fluorspar beneficiation process plant
Acid grade fluorspar which is in great demand by the chemical and aluminum industries, must contain at least 97.5% CaF2 with not more than 1.5% SiO2 and 0.5% Fe2O3. Often the Silica is limited to 1.2% with penalties starting at 1.0% SiO2. These limitations on grade and impurities require extremely close mill control, particularly through flotation where selectivity and high recovery is essential.
Over 95% of all acid grade fluorspar is processed by flotation through Sub-A (Fluorspar Type) Flotation machines. These machines, of the cell to cell type, are designed special for fluorspar with a high degree of flexibility essential for selectivity and multiple cleaning of concentrate. Middlings and clean tailings often must be completely isolated from the separate cleaning steps and diverted to the proper part in the milling circuit for most economical and efficient retreatment.
The flowsheet illustrated above is typical for the average Sub-A Fluorspar Flotation mill treating up to 100 tons of mine run ore per 24 hour day. Actual flotation conditions and equipment requirements should always be determined by having a comprehensive test made on the ore before proceeding with any fluorspar operation. Fluorspar ores may be quite complex, particularly when associated with lead and zinc sulphides, barite, calcite, iron oxide, and siliceous impurities. For this reason, a laboratory flotation test should be the first step in establishing a flowsheet.
For the average small mill treating up to 100 tons of ore a day, primary crushing is usually adequate and very economical. Larger tonnage will require primary and secondary crushing for maximum efficiency in size reduction and subsequent ball milling.
Fluorspar ores usually require grinding to 48 or 65 mesh to liberate the calcium fluoride from the gangue impurities.
Ball mill grinding with a Steel Head Ball Mill in closed circuit with classifier is the general practice. In larger plants, particularly when fine grinding is necessary, thickening of the classifier overflow is necessary to maintain proper density and feed regulation to flotation. This thickening step on fluorspar ores containing sulphides is usually between the sulphide and fluorspar flotation circuits. Reagents used for selective flotation of lead and zinc then can be rejected in the thickener overflow water.
Normally, conditioning at mill temperature willthoroughly film the fluorspar with reagent and makeit readily amenable to separation and recovery by flotation. Heating the pulp, even up to the boiling point,is often advantageous.
AnAgitator and Conditioner is ideal for fluorspar conditioning as circulation is positive and thorough reagentizing with a minimum amount of reagent is assured. Any frothing tendency is dissipated in the pulp through the stand pipe and adjustable froth collar.
Flotation of fluorspar must be extremely selective when producing acid grade concentrate. This selectivity is essential as the ratio of concentration is low, often up to 80% or more of the entire tonnage, and must be floated in the rough circuit. Cleaning by two or more stages of flotation must bring the rougher product up to acid grade and at the same time retain a high weight recovery with a minimum circulating load.
The Sub-A Flotation machine, the accepted standard in all fluorspar flotation plants, has been adapted specially for fluorspar treatment with provision for multi-stage cleaning and recirculation of middling products without the need of auxiliary pumps. Cleaner tailings may be conveniently removed at any point in the circuit. The flowsheet on the reverse side of this page shows one of the many possible cell arrangements used in treating fluorspar ore.
Thickening of fluorspar concentrates offers no special problem. Thickener capacity, however, should be adequate to handle the tonnage and have ample storage capacity during possible interruption in the filtering and drying sections. Fluorspar flotation froth has a tendency to build up on the thickener surface, but this can be taken care of by retaining rings near the overflow lip and by sprays so only clear water overflows the thickener. Thickened concentrates at 50 to 60% solids is removed by a Adjustable Stroke Diaphragm Pump, feeding by gravity to the filter.
Fluorspar is extremely rapid filtering even when ground fine, provided a non-blinding filter media is used. The rotary fluorspar type filter with stainless steel filter media, heavy duty oscillating mechanism, oversize valve and ports, and high displacement vacuum pump is standard for fluorspar flotation concentrates and will discharge a filter cake with as low as 6% moisture. In the event the filtrate is slightly turbid or contains solids, it should be diverted back to the thickener. For this reason a adjustable stroke diaphragm pump is often used in place of the conventional centrifugal filtrate pump.
Fluorspar flotation concentrates of acid grade must be dried to less than 0.5% moisture. Dust losses are kept to a minimum by providing a closed system with a cyclone to insure only vapor laden air discharging to the atmosphere. Enclosed screw conveyors, elevators and often air-born systems are used to transport the finely divided dried acid spar to the storage bins. Provisions should be made for handling efficiently the hot concentrate discharging from the dryer. The Standard Dryer is ideal for this purpose.
Fluorspar ores often contain appreciable amounts of sulphides in the form of galena, sphalerite, or both. These sulphides, when present, not only represent a valuable constituent of the ore, but also must be removed prior to fluorspar flotation to meet the market specifications for acid grade fluorspar.
If lead and zinc were present, the same flowsheet would apply to remove a bulk sulphide concentrate which could be subsequently refloated to produce the respective lead and zinc concentrates suitable for marketing.
The best approach to effectively produce separate lead and zinc concentrates should be established by test work. In some cases, selective flotation is indicated initially. This may be accomplished by removing a lead concentrate, then following this process by conditioning and flotation of the lead tailing to produce a zinc concentrate.
