Search Documents

By C. Guan

Since it was patented in 1942, the Crago process with fatty acid rougher flotation followed by deoiling and amine cleaner flotation has become the standard for phosphate flotation in Florida and North Carolina. This paper outlines the basic fundamentals involved in this proven technology. Included are the mineralogy of the flotation feed, the solution chemistry and the interfacial phenomena. The process variables affecting the performance of this technology are also discussed.

Jan 1, 2009

As opposed to a prescriptive-based design, a performance-based fire-protection system design allows the system planner the flexibility in choosing specific method, materials and design tools to achieve a preset goal. A different approach in fire protection and a powerful addition to existing method, the system has been around for many years, and has gone through numerous improvements. Despite its popularity in the building industry, it has not been introduced to the mining industry yet. As mines get deeper, increases in size and complication, mine fire safety system will also become increasingly challenging, a new and proven system should be able to great potential to eradicate many of the problems that are either unsolvable or cost-prohibitive with the traditional method. With increasing globalization and international exchange with other major mining countries a common language/practice also make the performance-based design method a desirable feature in mine fire safety system. This paper will first define a performance-based system, step-by-step design methods, likely impacts on fire-protection practices and possible challenges in implementing the system in US underground mines.

Jan 1, 2011

By G. F. Kuncir

The analysis summarized herein evaluates the Bureau of Mines Program for Advancing Mining Technology Oil Shale in terms of the economic benefits which may be derived for the Nation as a result of its successful research and development efforts. On the basis of postulated oil shale production scenarios, benefit-cost ratio indicators are computed for the program elements as they apply to surface, underground, and modified in situ oil shale mining systems. Eight value indicators are used to measure the Program's beneficial impact on the future oil shale industry in relation to planned research and development costs. These are: capital cost savings, operating cost savings, labor savings, environmental quality maintenance cost reduction, reduction in production costs, byproduct production benefits, total production benefits, and additions to recoverable reserves. Program costs are aggregated from 1974 through 1987 and major benefits are accrued from 1982 through 2040. Both cost and benefit streams are discounted to mid-1975 present value.

Jan 1, 1977

By Russell Barr

Strategic alignment throughout a mining organization is critical to effective strategic execution and good business performance. The alignment of strategic goals and objectives in the different operations and functions in a large global operation is a challenge facing executive management in light of legacy organizational structure and processes. An additional confounding factor in strategic alignment of acquisition and new project expansion is that mining projects and operations are frequently multi-cultural in terms of country origin as well as organizational culture. This brings together many cultures to accomplish the project or complete an acquisition (Financial Times, 2006). The mining industry presents numerous challenges to executives and mine management by the nature of the operation and organizational processes used in running it. For example, there seems to be an inherent conflict between the individual goals of the components that constitute a mineral processing operation (the mine, the mill, the smelter and the refinery), and lowest cost production of the final product. Managers of these component operations have specific objectives passed down from the corporate executives. Those site managers tend to be compensated based on throughput and perform to this main metric. This tendency to maximize throughput for an individual operation can contradict the general process flow of a best-fit solution to producing the final product at the lowest cost. In an effort to become more aligned, firms are increasingly using balanced scorecard derivatives at great cost and effort (whether internally or externally sourced) in an attempt to help with strategic alignment. Balanced score-cards have been effective in providing a good strategic scorecard to companies with some measure of strategic alignment. Many firms make this effort of using balanced scorecards. However, most firms are not strategically aligned or, at best, are sub-optimized, in terms of strategic alignment (Kaplan and Norton, 2006). The tendency in mining operations is to focus on operating individual components as optimally as possible and stretching existing assets to the best possible production rates. ?Making tons? is the common operational focus. Numerous factors underpin this focus and can affect the ability of the operation?s ability to align properly with the corporate strategy and strategic goals. Critical factors are legacy organizational structure, culture, processes and performance measurement systems. Changing anyone of these factors to affect better strategic alignment requires significant attention to managing change.

Jan 1, 2008

By B. Slaker

The ability to detect and quantify ground movements in underground mine workings is of the utmost importance to the safety of miners and continuity of operations. Rib sloughage is one of the most common forms of underground deformations. Remote sensing techniques, such as laser scanning, can be used to quantify this rib sloughage, by recording precise, time-lapse point clouds of mine workings. In order to determine sloughage volumes and locations, a total of 11 laser scans were performed 41 days apart at a working section of an underground limestone mine. During this period, significant sloughing and scaling occurred, allowing for differentiation between the initial and final point clouds.

