at the experimental site was also modelled. Agreement in the waveforms was not sufficient for conclusive statements to be made about the validity of the source model. However, the modelled waveforms show an encouraging correspondence with some measured waveforms. This gave confidence that the model could be used to examine wave propagation around the tunnel, and allowed a scaled source to be developed for forward analysis of the main experiment.

Jan 1, 2001

By Miller F. S

The famous Golden Mile of Kalgoorlie in Western Australia was, between 1888 and 19ft, one of the richest and most important gold camps in the world. Annual production, which reached a peak of nearly 1,300,000 oz. in 1903, however dropped to below 400,000 oz. after 1920.Depression seized the camp, and many mines closed down. The output from this old field is now steadily mounting under the stimulus. of a higher price for gold. In 1935 300,000 oz. of precious metal were recovered and the future is bright. Total production from the field is 21,400,000 oz., whereas something like &#163;23,600,000 have been paid in dividends.The Golden Mile has been studied by many geologists. Their somewhat diverse findings were published for the most part in the Australian technical press which, unfortunately, is not sufficiently represented in foreign libraries.The following article is based on an eighteen months&apos; study in the district for Gold Mines of Kalgoorlie Ltd., preceded by four years&apos; experience with almost identical rock types in the, Porcupine district of Ontario. The excellent record of observed fact left by previous invesigators, augmented by talks with observant mining men...

Jan 1, 1937

By L. W. Higley

It is estimated that well over 20 million lb of chromium and 8 million lb of nickel, plus other metal values, are diverted annually to furnace dusts, mill scale, and contaminated swarfs in the domestic stainless and other specialty steelmaking industry. Little inroad has been made toward the in-plant recycling of these wastes to reclaim the metal values. The Bureau of Mines has devised a process for recovering about 90 pct of the chromium and molybdenum and well over 90 pct of the nickel and iron from stainless steelmaking furnace dusts, swarf, and mill scale. Other metals such as manganese are also recovered. Initial laboratory results were confirmed and expanded to include a wide variation of stainless steelmaking waste mixtures. Later experiments have included specialty steelmaking wastes other than from stainless steelmaking. Significantly, the ease of technology transfer has been demonstrated by a number of industrial-sized (19-ton) commercial electric furnace heats, in which the pelletized wastes represented some 10 to 20 pct of the furnace charge, replacing part of the normal scrap charge. Results of these successful large-scale trials are presented. Some information is also presented on incorporation into the pellets of chromium-bearing particulate wastes originating from other than steelmaking production. Cost evaluation indicates the process is economically attractive.

Jan 1, 1982

By U. B. Kodukula

Attempts are made to evaluate the feasibility of using a water-injection cyclone for the concentration of iron ore fines. The separation characteristics of clay fines, containing large amounts of alumina and silica, from the iron values are discussed. The effects of critical variables, such as injection water rate, truncated cone diameter, spigot opening and vortex finder diameter, on the performance of a water-injection cyclone in terms of iron, alumina and silica grades; iron recoveries; and alumina and silica rejections are discussed. A set of 16 full-factorial experiments (24) was carried out to gain an understanding of the individual and interaction effects of the variables. The results indicate that classifying iron ore fines containing 63.0% iron, 2.5% alumina and 3.5% silica through a water-injection cyclone can produce concentrates with iron grades greater than 66.0% and alumina and silica contents less than 1.5% and 2.0%, respectively, at different study conditions. The achievable iron recovery values are greater than 88%, and alumina and silica rejections are more than 80% and 50%, respectively, within the specific study conditions. While the truncated cone diameter, injection water rate and spigot diameter influences the alumina rejection to a considerable extent, it is mostly influenced by the vortex finder diameter. The iron recovery and silica rejection values are mostly influenced by spigot diameter and truncated cone diameter. The effect of injection water on the concentration characteristics depend on the prevalent truncated cone diameter and spigot diameter conditions.

