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By S. Serata

Background Results of the above study in the August 1991 issue of Mining Engineering offer valuable lessons in the solution of cutter-roof problems. The original study plan was initiated by the discussion authors to solve the problems using the "stress control method" of mining (Serata 1976,1982; Serata, Carr and Martin, 1984; Serata and Gardner, 1986; Serata, Gardner and Preston, 1986; Serata, Gardner and Shrinivasan,1986; Serata and Kikuchi, 1986; Serata, Preston and Galagoda, 1987) However, the plan and the planner were changed to the arrangement reported in the paper. The change was considered reasonable at the time due to the mine engineers' uncertainties about the stress control method. Consequently, the basic principle of the study was shifted from the original stress control method to the method described in the paper, which will be called the "yield pillar method" for the purposes of this discussion. The paper convinces the reader that the yield pillar method fails to solve the cutter-roof problems. This doesn't mean that the stress control method also fails. Actually the contrary is true, as discussed below. Limitation of the yield pillar method The paper illustrates clearly how poorly the yield pillar method performs in solving the problem. The reason for this failure is the lack of the protective stress envelope needed to stabilize the cutter roof. Unfortunately, the protective envelope cannot be formed properly without utilizing the stress control method of mining. Changing the pillar size does not make much difference in the roof stability.

Jan 1, 1992

By J. H. Tu

The technical paper correctly points out that the kriging variance is not a good measure of the uncertainty of the estimated (i.e., kriged) value of individual blocks. The au thors claim that their proposed jackknife method, which is a rekriging of each block by eliminating, in turn, one sample from the original sample set and then taking the average of the rekriged estimates, not only gives good block estimates, but the resulting jackknife kriging standard deviation is a useful indicator of the "true uncertainty associated with block estimates." However, they immediately abandon the idea of using the block-by-block standard deviations, reasoning that these standard deviations are not independent and that there is no easy way to utilize them. There may be another reason for not using them. The jackknife standard deviations for individual blocks given in their example are mostly in the range of 0.004 to 0.005 oz/st (0.14 to 0.17 g/t) with only one block having a high value of 0.012 oz/st (0.41 g/ t). These individual block standard deviations are as low as the jackknife standard deviation for the mean grade of the entire shape, i.e., 0.0041 oz/st (0.14 g/t). Do they represent the "true uncertainty" of the individual block estimates? Could the authors explain this? In a global shape consisting of a large number of blocks, any given sample will affect the kriged estimate of only those few blocks within its vicinity. This is the rationale for the authors' selective rekriging, making the jackknife algorithm more efficient. On second thought, why not do away with jackknifing altogether? Just cumulate and normalize, if necessary, the kriging weights of each sample used during the ordinary block kriging process, and then compute the global variance from these kriging weights and their respective sample grades? After all, isn't the global mean grade nothing but the weighted average of the samples used in the estimation? Reply by G.S. Adisoma and M.G. Hester The jackknife is one of the many tools in a practitioner's toolbox to solve estimation problems. The strengths of the technique lies in its simplicity, i.e., it uses the concept of mean and standard deviation and the fact that it can be easily combined with other tools, in this case kriging. Because the jackknife kriging (JK) estimate is also the mean of the pseudovalues, the JK standard deviation is attractive just as the standard deviation of the mean explains the variability of the data. The difference is that the pseudovalue calculation in jackknife kriging uses the ordinary kriging (OK) weighting scheme instead of simple arithmetic averaging. The data used to illustrate the jackknife technique in the paper consist of high values that are roughly three times the low values. The resulting JK estimate of the block grades show that the highest estimate is roughly twice the grade of the lowest estimate. The contrast between the low and the high estimate is more evident in the JK estimate than in the OK estimate, even though the mean grades of the blocks for the two estimates are very similar. Nonetheless, in this paper, we are concentrating more on the need for a more realistic measure of uncertainty, or precision, for the estimate. Unlike its OK counterpart, the JK standard deviation of the blocks clearly reflects the original data variation. The highest JK standard deviation of the blocks is three times its lowest value. This follows our intuition that, when the samples used to estimate a block is more variable, the resulting estimation variance (or standard deviation) should be higher than the case where the samples are more uniformly valued. However, block-by-block standard deviation or variance is of little practical value in reserve estimation and classification, as well as in mine planning. One is usually more interested in quantifying not the variance of the individual block estimate, but the uncertainties associated with a much larger dimension, such as the minable reserve. Thus, the thrust of the paper is to find a simple way to obtain a single estimation variance or standard deviation associated with the reserve grade estimate. The discussion by J.H. Tu did not mention how one would obtain the global variance from the OK weights and the sample grades. As a technique that offers a data valuebased measure of uncertainty for its estimate, the "leave-one-out" jackknife fills this need nicely through the block kriging shortcut approach described in the paper. Note: The first column and the last two columns of Table 3 in the paper should have contained a single number each, namely, an OK estimate of 0.0317, a JK estimate of 0.0333 and a JK standard deviation of 0.0041 oz/st, respectively, for the shape, as are obvious from the text. ?

