Search Documents

By Y. M. Zhao

Reutilization of waste printed circuit boards (PCB) can not only prevent potential environmental pollution, but also achieve recycling of secondary resources. Study on recycling of discarded PCB belongs to the front edge of environmental pollution prevention and resource recycling. In order to effectively crushing waste PCB, and solve the problem of secondary pollution of fugitive odor and dust during the crushing process, a wet impacting crusher was adopted to achieve comminution liberation of PCB in water medium. 96.79% of the crushed yield was distributed in ?2mm, and metal took 97.18% in this grade of production. Size distribution of PCB particles was characterized double peaks at the sizes of ?1+0.5mm and ?0.074mm. A double Rosin-Rammler function, is called DRR function, is put forward to describe the size distribution of PCB particles. The DRR function of cumulative weight percentage oversize of PCB was non-liner fitted by self-defining DRR function of the Origin 7.0. The evaluation indexes of the fitting results are the correlation coefficients R2 of 0.9994 and residual E of 0.18671, those show that DRR function is perfectly accurate. Based on the research the mechanical separation and material properties of PCB, a wet process of reclaiming metals from PCB using a new style separator ? tapered diameter separation bed (TDSB) in water medium was designed, and the related separation mechanism was studied. The fluid field of the tapered diameter separation equipment was simulated by the Ansys Flotran software, the variation laws of the velocity in each position in the fluid field as well as the equation of axial distribution sectional max velocity were analysed according to the simulation result, the stress change and motion trajectory of the pellet in the certain fluid field were analysed. The results show that the expanding tubular structure of the varied diameter separation bed indicates gradually decreasing velocity of water flow along the axis, assuming a certain angle in the bed, as a result of the gravity?s influence, and there was corresponding change to the distribution of the flow velocity, thus the whole formed a special fluid field where velocity variations exist in different positions. When the material passed through the fluid field along with the water flow, as a result of the difference to the particle density, the granularity and the shape, there would appear the difference to the motion trajectory as well as the movement distance under the action of the fluid: The lower density materials discharge along with water flow, but the higher density materials would be kept in bottom of the bed and discharge through material mouth. Therefore, light and heavy products were separated effectively according to the density. Using tapered diameter separation bed to process the ?1mm size crushed material of waste printed circuit boards, with the discharge of water 5.5m3/h, feed-rate of the material 250g/min and with an obliquity 35°, product with 93.92% in integration efficiency and 93.73% in recovery ratio of the metal were obtained by the primary separation. Compared with other mechanical methods, the range of particle size was broaden using the TDSB to recovery metals from PCB, and lower limit size of separation almost approach to 0. An innovative process of ?wet impact crushing+ classifying+ tapered diameter separation? recycling metals from waste PCB was developed.

Jan 1, 2014

By J. A. Halt, E. Larsen

"The Sydvaranger magnetite mine closed in 1996 due to low iron ore prices and was reopened in 2009. The flowsheet has remained more or less as in the old plant, based upon two grinding steps followed by magnetic separation. The magnetite is free of harmful elements and should be attractive on the high quality market. This requires a concentrate containing less than 2% SiO2, which previously was achieved by amine flotation. Amine was preferred because it could be used with sea water. The use of amine is now restricted by the Norwegian Climate and Pollution Agency. Therefore, testwork to replace cationic with anionic flotation was done. Activation of quartz with Ca or Mg ions prior to anionic flotation was shown by research from 1960 to 1980 to be effective under laboratory conditions. The old research results have been followed up in order to make a commercial process for Sydvaranger. Calcium activation works excellently with all Sydvaranger ores, including the ones containing amphibole minerals, but Ca activation requires the high and not environmentally friendly pH of 12.4. Magnesium activation at pH 10.5 is a little more challenging, because the maximum concentration of MgOH+ is about 100 times lower than that of CaOH+. Fortunately, the major impurity in the main Bjornevatn orebody is quartz, which responds well to Mg activation. The process does not appear to depend much upon water quality, as the silicates may already be activated during grinding and magnetic separation in recycled water. The new process is environmentally friendly, as only well-known household reagents are used; oleic acid or tall oil, dextrin, sodium hydroxide and magnesium chloride at pH 10.5.IntroductionThe processing plant of Sydvaranger Gruve AS (SVG) is located in the port town of Kirkenes in the northernmost part of Norway, 10 km by rail from the high-grade deposit of Bjornevatn. This plant has been able to beneficiate magnetite ore to up to 99.5% purity, as described by Sandvik and Malvik (2001). Using the magnetic properties of the ore and making the most out of its enviable liberation characteristics, the original plant consisted of ball mills and 600-mm magnetic separators. The plant was shut down due to market conditions in 1996 but reopened in 2009."