Conditioning of the classifier overflow is required if sulphidization is employed to effect flotation of oxidized lead. A second stage conditioning of the thickened lead tailing, after repulping with fresh water, is required for flotation of the fluorspar. Heating of the pulp at this point is often advantageous.
The lead and fluorspar are recovered by Flotation of the cell-to-cell type, permitting maximum recovery and grade of concentrate. Wide acceptance of machines is well verified when considering that over 95% of all acid grade fluorspar is processed in the Sub-A Flotation Machine. Flexibility of these machines is of prime importance where such high specifications must be met. Multiple cleaning, always necessary in acid grade fluorspar plants, can be performed without the help of pumps.
Both concentrates are thickened and filtered. The thickenedlead concentrate is filtered on the Disc Filter. Thickened fluorspar concentrate, at approximately 60% solids here, has a high filter capacity of approximately 2000 pounds per sq. ft. per 24 hours. The Fluorspar Filter with its stainless steel filter media, is especially designed for this application.
The Standard Dryer effectively dries the filtered fluorspar concentrate to less than 0.5% moisture, as required for marketing. An elevated temperature in the dryer can also be used to burn off small amounts of sulphur and lead.
A screw conveyor and bucket elevator as employed to transport the dried fluorspar to the concentrate storage bins. Bins can be conveniently discharged into rail road cars for shipment, while the filtered lead concentrate may be marketed as produced, without drying.
While many ores respond to the same general pattern of treatment, each ore is an individual problem.Such is the case of this fluorspar ore which is characterized by the presence of a portion of the fluorite inextremely close association with calcium carbonate andsilica and containing appreciable clay.
High acid grade fluorspar concentrates are difficult to obtain from this class of ores by flotation with an ordinary -65 mesh grind. The concentrates, in this study, are currently being used for production of hydrofluoric acid and synthetic cryolite. Market requirements demand that the calcium carbonate content be reduced to an absolute minimum. Moreover, the future productionnow in demand, is desirable. This study deals with a flowsheet designed to achieve high recovery of acid-grade fluorspar in an economical manner.
The typical fluorspar flotation flowsheet normally consists of stage grinding by ball mill in closed circuit with a mechanical classifier followed by conditioning of the pulp either with or without steam in the presence of reagents followed by Sub-A Flotation with three or more cleaning steps by reflotation. This particular ore does not, with the normal flowsheet, produce an acid grade concentrate of 97.5% CaF2 with less than 1.5% SiO2.
The ore being studied is crushed underground at the mine and partially beneficiated by the heavy media process. This washed ore is further crushed at the mill. Soda ash is added to the primary grinding mill which is in open circuit with a duplex Spiral Classifier. The classifier is in closed circuit with the secondary grinding mill and the classifier overflow, which is all 65 mesh, is pumped by a SRL Pump to the Conditioner where the following reagents are added:
Reagent Amount, Pounds per ton
Na2Si03 (Optional) 0.2
Soda Ash 2.0
Oleic Acid up to 2.0
The conditions presented by this particular ore illustrate the importance of complete laboratory investigations as a great many different combinations of treatment were required to develop the final flowsheet. The deviations from the standard fluorspar flowsheet were first substantiated by locked cycle batch laboratory tests followed by a small tonnage pilot plant run to verify the laboratory results before final recommendations were made.
The rougher flotation circuit produces a final tailings while the rougher concentrate is subjected to the first cleaning stage. A 6 cell Sub A Flotation Machine, cell to cell type, is used for the rougher flotation and 6-cell Sub-A Flotation Machines are also used for the three cleaning steps.
Tailings from the first cleaners are pumped to a Morton 2-stage Cyclone for the removal of clay slimes. The ability to add clear water for washing in the classifier makes the Morton Cyclone particularly useful at this point in the flowsheet. The slimes go to final tailings and the cyclone sands, at high density, are reground in a Regrind Mill which is in closed circuit with a Hydro-Classifier. The regrind is to 325 mesh and the hydro-classifier overflow returns to the first cleaner cells for reflotation. Reagent sodium silicate is recommended to aid classification.
Concentrates from the first cleaners go to the second cleaner cells where further up-grading takes place.
The middlings (tailings) from the second cleaner cells go to the hydro-classifier in the re-grind circuit. The concentrates from the second cleaners advance to the final cleaners. Tailings from the final cleaner cells are returned to the second cleaners and the final, high grade concentrates are filtered, dried and shipped to market.
The concentration of fluorspar ores for the production of acid grade concentrates is accomplished by the use of combinations of reagents such as pH regulators, depressant and fluorspar promoters. The reagents commonly used are as follows:
Factors of simplicity, initial low plant cost, together with flowsheet flexibility for maximum results on a difficult ore were basic considerations in the design of this 125-ton per day Fluorspar Flotation Mill. The design proved successful and accomplished the desired metallurgical results, with low capital expenditure and operating costs.