Jan 1, 1998

By Jan C. Still

The Haines-Klukwan-Porcupine subarea of the Juneau Mining District (JMD) is located in the northern portion of the JMD and consists of the land between the Alaska-British Columbia boundary, the Chilkoot Valley, Glacier Bay National Park boundary, and extending south to the Davidson Glacier and the Chilkat Peninsula and Islands (fig A-I). This area was studied as part of the larger JMD study in a cooperative effort between the Alaska Division of Geological and Geophysical Surveys (ADGGS) and the Bureau during 1985 to 1988. Prior to this, during 1984, the ADGGS and the Bureau cooperated in a study of the Porcupine mining area and during 1981 to 1983 conducted separate studies in the area. Field work was conducted by foot, boat, truck, and helicopter. Figure A-I shows the Haines-Klukwan-Porcupine sub-area and shows the areas covered by previous Bureau and ADGGS studies. PHYSIOGRAPHY AND CLIMATE The physiography of the subarea ranges from gentle to rugged. Glaciers formed the major features in the sub-area and left "U" shaped steep walled valleys and rugged mountains. The Chilkat, Chilkoot, Klehini, Tsirku, and Takhin are the major rivers in the subarea. The higher mountains are glacier clad and some valleys still harbor glaciers that reach the valley floors. Lush forests and dense brush predominate up to timberline at about the 2,000-foot elevation. Low point in the subarea is at sea level in the area around Haines while the high point is 7,434 feet at: Mount Henry Clay in the Porcupine mining area.

Jan 1, 2012

By M. B. Ramachandra Roa

This paper records the results of the earth resistivity surveys made in the Godavari River in connection with the Ramapadasagar project. After describing the topographical and geological features of the area, the results of the investigation at the several sites have been mentioned. Further, a detailed account has been given of the resistivity methods adopted, the apparatus used and the interpretation of the curves. The resistivity measurements carried out over the water course have been specially outlined, and a review of the verification of the electrical indications by drilling in a number of cases, has also been presented. Introduction The Government of Madras has under consideration the Ramapadasagar project which is estimated to cost some Rs.660,-000,000 (about two hundred million dollars). The most essential feature of this project is the construction of a high dam across the Godavari near Polavaram for impounding the waters of this great river of South India, whose maximum discharge is 2 million cu secs in flood time. It is estimated that when the proposed reservoir is constructed, about one million tons of rice could be produced annually under irrigation. To those who have had any acquaintance at all with the recent distressing famine conditions which India has passed through and is still undergoing, the Lounteous effects which this project will bring about to the country in the future, are obvious. In the region of the proposed reservoir, the river flows for several miles through a hilly tract, hemmed in between flanks of hard, solid rock; but the river bed is mostly sandy. A few bore holes put down in this sandy bed touched rock at a depth of 180 to 200 ft. The river bed is about jo ft above the mean sea level (m.s.l.) and the bedrock was found at —125 to —160 ft m.s.l. The crest of the dam is proposed at +162 ft m.s.l. The proposed dam mould therefore be about 120 to 130 ft above the river bed, and the foundations would have to reach down nearly zoo ft below the bed at the deepest part. A number of alternative sites were proposed for investigation, and it became necessary to quickly ascertain the bedrock topography at these sites. Dr. M. S. Krishnan,* who was consulted in the matter, furnished a detailed account of the geological features of the region and suggested a geophysical investigation which the writer carried out from February to April 1944, employing solely the electrical resistivity methods. The total length of the survey lines for the electrical measurements amounted to 62,600 ft and the number of resistivity determinations was about 2500. Altogether nine alignments for the dam were inten-

Jan 1, 1949

By J. B. Gwiazda

S.S. Peng The highlights of this paper are: The shield supports with lemniscate links will exert a tangential force on the roof strata, causing unnecessary and harmful roof movements, thus damaging the roof structure. The principal source of the tangential force is the horizontal movement of the point O, the pivotal joint between the canopy and the caving shield. When the support height is changed during setting, the point O will move along a curve upward and rearward toward the gob. As a result, it produces a vertical movement, causing the roof to move up a distance of t d, and a horizontal movement toward the gob, creating vertical fractures in the roof. The major purpose for employing the lemniscate- link system is to protect the hydraulic legs against the action of the bending moments induced by the tangential force that causes deformation or destruction of the legs. Based on these two points, the author suggested that, in order to eliminate the harmful effects of the lemniscate-link system and strengthen its advantageous features, a shield support with a new design of that mechanism has been developed. With the new design, the conventional rigid lemniscate links are replaced by hydraulic rams. When the pressure inside the rams reaches a certain level, the ram will yield.

