Search Documents

By S. H. Lorain

This paper, which describes lode gold mining in Idaho county, Idaho, is one of a series on western mining districts being published by the Bureau of Mines. It describes mining and milling practices, gives costs, and discusses transportation, climate, power, labor, and other conditions that affect them. The materiel upon which the paper was based was gathered during a field survey in the summer of 1937.

Jan 1, 1938

By Y Li, Z Fu, D J. Williams, E Topal

Mine waste rock is unavoidable in the recovery of ore by open pit mining methods. Effective management of waste rock movement covers the removal and placement of the waste rock, generally in a surface dump in the most cost-effective way while satisfying the long-term environmental expectations. Since the 1960s, there has been significant research carried out into planning of ore movement from open pit mines such that the overall net present value (NPV) of the project can be maximised. Currently, several software packages are available to achieve this task as a common practice within the mining industry. However, planning of mine waste rock movement is still achieved manually, as the available software does not consider optimal waste rock movement and dump design. This paper describes a tool to optimise waste rock placement, which minimises the haulage cost while creating environmentally friendly waste rock dumps through selective placement and encapsulation of the reactive waste rock in the dump. The embedded methodology in this tool has been developed based on the mixed integer programming (MIP) modelling technique to optimise the waste rock movement. The methodology provides an easily-manipulated and fully-automatic tool for professionals to complete waste rock dump scheduling in four main steps: data input, model construction, model computation and solution retrieval. The application of the new tool into a real case project in this paper shows that it can determine economically and environmentally optimised waste rock dump schedules in an acceptable time frame.CITATION:Fu, Z, Li, Y, Topal, E and Williams, D J, 2015. A new tool for optimisation of mine waste management in potential acid forming conditions, in Proceedings Tailings and Mine Waste Management for the 21st Century, pp 333–342 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 27, 2015

By I. Y. Motovilov, S. A. Telkov, M. B. Barmenshinova, Z. S. Abisheva

As a rule, removing waste rock at the beginning of the process improves the conditions for the flotation of minerals, and also reduces the operating costs of crushing, grinding and flotation. Lead-zinc ores in the Shalkiya deposit of the Republic of Kazakhstan are classified as difficult for enrichment, but taking into account their significant reserves, they are one of the main potential ores for the production of lead and zinc. Under industrial conditions, for processing ores of the Shalkiya deposit, a selective flotation concentration scheme with the production of lead and zinc concentrates is recommended. However, according to the recommended scheme, it is not possible to obtain high-quality concentrates of lead and zinc, with high metal recovery. The paper presents the results of studies of preliminary gravitational enrichment of lead-zinc ore in the Shalkiya deposit. Based on the fractional analysis for the particle size range of -40+8 mm, Henri- Reingard enrichment curves were constructed, separation density (2730 kg /m3) was established to isolate the light fraction with the lowest possible lead and zinc content, and gravity enrichment indices (0.7-0.71) were determined. It was found out that when using gravitational enrichment of the original large-sized ore in the Shalkiya deposit, tails are released at the beginning of the technological process with a yield of up to 20% with the extraction of silica up to 30%, calcite up to 11% and carbonaceous matter up to 12%. As a result of the studies, it was found that the beneficiation of crushed lump ore could be conducted in heavy suspensions. One of the positive aspects of the use of gravitational enrichment is the reduction of energy costs for grinding the heavy fraction, which was experimentally proved in this article. Keywords: Lead-zinc ores, Shalkiya deposit, fractional analysis, gravitational enrichment, floating fraction, sunken fraction, separation density

Jan 1, 2020

By Joseph Kuti

Longwall mining is the term used for the underground extraction of a wide and deep panel of coal. Such a panel may be blocked out by one or more entries on either side ("head" and "tail") of a longwall face. If all the entries are pre-driven, the panel extraction is termed longwall retreat mining. With one exception, all the longwall faces in the United States are of this type. (1) prior to the general decline in the rates of production with Continuous Miners, the development of the multiple entries generally kept ahead of the longwall extraction rate. This is no longer the case: longwall extraction rates far exceed the rate of development. In addition to this problem, in newly established mines, the development of a haulage and bleeder system based on multiple entries results in a lead-time of three years or more before longwall mining can be applied. This is one reason why earlier projections and much of the optimism attached to longwall mining as a principal "savior" in the energy crisis should be viewed with some caution.