Following numerous laboratory tests, a flowsheet was developed that gives flexibility to handle the several types of fluorspar ores. Two stage open circuit crushing, with the average ore ground to 100 mesh,gives maximum results. Fine grained ores with some sulphides require secondary classification and a sulphide flotation stage. Due to character of most fluorspar ores heating the pulp gave improved results, and necessitated the installation of a boiler to provide hot dilution and make up flotation water for five stages of cleaning and recleaning. A Apron Feeder controls the feed from crude ore bin to jaw crusher while a wedge bar grizzly ahead of the jaw crusher removes the fines from the crusher feed. A 2x4 Dillon Screen removes the fines ahead of secondary crushing. An adjustable stroke ore feeder controls feed to the 5x8 Steel Head Ball Mill, and the spiral classifier discharge is pumped direct to flotation section or to hydroclassifier for secondary classification, depending on requirements.
The machinery was located for accessibility, ease of operation, minimum loss of floor space, resulting in reduced size of mill. The crude ore bin was constructed of natural timber on the site, on a steep slope, reducing expense of excavation and construction. An 8 clear opening rail grizzly prevented oversizegoing into bin.
The buildings for crushing section and mill are of light steel construction with corrugated sheet metal on walls and roof. The frame work and trusses lightweight for buildingsupport only and provided without insulation, because of mild climatic conditions. Account of heavy snowfalls the roof slopes are all of quarter pitch.
Launders on cleaning stages are made so that flows can be changed to regulate number of cells required, depending on the type ore being treated. Wood platforms and walkways of 2 spaced lumber are used in flotation sections, while piping between machines is carried below the floor.
All electric lighting and power wiring with ample reserve are in rigid conduit with flexible connections to motors; and motor controls are mounted on wall panels with stop and start push button stations located within sight or near each motor. Fluorescent lighting is provided over flotation section, as it gives operators better visual control of the flotation operation.The Rotary Dryer is lined with fire brick at discharge (burner) end.
With depletion of high-grade deposits, production must depend upon low-grade deposits that are highly contaminated with impurities which may be silica, calcite, barite, iron oxide, and sulphides such as pyrite, galena, and sphalerite, in close association. The flotation problem is largely one of impurity removal. The sulphide minerals are generally floatedfirst, and then the fluorite is floated from the silica, calcite, and other impurities.
Oleic acid or various mixtures of oleic and linoleic acids with soda ash and sodium silicate as silica depressant and slime controller, and quebracho to depress calcite, are the common reagents for fluorspar flotation. Sometimes pre-sulphide flotation with xanthate and a frother is necessary to remove sulphides and, often, heating the pulp to boiling temperature is advantageous in effectively depressing the silica, calcite and other associated minerals in the cleaning stages.
[/fusion_builder_column][fusion_builder_column type=1_1 background_position=left top background_color= border_size= border_color= border_style=solid spacing=yes background_image= background_repeat=no-repeat padding= margin_top=0px margin_bottom=0px class= id= animation_type= animation_speed=0.3 animation_direction=left hide_on_mobile=no center_content=no min_height=none]Geology-Where-are-Fluorspar-Deposits
This report is the fourth in a Bureau of Mines series describing the sodium fluoride-lignin sulfonate-fatty acid process of froth flotation separation of fluorspar from complex ores containing fluorspar, barite, calcite, and quartz which was developed and patented by Clemmer and Clemmons of the Bureau of Mines. At the Tucson (Ariz.) Metallurgy Research Laboratory the ores of Arizona were studied; and at the Tuscaloosa (Ala.) Metallurgy Research Center, the ores of Kentucky, Tennessee, and Illinois were studied.
The sodium fluoride-lignin sulfonate-fatty acid process is applicable to a variety of ores of different grades and mineral association for recovery of fluorspar from associated gangue materials; it has been shown to be practicable in continuous pilot plant operation as well as laboratory-scale flotation tests. This report deals with the application of the process to a complex calcareous fluorspar ore from Illinois and presents the results of laboratory batch flotation tests and continuous pilot plant flotation tests for recovery of the fluorspar in the ore.
The largest use of fluorspar is in the production of hydrofluoric acid in which no satisfactory substitute for acid-grade fluorspar is known. A prospective new outlet for hydrofluoric acid is in its addition to the oxidizer of the Atlas rocket, which will significantly increase the booster performance. The second major use of fluorspar is as a flux in the manufacture of basic open hearth and basic electric furnace steels in which no suitable materials are available to replace metallurgical-grade fluorspar, A third use of fluorspar is in the manufacture of glass and ceramic products. The specifications and prices of the various grades of fluorspar are listed in appendix A.
The complex fluorspar ore used in the investigation was from the fluorspar district near Cave-in-Rock, III. ; a 14-ton sample of ore was obtained from the Minerva Co. Crystal mine located about 5 miles west of Cave-in-Rock.
Petrographic examination showed that about 38 percent of the fluorspar reporting to the minus 48- plus 65-mesh fraction contained inclusions and that about 24 percent of the fluorspar was locked in the minus 325- plus 400-mesh fraction. However, the carbonate and quartz crystals locked in the fluorspar mineral were extremely small; in a concentrate analyzing 98.0 percent CaF2, 30 percent of the fluorspar grains were locked. The petrographic analysis revealed that no appreciable benefit to mineral liberation would be achieved by crushing finer than 65 mesh.
The primary carbonate in the ore was calcite with a considerable quantity of dolomite. The silica present was reported as quartz. Other materials consisted of 1.3 percent sphalerite and minor amounts of barite and galena. A chemical analysis of the sample is shown in table 1.