Jan 1, 1987

By P. A. Cornejo

Blastings always generates induced vibrations through seismic wave, these are lost energy from the blasting and therefore causing the disturbance of slopes and structures near the blast source. Most of the components of an open pit mine have to reach a physical stability, especially during its closure process. Aiming to develop a vibration prediction model using the historical records from an open pit mine located on Peru, this research has two main focuses depending on the distance from the blast source to the structure: to protect the final slopes and to protect the villagers houses. Both are based on the Devine model of the Square Root Scaled Distance (SRSD) to predict the Peak Particle Velocity (PPV). The Mid field model, where is found mainly final slopes from the pit, limited by the failure criteria of McKenzie based on the geo mechanical properties of the rock mass. The Far field model ranges distances where are located the most community households, limited by the German rule DIN 4150 for highly sensitive structures. Both scenarios resulted on a distribution abacus to limit the weight of used explosives to get the best operational advantage and structure safety.

Mar 2, 2022

By D. Nilsson, B. Aaro

To find out if there are any potential savings in "in-the-wall haulage for open pits," the Swedish companies ASEA and Kiruna-Truck in 1984 gave us financial support to study this solution in more detail. In the study, the use of in-the-wall haulage was studied for three hypothetical open pits with different sizes and shapes. Annual ore production rates also varied. A part of this thorough study was published by the Swedish Mining Research Foundation (see Report F 8444, "The Use of Electric Trucks and In-the-Wall Haulage in Open-Pit Mining," in Swedish with an English Summary). Although in-the-wall haulage is of interest in some open pits, the local terrain is important for profitability. The authors do not think in-the-wall haulage is of any major interest for the mining industry. The following is a summary of some of the author's findings. To place the haul road in the wall is not of interest in open pits with declining metal contents in the bedrock. In such mines, the volumes from the haul roads will not yield revenues. The cost per m3 for an underground haul road is much higher than for a haul road in the pit. This means that it is only of interest to place the haul road in the wall when mining gets deeper than about 50 m (165 ft). In-the-wall haul roads will reduce flexibility in the pit, and it will make necessary the use of smaller equipment with lower productivity and higher costs per ton. As Hustrulid, et al. show (Table 4), excavation savings of material hauled are very low, $0.10 to $0.25/t (0.09 to 0.27 per st). The extra operating cost, due to lower productive equipment, will normally be much higher and thus destroy the whole idea with in-the-wall haulage. If electric trucks are profitable for in-the-wall haulage, it is normally also profitable to use electric trolley assist for trucks on the haul road in the pit. But the profitability of using electric power is different in different countries, and depends on the relation between the cost for electric power and diesel fuel. In the US, diesel fuel is inexpensive compared with electricity, but in a country like Sweden, diesel fuel is much more expensive. In most open pits, the trucks have to move a considerable distange from the loader until they reach a final haul road in the wall. A trolley line along such temporary haul roads will be exposed to flyrock. It is normally less expensive to perform rock support from the open pit than from underground ramps in the wall. In Fig. 8, Hustrulid et al. gives the impression that the underground haul roads will be very close to each other. This is seldom the case. Figure 1 shows a haul road in the wall. The haul road passes through each cross-section only once. Arranging a reliable dewatering and rock support system from only one underground ramp is probably impossible. Many more drifts are therefore necessary. In our report, we also studied what would happen if the final pit slope is increased by 5°, using underground drifts in the wall. Our conclusion was that, the extra cost for drifts, rock support, pumps, etc. destroys the whole idea, and that it was better to accept a higher stripping ratio with the haul road in the pit and to use conventional low-cost open pit equipment. Finally, we think it is would be difficult to try to reduce the safety factor when determining the slope angle by moving the haul road in the wall. Minimizing the risk for pit wall collapse is also important with in-the-wall haulage, primarily because men and equipment will be working in the pit, but also to guarantee many accesses between the pit and the haul road in the wall.