Jan 1, 1978

By M. X. Jia

This paper discusses two mineralogy characteristics that may affect the flotation of sulfide ore. The first one, process mineralogy study on a Cu-Ni sulfide ore shows that valleriite and tochilinite are incline to form on chalcopyrite, pyrrhotite and serpentine surface, and the further guess is that the formation of valleriite and tochilinite may also causes the serpentine transform to talc, so the serpentine surface is a phase of talc or a transition phase from serpentine to talc and this enhances the flotability of the gangue minerals. The second discussion on the flotation behavior of the sulfide minerals from the point of view of the mineral semiconductor character.

Jan 1, 2014

By Dezhou Wei, Cong Han, Shuling Gao, Wengang Liu, Yanbai Shen

The recycling schemes of copper mineral to a copper raw ore produced in China is investigated in this paper, of which is aimed at copper concentrate‘s grade is over 26.00% and recovery is over 90.00%. The primary copper mineral the raw ore is chalcopyrite and the main gangue minerals are quartz, orthoclase, albite, clinochlore and phlogopite. There are four useful metal elements contained in the copper raw ore, which are Cu, Pb, Zn, Au and Ag, and the grade of each is 3.75%, 0.01%, 0.11%, 0.31 and 483.5 g/t. Two floating recycling schemes and its corresponding flotation reagent system were determined. The results show that calcium oxide, sodium carbonate and sodium sulphide were used as regulators, sodium n-butylxanthate was used as collector and terpenic oil was used as frother in two floating recycling schemes are effective. The open-circuit test’s results show that the copper mineral contained in raw ore can be effective collected, and both concentrates gotten through two schemes are up to grade.

Jan 1, 2014

By D. Krebs, Tero Korhonen, X. Yang, A. Kirpala, N. Heino

At present, over 85% of the world’s rare earth oxides are produced from the bastnäsite ores, found mainly in China. Meanwhile, the ion adsorption ores are the most important sources for heavy rare earth production. As the demand for rare earth materials in the world increases, other forms of rare earth deposits will become new potential resources of rare earth production in the future. In this paper, flotation performance of a steenstrupine ore from the Kvanefjeld deposit for rare earth recovery was studied through laboratory bench testwork. The effects of various factors including grinding size and collector dosage on recovery and selectivity of the selected rare earth elements, La, Ce and Y, were investigated. A rare earth concentrate with the grades of La 2.8 wt%, Ce 5.1 wt% and Y 1.0 wt% at recoveries of 80.5% La, 79.8% Ce and 55.3% Y was obtained in optimum conditions. Further potential to improve the efficiency of REE flotation was discussed. Keywords: Rare earth elements, flotation, steenstrupine, Kvanefjeld deposit

Jan 1, 2020

By Jonathan Ludlow

This document is the final report of a study aimed at defining the relative contributions of various sources of respirable airborne dust on longwall faces. In order to define the contribution of various primary sources, continuous dust level data were recorded on five longwalls for an average of five shifts at each. By analysis of these data with records of face activity, it was possible to determine the overall contribution of each of the major elements of the longwall system to overall respirable dust exposure at various face locations. Average levels were determined for roof support generated dust, conveyor and intake dust and for dust levels during various shearer activities. When average levels are compared between mines, a significant degree of consistency is noted. The next phase of the study was aimed at differentiating the various dust sources associated with the shearer. To this end 'fly cuts' varying from 6 in. to 30 in. were closely monitored as were dust levels at operators' positions over the various segments of longwall cycle. These data indicated that the clearance ring of the shearer drum as a disproportionately heavy contributor to shearer dust make. A statistical analysis of the effect of shearer haulage speed on dust make revealed that high haulage speeds are associated with low specific dust make.