Samples of the ore were prepared for flotation by dry crushing to minus 10 mesh followed by wet stage grinding to minus 65 mesh in a laboratory pebble mill, using Tuscaloosa city tap water that had about 45 parts per million equivalent calcium carbonate total hardness. Prior to flotation, the ground ore pulp was treated, at about 40 percent solids, in a mechanically agitated flotation cell with conditioning reagents and then with a collector. A rougher fluorspar concentrate was floated off and cleaned six times.
A series of preliminary flotation tests was made of the ore to determine the quantities of sodium fluoride and calcium lignin sulfonate necessary to produce the maximum recovery and grade of fluorspar. The quantities of sodium fluoride and calcium lignin sulfonate were varied from 2.0 to 8.0 pounds per ton of ore; the quantity of oleic acid was held constant at 0.48 pound per ton of ore. The grade of fluorspar concentrate was increased with the dosages of sodium fluoride and calcium lignin sulfonate and leveled off at 5.0 pounds per ton. The summarized results of flotation tests made to determine the effect of varying the quantities of sodium fluoride and lignin sulfonate are given in table 2.
The laboratory batch flotation studies were continued to determine the optimum quantity of collector needed to obtain the maximum grade and recovery of fluorspar. The pulp was conditioned (1) with 5.0 pounds of sodium fluoride per ton of ore. to disperse the pulp and clean up the mineral faces, (2) with 5.0 pounds of calcium lignin sulfonate per ton of ore to coat the surfaces of the gangue particles and render them hydrophilic, and (3) with various quantities of sodium oleate, as a collectors to concentrate the fluorspar. In most instances an acid-grade fluorspar concentrate was obtained. The rougher concentrate contained 94.3 percent of the total fluorspar in the ore at a collector dosage of 0.30 pound per ton of ore, but the mineral particles did not adsorb enough collector to sustain their flotation during the six cleaning stages. About 0.50 pound of collector per ton of ore appeared to be the optimum dosage; the grade and recovery of fluorspar were essentially constant with larger quantities. This indicated that large quantities of sodium oleate were adsorbed by the fluorspar mineral and not by the gangue materials. The summarized results of these tests are shown in table 3.
Another series of tests was made using various quantities of oleic acid as the collector while maintaining the quantities of sodium fluoride and calcium lignin sulfonate at 5.0 pounds per ton of ore. The tests revealed that the oleic acid was as selective as the sodium oleate in producing acid-grade fluorspar concentrates ; however, the fluorspar recovery was somewhat lower with the oleic acid because it did not disperse, as well. The optimum amount of oleic acid was 0.48 pound per ton of ore. The summarized results of these tests are shown in table 4.
Additional laboratory batch flotation tests were made using the data obtained in determining the optimum quantities of reagent. The minus 65-mesh pulp was conditioned at about 40 percent solids in a mechanically agitated flotation cell for 5 minutes with 5.0 pounds each of sodium fluoride and calcium lignin sulfonate per ton of ore for dispersion of pulp and retardation of gangue minerals. Sodium oleate, 0.5 pound per ton of ore, was then added as a collector; conditioning was continued for another 5 minutes. The rougher concentrate was floated and refloated (cleaned) six times to remove gangue minerals. A concentrate analyzing 97.8 percent CaF2 and accounting for a fluorspar recovery of 84,5 percent was obtained. The results of a selected test are presented in tables 5 and 6.
Based on the results of the laboratory batch tests, a continuous pilot plant with a capacity of about 150 pounds of dry feed per hour was assembled. The process included grinding, classification, conditioning, and flotation, as shown by figure 1.
The ore was reduced by jaw and roll crushers in closed circuit to minus 3/8 inch and stored in a bin. From the bin it was transferred by a constant-weight feeder to a rod mill operated at 60 percent solids. The rod mill operated in closed circuit with a vibrating screen to grind the ore to minus 65 mesh. The screen undersize (minus 65-mesh) passed to a hydroseparator for
removal of colloidal slimes. The hydroseparator overflow represented about 1.5 percent of the weight of the ore and a loss of less than 1 percent of the total fluorspar. The hydroseparator underflow, at about 40 percent solids, passed to a conditioner where sodium fluoride and calcium lignin sulfonate were added. The discharge from the first conditioner flowed to a second conditioner where oleic acid was added as the fluorspar collector. A retention time of about 9 minutes in each conditioner gave satisfactory results. The conditioned pulp then flowed to a bank of three rougher flotation cells where a rougher concentrate was floated. The rougher tailing flowed to a single cell operating as a scavenger to recover additional fluorspar. The froth from this cell was recycled to the last rougher cell; the tails flowed to waste. The rougher concentrate was cleaned nine times, and the middlings were circulated back to the first cleaner where they were removed and thickened in a bank of three hydrocyclones (parallel). The underflow from the hydrocyclones was sent to the first conditioner, and the overflow went to waste. An emulsion-type collector (made up of 17.7 parts oleic acid 1.3 parts sodium oleate, and 361.0 parts water) was added to the second rougher cell to aid the flotation of the fluorspar.