Jan 1, 2015

Heap leaching has become one of the prime technologies used for the recovery of precious metals from low-grade ores over the past ten years. The recovery process involves mining and stacking the ore on engineered leach pads, with or without crushing and/or agglomeration, followed by leaching with a low concentration sodium cyanide solution. The metal-bearing, pregnant solution is then processed by zinc precipitation or activated carbon absorption to extract the precious metals. The leaching process continues until all economic metal values are removed at which time the spent heap leach pads, the leached materials and excess process solutions require decommissioning. The management of these solutions may present a potential challenge for successful and environmentally sound decommissioning due to their composition, volume or continuity of generation from the spent ore even though active leaching has been terminated. The principal objectives of solution management in decommissioning spent heap leach pads are to meet all operative environmental regulations in a cost-effective manner and to leave the pads in a condition where they do not present a long-term liability either to the operator or to the environment. This involves the reduction in volume of the residual process solutions and/or treatment of these solutions to where they meet applicable water quality criteria. In the past, the environmental and economic aspects of decommissioning and solution management have been overlooked or under-estimated during the planning of mining operations. As a result, there is a potential for long-term liabilities and costs to be recognized only at the end of active leaching, a time at which operating revenues and cash-flow are minimal. Therefore, the decommissioning process requires addressing at the mine planning stage and agencies now require that a plan for solution management after the cessation of operations be submitted prior to initiation of mining. This paper discusses aspects of process solution management during decommissioning, both in terms of quantity and quality, and how potential issues may be recognized and planned for in the initial stages of heap leach design.

Jan 1, 1992

By Charles H. Greene

This paper will describe briefly the steps taken in the development of a Central Florida phosphate ore deposit from acquisition to mining. Further, the design and early operation of the mining and beneficiation facilities will be depicted. During the 1950's, with confidence in the continually increasing requirement for phosphate fertilizer materials essential for producing food, I'. R. Grace & Cc acquired the Hookers Prairie tract containing approximately 45 million metric tons o. (50 million short ton~ of potential phosphate rock product. In 1973, at a point when the Grace captive need for rock was being greatly enlarged by a planned doubling of its Bartow phosphate chemicals plant capacity, the decision was made to expand existing phosphate rock production by building a new mine and plant. The initial annual capacity was to be 2.3 million tons (2.5 million st); however, subsequent investigation pointed to a higher than estimated dragline operating factor 'and the forecast of annual capacity was set at 2.5 million tons (2.8 million st).

Jan 1, 1977

By B. Simonian

"The Mount Milligan mine, located in central British Columbia, began operation in September 2013 with a design nameplate capacity of 62,500 dry metric tonnes per day, using a conventional SABC grinding circuit followed by bulk sulphide flotation. Over the last few years, many changes have been implemented in order to debottleneck and optimize the comminution circuit to meet and exceed the nameplate design capacity as well as improve the metal recovery of the flotation circuit. This paper discusses changes, grouped into mechanical modifications, projects, and operational changes. The implementation of these changes as well as ongoing challenges and potential solutions will also be discussed.INTRODUCTION The Mount Milligan mine began production in the fall of 2013. It is a conventional truck-shovel open pit gold and copper mine and concentrator with a 62,500 tpd design capacity copper-gold flotation processing plant. The average head grades for the life of mine are 0.18% Cu and 0.39 g/t Au. At full capacity, the mine produces an average 400 dry tonnes per day of copper-gold concentrate containing approximately 21.5% Cu and 50-60 g/t Au. Current mine life is 21 years. This paper will discuss a selection of the operational, engineering, and administrative changes that have been made at the site since the publication of An Update on Reaching Nameplate Throughput and Continuous Improvement at Mount Milligan (Yue, 2017). PLANT FLOWSHEET CONFIGURATION Run-of-mine ore is crushed through a 60-110 gyratory crusher to a top size of seven inches. Up to 100% of the gyratory crusher product is diverted to the secondary crushing circuit by a moveable gate. The material diverted to the secondary crushing circuit is fed to a vibrating double-deck screen. The screen oversize is placed on a surge pile and reclaimed by four vibrating feeders to feed two 1250 kW standard head cone crushers in open circuit. The screen undersize is combined with the crusher discharge and returns to the conveyor feeding the coarse ore stockpile along with primary crusher product that by-passes the diverter gate."