Jan 1, 1983

By H. A. Baxter

THE Eustis concentrator is located about nine miles southwest of Sherbrooke, on the main line of the Quebec Central railway. The concentrator is served by a side track parallel to the main line, and all supplies for both the mine and mill are delivered at the mill yards. The concentrator treats 250 tons of copper-iron ore daily by preferential flotation, whereby, first, the copper sulphides are floated and the iron pyrite is depressed, and, second, the iron pyrite is reactivated and floated, resulting in an iron pyrite concentrate and tailing, the latter going to waste. During 1934, 71,709 tons of ore were treated, averaging 3.712 per cent copper and 32.1 percent iron, with a small amount of silver and a trace of gold. The recovery was 93.77 percent of the copper, and 70.43 percent of the iron as iron concentrate. The copper concentrate produced averaged 25.623 percent copper and the iron-pyrite concentrate averaged 42.13 percent iron and 49.173 percent sulphur. HISTORY The first mill at the Eustis mine was built in 1904 (1). For a few years previous to this rime, some jigs had been operated to concentrate the fines from the mine, the coarse ore being hand-sorted. Both the fines and sorted lump ore were shipped to acid plants. The building of the mill was based on test work carried out by Professor Robert H. Richards at the Massachusetts Institute of Technology. The method of concentration consisted in careful classification in a Richards classifier, which made six different sized products. These products went to Wilfley tables, where a concentrate, middling, and tailing were made. The middling was returned to the rolls. Crushing was done by a jaw crusher and two sets of rolls, resulting in a screen size of about five-mesh.

Jan 1, 1936

By David E. Nicholson

In a Bureau of Mines investigation of the dynamic compaction of mine hydraulic backfills, two basic forms of mechanical vibration inducers were used and compared for effectiveness of compaction. Internal compaction, using concrete probe vibrators of various sizes, and external compaction, using plate vibrators} were studied and evaluated for use in underground stopes. An aluminum plate, which was light and displaced sufficient volume to remain afloat in fluid fill, and a steel plate were both used in the external compaction tests. Efficiency of compaction was better in studies conducted with internal probe vibrators. Tests were conducted in an underground backfilled stope with three different-sized probe vibrators (2-, 4-, and 6-inch-diameter) to determine an approximate area of influence. Results showed the areas of influence to be approximately 3, 4, and 5 feet in diameter, respectively, for the three different probes. Physical properties of hydraulic slurries were determined; preferred grain-size distributions, water exudation properties, and water contents of hydraulic slurries are reported. The best compaction results were obtained with moderately well-graded, fully saturated backfills, with slow rates of water exudation, which were capable of remaining in a semifluid state during the compaction process.

Jan 1, 1967

By L. E. Djingheuzian

Introduction For many years the immense deposits of bituminous sands of northern Alberta have been attracting the attention of our governments, engineers and .scientists ( 6). Undoubtedly these bituminous sands represent one of the greatest potential resources of liquid fuels in the world. However, their commercial development has been retarded by the complexities of the technological problem and by the economic factors involved in the processing of the sands. The technological problem has been attacked from several angles ( 2, 3, 1, 11), and cold water separation of bitumen is one of them. To the ore dressing engineer, the cold water process is of utmost importance since, by the use of standard ore dressing machinery, it accomplishes most efficiently and economically the separation of bitumen from bituminous sands. The significance of using ore dressing equipment was especially enhanced by the fact that the scientific principles which underlie the design of standard machines were demonstrated in the cold water separation process with an extraordinary clarity. This greatly facilitated the analysing of limitations and advantages of different machines during-the pilot plant investigations and helped to complete the investigations practically in the allotted time.