The summarized results of a continuous flotation test are given in tables 7 and 8. The final fluorspar concentrate analyzed 96.4 percent CaF2 , a recovery of 90.0 percent of the total fluorspar in the ore. About 7 percent of the fluorspar was lost in the overflow from the cyclones.
The fluorspar concentrate was slightly below the specifications for acid-grade fluorspar; however, it meets all specifications for high-quality ceramic-grade fluorspar. It was possible to obtain an acid-grade fluorspar by introducing additional cleaners into the circuit; however, there was some sacrifice in recovery.
The laboratory batch and continuous pilot plant flotation tests demonstrated that the sodium fluoride-calcium lignin sulfonate-fatty acid method for selective flotation of fluorspar from complex calcareous fluorspar ore is an effective and practical means of producing high-grade fluorspar concentrates.
The flotation of the fluorspar in a continuous test in which the middlings were removed from the circuit, thickened, and returned for further conditioning produced a fluorspar concentrate analyzing 96.4 percent calcium fluoride, a recovery of 90.0 percent of the fluorspar in the ore. The fluorspar concentrate produced from a deposit near Cave-in Rock, III., meets all specifications for high-quality ceramic-grade fluorspar.
galena mining in ghana - binq mining
A profile of Mining in Africa with directories of companies, people, industry note that Africa's contribution to the world's major metals (copper, lead and zinc) is less South Africa, Ghana, Zimbabwe, Tanzania, Zambia and the DRC dominate
Water pollution at gold mines in Ghana. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 13. 5. Elevated blood lead levels among children living
Central Ashanti gold mine is located approximately 57km south-west of Obuasi and 195km north-west of Ghana's capital Accra. nearly three percent pyrite, minor arsenopyrite and lesser amounts of sphalerite, chalcopyrite, galena and rutile.
4 Jun 2008 study conducted on the effects of mining activitieson Obuasi and its are acidic, falling outside Ghana's Environmental Protection Agency and World Health was prepared to lead the debate on the kind of mining that would
10 Oct 2012 In an effort to ensure that resources from mining have the greatest As the World Bank's Country Director in Ghana, Yusupha B. Crookes, put it in project and lead mining specialist with the Bank's Oil, Gas and Mining unit.
lead zinc separation by flotation - froth flotation (sulphide & oxide) - metallurgist & mineral processing engineer
Sphalerite generally does not float well with xanthate collectors especially short xanthate collectors. So to be able to float sphalerite, it should be activated by a metal at a certain PH. Some of the metals that can do the job are copper and lead. At a pH between 8-9 lead can activate sphalerite to cause it flotation. So the reason is during the flotation of the two there may be lead re dissolution from galena which may activate sphalerite to cause it flotation. It may best if you are doing bulk flotation of the two.
Some of Pb-Zn beneficiation plants are having minerals in combination of Galena, sphalerite, Pyrite, Pyrotite, silica gangue, graphitic carbon, mica schist and traces of sulphide forms of silver, cadmium, bismuth and etc.
The depressed sphalerite in galena floatation is activated by Cuso4 solution in zinc conditioner while other minerals like pyrite, pyrohotite , silica are depressed with NaCN, Nasio3, lime for increasing pH.
To enhance selectivity, a combination of de-activator (usually sodium cyanide) and depressant (usually zinc sulphate) is used. The de-activator "leaches" the activating ions from the surface while the depressant essentially inhibits the activation reaction (Cu(II) + ZnS =>CuS + Zn(II)).
Yes your question is good. Why to float Galena first. It is the requirement of smelters. You can float both as bulk and feed for pyrometallurgical operations. But for hydrometallurgy sphalerite is roasted and leached. Hence separate floatation. It is like taking starters before drink.
While so-called natural flotation does occur to some extent for most sulphides, by and large, there will be no significant flotation unless you add a collector (molybdenite the classic exception) and for a select group of sulphide minerals (sphalerite/marmatite the best known but also includes antimonite, pyrite, pyrrhotite and even pentlandite) flotation is significantly enhanced after activation.
The inability of sphalerite to readily float without activation is an important separation tool in the selective recovery of sulphide minerals and has been routinely exploited industrially since selective flotation was developed in the early 20th century (Broken Hill).
In passing the comments about gravity recovery techniques are worth pursing as well as noting that galena is over ground in milling circuits (SG effect in the classifiers - a marketing guy once told me that lead flotation concentrates were not coarser than 25 microns) and the use of screens is also worth examining.
It is easier (proven) to selectively float galena by depressing sphalerite than the other way around if you are trying to make two concentrates. Bulk (or combined Pb/Zn) concentrate is the quickest way into bankruptcy because of heavy penalties imposed by smelters which becomes a real issue if your concentrate happens to contain Gold and Silver. As he says this flotation regime is probably the best understood of all because it was the origin of flotation.
Really we (Mineral processing community as a whole) took too long time to explore the principle of fastest settling rate of particle falls in between 5 to 7 % solids slurry density. This was materialised in High rate thickeners (HRT) by providing auto dilutionin feed well design very lately.
The conventional approach for Pb-Zn sulphide flotation arises from the chemistry of the mineral system: general chemistry, electrochemistry, and surface chemistry. What has been done over the years (since its discovery around a century ago) is the refinement of the approach mostly in terms of costs. The day that a multi-billion dollar Pb-Zn sulphide resource is discovered and found not amenable to the conventional approach is the day on which focused and intensive research for an innovative approach will be initiated.
