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By Guillermo Lazcano Ríos, Jorge Parvez Aguirre, Alejandro G. Díaz Berríos, Felipe Rodriguez Alveal

Codelco's Chuquicamata Underground Division (DCHS) faces issues due to collapses in its subsidence streets, which hinders the advancement of the subsidence front and leads to the loss of areas and reserves. To address these issues, DCHS and Orica have developed a new excavation methodology based on pre-loading explosives in high drawbell. This methodology involves creating a high drawbell extending from the floor of the production level to the ceiling of the subsidence level, divided into two phases: upper and lower. A resin stemming is used to separate these phases. The excavation strategy adjusts to variables such as reception volume, fluffing, and excavation geometry, involving vertical pre-loading of the explosive and the use of a separating element between the phases. The test aims to validate the new excavation methodology and the retaining element in the rock mass, evaluating aspects such as drawbell geometry, impact on the rock mass, damage to the retaining element, and fragmentation. Wireless electronic detonators are used in the upper phase to improve safety by eliminating the need for physical connections, while traditional electronic detonators are employed in the lower phase. Low-density bulk emulsion is applied to minimize the impact on the Crown pillar and extraction point, and higher-density explosives are used above the subsidence level to ensure good caving. Results showed that the blast achieved 86% conformity with the theoretical design. Resin stemming allowed for 100% effective detonation in the upper columns, and the fragmentation obtained was 5.44 inches, which was less than expected. No damage to the Crown pillar was observed, and the explosive quality met standards. This technique has proven to be effective and viable, suggesting the possibility of eliminating the subsidence level in future levels, resulting in savings, and providing greater stability to the rock mass.

Jan 21, 2025

By R. R. Schehl

The Bureau of Mines analyzed the structure of a high-grade metallurgical coke, carbonized at different temperatures in the range 500° to 1,000° C, through X-ray diffraction studies. Growth in the crystallite size with increasing coking temperature was calculated from profile analysis and compared with values obtained from the radial distribution function. The carbonization temperature appeared to have little influence upon the stack height, whereas the layer diameters more than doubled in size between 500° and 1,000° C. These results are comparable to those obtained from other studies of poorer quality cokes.

Jan 1, 1974

By D Oliveira

The use of an equivalent continuum for a rock mass is not always suitable for situations where the failure is structurally controlled by discontinuities as in the case of wedges in the tunnel roof. In these instances, discontinuum approaches are usually preferred. Rock joints that are filled with soft infill are likely to be the weakest planes in a rock mass, having a dominant influence on its overall shear behaviour. In this case, the joint material model adopted for the discontinuities should be able to describe important mechanisms, such as asperity sliding and shearing, post-peak behaviour, asperity deformation, and the effect of the soft infilling. The latest version of a soil-infilled joint model is discussed here. It describes more comprehensively than previous models the occurrence of dilation and compression with lateral displacements, and also represents the hardening mechanism related to asperity interference as observed in the laboratory that cannot be readily captured by the existing joint models. An analytical approach for the analysis of rock wedges structurally controlled by soil infilled joints is presented.The analytical approach presented in this paper for assessment of symmetric rock wedges in the roof of a tunnel demonstrates, that the peak shear strength of the joint may not be fully mobilised before detachment of the wedge. The analysis verifi ed that an overestimation of the factor of safety may be observed when an elastic-perfectly-plastic joint model was assumed, whereas the computed maximum factor of safety was considerably less when adopting an appropriate soil-infilled joint model.