Feb 27, 2013

By S. Vujic

The exploitation of stone (trachyte) at the ?Srebro? open pit mine, located on the northern slopes of the Fru?ka Gora National Park (Serbia) central area begun in 1964. Stone exploitation in this locality does not endanger the environment in an ecological sense, does not pose a threat to plants and animals, and apart from the terrain configuration changes, it did not leave behind permanent consequences to the environment. Projects for permanent cessation of mining operations and for land reclamation and spatial arrangement were completed in these circumstances. The paper gives a short review of the condition at the open pit mine; furthermore, the key problems and basic demands are presented, together with a concept of technical and biological land reclamation and the spatial arrangement of the mining complex. Keywords Open pit mine, land reclamation, spatial planning, spatial arrangement, National Park, Fru?ka Gora, Srebro.

Jan 1, 2011

By E. A. Lowe, J. W. Smith, C. W. Overton

"The operations of Bethlehem Copper are situated in the south-central interior of British Columbia at an elevation of 4900 feet above sea level. Main road and rail transportation systems pass through Ashcroft in the Thompson Valley, which is 28 miles by road from the minesite. The operation was the first low-grade open-pit porphyry copper mine in British Columbia.The minerals are chiefly chalcopytite and bornite, while the rock is principally feldspar and quartz. In the outer fringes of the orebodies some pyrite is present. Average grade of mill feed is 0.55% copper.History of Milling Operations Production commenced in December 1962 at a rate of 3,000 tpd in a plant that comprised two-stage crushing followed by rod and ball milling. Production ratings were increased to 6,000, 10,000 and 12,000 tpd by the addition of a further stage of crushing and more rod and ball mills. When the last expansion was completed tonnage throughput exceeded the 12,000 ton designed rate and was stabilized at 15,500 tpd.However, recovery did not meet expectations at the expanded rate because of coarse flotation feed. Capacity in the rod mill circuits exceeded the subsequent fine grinding capacity, resulting in poor liberation. Studies were initiated to improve recovery without decreasing throughput.Description of Operation Prior to Pebble Milling Crushing with primary, secondary and a closed tertiary stage proceeds at a scheduled 20.5 shifts per week. Fine ore screened on 1/2"" x 4"" apertures is stored in a 30,000-ton bin ahead of two 12.5' x 15' rod mills. Rod mill discharge joins secondary ball mill (one 10.5' x 14' and one 11' x 14') discharge for classification in a single 50"" cyclone which closes the secondary mills. Overflow from this cyclone is split between three third-stage ball mill-pump-cyclone circuits (two 11' x 14' and one 12.5' x 15'). Prior to addition of the two pebble mills the flotation feed was produced by this three-stage grinding system. Grinding criteria are shown in Table I."

Jan 1, 1975

By B. Radwanek-Bak

The assurance of future raw materials supply to the EU mineral industry has become, in recent years, one of the most important tasks of the EU Commission. It is focused on creation a new mineral policy within the EU zone. One of the starting points for this activity was identification of the mineral commodities that are critical for the sustainable development and technological progress. The paper presents the glance to this problem from the National Member level. Domestic mineral reserve base as a possible source of critical raw materials for the EU was shown. Poland cannot be considered as its resource base. The source of these raw materials are not only scarce in Poland, but also they are not produced, and their demand is now - and according to forecasts is going to be in the future - met by imports. The raw materials in question are not considered exactly critical for Polish economy, but they are of limited economic importance. On the other hand, their growth is a best opportunity and at once the need to the future sustainable development.