The need for more efficient thickening schemes, aiming at decreasing the specific thickening area (m2 per tonne per hour), has arisen from the drive to reduce costs and stay in business. As labour is a component of the operating costs of a mine-mill facility, the obvious efficiency driven solution has been to increase throughput via larger equipment along with inclusion of instrumentation and process control to reduce the intensity of human intervention per tonne per hour processed. Since the capital costs of a thickener (and cost of the building surrounding it or roof above it) are essentially dictated by the settling rate of the material, the obvious efficiency driven solutions have been to use chemical principles giving rise to the use of flocculants (along with pH control) and physical principles giving rise to the high efficiency thickener design which dilutes the feed material to the pulp density yielding the highest settling rate.
In short - innovation in mineral processing has arisen from challenges encountered and needs related to the development of mineral resources. In many cases, innovative solutions developed for a specific mine-mill site were largely ignored by the industry as a whole until driven to react to changing market conditions (e.g. column flotation).
Your remarks about the perception that innovation is so slow in mineral processing were valuable ones and deserved to be written and read. I believe that we need to go further than just mention the slowness - we need to find a convincing way for mine executives to allow innovation take flight when it is appropriate.
As for Caribou - it is a decent deposit. I believe a technological alternative to ultrafine grinding for lead zinc flotation and separation was investigated decades ago when Brunswick Mining & Smelting and Heath Steele Mines were struggling with flotation selectivity at regrind P80s of 35 um. I don't recall all the details but it was hydrometallurgically based and had been developed at the Research & Productivity Council in New Brunswick. Whether it would be cost effective (capital and operating) is a question not easily answered. I believe the Imperial Smelting Process can take a bulk Pb-Zn concentrate. However, facilities using this process were far away from New Brunswick and shipping costs become significant. At the end of the day, and no matter how old or new a technology is, costs remain the determining factor.
In response to the comment about the lack of adoption of new technologies by our industry, it is fair to say that this lies mainly at the feet of the bankers and financiers. Unfortunately for the technologists, mining projects are principally about making money, and since the company proposing the project generally doesn't have the money, or chooses to share the 'risk' by involving other investors, risk comes into play.
Bankers and investors are not willing to lose money, indeed they expect to make money, so during the inevitable due diligence, the nature of risk is a major issue and how it may be mitigated is a major theme.
Unless an 'innovative' flow sheet has been proven on the bench scale, pilot plant scale (continuous, with recycle streams and intermediate product management as well as reliable measure of metal recoveries, product grades as a function of feed grade are shown) and then on a larger continuous scale (demonstration plant where other issues such as engineering, design, materials of construction, safety and environmental) are convincingly demonstrated then it is unlikely that money from these sources will be attracted.
This is particularly the case for 'novel' hydrometallurgical flow sheets, where each scale of test work can introduce some real surprises that need to be resolved such as the formation of intermediate species, precipitates, etc. since our knowledge of the chemistry, kinetics and thermodynamics of these systems is so limited outside the copper and gold systems.
One of the best examples of the development of 'innovative' hydrometallurgical was the 'borate' leaching/EW flow sheet undertaken by Doe Run to replace lead and zinc smelting requiring a significant amount of research, testing, time (many years) and money (deep pockets).
Some other risk factors about innovative processes are the incidental higher inaccuracy on capital and operating costs estimates, schedule (time to build, commission and reach name plate throughput), design criteria on piping and material selection for long life, and scale-up parameters from laboratory/pilot plant to full plant. Also, where one can find experienced operators with in depth knowledge of the process it-self and the equipment used?
Incidentally, a likely consequence of the poor track record of hydrometallurgical processes build in the last decade to reach their business objectives (at the costs and within the schedule initially estimated) will be much higher hurdles for any innovative process in the mining and metallurgical industry to be financed by banks and investors.
What if the content of sphalerite is much higher than galena like say ore with 2% galena and 30% sphalerite? Should we still have to depress sphalerite first? What do you think is the best way to treat this ore? It also contains Cu 0.5%, Au 9.8 ppm, Ag 17 ppm.
The onward technology for desulphurisation or oxidation of spharalite concentrate was with fluidised bed calciner at 950-1100 Degree Celsius. The increase in galena content in zinc concentrate and in combination of spharalite, pyrite and pyrohotite (Pb-Zn-Fe) results in the formation of hard nodule which were un-reactive in the above temperature range and certainly these modules cant be in fluidised state because of higher size and higher density.
I see that there is no need to reiterate the mineralogical, metallurgical or surface chemistry underpinning differential flotation of galena and sphalerite, but you may be interested to know that the traditional galena float with sphalerite depression followed by sphalerite activation and flotation is not the only approach in use. The Zinkgruvan mine in Sweden uses bulk flotation of galena and sphalerite followed by depression of the sphalerite to produce separate galena and sphalerite concentrates; however, it is still a variation on the traditional approach in that after the bulk float, it is still sphalerite that is depressed first.