Mar 8, 2011

By W. M. Ma, Daniel W. H. Su, K. Centofanti, Yi Luo, W. L. Zhong, Syd S. Peng

8.1 INTRODUCTION Under room and pillar mining if the extraction ratio is low, (i.e., < 50%) a pillar can support the overburden without collapse and there will be no surface subsidence whatsoever. In other words, if the pillar sizes are properly determined to support the overburden weight, the roof strata cannot cave and there will be no surface subsidence, provided the entry width is small and properly reinforced so that a continuous roof fall leading to piping or the total collapse of the roof does not occur. Under the prevailing mining conditions in the US if only development mining is em¬ployed in room and pillar mining (i.e., no pillar extraction), the extraction ratio is usually less than 50% and generally no surface subsidence will occur. If pillar mining is employed, surface subsidence may occur, depending on the extraction ratio and the size of the remnant pillars (or stumps) left in the gob. The higher the extraction ratio, the smaller the stump pillars left behind in the gob. As the size of the stump pillars becomes smaller, the possibility of their support¬ing the roof without collapse is smaller. If all stumps are crushed completely, the roof will collapse and surface subsidence will occur. Again, whether surface subsidence occurs, it depends on whether the remaining stumps can support the overburden or not. Therefore, since the weight of the overburden which the stump pillars have to support is proportional to the mining depth, a fixed¬size stump pillar can support the overburden at a shallow depth, but will collapse under greater depths. In this context extraction ratio is not the only parameter for determining whether surface subsidence will occur. Furthermore, during pillar mining a pillar can be cut away in many ways to finally reach the same extraction ratio. Therefore, the safest way to determine whether or not sur¬face subsidence will occur in room and pillar mining is to compare the strength of all stump pillars in the panel of interest with the weight of the overburden that the pillars have to support. If they are too small to support the overburden, surface subsidence will occur. Under normal mining practices the characteristics of surface subsidence profiles vary with methods of mining. In the US there are two types of coal mining methods, i.e., longwall and room and pillar mining. Longwall mining is indisputably a total extraction method, but different percentages of coal recovery are achieved in room and pillar mining. Recovery is much higher for sections with pillar extraction than those without pillar extraction. For those without pillar extraction, coal recovery increases with increasing entry width or decreasing pillar size. For those with pillar extrac¬tion, the total coal recovery depends on the amount of coal ex¬tracted from each pillar which is in turn dependent on the roof conditions and methods of pillar extraction. Bauer and Hunt (1981) analyzed the characteristics of the subsidence profiles due to the three types of mining methods in Illinois. Fig. 8.1 shows the range of locations associated with the following profile features: profile edge, maximum tensile curvature, maximum slope, maxi¬mum compressive curvature, and maximum subsidence. The scat¬ter of the data increases with decreasing percentages of extraction. The profile is much sharper for longwall mining and becomes milder as percentage of extraction decreases; for longwall mining the profile edge is always beyond the opening, while in room and pillar mining it may fall within. The location of maximum tensile strain moves outward, while that of the maximum compressive strain moves inward as the percentage of extraction decreases and the maximum subsidence decreases with decreasing extraction ra¬tio. The general concepts discussed in this chapter may not apply to the entry development sections between two longwall panels where roof caving in both panels are very severe and overlap each other. Consequently, it must be treated differently. 8.2 CASE STUDIES Kohli et al. (1982) described several case studies on surface subsidence due to room and pillar mining with pillar extraction. In one case, the seam was 6 ft thick and ranged from 650 to 800 ft deep. The panel was 720 ft wide (Fig. 8.2) and was developed with seven entries. Each entry was 18 ft wide and the pillars were driven at 88 ft center-to-center between entries and crosscuts. Dur¬ing retreating, two additional pillars (pillars J and K) were devel¬oped, two blocks ahead of the pillar line, to take the barrier pillar, 160 ft wide, between panel No. I and No. 2. Pillar extraction began by taking a central slice across the whole panel, beginning from pillar 0 through J. After that, each pillar was sliced by 6 wing-slices, 3 on each side, leaving 4 stumps, 2 on each side. One of each pair was triangular shaped approximately 6 ft wide by 26 ft long. The other was trapezoidal, 26 ft long by 16 ft and 10 ft wide at the upper and lower bases, respectively. This resulted in a total extraction ratio of 85% within the panel. Obviously the stump pillars were not large enough to support the overburden and col¬lapsed completely. The final subsidence profiles were rather smooth, just like those under total extraction (Fig. 8.3). The max¬imum subsidence on Lines A and B was about 3 ft for an average mining height of 6 ft, but the locations of the maximum subsidence were skewed toward panel No. 1. Along Line D, the central por¬tion was flat-bottomed, but the maximum subsidence was only 2 ft, mainly because Line D was only two blocks from the right-side edge. Notice in every profile there was heaving of as much as 0.3 ft, beyond the edges above the solid coal. In another case (Fig. 8.4), the seam was 6 ft thick under 650 to 850 ft cover. Three panels, 710, 400, and 350 ft wide, respec¬tively, were extracted side by side. The entries were 18 ft wide and the pillars at 88 ft center-to-center between entries and crosscuts. Pillar extraction was practiced the same way as described previ¬ously. According to the mine maps, there was no barrier pillar left between the adjacent panels. Thus after extraction the three panels formed a common gob of more than 1460 ft wide. This was re¬flected in the development of the subsidence profile along the county road. After Panel No. I was mined, the maximum subsi¬dence of 1.2 ft occurred at Monument No. 1. This was increased to 1.6 ft after the extraction of Panel No. 2 and went to a further 1.8 ft after the extraction of Panel No. 3. The final maximum subsidence of 4.3 ft occurred at Monuments No. 6 through No. 8, while the remaining portion of the county road showed rebound. However it must be noted that the subsidence profiles were rather bumpy and fluctuated, perhaps an indication of the irregular stump pillars that were left and crushed at various time after mining. Moebs (1982) surveyed the subsidence profiles of four room