Jan 1, 1952

By George H. Holmes

This paper is one of a series being prepared by the Minerals Division of the Bureau of Mines on mining and milling methods and practices at various mines in the United States. It describes such methods at the Caselton mine of the Combined Metals Reduction Co. The Caselton mine and mill are in secs. 28 and 29, T. 1 N., R. 67 E., M. D. B. & M., Lincoln County, Nev., about 10 miles by paved highway south- west of Pioche, Nev. (fig. 1). the history of the district and company This circular summarizos operations, describes physical features, labor and living conditions, the geology, ore deposits, and the ore. Prospecting and exploration methods are outlined, and shaft sinking, drifting and crosscutting, and raising are covered under development methods. In the soction on mining mothods, stoping and slushing are discussed, following which are descriptions of underground transportation, hoisting, ventilation, drainage and sampling, and surveying. The surface and crushing plants are described, and milling practices are presented in dotail with explanation of the filter plant and assay office. Safoty practices and the company organization are described in the concluding sections of the circular.

Nov 1, 1950

By G. J. C. Young

In 1986 the city of Johannesburg celebrates its first century of existence. The fact that the city was founded to exploit the gold deposits which underlaid it, prompts this review of the gold process development in the mining industry of South Africa. Clearly a complete history of, the gold recovery process to date is not called for and it is proposed therefore to outline the most significant developments which have taken place since the middle 1930's, that is to say the last 50 years or so. Prior to this period the industry had seen the development of comminution processes employing cone crushing in closed circuit with vibrating screens followed by grinding in rotating mills and the gradual disappearance of stamp mills. The all sliming process was well established as a viable route and continuous dewatering in thickeners, continuous leaching and filtration were serious rivals for the earlier batch processes. Continuous processes for gravity concentration had also been developed and the recovery of gold from pregnant solutions employed cementation on zinc dust as opposed to shavings. Thus at the beginning of the half century under review the industry had adopted, or started to adopt, many of the processes still accepted as modern practice. However in some areas little change was evident this was especially so in the smelthouse where some thirty more years were to elapse before the hand stoked coal fired calcine and reverbatory furnaces were replaced by more environmentally acceptable electric furnaces. The most significant feature in recent history has been the mining of progressively lower grades of ore and the retreatment of tailings from earlier operations, much of the latter comprising tailings only a decade or even less in age. Notable among these operations are those on the East Rand near Springs in the Transvaal, in the Orange Free State at Welkom and more recently, massive reclamation of tailings close to the city centre itself. The carbon in pulp (CIP) process has been much publicised in recent years and all new plants erected during the 1980's have adopted this route for gold recovery. However the major proportion of tonnage milled in the industry is still subjected to the more traditional process of liquid solids separation by vacuum filtration. The discussion that follows will centre largely on the developments in the traditional process route, which for the sake of completeness is listed below: Comminution Dewatering Leaching Liquid solids separation Gold recovery Smelting