Jan 1, 2011

By H. B. Carroll

The performance of commercially available portable instrumentation for detecting nitrogen dioxide, sulfur dioxide, and hydrogen sulfide was evaluated. Length-of-stain detector tubes from several manufacturers were included in the study. The instruments were evaluated in atmospheres with 50-and 80-percent relative humidities at 75° F and containing test gas concentrations of 0.1, 0.25, 0.50, 1.0, 1.5, and 2.0 times the threshold limit values. The results from the gas analyzers were correlated with the known contaminant concentration by calculating and plotting an estimated line of regression, confidence limits, and correlation coefficient for each detector system. An increase in the relative humidity from 50 to 80 percent at /5° F changed the instrument response to the contaminants. The uncertainties in detection precision reported are for the observer-instrument system. The main difference or inconsistency in toxic gas concentrations determined with length-of-stain tubes resulted from the observer's interpretation of the results rather than from a predominant difference in the tubes.

Jan 1, 1975

By S G. Walters

The minerals industry is facing a productivity and investment crisis. The ‘Millennium Super Cycle’ from 2003–2011 was an unprecedented period of growth and investment resulting in increased throughput and development of lower grade resources to meet demand. During the boom quantity became more important than quality with throughput the key metric. This was accompanied by a general trend of decreasing feed grades across all commodities that was offset with higher production volumes.Current industry perception is that feed grade depletion is an unavoidable consequence of ore deposit geology and mass mining technologies for increasingly mature mining operations. In typical crush-grind-float operations value recovery only takes place at ~1–200 micron particle size during flotation, ie 3–4 orders of magnitude size reduction compared to primary feed. For increasingly low-grade deposits the cost of energy and capital intensity required to process and reject low value material at micron scale drives poor productivity. An alternative is to deploy a range of coarse (~10–100 mm) separation and rejection technologies. Grade Engineering® is an integrated approach to coarse rejection that matches a suite of separation technologies to ore specific characteristics and compares the net value of rejecting low value components in current feed streams to existing mine plans.Coarse rejection can be used on size distributions ranging from run-of-mine (ROM) to semi-autogenous grinding (SAG) discharge. Grade Engineering® is based on five rock based ‘levers’ linked to combinations of screening, sensor-based sorting and heavy media separation. These involve exploitation of preferential grade deportment to specific size fractions during breakage: differential blasting to condition grade by size at bench scale; sensor based sorting at truck and conveyor scale; and differential particle density.Grade Engineering® is being developed and implemented by a consortium of over 30 mining companies, equipment suppliers and research organisations. Emerging results from collaborative site activities demonstrate potential for generating significant value, which can reverse the trend of declining productivity due to feed grade depletion.CITATION:Walters, S G, 2017. Driving productivity by increasing feed quality through application of innovative Grade Engineering® technologies, in Proceedings Tenth International Mining Geology Conference 2017, pp 263–274 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 20, 2017

Newcrest Mining is one of the leading independent gold producers in Australia thanks in large part to its Cadia Valley operations. It is the largest gold and copper producer in New South Wales and one of the largest in Australia. This year Newcrest will expand its operations with a A$545-million expansion in the area. The Cadia Valley is approximately 250 km (155 miles) west of Sydney. The province presently includes the Cadia Hill opencut mine and the Ridgeway underground mine. With a significant mineralized system to the east of Cadia Hill and a mineral resource below Ridgeway, Cadia Valley has the potential to underpin Newcrest?s production for up to 30 years. In 2007, total gold production for the company in-creased to more than 50 t (1.6 million oz) and cash flow increased by 47 percent, according to the company?s annual report. The history of mining in the area dates back to 1851, but Newcrest?s history is young. In 1990, Newcrest commenced activities in the Cadia Valley. In 1992, Cadia Hill was discovered, with the discovery of Cadia East (1995) and Ridgeway (1996) following close behind. In 1998, the Cadia Hill openpit mine was commissioned. In 2002, the Ridgeway underground mine was commissioned and in 2007 the company?s board of directors approved an A$545 million development of the Ridgeway Deeps. The Ridgeway underground mine is located adjacent to Cadia Hill. The Ridgeway orebody has no surface expression but lies approximately 500 m (1,640 ft) below the surface and resembles an inverted teardrop. It measures approximately 450 m (1,476 ft) by 250 m (820 ft) and extends in excess of 850 m (2,788 ft) in depth. Ridgeway is a sublevel caving operation. Crushed ore from underground is delivered by conveyor to a surface stockpile, adjacent to the Ridgeway concentrator. Using similar processes to the adjacent Cadia Hill concentrator, gold and copper are recovered in a conventional flotation circuit to produce a copper concentrate containing elevated gold levels. Ridgeway concentrate is combined with the Cadia Hill concentrate before it is pumped to the filtration plant in