Jan 1, 1992

By Pierre Marcotte

Jackleg rock drills expose their operators to high hand-arm vibration levels. To address this problem, an antivibration handle has been developed. The antivibration properties of the handle were achieved through a suspension based on a hinge design, achieved with a torsion spring and o-rings. To provide some bench marking for this new handle and to follow the evolution of its performance over time, a test bench was developed to characterize the vibration emission values of jackleg drills under controlled operating conditions. Testing of four antivibration handles on the test bench have shown reductions of the overall weighted acceleration level in the order of 54 % to 60 %, with most of the attenuation occurring in the zh-axis at the rock drill percussion frequency.

Jan 1, 2008

In order to assist in the building up of the industries and commerce of France, by cooperating with the Congres General du Genie Civil, the four Founder Societies have appointed the following committee: E. G. Spilsbury and Arthur S. Dwight representing the American In-stitute of Mining and. Metallurgical Engineers; George F. Swain and Nelson P. Lewis to represent the American Society of Civil Engineers; Charles T. Main and George W. Fuller to represent the American Society of Mechanical Engineers; Lewis B. Stillwell and Andrew M. Hunt to represent the American. Institute of Electrical Engineers.

Jan 12, 1919

By W. W. Agey

Under terms of cooperative agreements, the Federal Bureau of Mines Inter-mountain Experiment Station assisted the General Services Administration (GSA) in operating the manganese-ore purchasing depots at Deming. N. Mex., and Wenden, Ariz. A resident metallurgist stationed at the depots directed sampling of the ore lots for assaying, took samples for amenability testing, calculated acceptances for payments to shippers, and served as technical adviser to the depot manager. The Intermountain Experiment Station at Salt Lake City, Utah, tested the amenability of the ores to determine how much of the manganese could be recovered in concentrates assaying a minimum of 40 percent manganese (Mn) and containing a maximum of 16 percent iron (Fe), 15 percent silica (Si02) plus alumina (AI203), 0.3 percent phosphorus (P), and I percent copper (Cu) plus lead (Pb) plus zinc (Zn). All chemical analyses of the ores and test products were done at Salt Lake City. In all, 567 ore lots submitted to the Deming depot were tested by flotation to determine their amenability to concentration into specification-grade material; 421 lots were amenable, and 146 were unamenable, primarily because of high copper content or excessive silica or iron in the flotation concentrate. Chemical analyses totaling 42,700 were made on the ores submitted to the Deming depot and on the products of the amenability-test program.

Jan 1, 1959

By P. J. Mackey

Over the course of Dr. Robertson’s career, the ferrous and non-ferrous plants have seen enormous changes in technology and increases in plant capacity, essentially amounting to a “technological revolution”. In iron and steel, the “mega” blast furnace of some 6,000 m3 working volume is now standard (~10,000 tonnes (mt) of pig iron/day). Similar huge changes in process technology and plant size have occurred in the non-ferrous industry. As an example, the fuelfired reverberatory furnace, once the mainstay of the copper industry, has disappeared - replaced by large capacity flash and bath smelting technologies. The energy consumption per unit mass of metal produced has also been reduced considerably. Our understanding of the thermodynamics and mechanisms of metallurgical reactions, a field to which Dr. Robertson has significantly contributed, has made great strides. This paper reviews these changes with particular reference to the copper smelting industry, providing also comments on expected future trends.