The relative galena and sphalerite contents do not change the surface chemistry and tend only to have an impact in cases where the galena content is higher than the sphalerite content, e.g. at Cannington where if the lead feed grade relative to zinc is too high, then too much of the faster floating zinc with lower iron content is recovered to the lead concentrate, making zinc target concentrate grade difficult to achieve.
In the case of your project, differential flotation to produce separate copper, lead and zinc concentrates would be the likely first choice of flow sheet. Provided that you can make saleable grades in all three products at reasonable recoveries, payment terms and markets are better for individual concentrates than bulk concentrates, payable gold and silver in copper concentrate is better than the other concentrates and payable silver in lead concentrate is better than in zinc concentrate.
If investigating flash flotation routine for copper, you should play close attention to the REDOX at which the collector is added. Even with a thionocarbamate collector, there is a threshold REDOX below which it will not interact with the copper minerals. This will be important for scalability from the laboratory to commercial plant.
With respect to floating sphalerite first than galena, success will likely be dependent on the pulp chemistry and the susceptibility of the galena to oxidation. The more susceptible to oxidation the galena in the feed is, the more difficult it will be to recover it as the last valuable mineral.
Your quoting Zinc gruvan, Sweden process, bulk floatation followed by galena - sphalerite conventional floatation was good example of simple and lucid mineral processing practise. A different approach, interesting!
Please let me know the leverages or advantages achieved in reduction in OPEX ( reagents and power consumptions) as well as capex ( i mean reduction in cell volume or residence time, shed area if any) if possible.
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beneficiation - an overview | sciencedirect topics
Thermal beneficiation is the use of combustion to reduce the level of carbon in the ash. Thermal beneficiation also eliminates ammonia issues and can improve fineness and uniformity. Successful thermal beneficiation technologies have been commercially deployed since 1999 (Keppeler, 2001). This technology produces more than a million tons of marketable fly ash per year in the eastern United States. There are two technologies that can be considered proven: the first is PMI's Carbon Burnout (CBO) system, based on dense phase fluidized bed combustion; and the second is SEFA Group's STAR technology, based on dilute or entrained fluidized bed combustion.
The ability of thermal beneficiation to improve ash quality is truly impressive. It is a proven, highly flexible technology that can operate on a variety of ash types with a very wide range of carbon concentrations and sizes. It produces an ash that is low or even free of carbon. It also eliminates ammonia from fly ashes impacted by nitrous oxide controls or opacity treatments. The process may improve fineness by eliminating coarse carbon and liberating ash trapped within.
Thermal beneficiation is a combustion process and may require additional air emission permitting. If not integrated into the power plant, it will also require its own emission control system. It is by far the most expensive of all the technologies considered. A facility can cost tens of millions of dollars, which suggests that it would be more attractive for larger power plants with access to large and stable markets. The construction of a thermal beneficiation facility may require significant plant modifications and systems integration; however, it does not specifically target ash fineness and uniformity.
Dry beneficiation has two important advantagessaving water, a valuable resource, and no tailings pond and subsequently, no leaching of the trace/toxic elements into ground water. In dry beneficiation of coal, coal and mineral matter are separated based on differences in their physical properties such as density, shape, size, luster, magnetic susceptibilities, frictional coefficient, and electrical conductivity . Dry beneficiation gives a clean coal as well as reduces some of the polluting elements associated with minerals. It cannot remove the inorganic matter in coal present as salts resulting from the marine environment during coalification. Azimi etal.  evaluated the performance of air dense mediumfluidized bed separator in removing trace elements, such as Hg, As, Se, Pb, Ag, Ba, Cu, Ni, Sb, Co, Mn, and Be. Their study revealed the association of Pb, Ag, Ba, Cu, Mn, and Be with ash-forming minerals. Elements such as As, Se, and Sb showed some organic bonding. High rejection of Hg was achieved through dry beneficiation of coal where Hg is mostly associated with pyrites.
Beneficiation of copper ores is done almost exclusively by selective froth flotation. Flotation entails first attaching fine copper mineral particles to bubbles rising through an orewater pulp and, second, collecting the copper minerals at the top of the pulp as a briefly stable mineralwaterair froth. Noncopper minerals do not attach to the rising bubbles; they are discarded as tailings. The selectivity of the process is controlled by chemical reagents added to the pulp. The process is continuous and it is done on a large scale103 to 105 tonnes of ore feed per day.
Beneficiation is begun with crushing and wet-grinding the ore to typically 10100m. This ensures that the copper mineral grains are for the most part liberated from the worthless minerals. This comminution is carried out with gyratory crushers and rotary grinding mills. The grinding is usually done with hard ore pieces or hard steel balls, sometimes both. The product of crushing and grinding is a waterparticle pulp, comprising 35% solids.
Flotation is done immediately after grindingin fact, some flotation reagents are added to the grinding mills to ensure good mixing and a lengthy conditioning period. The flotation is done in large (10100m3) cells whose principal functions are to provide: clouds of air bubbles to which the copper minerals of the pulp attach; a means of overflowing the resulting bubblecopper mineral froth; and a means of underflowing the unfloated material into the next cell or to the waste tailings area.