Jan 1, 1990

By M. Hermanus, N. Coulson, N. Pillay

"SynopsisThis paper assesses the effectiveness of the Mine Occupational Health and Safety Leading Practice Adoption System (MOSH) and its potential to improve mine health and safety in South African mines. Developed by the Chamber of Mines, which represents the majority of the country's large scale mining employers, MOSH was devised to accelerate progress towards achieving health and safety milestones, which were set by tripartite agreement in 2003. The paper documents and builds on the findings of a study conducted by the Centre for Sustainability in Mining and Industry (CSMI) in 2011 that evaluated MOSH strategy, structures, and process of implementation.The study found that MOSH operated across the mining sector, was directed and dominated by experts and, despite best efforts to include other stakeholders, was led by employers. Statutory worker health and safety representatives and structures were not integrated into the complex change process developed by MOSH. The Mine Health and Safety Inspectorate (MHSI) and organized labour were ambivalent about direct involvement in MOSH and preferred regulatory measures to enforce the participation of mines. MOSH interventions were not targeted at mines with a poor health and safety record, and MOSH lacked a baseline from which to track impacts on sector-wide health and safety performance. The leading practices most widely adopted by mines were designed to improve, rather than fundamentally alter, existing practice. Although the depth of engagement with MOSH among stakeholders and on mine sites varied, mining companies, labour representatives, and the Mine Health and Safety Inspectorate (MHSI) saw the programme as significant.IntroductionIn 2011, the Chamber of Mines commissioned the Centre for Sustainability in Mining and Industry (CSMI) to evaluate the progress and effectiveness of the rollout of the Mine Occupational Health and Safety Leading Practice Adoption System (MOSH). CSMI, based in the School of Mining Engineering at the University of the Witwatersrand, is an independent research centre conducting health and safety research relevant to all major stakeholders in the mining sector. The purpose of the evaluation was to identify the factors that enabled or hampered the adoption of new technologies or practices. The timing of the evaluation, two years before the deadline for the 2013 safety and health milestones, allowed time for adjustments to be made. This paper summarizes the results of the evaluation and includes further analysis and discussion of the nature of the programme and the conditions that act to advance it or it hold back."

Jan 1, 2015

By G. Sachs, Jr. Brown W. F., Ebert L. J.

It is a generally recognized fact that a steel quenched from the austenitic range and tempered to yield a structure composed of tempered martensite becomes increasingly more ductile with increasing tempering temperature. However, the conventional tensile test fails to yield any value that is sensitive to differences in behavior on tempering among the various low-alloy steels or among different heats of a particular steel. The notched impact bending test, on the other hand, has been recognized for a considerable length of time as capable of revealing differences in impact properties among various steels tempered to 'the same strength level (i.e., the same tensile strength in a conventional tensile test). However, this test is very complex and its significance has not been clarified up to the present time. Of its three characteristic features—the high speed, the notch, and the bending—the notch has been found by Davidenkov2 to be the most potent embrittling factor. Recent investigations on low-alloy steels have shown static notch-bar tensile tests to be more lucid and to evaluate the metal properties in much the same manner as impact tests. In an attempt to analyze the fundamental nature of this test, it was found possible to separate the effects of the two basic characteristics of the stress pattern introduced by the presence of a circumferential notch, the stress state and the high initial stress concentration at the notch bottom. The major conclusions drawn from these previous investigations were as follows: I. The magnitude of the triaxiality at fracture depends primarily on the initial notch depth. Its effect is a great reduction in the average ductility (contraction in area at fracture in the notched section) as compared with that observed in uniaxial tension. This triaxial stress state occurs at a point in the interior of the notch specimen. It determines the notch ductility if the notch is mild and/or the metal possesses a considerable ductility. On the contrary, brittle specimens probably fracture on the metal surface (at the notch bottom) at which point biaxial stress state is present. This biaxiality also considerably reduces the ductility of the metal. For small notch ductilities the notch strength and the notch ductility become very sensitive to embrittling factors. This is contrary to the behavior. of the ductility values derived from a conventional tensile test, which are comparatively insensitive.

Jan 1, 1947

The relationships between Aboriginal Traditional Owners and mining companies have often been mired in dispute, particularly since the High Court Mabo decision and the Native Title Act of 1993. In the Pilbara region, Rio Tinto Iron Ore (RTIO) impacts the traditional country of 12 Aboriginal Traditional Owner/language groups. Over the past 12 years the company and the Traditional Owners have negotiated and settled several significant land use agreements including the Yandi Land Use Agreement over the Yandicoogina mine in 1997, between Hamersley Iron (now Rio Tinto Iron Ore) and three language groups û the Bunjima, Nyapaili and Innawonga people and the Eastern Guruma Indigenous Land Use Agreement (ILUA) of 2006. These substantive agreements are now being followed by a new generation of region wide ILUAs which are close to conclusion. Arising from these regional agreements is a new framework for engagement which has been informed by the implementation of the previous agreements in the Pilbara and across the Rio Tinto group of companies. They are driven by both a company desire to have a consistent approach to engaging with Traditional Owners and the Traditional Owners wanting to have a similarly transparent and consistent standard of performance by the company in implementing the agreements which can deliver improved employment, cultural heritage engagement and direct and indirect benefits for the life of mining by the company. During the past three years the growth in the Pilbara, Western Australia brought new opportunities with people flocking to the region for employment flowing from the mining boom. Housing was at a premium and FIFO (fly-in, fly-out employment arrangements) was developed as a method to manage the general growth and demand for skills. Despite the recent down turn in the world economy since September 2008, which has reduced demand for minerals, the Pilbara continues to enjoy high levels of employment and as a result there is still significant pressure on housing and services provision. Although there has been vast economic growth in the Pilbara not only by mining companies but also oil and gas companies, the 2006 census data indicated that the direct benefits of improved employment, business development and engagement in the mainstream economy continued to exclude Aboriginal people. Indigenous people constitute a significant proportion of the population in the Pilbara region, some 6000 people û currently 14 per cent of the population rising to 18 per cent by 2015 û yet they remain at the bottom of the socioeconomic scale. More than 40 years of mining has returned relatively little to this group of people who are the permanent residents of the region and are connected through their traditional culture to this country (Taylor and Scambary, 2005). Ensuring local Aboriginal participation in the mainstream economy has been a challenge for resource companies, Traditional Owners and local residents in the Pilbara. The case study explores the relationship between a mining company û Rio Tinto Iron Ore and the Traditional Owners of the Pilbara. It examines two projects that have been developed to engage Indigenous people in the development of their region and through regional representation and business development to build a sustainable future for themselves and their future generations and at the same time ensure social cohesion to support social licence to operate. The case study also examines the process that was followed by Rio Tinto Iron Ore and its operations in undertaking a regional agreement for land access, participation and the integration of local Indigenous people into their business and the local economy.