Jan 1, 2014

By J. L. Feng, F. Zhang, S. K. Hu

"Cracking is one of the most serious damages in tunnel linings, which is often observed by in-situ inspection to tunnel engineering. It is difficult to better characterise for the deformations from continua to discontinua by the traditional continuum mechanics due to emerging of crack initiation, crack propagation, and coalescence, at several scales. In this paper, the process of failure that was from micro damage to macro crack in concrete tunnel linings under experimental load conditions, was quantitatively simulated by the combined finite-discrete element method (FDEM). Comparison of the simulating results with that obtained by the full-scale test of concrete lining under different loading conditions in the literature was respectively good agreed and shows that such FDEM technique is appropriate to model the continuousdiscontinuous problems. The present study indicates that the FDEM technique could be benefit to quantitatively evaluate the stability and damage level of tunnel lining system. INTRODUCTIONMost damage to tunnel linings is caused during the construction phase, which can emerge in a form of single macro cracks or even large-scale spalling. The cause of such damage is mostly the occurrence of, either impermissibly large contact stresses resulting from high thrust loads combined with production and installation tolerances in the segments resulting in a geometrical lack of fit (Maidl et al., 2012; Transportation Research Board, Basic Research and Emerging Technologies Related to Concrete Committee, 2006), or unallowable tensile/shear stresses appearing in the tunnel linings under complicated loading conditions (Ansell, 2010). Such a macro failure phenomenon may also be regarded to as a result of growth, coalescence, initiation of microcracks in tunnel linings due to the complicated loading action, and even potentially lead to an instability or collapse of the local or entire tunnel system. Much work has been devoted to investigate for the characters of micro and macro cracks, the influence factors of damage patterns, the settlement of a tunnel, the width, length and density of lining cracks, especially from classifiedly and qualitatively (Wang, 2010; Huang et al., 2013). However, the safety evaluation of tunnel is definitely required to not only inspect in situ, but further also quantify computing and analyzing for tunnel lining systems, so as to provide a global viewpoint of process of the tunnel deformation, which aims to help for subsequent decisions, e.g. controlling traffic flow or immediately repairing task. Additionally, the design of lining thickness and its structural analysis that is mostly determined by previous experience and similar construction examples is also needed to be improved."

Jan 1, 2015

By Kazuteru Tozawa

Surface tension, density and viscosity coefficient of acidic copper sulfate solutions simulating electrolyte solution for electrolytic production of copper have been measured and are summarized as a function of the solution composition and temperature. Nickel sulfatc is taken to be a representative coexisting sulfate that affects the physical properties of the copper sulfate electrolyte solution. The sulfates in the solutions, including sulfuric acid, increase the physical properties under consideration. Gelatin, representing organic additives for electrolyte solution, was found to sensitively affect the surface tension of the solution at low concentration of up to 200mgIL. The surface tension was markedly lowered by the addition of gelatin and the increasing concentration of gelatin led to a more pronounced decrease in the observed values. The surface tension gradually increases with time after addition of gelatin and the increase is enhanced by higher acid concentration and higher temperature. The observed change in the surface tension is considered to reflect the preferential adsorption of gelatin to the bubble-solution interface and the decomposition of gelatin due to hydrolysis to form various molecules having smaller molecular weights.

Jan 1, 2003

By B. R. Stewart

The exclusion of coal fly ash from regulation as a hazardous waste has led to increased interest in returning ash to the coalfields for co-disposal. A multi-year column study used coal refuse with high potential acidity (4% total-S), and two alkaline ashes in various bulk-blends. The unamended refuse acidified rapidly, producing acidic (pH 1.7) leachates, high in dissolved metals. Treatments blended with sufficiently high amounts (20 to 33%) of alkaline ash produced alkaline (PH >8.4) leachates with low metal levels, while lower ash blending rates acidified over time to produce metalliferous leachates. Examination of solid materials indicated that pyrite oxidation was occurring to some extent in all treatments. A regression equation (r = 0.95) was developed to estimate time to net acid release for mixtures of these materials. Ash alkalinity and refuse potential acidity must be balanced to insure water quality for co-disposal environments, or the ash materials should be isolated away from acidic drainage waters.