Jan 1, 1985

By A. Bakalarz, M. Duchnowska

"The copper sulfide ore from the Legnica-Glogow Copper Basin in Poland, which is processed by KGHM Polska Miedz S.A., consists of three lithological fractions: dolomitic, sandstone and shale. The copper ore, depending on the mining area, contains all fractions in different ratios. The lithological fractions differ in the particle-size distributions of their useful minerals. In the dolomitic part of the ore, most of the metal sulfide particles are 50-200 microns in size, while those in the sandstone and carbonaceous shale are 30-200 and 5-40 microns in size, respectively. This complex composition results in different mineralization, physicochemical properties and, in particular, upgrading selectivity. Due to the specific nature of the examined copper ore, two of the crucial parameters determining the beneficiation efficiency are size and liberation of sulfide particles. We analyzed the influence of grain size of feed on scavenging selectivity in one of the KGHM flotation plants and performed evaluation and analysis of the upgrading results of industrial beneficiation of the ore, containing mostly dolomite and shale. The upgrading efficiency was estimated using a selectivity indicator obtained by means of the Fuerstenau upgrading curve, which was then correlated with the particle-size distribution of the flotation feed.IntroductionFroth flotation is used extensively in commercial ore treatment flotation plants to separate minerals from gangue and to separate minerals from each other. The floatability of ores and their benefication selectivity are determined by many factors. One of them is particle size. The effect of particle size on flotation performance has been widely studied by many authors (Gaudin et al., 1931; Trahar, 1976; Trahar and Warren, 1976; Lynch et al., 1981; Pease et al., 2004). Particle size also plays a critical role in the probability of particle collision and attachment to bubbles (Solari and Gochin, 1992; Schulze, 1993), kinetic flotation (Taggart, 1956; Ahmed and Jameson, 1985; Debacher, 2002; Brozek and Mlynarczykowska, 2007; Potulska 2008), collector consumption (Sutherland and Wark, 1955; Vianna et al., 2003) and mechanical entrainment (Lynch et al., 1981; George et al., 2004; Konopacka, 2005). Most of the results have shown that there is an optimum particle size range for a given flotation system and that the simultaneous optimization of the recovery and concentrate grade of these particles in conventional flotation equipment is extremely difficult."

Jan 1, 2015

By M. L. Wouden

The Bureau of Mines investigated pegmatite and granite formations at Copper Mountain, Fremont County, Wyoming, for recovery of tantalum and niobium. Sampling of surface exposures identified about 1.27 mega tonne of pegmatite resource material ranging between 0.02 and 0.04% Ta20, and 0.007% Hb20,. Beneficiation studies were conducted involving coarse-grained and fine grained pegmatite samples. Over 80% recovery of the coarse-grained tantalite was achieved in a concentrate averaging over 45% combined tantalum-niobium oxide. Lower concentrate grades and recoveries resulted from the treatment of fine-grained pegmatite.

Jan 1, 1991

By J-P Franzidis

Mill operators have traditionally been forced to rely on grade-recovery curves and overall flotation rate constants when analysing proposed improvements to plants. This has led to loss in money and time when a proposed change that showed potential in laboratory scale tests failed to show improvement in the plant. Frequently, this is a result of the operators being unable to gain insight into how and why the proposed change affected the flotation performance. This is due to the classical models not including any description of the subprocesses of flotation and their individual contribution to the overall flotation performance. A new flotation model has been developed as part of the AMIRA P9 project that separates the flotation process into subprocesses and includes the contribution of both the pulp phase and froth phase to the overall flotation recovery. It is not possible to determine the individual components of this new model on traditional pilot plants as they lack the required flexibility and scale. Therefore a new pilot plant has been developed known as the Floatability Characterisation Test Rig (FCTR) that has the flexibility and features required for the determination of the flotation subprocesses. The flexibility and features of the FCTR also make it an ideal tool for investigations into many different aspects of flotation research. These include the effect of reagents, grinding, and power on the flotation performance. This paper will describe the methodology for using the FCTR in these fields of research and highlights the advantages of using the FCTR over traditional methods.