Jan 1, 2008

By K. G. Thorne, L. J. Cabri, L. R. Fyffe, D. R. Lentz, D. Hoy

"Mineralization at the Main zone of the Clarence Stream gold deposit in southwestern New Brunswick occurs within a parallel series of quartz veins that occupy a NE-trending, steeply north-dipping, brittle-ductile shear zone related to the nearby terrane-bounding Sawyer Brook fault. This deformation zone cuts volcaniclastic sedimentary and volcanic rocks of the Silurian Waweig Formation, and is intruded by mantle-derived East Branch Brook gabbroic dikes and a number of felsic dikes. Intermittent shear-zone reactivation during local magmatic activity induced heterogeneous deformation and alteration of the country rocks and dikes. The terminal phase of shearing is associated with narrow pegmatite-aplite dikes that grade laterally into granophyric granite, and into an auriferous quartz vein. Gold, aurostibite, and gold-antimony intergrowths, predominantly with vein quartz, are associated with a low ƒO2 sulfide mineral assemblage (pyrrhotite, arsenopyrite, and berthierite) that has a sulfur isotopic signature consistent with a magmatic source. Gold saturation and deposition occurred episodically between 300° and 360°C at low activity of sulfur, predominantly in response to pressure reduction during brittle failure combined with the effects of decreasing temperature and sulfidation reactions during fluid-wall-rock interaction. Geochronological studies demonstrate that the timing of the emplacement of the pegmatite-aplite dikes and associated auriferous veins overlap with that of the Early Devonian Magaguadavic granite. Moreover, the composition of the dikes is consistent with their derivation from this granite as late fractionates. The Main zone of the Clarence Stream deposit is therefore interpreted to be intrusion-related, with the Magaguadavic granite being the ultimate source of the gold-bearing fluids. La minéralisation de la zone principale du gisement aurifère de Clarence Stream dans le sud-ouest du Nouveau Brunswick occupe une série de veines de quartz parallèles dans une zone de cisaillement fragile-ductile de direction NE et à pendage abrupt vers le nord apparentée à la faille de Sawyer Brook avoisinante qui délimite deux terranes. Cette zone de déformation recoupe les roches volcaniclastiques et volcaniques de la Formation de Waweig d’âge Silurien et est recoupée par les dykes de gabbro d’East Branch Brook, d’origine mantellique, ainsi que par divers dykes felsiques. La réactivation intermittente de la zone de cisaillement durant l’activité magmatique locale a entraîné la déformation et l’altération hétérogènes des dykes et de l’encaissant. La phase finale du cisaillement est associée à de minces dykes d’aplite-pegmatite qui passent latéralement à des granites granophyriques, puis à des veines de quartz aurifères. L’or, l’aurostibite et les intercroissances d’or et d’antimoine principalement associées aux veines de quartz sont associés à un assemblage minéral de sulfures à faible ƒO2 (pyrrhotite, arsénopyrite et berthierite) présentant une signature isotopique du souffre compatible avec une source magmatique. La saturation en or et la précipitation a eu lieu épisodiquement entre 300° et 350°C dans des conditions de faible activité de souffre, principalement en réponse à la baisse de pression durant la rupture fragile ainsi qu’aux effets de la baisse de température et aux réactions de sulfuration durant l’interaction entre le fluide et l’encaissant. Les études géochronologiques indiquent que l’âge de la mise en place des dykes de pegmatite-aplite et des veines aurifères qui leur sont associées coïncide avec le granite de Magaguadavic, d’âge Dévonien Précoce. De plus, la composition des dykes est compatible avec une origine provenant de ce granite sous la forme de phases fractionnées tardives. Le gisement principale (Main zone) de Clarence Stream est ainsi considéré comme étant associé aux intrusions, le granite de Magaguadavic étant la source ultime des fluides aurifères."

Jan 1, 2008

Mining men and State officials in their efforts to make mining safer are confronted from time to time with various complicated questions. In recent gears these questions have tended to increase with the use of mechanical appliances. Twenty-five years ago electricity had been used but little in mires; today there is practically no other means of transmitting power employed in American mines, except that in metal mines compressed air is used, chiefly for drilling and for operating small portable hoists. In coal mines there has been a great increase in mechanization, which tends to concentrate the work of production in small areas and to increase the local difficulties of ventilation; also, it has intensified locally the hazard from coal dust as well as firedamp in bituminous and subbituminous mines. The answers to the questions that arise are not always obvious, as is shown by the fact that even able, experienced men differ in their opinions of the best, way to overcome a particular mining hazard. The difficulty of .solving the problems is also demonstrated by the varying provisions in mine-safety rules, both in official regulations and in the rules of those mining companies that are endeavoring to do more than follow the letter of the law. Such differences of opinion as to the best way of meeting a difficult situation are often found within the same mining organization, and, similarly, varying opinions are held by different members of the Bureau of Mines staff.

Jan 1, 1937