Jan 1, 2008

In North-West Nelson the Riwaka Complex, a linear basic intrusive complex, is exposed for approximately 45 km. Some 500 km south, across the Alpine Fault, a linear zone of similar but less well known basic intrusives that may be correlatives of the Riwaka Complex is exposed discontinuously for about 80 km. The Complex was emplaced at several centres along a major tensional fracture zone that developed in upper Carboniferous to Lower Permian time. This fracture zone probably formed in response to the development of a west dipping subduction zone that gave rise to the Permian volcanic arcs in east Nelson and western Southland. The Riwaka Complex contains three rock suites. The Fractionation Suite, characterized by cyclic mineral layering and small scale cumulus banding is the largest and is divided into the Ultramafic, Hornblende and Plagioclase Series, each of which is subdivided into several units. The Cumulus Gabbros represent a restricted but important suite of unusual composition, while the third suite is the late phase Diorite Suite. Except in the lower units of the Ultramafic Series sulfide mineralisation is widespread. Pyrrhotite is the dominant sulfide with lesser chalcopyrite pentlandite and pyrite. Four significant zones of sulfide accumulation are recognised and discussed in some detail. Evidence suggests that the primary magma from which the Riwaka Complex was derived was sulfur deficient. It is concluded that the sulfur content of the Riwaka Complex was derived from assimilation of sulfur rich wall rocks during intrusive processes.

Jan 1, 1980

By Eugene T. Sheridan

The 24th annual survey of the peat industry by the Bureau of Mines, United States Department of the Interior, revealed that 76 producers in 20 States reported commercial production of peat in 1957. In addition to the States that reported production in 1956, Colorado, Idaho, and Wisconsin produced peat in 1957 Forty-four producers reported humus production; 21, reed-sedge peat; and 22, moss peat production. This total exceeds the number of active plants but several producers recovered more than one type of peat. Fifty percent of the total production was reported as reed-sedge peat, 36 percent as humus, and 14 percent as moss peat. About one-sixth of the total production was raw peat, with no preparation other than being air-dried. The remainder consisted of processed peat prepared for use by cultivation, shredding, and/or kiln drying.

Jan 1, 1958

By J. Richter, D. Schulze

J. Richter and D. Schulze (Deutschen Akademie der Wissenschafte zu Berlin)—Introduction. In a recent paper R. G. Garlick and H. B. Probst reported on experimental results of investigations of room-temperature slip in zone-melted tungsten single crystals.17 In the paper it was pointed out that slip lines could be detected with an optical microscope only after about 10 pet plastic strain of the crystal. In our investigations we could observe slip lines in tungsten single crystals of defined orientations already with an elongation of 0.02 to 0.03 pet.18"9 Experimental Technique and Results. The tungsten single crystals were produced from 4 to 5 mm sintered rods* as starting material by electron- beam zone melting. With single-zone-pass melting a speed of 0.5 mm per min was used, while three-pass melting could be handled at a speed of 3 mm per min. Crystal axis orientations were located in the shaded area of the stereographic triangle, see Fig. 6. Orientation 1 is located at about 18 deg and orientation 2 at about 5 deg on the great circle from [011] to [ill]. This corresponds approximately to the orientations P and M referred to by Garlick and Probst. Our single crystals had a length of about 70 mm and were profiled solely by electrolytic means, leaving grip ends at both ends of the crystals. Between the latter, a cylindrical crystal of 45 to 50 mm length and about 2.4 mm diam was obtained. The crystals were deformed at room temperature in a commercial tensile-testing machine to obtain a strain of 0.02 to 5 pet. The average strain rate e = eP/t for all tensile tests amounted to approximately 10 -6 to 10-5 per sec, with ep representing plastic strain and t the duration of applied stress. In Fig. 7 a crystal is shown elongated by 0.03 pet. The picture was obtained by means of an optical microscope. Figs. 8 and 9 give electron-microscopic pictures of slip lines of crystals elongated by 0.07 and 1.1 pet, respectively. For crystals with the orientation 1, see Fig. 6, only the two main slip systems (101)[111] and (110)[111] were found. With the crystals of group 2 the two main slip systems are again evident, with