Jan 1, 2019

By H. Kenworthy, M. M. Fine

The progress made to date in the mineral dressing of complex Arkansas titanium ores is reported in this paper. Concentrates of rutile, a dioxide of titanium, were produced by treating a submarginal ore from the Magnet Cove Rutile Co., Inc., Magnet, Ark. The rutile recovery at present is 45 to 50 pct of the total TiO2 with the possibility of increased recovery with additional research. The work reported herein is part of a general study which will later include results of investigation of other titanium ore bodies in the Magnet, Ark., area. The fragile character of the rutile minerals, the softness of the gangue, the occurrence of pyrite and leucoxene, and the considerable proportion of extremely small size rutile in the presence of montmorillonite- and illite- type clays in the Magnet Cove open-pit mine are the important factors in the mineral dressing of this submarginal rutile ore.

Jan 1, 1949

At Pasminco Mining's South concentrator. which was commissioned in 1952. an expansion. modernisation and efficiency improvement program commenced in 1986 with a coarse ore stockpile and grinding circuit expansion. the replacement of small flotation cells and the upgrading of reagent mixing and metering equipment. The program continued with the installation of a TDC 3000 based process control system (1988) and concentrate thickeners (1990). By mid-1991 a further grinding expansion and the installation of zinc flotation feed regrind ball mills and zinc recleaner column cells will be complete. leaving the use of ceramic disc fillers and development of process control strategies as probably the remaining major avenues for improved process efficiency at present.

Jan 1, 1991

By Dudley L. Davis, E. C. Peterson

INTRODUCTION THE Darwin District is 30 miles east of Olancha which is 220 miles north from Los Angeles via U. S. Highway No. 6. The ore deposits occur in the Darwin hills that have been elevated above the Darwin Plateau by typical basin and range-type faults. Some of these faults have been active quite recently as shown by the displacement of Quaternary (?) basalt flows and the rejuvenation of erosion in Darwin wash. The ore bodies occur as fissure fillings and replacement of favorable beddings in a thick series of Paleozoic limestones, dolomites, shales and quartzites. Rich oxidized lead-silver ores were discovered in the early seventies. By 1875, Darwin had a population of about 5000, and by 1880 several small mills and smelters had been built. Early exhaustion of surface ores, the isolated location of the district, and fluctuation of metal prices caused intermittent operations until World War II. The district production is estimated at about $3,000,000 prior to 1900 and perhaps $4,000,000 between 1900 and 1945 when Anaconda purchased the major producing mines. For the past two years, daily production has averaged 75 tons of direct shipping and 150 tons of milling-grade ore. At present, 300 tons of mixed oxide and sulphide ore are treated by flotation, the oxidized lead minerals being activated by addition of sodium sulphide after galena and sphalerite have been recovered from the circuit. Extraction averages 85 pct of the lead, 80 pct of the silver, and about 33 pct of the zinc in a bulk concentrate. The mill is designed so that minor changes permit selective flotation of pyritic lead-zinc ores encountered at depth. GENERAL GEOLOGY The oldest rocks in the Darwin hills are a series of Paleozoic limestones, dolomites, shales and quartzites probably Pennsylvanian in age. These rocks have been intruded on the west by the Coso granite batholith and on the east by a granodiorite stock. Numerous sills and dikes grading from orthoclase granite to gabbro are exposed in the underground workings. The sedimentary rocks have been considerably folded and faulted, the most prominent structure being a northwest pitching anticlinal fold, the crest of which lies just west of the ridge line of the Darwin hills. The east limb of this fold has been intruded by the granodiorite stock and east of the stock the sediments show several closely spaced anticlines and synclines. The minor flexures on the west flank of the major fold are structurally important in localizing the ore in the Defiance and Essex mines of the Darwin group. There are three principal systems of faulting and fissuring, most of which are post-intrusive and premineral in age, with minor amounts of postmineral movement. The Darwin tear fault is the largest, most persistent fault in the district. It strikes N 60 to 95°W, dips steeply south and is traceable for about 1o miles. It is a normal