Jan 1, 2009

By Ebert L. J., G. Sachs, Jr. Brown W. F.

It is a generally recognized fact that a steel quenched from the austenitic range and tempered to yield a structure composed of tempered martensite becomes increasingly more ductile with increasing tempering temperature. However, the conventional tensile test fails to yield any value that is sensitive to differences in behavior on tempering among the various low-alloy steels or among different heats of a particular steel. The notched impact bending test, on the other hand, has been recognized for a considerable length of time as capable of revealing differences in impact properties among various steels tempered to 'the same strength level (i.e., the same tensile strength in a conventional tensile test). However, this test is very complex and its significance has not been clarified up to the present time. Of its three characteristic features—the high speed, the notch, and the bending—the notch has been found by Davidenkov2 to be the most potent embrittling factor. Recent investigations on low-alloy steels have shown static notch-bar tensile tests to be more lucid and to evaluate the metal properties in much the same manner as impact tests. In an attempt to analyze the fundamental nature of this test, it was found possible to separate the effects of the two basic characteristics of the stress pattern introduced by the presence of a circumferential notch, the stress state and the high initial stress concentration at the notch bottom. The major conclusions drawn from these previous investigations were as follows: I. The magnitude of the triaxiality at fracture depends primarily on the initial notch depth. Its effect is a great reduction in the average ductility (contraction in area at fracture in the notched section) as compared with that observed in uniaxial tension. This triaxial stress state occurs at a point in the interior of the notch specimen. It determines the notch ductility if the notch is mild and/or the metal possesses a considerable ductility. On the contrary, brittle specimens probably fracture on the metal surface (at the notch bottom) at which point biaxial stress state is present. This biaxiality also considerably reduces the ductility of the metal. For small notch ductilities the notch strength and the notch ductility become very sensitive to embrittling factors. This is contrary to the behavior. of the ductility values derived from a conventional tensile test, which are comparatively insensitive.

Jan 1, 1947

By Jan C. Still

The Glacier Bay subarea is that part of the Juneau Mining District (JMD) within Glacier Bay National Park and Preserve (GBNP). It is the western portion of the JMD and consists of the land between the Fairweather Range crest and the Takhinsha and Chilkat Ranges crests (excluding that land east of Excursion Inlet.) Figure B-1 shows the area. It is bounded on the north by the Alaska-British Columbia International boundary, to the south by Cross Sound and Icy Strait, and contains Glacier Bay. During 1975 -1977 the Bureau conducted a detailed study of the occurrences, prospects, and deposits in GBNP (B_4).2 This work was in cooperation with the U.S. Geological Survey (USGS) under the provisions of the Wilderness Act (Public Law 88-577). As a follow up to this earlier Bureau work, limited studies were conducted, during 1985 and 1986, as part of the JMD study. This report covers the most important aspects of the earlier studies and the 1985 and 1986 studies. ACCESS The Glacier Bay area may be reached through Gustavus airport, a paved strip which is serviced by scheduled jet airliners in the summer and air taxi services year-round. The airport is connected to Bartlett Cove (GBNP headquarters) by an all-weather road. No other roads exist in the area except a skid road to the LeRoy Mine and mill in the Reid Inlet area. Much of the Glacier Bay area is accessible through the salt water inlets and fjords which are part of Glacier Bay. Cruise ships, charter boats, and fishing boats visit these inlets and fjords during the summer months. Elevations within the area range from sea level to over 15,000 feet in the rugged Fairweather Mountain Range. More remote areas are accessible by small boat, foot, or helicopter. Glacier ice is a year-round hazard to navigation in Glacier Bay, and during the winter months some of the more northerly inlets are frozen.