Jan 1, 1965

By F. J. Perzak

The Bureau of Mines conducted a flammability study of plastic ventilation ducts. Six rigid 16-in-diam fiberglass-reinforced plastic ducts from five manufacturers were examined using laboratory-scale, large-scale, and full-scale fire tests. The laboratory-scale tests included the American Society for Testing and Materials E162 radiant panel test, the E2863 oxygen index test, and the standardized small-scale flammability test (S3FT) developed by the Mine Safety and Health Administration (MSHA), U.S. Department of Labor, and replicated by the Bureau of Mines. Ten-foot lengths of the same ducts were also subjected to a large-scale vertical fire test, and 20-to 30-ft lengths were tested in a full-scale in-mine fire test at the Bureau's Lake Lynn Laboratory mine. The data from the laboratory-scale radiant panel test and oxygen index test indicated that the ducts were fire-resistant, However, five of the six ducts failed the S3FT and vertical fire tests, and four of the ducts failed the full-scale in-mine tests with flame propagation rates that ranged from 3 to 10 ft/min. The results indicate the need for improved fire-resistant plastic ducts. The good agreement between the S3FT re-sults and the vertical and full-scale tests show that the S3FT provides a good indication of the flammability of fiberglass-reinforced plastic ducting under realistic fire conditions.

Jan 1, 1987

By W. L. Dumoulin

A RATHER detailed description of the entire plant and leaching process was given in a paper recently presented to the Institute,1 so this paper will cover briefly only the crushing practice of the New Cornelia Copper Co. for the year, 1918. The ore, which is mined by steam shovels and loaded in side-dump cars, passes through two crushing plants. These crushing plants were constructed to crush the ore required by a leaching plant of 5000 tons daily, capacity, during a crushing period of 16 hours. The primary plant reduces to less than 3 in. (76 mm.) and the secondary plant, to ¼ in. (6.35 mm.), which is sufficiently fine for satisfactory percolation of solution through the ore in the leaching tanks, and to give good extraction. The ore is then conveyed to the leaching tanks by a system of 28-in.. (71-cm.) belt conveyors. On the way to the leaching tanks, it passes through an automatic sampling plant, where a 1-per cent. sample is taken. From the primary crushing plant, the ore is conveyed by two 36-in. (91-cm.) belt conveyors to a 10,000-ton steel storage bin, with a reinforced-concrete flat bottom. The ore discharges from the bottom of this bin onto four 20-in. belt conveyors, which deliver it to the four units of crushers in the secondary crushing plant. There is no storage bin between the mine and the primary crushing plant. The ore breaks very coarse, is very hard, and contains a great many boulders, some as large as 4 ½ by 4 1/2 by 10 ft. (1.3 by 1.3 by 3 m.). Jams form in the bowl of the coarse crusher in the primary crushing plant, as a consequence, and are freed by means of an immense steel hook operated from a 40-ton electric traveling crane.

Jan 8, 1919

By P F. Lauricella, D Hendri

In the Mesel sediment-hosted disseminated gold deposit, the spatial distribution of hydrothermal alteration is controlled by steeply dipping faults and shallowly dipping lithologies. Hydrothermal alteration of carbonate rock is characterised by decalcification, dolomitisation and silicification. Intense argillic alteration occurs in a laccolithic andesite intrusion adjacent to decalcified and silicified carbonate units. Sulphide minerals (pyrite, realgar, orpiment, stibnite and cinnabar) accompany hydrothermal alteration and precious metal mineralisation. The trace elements Ag, As, Sb and Hg are zoned spatially throughout the deposit. Mesel contains a mineable gold reserve of 7.8 million tonnes at 7.3 g/t Au. Gold mineralisation is preferentially localised in carbonate rocks adjacent to and beneath the andesite intrusion. The ore body consists of a central, near-vertical, high grade 'feeder' system flanked by tabular, inward dipping, moderate grade 'ouflow' zones. Mineralisation is controlled predominantly by two steeply dipping faults with high grade (=10 g/t) gold zones centred about structural intersections and fault splays. Gold is submicron size and locked in very fine grained (<10 microns), disseminated and fracture-controlled arsenical pyrite. The Mesel deposit developed where high angle faults guided ascending, sulphidising fluids into reactive carbonate stratigraphy adjacent to and beneath the andesite intrusion and during subsequent migration of the auriferous fluids along 'outflow' channels. Early decalcification and dolomitisation enhanced porosity and permeability prior to silicification and sulphidation. Gold deposition occurred late in the genesis of the deposit.