Jan 1, 1947

By Dudley L. Davis, E. C. Peterson

Introduction The Darwin District is 30 miles east of Olancha which is 220 miles north from Los z4ngeles via U. S. Highway No. 6. The ore deposits occur in the Darwin hills that have been elevated above the Darwin plateau by typical basin and range-type faults. Some of these faults have been active quite recently as shown by the displacemerit of Quaternary (?) basalt flows and the rejuvenation of erosion in Darwin wash. The ore bodies occur as fissure fillings and replacement of favorable beddings in a thick series of Paleozoic limestones, dolomites, shales and quartzites. Rich oxidized lead-silver ores were dis'Overed in the early seventies. BY 1875, Darwin had a population Of about 5000, and by 1880 several small mills and smelters had been built. Early exhaustion of surface ores, the isolated location of the district, and fluctuation Of metal prices caused intermittent operations until World War 11. The district production is estimated at about $3,000,000 prior to 1900 and perhaps $4,000,000 between 1900 and 1945 when Anaconda purchased the major producing mines. 'or the Past two Years, daily production has averaged 75 tons of direct shipping and 150 tons of milling-grade ore. At present, 300 tons of mixed oxide and sulphide ore are treated by flotation, the oxidized lead minerals being activated by addition of sodium sulphide after galena and sphalerite have been recovered from the circuit. Extraction averages 85 pct of the lead, 80 pct of the silver, and about 33 Pct Of the zinc in a bulk concentrate. The mill is designed so that minor changes permit selective flotation of pyritic lead-zinc ores encountered at depth. General Geology The oldest rocks in the Darwin hills are a series of Paleozoic limestones, dolomites, shales and quartzites probably pennsyl-vanian in age. These rocks have been intruded on the Rest by the Coso granite batholith and on the east by a granodiorite stock. Numerous sills and dikes grading from orthoclase granite to gabbro are exposed in the underground workings. The sedimentary rocks have been considerably folded and faulted, the most prominent structure being a northwest pitching anticlinal fold, the crest of which lies just west of the ridge line of the Darwin hills, The east limb of this fold has been intruded by the granodiorite stock and east of the stock the sediments show several closely spaced anticlines and synclines. The minor flexures on the west flank of the major fold are structurally important in localizing the ore in the Defiance and Essex mines of the Darwin group. There are three principal systems of faulting and fissuring, most of which are post-intrusive and premineral in age, with minor amounts of postmineral movement. The Darwin tear fault is the largest, most persistent fault in the district. It strikes N 60 to 75°W, dips steeply south and is traceable for about 10 miles. It is a normal

Jan 1, 1949

By G Zhang, I Glacken

Multiple indicator kriging (MIK) has been used in the minerals industry for some decades. As one of the non-linear estimation methodologies, MIK has advantages related to resolving multiple or mixed populations, high variability data and strongly skewed distributions. Applying the MIK methodology requires a lot more effort when compared to other non-linear methodologies, for example localised uniform conditioning. This is due not only to the variogram modelling for multiple indicators, but the validation is also very important because the panel model is fundamental for follow-up work, including the generation of recoverable resources at the local or SMU scale. Apart from this, only a few reported case studies of MIK have documented post-processing of the estimated modelled distribution, representing the probability distribution at point scale at un-sampled locations. The MIK post processing should include the change of support from the point scale to the panel scale, and then the extraction of quantile values from the panel conditional cumulative distribution function (CCDF) for localisation of grades into the SMUs within each panel, based on a ranking estimate marking the SMUs from high to low. In this paper, we present a case study on a gold deposit with the MIK method with full-post processing implemented for recoverable resources. The MIK point estimate was carried out with an indirect lognormal change of support from point to panel scale, and localisation was done with a custom script. The localised uniform conditioning (LUC) method was also applied for comparison purposes. The comparison shows that MIK has advantages compared to LUC when comes to capturing the high-grades, especially when mixed or varying anisotropy is present within the mineralisation.