Jan 1, 1997

By S. Combrink, N. Naude

Hydrocyclones are widely used in wet mineral classification applications and can be notoriously prone to wear. The methods currently used for quantifying hydrocyclone wear profiles are mostly twodimensional. In this paper it was attempted to quantify and monitor wear in three-dimensions, using a micro-CT scanner (Nikon XTH 225 ST). Hydrocyclone wear and, more specifically, spigot wear was investigated on a Multotec HC250 (Dc 250 mm, barrel diameter) hydrocyclone. The hydrocyclone is lined by a vulcanised natural rubber compound within a metal shell, with a thermosetting polyurethane spigot and vortex finder. Tests were conducted on the hydrocyclone in a pilot-scale recirculating test setup. The wear rate of the spigot was quantified by measuring the volume change within the spigot over 122 days in operation, using the micro-CT scanner. After 122 days in operation (recirculating feed materials) the wear in the hydrocyclone led to only a 13.67% loss of spigot volume. The relatively low wear rate could be related to the fact that material recirculation led to particle smoothing and a reduction in size, thereby decreasing the wear rate. This also shows the enormous influence that particle shape and size have on wear. The research was conducted with the support of Multotec, NECSA and the University of Pretoria. Keywords: Wear, hydrocyclone, micro-CT

Jan 1, 2020

By Yujing Jiang, Xuezhen Wu, Bo Li, Tomomi Iura, Gang Wang

"Yielding bolt has been widely used for rock reinforcement in mining and civil engineering in high stress conditions. However, the interaction mechanism of yielding bolt and matrix mass is still not clear at present, and no analytical model is available to predict its reinforcement effect quantitatively. In this paper, the mainstream yielding rock bolts were reviewed firstly. After that, a coupling model is proposed to account for the interaction between yielding bolt and matrix mass. Based on the plane strain axial symmetry assumption and the incremental theory of plasticity, equilibrium equations and compatibility equations of matrix mass, and also the response of yielding rock bolt are deduced theoretically. The proposed method was programmed in Visual Basic development environment, and a semi-analytical solution for the coupling model is achieved. The reinforcement mechanism of yielding bolt in conventional tunnelling is clearly demonstrated through an illustrative case study. The reinforcement effect of yielding bolt is estimated quantitatively, and the distribution of axial force along the bolt is presented. In addition, the validity of proposed method is verified by numerical simulations. INTRODUCTIONHigh stress in surrounding rock mass can cause serious stability problems such as massive squeezing in soft rock and rock burst in hard rock. It is observed that many conventional rock bolts failed when experiencing large displacement, which implies that they are too stiff to sustain large deformation and dynamic conditions (Stillborg, 1994; Hoek et al., 1995). Yielding bolt, which can also be called energy-absorbing bolt, has been widely used for rock reinforcement in mining and civil engineering for these conditions (Neugebauer, E., 2008). This paper focuses on the static reinforcement effects of the yielding bolts with analytical method and numerical simulations."