Jan 1, 2012

By E Amini, N Beaton

Application of ore pre-treatment technologies such as Grade Engineering and in-pit ore treatment require comprehensive analysis to show proven value is being generated by switching away from the current practices. Considering most mining operations last several decades; the pre-treatment technologies must be investigated over life of mine (LOM) to truly show value was generated or lost. Therefore, a value chain flowsheet model has been usually developed to run those analysis/simulations over LOM using a set of defined ore characteristics and the operating equipment capacities. The Integrated Extraction Simulator (IES) has been used for this type of LOM simulations as it has the necessary flexibility to accommodate the whole mining value chain flowsheet in one platform and provide essential computation power to run mass simulations for the given scenarios using the resource block model. This allows to forecast the value chain response over the LOM for all given scenarios in a feasible amount of time. The traditionally developed flowsheets for the analysis/simulations are usually limited to the mineral processing plant and were developed based on a single audit conducted on the plant in a very short time, e.g. several hours. The process plant is being operated with a certain ore at a certain condition during the short audits which would not reveal the true capacity of the operation equipment when ore characteristics changes. Therefore, it would be impossible to define the true capacity of the equipment and identify the constraints along the value chain. In this study, a methodology was developed to identify the constraints along the value chain and implement those in the IES flowsheet using process information. Also, the equipment models were fine-tuned to represent the true capacity of the equipment in the flowsheet using the same data set. The methodology was tested over five process plant in two operations and showed the performance forecast was significantly improved over the traditional methods being used so far. Then those high-fidelity flowsheets were used to evaluate the ore pre-treatment technologies proposed by the relevant resource company experts.

Nov 10, 2020

By R. D. Hagni

Introduction The ore deposits of the Viburnum Trend in the Southeast Missouri Lead District constitute the world's largest lead producing district. Six or seven mines currently are in production in the Viburnum Trend. The principal minerals are few in number, simple in composition, and form simple superposition and replacement textures. The economically valuable minerals in the typical lead-zinc ores of the Viburnum Trend are the sulfides, galena, sphalerite, chalcopyrite, and siegenite. The most abundant gangue minerals are dolomite, marcasite, pyrite, calcite, quartz, and dickite. Four concentrates, lead, zinc, copper, and cobalt are produced from the ores, and they contain an interesting suite of trace elements, of which silver and cadmium are recovered at the smelter and refinery.1 Emphasis is given here to the variations in directions of ore fluid flow based upon interpretations of the positions of orebodies due to the ore fluids being impeded by Precambrian basement relief and upon patterns of ore mineral/metal zoning. The Precambrian basement relief and Lamotte Formation pinchout has exerted an important control upon the specific locations of orebodies in the Southeast Missouri Lead District. The locations of some orebodies along one side of Precambrian knobs or ridges strongly suggests that they are positioned on the side from which the ore fluids flowed. Other orebodies are located in positions that suggest that they formed where the ore fluids passed around or between Precambrian knobs. Patterns of mineral and/or metal zoning have been used in many mineral districts to interpret the directions of ore fluid flow. The mineral/metal zoning pattern correlates closely with that of the mineral paragenesis. The earliest deposited minerals tend to be deposited nearest to the source, the later deposited minerals next closest to the source, and the latest deposited minerals furthest from the source. In the Southeast Missouri Lead District, it is interpreted here that the early-deposited copper-rich ores are closest to the source of the ore fluids, the subsequently-deposited zinc-rich ores are next closest to the source, and iron-rich ores are furthest from the source.

Jan 1, 2006

By A. V. Reddy, C. Beckermann

"The DC continuous casting of a grain refined AI-4.5%Cu round ingot with realistic boundary conditions is numerically simulated. Of particular interest are the effects of the transport of free solid in the form of equiaxed grains on the macro segregation patterns and the final grain size distribution. The study uses a previously developed two-phase model that incorporates descriptions of heat transfer, solute redistribution, melt convection, and solid transport on the system scale with microscopic models of solutal undercooling, microsegregation and grain impingement. The numerical results indicate that grain transport inhibits the formation of inverse segregation in the ingot shell and promotes negative centerline segregation. This is explained by the reduction in the solid coherency next to the mold and the advection of solute-lean grains towards the centerline. Comparisons with the experiments of Finn et al. [5] illustrate the need for considerable improvements in the model. IntroductionMacrosegregation continues to be one of the concerns in DC continuous casting of aluminum ingots. The variation in composition over the ingot cross section results in the need to adjust alloy composition in order to meet mechanical property specifications in finished products. The extent of macrosegregation is greatly affected by the casting conditions, such as casting speed, superheat, cooling rate, mold size, alloy composition, and grain refining practices. Thus, it is of great interest to develop a comprehensive mathematical model to simulate the casting process, enabling the selection of conditions that minimize macrosegregation while maintaining a high productivity."