May 24, 2023

By J Lombardi

OZ Minerals' Prominent Hill copper-gold concentrator is located 130 kilometres south east of the town of Coober Pedy in the Gawler Craton of South Australia. The concentrator was built in 2008 and commenced commercial production in early 2009. The Prominent Hill concentrator is comprised of a conventional grinding and flotation processing plant with a 9.6Mtpa ore throughput capacity. The flotation circuit includes six rougher cells, an IsaMill for regrinding the rougher concentrate and a Jameson cell heading up the three stage conventional cell cleaner circuit. In total there are four level controllers in the rougher train and ten level controllers in the cleaning circuit for eighteen cells.Generic proportional – integral and derivative (PID) control used on the level controllers alone propagated any disturbances downstream in the circuit that were generated from the grinding circuit, hoppers, between cells and interconnected banks of cells, having a negative impact on plant performance. To better control such disturbances, FloatStar level stabiliser was selected for installation on the flotation circuit to account for the interaction between the cells. Multivariable control was also installed on the five concentrate hoppers to maintain consistent feed to the cells and to the IsaMill. An additional area identified for optimisation in the flotation circuit was the mass pull rate from the rougher cells. FloatStar flow optimiser was selected to be installed subsequent to the FloatStar level stabiliser. This allowed for a unified, consistent and optimal approach to running the rougher circuit. This paper describes the improvement in the stabilisation of the circuit achieved by the FloatStar level stabiliser by using the interaction matrix between cell level controllers and the results and benefits of implementing the FloatStar flow optimiser on the rougher train.

Aug 8, 2011

By D. W. Frommer

As one phase of a general problem of recovering sulfide slimes, mineral-dressing research was conducted on flotation tailing from the St. Joseph Lead Co., Bonne Terre, Mo., mill to recover a significant portion of the residual, finely divided galena. Production of a high-grade concentrate was not part of the immediate objective. Only a reasonable ratio of concentration of lead sulfide slime was sought, since the problem of upgrading by mechanical or extractive processes would be considered in the later stages of this study. The results show that, from a feed material averaging 0.20 percent lead (about 70 percent finer than 400-mesh), a recovery of up to 45 percent could be effected by flotation in rougher concentrate analyzing up to 6.5 percent Pb. These results were obtained by using larger quantities of various xanthates and sodium sulfide than used customarily in mills. Sodium sulfide, which is ordinarily detrimental to flotation of galena, was beneficial in this case.

Jan 1, 1959

By Foster Fraas

In the separation of a component from a mixture , two fractions are obtained , one containing a high - percentage composition of the component and the other a low - percentage composition . The high - percentage fraction is frequently called the concentrate and the low - percentage fraction the tailing . In many separation processes the recovery is not complete , and the composition of the concentrate decreases as the recovery of the component, which is concentrated , increases . It is this particular class of separations to which this paper is confined .

Oct 1, 1942

By Todd Kilduff, Paul Wilkinson

Microtunneling is rarely a straight forward “business as usual” endeavor. More often than not, these projects take twists and turns that require on the fly, out of the box solutions to get the project completed. This paper is a case history of a 63-inch diameter, Permalok, microtunneling project, extending nearly 700 linear feet under a hill forming the left-hand abutment to an earth embankment tailings dam. The job had three major and unexpected challenges that required unorthodox solutions. Challenge #1— Face Configuration: The Santiago Formation, a cemented sand formation with widely varying strengths, presented a challenge to configuring the face with regards to the number and length of pics and cutters. Subtle differences in type and size of the tooling had a big impact on the performance of the machine. Challenge #2—Eccentric Loading: Issues at the launch shaft with the Permalok pipe adaptors caused eccentric loading on the casing pipe. This issue resulted in overstressing and breaking of several pipe joints that ultimately led to seizure of the tunnel. Challenge #3—Back-tunnel Excavation and MTBM Recovery: There were many detailed considerations into excavating a recovery tunnel to retrieve the MTBM, especially being under a high hazard earth embankment dam.