Jan 1, 2015

By H Ishikawa, H Zhou, M Spagnolo, R Halomoan, S Palaniandy

The Karara Mine concentrator commenced operation in November 2012 to produce high-grade magnetite concentrate. The circuit consists of three stages of comminution plus magnetic separation at each stage for early gangue rejection. High-Pressure Grinding Rolls (HPGRs) are used for fine crushing, followed by ball mills and TowerMills in secondary and fine grinding duties respectively. In recent times, the plant throughput rate has had to exceed design by 18% to maintain the targeted concentrate productivity. Meanwhile, the ore hardness has increased by 25% to 15 kWh/t, leading to changes in the circuit operation and design. These changes led to coarser and harder feed to the TowerMill circuit that affects the subsequent separation processes. Nippon Eirich (the TowerMill manufacturer) has partnered with Karara Mining Limited (Karara) to optimise the TowerMill circuit operation to achieve a circuit product with an 80% passing (P80) of 35 &#181;m. Several activities, including a TowerMill circuit audit and performance evaluation, multi-component analysis, modelling and simulation, were carried out as part of the optimisation exercise. The circuit audit results revealed that the TowerMill circuit feed size F80 had increased and ranged from 67–75 &#181;m. The circuit product P80 ranged between 42–55 &#181;m. The average circuit specific energy was 3.4 kWh/t. Moreover, the differences in density between magnetite and gangue minerals had affected the hydrocyclone performance where the efficiency curve exhibits coarse end deflection indicating that low-density coarse material was bypassing to the overflow stream. A multi-component analysis was carried out to further investigate the coarse end deflection of the hydrocyclone efficiency curve. The analysis showed that the magnetite and gangue minerals had different cut sizes of 32 and 64 &#181;m respectively. Furthermore, 33% of the magnetite (below 35 &#181;m) was being recycled back to the mill due to the density effect. Based on these findings, several recommendations for process improvement were identified and implemented including:• converting to a bottom feeding configuration to improve the grinding efficiency in the mill• additional grinding power in the fine grinding duty to compensate for the increase in plant throughput and ore hardness• improved classification efficiency by converting to smaller size hydrocyclones. Simulation studies carried out to evaluate the impact of these implementations had indicated that a P80 close to 35 &#181;m with a steeper size distribution could be achieved by using the bottom feed configuration with additional grinding power. The TowerMill circuit was converted to bottom feed configuration and the performance was evaluated. The operating work index, size specific energy and circulating load were 30% lower than the top feed configuration, indicating that the bottom feed configuration had a higher grinding efficiency. This paper discusses the outcome of the circuit audit, multicomponent analysis, circuit simulations, improvement strategies and quantification of the improvements. CITATION:Palaniandy, S, Spagnolo, M, Halomoan, R, Zhou, H and Ishikawa, H, 2017. Fine grinding circuit process improvement at the Karara mine concentrator, in Proceedings MetPlant 2017, pp 101–120 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 11, 2017

By C J. Unger, A M. Lechner, A Walton, V Glenn, M Edraki, D R. Mulligan

In Australia, responsibility for abandoned mines can be unclear. With a few exceptions, state and territory governments have become responsible for abandoned mines on Crown land and private landholders are responsible for abandoned mines on freehold land. Abandoned mines pose risks and opportunities across the full suite of sustainability themes (environmental, human health and safety, socio-economic and community development) after mining ceases.Across Australia, there are more than 50 000 abandoned mines recorded, ranging in size from individual shafts to large polluting open cut mines. Addressing the safety and environmental impacts of these sites, as well as the socio-economic development opportunities of significant rehabilitation projects, requires the implementation of effective policies and programs. Qualitatively defined graded scales or ‘rubrics’ have been used for many years in student assessment and increasingly in the performance assessment of other sectors (Davidson, 2005). While risk-based maturity models have been used for safety prioritisation (Westrum, 1993; Hudson, 2007; DRET, 2008; Hancock, 2010) and evaluation of organisational development (Esteves et al, 2010), this paper is the first to adapt maturity models for the evaluation of abandoned mine rehabilitation programs in Australia.In this paper, we used indicators based on the national policy for abandoned mines to assess the maturity of abandoned mine programs. Fourteen elements of maturity were identified (Unger et al, 2012) under the five chapters of the national policy for abandoned mines (MCMPR and MCA, 2010). Jurisdictions were then evaluated and ranked on their maturity on the basis of information that was web-accessible. Information transparency of abandoned mines policies and programs is in itself one element of the maturity chart. Each jurisdiction in Australia was compared with one another and with a Canadian jurisdiction; the British Columbia (BC) Crown Contaminated Sites Program (CCSP), which represents leading practice (Unger, 2009; Unger et al, 2012).Our research found that in some jurisdictions in Australia there is little or no information on abandoned mine programs or individual abandoned mine sites on government websites, while other jurisdictions provided very detailed program information on planning, funding and implementation. Each jurisdiction in Australia, for which web-based information was available, had some elements where they were more mature than other Australian jurisdictions. For some elements, maturity ranked equivalent to the BC CCSP, but most Australian jurisdictions were less mature. There was only one element where a single Australian jurisdiction ranked higher than the BC program in maturity. Legacy mine policies, programs, priorities and funding information via departmental websites are easier to find in some jurisdictions, indicating progression toward greater transparency and more mature programs. A systematic approach to monitoring and evaluating abandoned mines programs is essential for accountability as it can demonstrate liability reduction over time and continual improvement. The maturity model approach can provide both a measure of progress over time as well as a tool to support the implementation of the national policy for abandoned mines (MCMPR and MCA, 2010).CITATION:Unger, C J, Lechner, A M, Walton, A, Glenn, V, Edraki, M and Mulligan, D R, 2014. Maturity of jurisdictional abandoned mine programs in Australia based on web-accessible information, in Proceedings Life-of-Mine 2014 , pp 467–480 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 16, 2014