Jan 1, 1994

By S. Banisi, A. Ghasemi, A. Fazeli, M. Rezaei, M. Maleki-Moghaddam

The efficiency of high capacity SAG mills which consume large amounts of energy depends on the grinding and slurry discharge systems. A 9.7 m × 4.8 m SAG mill with radial pulp lifters is used at the Sarcheshmeh copper complex grinding circuit. The process audit indicated that the pulp discharge system was not efficient, judging by the amount of slurry flow back and carry over. These events limited the discharge rate and increased the wear rate of the lifters and grates. With the objective of finding a new design for pulp lifters, a 3D DEM software (KMPCDEM ©), along with the plant operational data and photographic evidence was used. The KMPCDEM © software was used to compare the efficiency of the radial and curved pulp lifter design based on the material flow. The slurry flow was analysed in terms of ease of discharge and lifter wear rate. It was found that changing the radial pulp lifters to curved pulp lifters could improve the discharge efficiency,and decrease wear rate of the pulp lifter in particular at the base of the pulp lifters. The simulation results showed that if curved lifters with a radius of 0.3×mill radius are used, the amount of slurry flow back and carry over will be negligible. It is predicted that by installing the curved pulp lifters, the capacity of the mill increases by 10% and the pulp lifter service life improves by at least 20%. Keywords: SAG mill, radial pulp lifter, DEM,curved pulp lifter, Sarcheshmeh

Jan 1, 2020

By L. F. Mondolfo

R301 is a clad aluminum alloy, composed of a core of a duralumin-type alloy clad with a magnesium silicide alloy. It differs from other well-known clad alloys in that the cladding and the core respond to the same heat-treatments. and therefore the cladding contributes appreciably to the mechanical properties of the material. In Table I are reported the compositions of the core and cladding alloy for the material used for the experiments. Table i.—Chemical composition of R~OI Used for the Experiments Element Core Cladding Cu........................ . 0.01 Fe......................... . 0.42 Si......................... 0. 0.71 Mg........................ . 0.99 Mn ........................ . 0.54 In clad aluminum alloys where the core contains copper and the cladding does not, at elevated temperatures the copper tends to diffuse from the core through the cladding. This fact is well known and many investigations have been conducted to determine the features of this diffusion. Most of those investigations cover materials clad with pure aluminum; the present investigation covers a material clad with an alloy. Since the cladding thickness on R~OI varies with the gauge, as shown in Table 2, several gauges were investigated. Material and Procedure he material used was sheet of the following gauges: o.o2~, 0.040, 0.064, 0.102, 0.170 Cut in pieces approximately 3 by 5 in., it was in the cold-rolled state and was soaked for increasing times at the following temperatures: glo°F., 940°F., 970°F. A Lindberg cyclone furnace was used for the experiments. The procedure was as follows: The furnace was started empty with controlling pyrometer set at the determined temperature. After the furnace had reached that temperature, it was run for some 20 to 30 min. empty, in order to allow the temperature inside to become stabilized. Then the first samples were introduced. A load thermocouple, with recording pyrometer, was connected with the sample of the heaviest gauge. About 10 min. was required for the material to reach temperature and the soaking time was counted from that point. After zoo min. soaking, the second group of samples was introduced without stopping the furnace. After 200 min. more, the third group of samples was introduced, and so on, until the last sample had been introduced. Every time the furnace was opened and new samples were added, a drop of temperature was registered by the thermocouple connected with the material

Jan 1, 1946

By L. F. Mondolfo

R301 is a clad aluminum alloy, composed of a core of a duralumin-type alloy clad with a magnesium silicide alloy. It differs from other well-known clad alloys in that the cladding and the core respond to the same heat-treatments. and therefore the cladding contributes appreciably to the mechanical properties of the material. In Table I are reported the compositions of the core and cladding alloy for the material used for the experiments. Table i.—Chemical composition of R~OI Used for the Experiments Element Core Cladding Cu........................ . 0.01 Fe......................... . 0.42 Si......................... 0. 0.71 Mg........................ . 0.99 Mn ........................ . 0.54 In clad aluminum alloys where the core contains copper and the cladding does not, at elevated temperatures the copper tends to diffuse from the core through the cladding. This fact is well known and many investigations have been conducted to determine the features of this diffusion. Most of those investigations cover materials clad with pure aluminum; the present investigation covers a material clad with an alloy. Since the cladding thickness on R~OI varies with the gauge, as shown in Table 2, several gauges were investigated. Material and Procedure he material used was sheet of the following gauges: o.o2~, 0.040, 0.064, 0.102, 0.170 Cut in pieces approximately 3 by 5 in., it was in the cold-rolled state and was soaked for increasing times at the following temperatures: glo°F., 940°F., 970°F. A Lindberg cyclone furnace was used for the experiments. The procedure was as follows: The furnace was started empty with controlling pyrometer set at the determined temperature. After the furnace had reached that temperature, it was run for some 20 to 30 min. empty, in order to allow the temperature inside to become stabilized. Then the first samples were introduced. A load thermocouple, with recording pyrometer, was connected with the sample of the heaviest gauge. About 10 min. was required for the material to reach temperature and the soaking time was counted from that point. After zoo min. soaking, the second group of samples was introduced without stopping the furnace. After 200 min. more, the third group of samples was introduced, and so on, until the last sample had been introduced. Every time the furnace was opened and new samples were added, a drop of temperature was registered by the thermocouple connected with the material