Jun 13, 2021

By E Amini, M Scott

Grade Engineering® involves the planning, integration and operation of coarse preconcentration techniques to reject uneconomic material as early as possible in the mining value chain to improve the quality of ore delivered to mineral processing activities. These preconcentration techniques include screening for natural and induced deportment of grade to finer size fractions during blasting and coarse breakage, sensor-based measurements of grade and mechanised sorting at particle or bulk scales and coarse gravity separation. Over the last decade, the Cooperative Research Centre for Optimising Resource Extraction (CRC ORE), with the support of its participants, has developed a framework for the holistic evaluation of economic and production impacts attributable to these preconcentration techniques. The framework combines the characterisation of preconcentration responses, geostatistical modelling, process modelling and simulation of downstream and upstream production impacts and the optimisation of strategic mine plans to quantify changes in net present value due to different preconcentration strategies. This framework has been applied to evaluate preconcentration techniques at numerous metalliferous operations and can also be used to quantify changes in the environmental metrics for energy, water and steel consumption and emission, tailings and waste rock production. This paper collates the results of these Grade Engineering® studies to show the range and magnitude of economic and environmental impacts attributable to preconcentration. The studies have shown that rejection of uneconomic material prior to mineral processing activities can have a pronounced improvement in energy, water and comminution steel productivity, emission intensity and the profitability of metalliferous mining.

Nov 10, 2020

By M. Riddell

This paper describes the incipient heatings that occurred below ground in the High Main and Main seams at the mine between the years 1963 and 1966. The possible causes, effects and activities associated with such heatings are discussed in an attempt to establish the precautions necessary to minimize such occurrences. The assessment of as many "norms" as possible is made to enable the earliest detection and level of activity to be determined.

May 23, 1905

By Wickham P

Digital computer based control systems are used to implement control and monitoring over most unit operations in Mount Isa Mines Limited lead-zinc concentrator and copper concentrator. Implementation has been gradually devel- oped and expanded since the first process computer installation in 1969. Operations currently using high level control logic include crushing, grinding, on-stream analysis of flotation products, flotation reagent additions, and tailings thickening. Details of gains in metallurgical efficiency for each system are considered.

Jan 1, 1978

This work describes preliminary laboratory testwork to study the relative leaching response to cyanidation and thiourea leaching of an oxidised low grade gold ore. The potential of an alternative lixiviant like thiourea to commercial processing of mine dumps is discussed by specific reference to selected case histories. It would appear that on the basis of the available published information, there could well be a case advanced for the acid thiourea leaching of sulphide concentrate tailings being subjected

Jan 1, 1994

By Richard H. Otto

Three longwall mining methods practiced in foreign countries are described that have potential for increasing underground recovery and productivity in thick coalbeds of the Western United States. The methods are multislice long-walling, longwall caving, and high-face longwalling. Foreign practice and possible application in the United States are discussed.

Jan 1, 1979

By Stanley G. Butcher

"Introduction Many operators in western Canada, with a background in the minerals industry, talk of ""coal beneficiation"", whereas the coal industry generally uses the term ""coal preparation"" . The sophisticated cleaning circuits and thermal drying units found in western Canada makes the term beneficiation particularly appropriate. This chapter uses standard coal preparation terminology(l,2J and assumes general knowledge of coal preparation unit processes and equipment(3,4,5). The geologic setting and mining methods of western Canada, described elsewhere in this volume, present for preparation run-of-mine coals which have particular problems. As a result the cost of preparation is a higher proportion of the total minesite production costs than in other coal fields . Preparation requirements are determined by the types of coals being produced and the markets in which they are sold. Western Canadian coals have a low sulphur content and do not require cleaning on this account. Prepared coal from western Canada has to be shipped vast distances -to the Pacific Rim countries and Ontario in particular -and the savings in transport costs justify the high costs of preparation. Factors which dictate selection and operation of the preparation processes are: i) Market specifications and coal rank. ii) Size distribution of particles and particularly the liability of coal and refuse particles to degrade during handling and processing. iii) Quantity and characteristics of the associated refuse. iv) Washability characteristics (the distribution of particles by specific gravity within each size fraction). v) Surface characteristics of the fines. vi) Variability of the raw coal. Where the product (coal or discard) has a high fines content special consideration must be given to dewatering, the removal of surface moisture, to give acceptable handle ability characteristics. In the severe winter climate of the western foothills and mountains thermal drying is necessary to prevent freezing of washed coal in stockpiles and railcars."

Jan 1, 1985