By A. W. Fahrenwald

As an aid to studies of grinding, a batch laboratory classifier and elutriator has been developed at the U. S. Bureau of Mines Metallurgical Field Office at Moscow, Idaho, in cooperation with the University of Idaho. It is designed for use in conjunction with other batch laboratory apparatus, such as ball mills and flotation machines. It is shown in Figure 1. Description and Operation of Batch Laboratory Classifier The classifying column is a cylinder 4 feet 2 inches long, of 5 3/4-inch inside diameter, and terminates at its top in a cone or funnel 18 inches in diameter at the overflow elevation. The pressure chamber is of the same diameter as the classifying column, and has a conical bottom; it is provided with two openings, one in the cylindrical portion through which water is admitted, and the other, a small drain hole, in the apex of the conical bottom. The two compartments are bolted together and communicate with each other through the medium of a constriction plate provided with 3/16-inch holes spaced 1/2 inch apart. Constriction plates having any other percentage of opening may be substituted as the needs demand water is admitted to the compartment below the constriction plate at constant pressure. The volume of eater admitted is controlled with a valve. It should be mentioned that the classifying column at the bottom, just above the constriction plate, is glass.

Jan 1, 1929

By G. Demi

Chromium ore exploitation has commenced in 1948 in Albania and there are produced about 24.3 million tons up to the end of 2006. Last years the prices and need were increased notably. Main export of chromium ores and concentrates was accomplished respectively 114 500 ton and 13 000 ton in People Republic of China during 2006 year. Location, status, reserves and grade of chromium deposits of Albania are presented in the paper. Chromium ore enrichment is related with their composition and Cr2O3 content, as well as with market requests. Actually, high grade chromium ores with Cr2O3 content over 38% ~ 40% are directly exported or treated in metallurgy, but last time have been exported and chromium ores with 28% ~ 30% Cr2O3 content for the reason that dressing plants can not be operated for investment lack. High grade chromium ores have been treated in heavy media separation plant before 1990. It is presented flow sheet, figures and necessity to put in operation HMS plant in the paper. Actually, low grade chromium ores with content 16% ~ 20% Cr2O3 are not enriched for the lack of investments, but before 1990 they have been concentrated in dressing plants producing chromium concentrates grade 48% ~ 50% Cr2O3. In the paper are shown technological schemes and industrial figures during more than 20 years operation and possibilities to reconstruct these mineral processing plants. Only chromium mineral processing plant of Bulqiza is in operation using as raw material its old existing tailings. The flow sheet and figures show that are possibilities to increase recovery and grade of concentrate in the plant. Separation and concentration of slimes in suitable equipment gives visible improvement of Cr2O3 recovery. Several tests have been carried out with flotation and magnetic separation methods for enrichment of low grade chromium ores. Flow sheet and figures of mineral processing ores content olivine as gangue mineral are interesting, because combination of gravimetric with magnetic methods is necessary in this case. In the paper are presented the tests and good results of low chromium ore pre concentration. Metallurgical treatment of chromium ores to produce Ferro chrome, as well as technological process, figures, consume of electrical power and materials in Elbasani and Burreli furnaces are shown and necessary investments to treat in metallurgy all produced chromium ores in Albania are considered.