Jan 1, 1946

By Ahmad Ghahreman, Denver Cowan, Yeonuk Choi

"Previous testwork by the authors has indicated that both granular coconut shell based activated carbon and a new microporous carbonaceous bead, Lewatit® AF 5, are highly effective in improving the oxidation kinetics and decreasing the elemental sulphur yield when assisting the atmospheric oxidation of pyrite. While the two carbon catalysts show similar results and improvements to the pyrite oxidation, there are differences between them. For example, the activated carbon assisted oxidation showed faster kinetics than the AF 5 assisted oxidation, while elemental sulphur seemed to remain adhered to the surface of the AF 5 after oxidation when much of the elemental sulphur formed during the activated carbon assisted tests was found in the oxidation residue. The objective of this paper is to investigate the differences between the two catalysts through the use of scanning electron microscopy, photography, Raman spectroscopy, surface chemistry analysis, and Brunauer-Emmett-Teller (BET) surface area analysis. With this investigation the authors should be able to achieve a better understanding of the surface chemistry of the two carbon catalysts (i.e. functional groups, charges, acidity etc.) and help explain the strengths and/or the weaknesses of the catalysts when applied to the atmospheric oxidation of pyrite.INTRODUCTION Previous research conducted by the authors, indicated that both a granular coconut shell based activated carbon (activated carbon) and a microporous carbonaceous bead, Lewatit® AF 5 (AF 5), acting as a catalyst enhanced atmospheric pyrite oxidation in ferric sulphate media by approximately 40% when compared to unassisted oxidation and that elemental sulphur yields were increased from 23% to approximately 69 and 63% when assisted by AF 5 and activated carbon respectively (Cowan, Jahromi, & Ghahreman, 2017). The study indicated that although total pyrite oxidation assisted by the two catalysts was similar, differences were noticed in the impact that the two catalysts had on the outcome of the oxidation. Kinetic curves in the Cowan et al. study indicated that atmospheric oxidation of pyrite seemed to reach completion after 48 hours of oxidation when assisted by activated carbon. A similar trend was noticed in the kinetic curves for oxidation when assisted by AF 5, however the oxidation curve did not reach completion until approximately 96 hours (Cowan et al., 2017). This observation indicated that the kinetics of pyrite oxidation are enhanced by both catalysts, oxidation kinetics are much faster when assisted by activated carbon rather than AF 5."

Jan 1, 2017

By R. Kuyvenhoven, A. Parbhakar-Fox, X. Prieto, J. Quevedo, L. V. Danyushevsky

This paper presents the results of a study focussed on determining the distribution of cobalt in different types of pyrite of the La Estrella IOCG deposit in Chile. In the IOCG mineral deposits, where copper is the main product, the quantification and distribution of cobalt is required to correctly define the technical and economic viability of recovering cobalt as a by-product. In the La Estrella mineral deposit, seven alteration and mineralisation events were recognised through both optical and automated electronic mineralogical analysis. Of these seven events, five different types of pyrites were identified, each with unique characteristics of the pyrite crystals. Through advanced elemental analysis of these five types of pyrites, it was determined that the concentration and distribution of cobalt within the pyrite crystal is dependent on the mineralisation event. The results indicate high cobalt concentrations in the early pyrite mineralisation events, associated with relatively fine pyrite grains, and cobalt generally occurring in a solid form in the pyrite core. The late pyrite mineralisation event exhibits the lowest concentrations of cobalt, with cobalt predominantly present along the outer periphery of the pyrite. These results are relevant from a metallurgical point of view because the preferential location of cobalt within the pyrite crystal will influence the maximum metallurgical recovery of cobalt. For this, a cobaltiferous pyrite concentrate must be formed which, according to the characteristics of pyrite, must be processed with the optimal solubilisation method. The overall objective is to ensure both good cobalt recovery and a high concentration of cobalt in the leach solution resulting from pyrite leaching. Keywords: Cobalt, pyrite, IOCG, copper, metallurgical recovery, solubilisation, Chile

Jan 1, 2020