Jan 1, 2000

By H. S. Chiang, Y. Luo, S. S. Peng

"IntroductionThe way in which the particle sizes distribute in the airborne coal dust plays an important role in the transportation of the dust in the mine air and human respiratory systems. Though the particles of a size less than 10 microns could be inhaled, the real distractive portion of the dust is that the sizes range from 1 to 5 microns since only they can reach the lungs and be retained there. Those with an equivalent diameter larger than 5 microns will settle on the membranes of the branchial tubes(1) while those with diameter less than 1 micron can either be exhaled out of the lungs or the settled quantity is so insignificant that does very little harm to the lungs. In the transportation process, the finer dust will fly farther away from the sources than the coarser ones. Thus once the fine dust is generated and becomes airborne, it will spread out to a larger area.In this paper, the coal samples collected from two longwall faces were analyzed statistically for size distribu¬tion. A mathematical model is proposed and the effects of different conditions on the dust size distribution were analyzed using the mathematical model.Sampling and Statistical AnalysesSampling Instrument and ProcedureThe instrument used in the sampling were Mine Safety Appliance (MSA) personal sampling units equipped with 10-mm nylon cyclones, pre-weighed MSA PVC filters and constant-flow pumps.Two longwall operating faces were selected for the dust sampling. The two faces were located in two coal mines ex-tracting Pittsburgh seam. The double-ended ranging-drum shearer applying the unidirectional cutting sequence was employed. The two-leg shield supports were used for face support. The width of the two faces was about 215 meters.In each of the two faces, four sampling stations were set up. The stations were located at the cross-cut immediately outby the headentry side, the 15th support (25 m), 65th support (107 m) and 111th support (182 m) from headentry T-junction, and numbered as #1, #2, #3, and #4 station, respectively (Figure 1). At each sampling station, a group of three sampling units was hung at proper position. At cross¬cut (#1) station, the units were hung about one foot under the roofline in the center of the cut (Figure 2a), and at the stations in the face, the units were position near the hydraulic legs and one foot below the support canopy (Figure 2b).One complete set of 12 samples was collected for each day. The sampling were performed for five consecutive days. All the samplings were conducted in the morning shifts. The time duration of the sampling for each day was about five hours. During that time, the shearer might complete 5 to 7 mining cycles."

Mar 1, 1989

By Oliver Hall

This paper sketches the mining at Noranda as it would be sketched to a visiting engineer. It does not follow quite the usual form, but is presented somewhat along the lines of a conversation. TO give the mining at Noranda its proper perspective, it is necessary to sketch the whole picture and outline: 1. History 2. Metallurgical Problems 3. Exploration 4. Mining Methods 5. Mining Details 6. Filling 7. Pillar Extraction 8. Costs HISTORY Noranda is a relatively young enterprise, with much of the personal character of its founders woven into it. The name Noranda is an abbreviation of Northern Canada and the name has been a happy one for an enterprise based on northern resources, tenacity and courage. The discovery credit goes to Ed. H. Horne. This intrepid prospector returned again and again to his sulphide discoveries on the shores of Osisko lake and, in 1922, after ten years of prospecting, succeeded in interesting S. C. Thomson and H. W. Chadbourne in his discoveries. The early exploration by L. K. Fletcher, from 1922 to 1925, for Thomson and Chadbourne, indicated promising, but somewhat perplexing, ore conditions. Mr. E. Hibbert and Mr. H. L. Roscoe rook on the task of preparing the mine for production and building a smelter in 1926, and this difficult task in this isolated area made steady progress and the smelter was blown-in in December, 1927.

Jan 1, 1937

By M. H. Kline

Round Mountain is one of a number of locations investigated by the Bureau of Mines in its search for manganese deposits during the early phases at the strategic minerals program. The Suiter mine, on Round Mountain, had provided some 5,000 tons of manganese concentrates during the period 1917 to 1919, and early in 1941 prospecting by the Virginia Hardwood Lumber Co. in this same area had discovered manganese on what later became known as the Hall Pit. These promising indications encouraged the Bureau to begin an exploration project in search of additional reserves of manganese on Round Mountain in January 1942 (fig. 1).

Jan 1, 1948

By C. Edwards

ABSTRACT The versatility and specificity of rare earth elements (REEs) have led to their use in an everincreasing variety of applications in new technologies; consequently, demand for REEs has increased significantly. However, separating REEs safely and effectively is a complex and expensive process. This paper presents a review of REE mineral processing technologies, providing an update on current capabilities in REE mineral processing and REE extraction from the major commercially valuable REEbearing minerals such as bastnaesite, monazite, and xenotime, as well as the ion-absorption type of REE deposits.

Jan 1, 2013