Jan 1, 1995

By Ed Cope

The newly discovered South Arturo deposit is a Carlin-type, breccia-hosted, oxide gold deposit located on the Dee Mine property at the north end of the Carlin Trend in Nevada. The Dee Gold Mine was discovered in 1981 by Dee Delamar, a Cordex geologist. The property was put into production by Rayrock Mines in 1984. During its 16-year mine life, the Dee Mine produced about 18.6 t (600,000 oz) of gold from oxide ore at an average production rate of about 1.24 to 1.55 t/a (40,000 to 50,000 oz/year). In October 1997, Bar-rick Gold Exploration acquired the right to earn a 60-percent interest in the Dee property from Dee Gold Mines through an exploration farm-out agreement. Barrick Gold Exploration began exploration activities on the property in 1998 and vested in its 60 percent operating interest in 2004. However, the discovery hole at South Arturo was not drilled until July 2005 from atop the historic mine dumps located 300 m (985 ft) south of the Dee openpit. The discovery hole (BD-72) encountered a cumulative 75.6 m (248 ft) of oxide mineralization grading 1.9 g/t (0.055 oz/st) gold beneath 260 m (855 ft) of dump material and post mineral Tertiary Carlin Formation volcanic and sedimentary rocks. Oxidation within the deposit extends to depths greater than 610 m (2,000 ft). The deposit has passed the prefeasibility test and is presently in the feasibility stage. Exploration is on-going to enhance or enlarge the known deposit. The current resource for the deposit stands at 40.4 t (1.3 million oz) of gold. The South Arturo discovery is an example of an historic mining property yielding a new discovery through the diligent application of area selection criteria of the model type and persevering when early results were generally negative. Barrick?s original exploration pro-grams on the property were focused on the discovery of deep, high-grade sulfide deposits that would likely be mined by underground methods. As it turns out, South Arturo is mostly oxidized and will be amenable to mining by openpit methods

Jan 1, 2008

The VoD of non-ideal explosives depends on the charge diameter and the thickness, strength and sound velocity of the confiner. VoD measurements are made to evaluate explosives for development, quality control and prediction of field performance. In testing, the container is part of the explosive system and affects the measured VoD. Pipe testing in cardboard tubes delivers the best guide to unconfined critical diameter and minimum VoD range, but the thickness of the pipes affects the result close to critical diameter, even in emulsions. PVC pipes are not good for unconfined test work because they provide relatively strong confinement. Care must be taken to select pipes with constant relative pipe area across the range of diameters. Where pipes are used to evaluate the effect of confinement, the ratio of the wall thickness relative to the thickness of the detonation driving zone in the explosive needs to exceed 0.5 to create full confinement and the ratio of the detonation velocity to the confiner wave speed must be considered. Estimates for critical thicknesses are provided from Non-Ideal detonation models. The relationship between confinement modulus and VoD is discussed and indexes are suggested for quantifying the explosive performance.

Jan 1, 2007

By William Pitman

Ground instability and subsequent rock falls are a major concern in underground mines. As a result, several mining companies are conducting risk assessments of the ground support installed in their mines. Risk assessment is now commonly used to provide mine operators with a systematic process to assess the strategic risks that could lead to injury to personnel or loss of production, and evaluate possible risk control measures. Risk assessments can be carried out to identify major hazards, assess the likelihood of occurrence and consequences of each hazard and estimate the level of risk. This process allows assessment of possible remediation actions to minimize the risk and likelihood of occurrence or mitigation of the consequences to a level that is reasonable and practical. This paper presents two practical mine applications of qualitative risk assessment of ground support to highlight the benefits of this process. Yearly or bi-yearly risk assessment of mine excavations can assist mine operators in prioritizing resources and areas that require support remediation. In addition, it provides a mechanism for proactive response to reduce potential risks associated with personnel safety, equipment safety and environmental impact. It also provides a baseline against which changes in ground condition can be compared and quantified.

May 1, 2004