Jan 1, 2014

By Elaine R. Goetz

The eastern and southeastern regions of Ohio are part of Appalachian coal country, with a documented 1,300 miles of streams impacted by acid mine drainage. Since 1995, partnerships of local watershed groups, government agencies, private industries and universities have focused on the remediation of the affected waterways. The partnerships have spent $2.4 million annually in constructing and operating treatment systems such as wetlands, anoxic limestone drains, successive alkalinity production systems, steel slag leach beds, and lime dosers. These systems have been partially successful, and water quality has improved in the region, but no treatment has yet met the ultimate objectives of low cost, continual remediation. Our research group has investigated an alternative system for acid mine drainage treatment: electrolysis combined with pH adjustment to selectively precipitate recoverable, saleable metal products to offset the cost of treatment. We utilized a $0.2 million pilot facility constructed by the Wayne National Forest, in Nelsonville, Ohio, to process approximately 40 L/min of mine drainage from surrounding seeps. The pilot facility has multipurpose tanks that can be adapted to process acid mine drainage with different characteristics, such as high or low aluminum content or high or low carbonate concentrations. Sodium hydroxide was used for pH adjustment. Results indicate that saleable product can be generated with this treatment technology. In an initial, not optimized trial, one seep yielded 4.7 kg of dried 69 % iron oxide product from 10 hours of pilot facility operation, treating 23,000 L with electricity costs of $4.68. Dissolved iron concentrations in the water were reduced from 242 ± 18 mg/L to 2.8 ± 2.0 mg/L (99.0 % removal) while the initial acidity of 443 mg/L as CaCO3 was neutralized to an alkalinity of 160 mg/L as CaCO3. Iron oxide can be sold as pigment for an average of $1.47 per kg in the United States. Currently, approximately 80 % of all pigment used in the United States is imported, and there is incentive to purchase domestically recovered pigment. If the iron oxide from the pilot facility could be sold as pigment for $1.47 per kg, 43 % of the operating cost (raw materials, electricity and construction costs) of treatment could be offset by the sale of the product. Considering the cost offset, the electrolysis process shows potential as a sustainable low cost alternative for treating acid mine drainage.

Aug 1, 2013

By William H. Hallahan

The Friedensville zinc mine of The New Jersey Zinc Company is located about four miles south of Bethlehem, Pennsylvania in the Saucon Valley, an infolded and down faulted block of Cambro-Ordovician carbonate sediments surrounded by hills of pre-Cambrian gneiss and Cambrian quartzite except where breached by Saucon Creek at its entrance into the Lehigh River. The Beekmantown formation of lower Ordovician age is the host for zinc and limonite ores, the former occurring in the lower dolomitic portion thereof and the latter in the upper interbedded dolomite and limestone portion of the section. The Beekmantown is overlain unconformably by the Jacksonburg of middle Ordovician age and according to Miller (2) underlain conformably by about 2500 feet of Cambrian dolomites and 200 feet of the basal Cambrian Hardyston quartzite. Igneous rocks, younger than the Cambro-Ordovician carbonate rocks, are absent in the vicinity of the ore deposits. The major structure is a N60°E-trending saddle-shaped doubly plunging anticline overturned to the north. A syncline lies to the north of the anticline. A thrust fault of substantial displacement repeats the south limb of the anticline. The ore consists of sphalerite and pyrite, along with gangue dolomite and quartz, filling and replacing the sedimentary type matrix of a strata-bound solution collapse breccia in the dolomitic portion of the lower Beekmantown. This breccia was produced in pre-Jacksonburg time as a consequence of uplift, erosion, and development of a karst topography and subsurface drainage system in the Beekmantown. Deposits of both zinc and limonite are present on the vertical north limb, on the crest and on the gently dipping (25°) south limb of the Friedensville anticline; consequently, there is no preferred structural setting for mineralization. Rather this distribution supports the view that the folding is post-ore. Features confirming this view are the development of flow cleavage in the breccia matrix, the deformation of some of the sphalerite, and the fracturing of the pyrite mineralization therein. There is nothing to indicate that the age of mineralization is other than Ordovician and older than Taconic deformation. In the absence of intrusive igneous activity younger than the ore host, the writer suggests that submarine volcanic exhalations associated with the volcanic activity recorded in the eastern eugeosynclinal facies of the Ordovician may be the source of the metal ions.

Jan 1, 1968

By D. E. Ranta

Orebodies with intricate three-dimensional shapes (such as porphyry molybdenum deposits) are difficult to model using geostatistics or other computer-assisted methods. The strong trends in their spatial grade distributions occur as mathematically complex shapes and prevent the customary assumption of local stationarity that is required for accurate ore reserve estimation. "Zoned Kriging", a new method developed by AMAX, solves this problem directly by using the geologist&apos;s interpretation of the ore zones as a parameter in the kriging equations. The geological basis and methods used in the Climax ore zoning technique are reviewed and discussed relative to their application in zoned geostatistical modeling. The Zoned Kriging technique is described and demonstrated using data typical of mineral deposits.

Jan 1, 1984