Selective attachment of the copper minerals to the rising air bubbles is obtained by coating the particles with a monolayer of collector molecules. These molecules usually have a sulfur atom at one end and a hydrophobic hydrocarbon tail at the other (e.g., potassium amyl xanthate). Other important reagents are: (i) frothers (usually long-chain alcohols) which give a strong but temporary froth; and (ii) depressants (e.g., CaO, NaCN), which prevent noncopper minerals from floating.
Beneficiation of complex base metal sulfide ores is based on selective production of individual clean concentrates of copper, zinc, and lead. Sphalerite flotation through copper activation becomes complicated when other minerals such as pyrite can get inadvertently activated.
Adsorption density of cells of P. polymyxa was found to significantly higher on pyrite than on sphalerite irrespective of pH. Adsorption on sphalerite was the highest in acidic pH regions only (26), beyond which cell adsorption decreased steeply.
Flocculationdispersion behavior of pyrite and sphalerite was seen to be influenced by interaction with bacterial cells and their metabolic products as a function of pH, cell density, and bioreagent concentrations. For example, more than 90% of pyrite particles were observed to be flocculated and settled at pH 89 in the presence of bacterial cells, while sphalerite was preferentially dispersed. Similarly, interaction with EBP isolated from metabolites promoted selective flocculation of pyrite and dispersion of sphalerite. On the other hand, interaction with ECP was not very effective in separation of pyrite from sphalerite because the selectivity ratio was very poor. Pyritesphalerite separation can be effectively achieved through selective bioflocculation of pyrite and dispersion of sphalerite using either bacterial cells or bioproteins.
Pyrite can also be selectively depressed through bioflotation after bacterial conditioning. Flotation tests using 1:1 mixtures of pyrite and sphalerite indicated that prior bacterial interaction followed by xanthate conditioning and copper activation resulted in preferential flotation of only sphalerite, while pyrite was depressed.
Pyrite could also be similarly removed from galena because differential adsorption and surface chemical behavior of P. polymyxa cells as well as proteins and polysaccharides were also observed on pyrite and galena as well. Selective bioflocculation in the presence of either bacterial cells or extracellular proteins could selectively flocculate pyrite from pyritegalena mixtures. Galena was also found to be selectively flocculated after interaction with exopolysaccharides. Similarly selective flotation of galena along with efficient pyrite depression could be attained after interaction with extracellular proteins.
A. ferrooxidans have been used to demonstrate selective pyrite depression from a low-grade leadzinc ore. Both sphalerite recovery and zinc grade in the floated sphalerite concentrate were enhanced by bacterial cells in the absence of conventionally used cyanides .
Ore beneficiation refers to the selection and collection of higher-grade ore fragments or rejection of lower-grade fragments from ROM ore. The upgraded ore will have a higher grade, and therefore require a smaller-scale processing plant, perhaps with different technology, compared to ROM ore. Ore beneficiation is only worthwhile if the majority of the uranium is retained and the majority of the mass is rejected.
In the earliest times of uranium mining hand sorting was employed, based on visual appearance or simple gamma scanning. In more recent times, mechanical sorting based on physical, mineralogical, or radiometric characteristics are employed.
At the Cluff Lake uranium mine and mill in Canada, in phase 1 operations high-grade ore was fed to a gravity concentration plant with jigs and vibrating tables. The concentrate averaged over 30% U (Schnell and Corpus, 2000). The gravity concentrator rejects were stored and retreated later (see Section 6.5.1).
Flotation (sometimes spelled floatation) separates mineral grains that respond differently when air bubbles are forced through a suspension in water with chemical additives, causing certain minerals to rise with the froth, from which they can be collected in a concentrated form. It was used in some Canadian uranium mines in the 1980s (eg, Muthuswami et al., 1983) and has been investigated in India (eg, Singh et al., 2001). It is used at the Olympic Dam copperuraniumgold mine to separate sulfidic copper-bearing minerals (Alexander and Wigley, 2003) and only incidentally for uranium minerals; uranium is recovered from the reject stream of the flotation circuit. Some recent investigation of the technique for application to multimetallic ores containing uranium is described by Kurkov and Shatalov (2010).
Shatalov et al. (2001) report that the automated radiometric ore separating was widely adopted in former Eastern Bloc countries starting in 1955. Variants and improvements up to 2000 are discussed. Radiometric sorting has been used in recent times in Ukraine (OECD-NEA/IAEA, 2014, p. 426). Radiometric sorting trials at Ranger mine in Australia and Rssing in Namibia are reported by Schnell (2014), who also comments that [G]ravity separation has been applied to uranium ores in the past with some success, but this was associated with radiation issues (cf. Section 6.5.1, where it was used with very high-grade ore). Lund et al. (2007) mention radiometric sorters in use historically in Australia and South Africa.
A form of upgrading by physical means, rotary scrubbers, at the Langer Heinrich uranium mine in Namibia is described by Marsh (2014), who states that the rejected oversize Barren Solids will contain 4050% of the solids mass but only 510% of the uranium in the ROM feed.
When the price of uranium was high in the mid-2000s, there was relatively low interest in upgrading, rather treatment of low-grade ore was considered feasible and pursued (cf. Lund et al., 2007). However, with lower prices since 2010, more experimentation is being reported. For example, ablation can remove uranium-rich mineral crusts from some sandstone ores (Coates et al., 2014) with some ore types; Scriven (2014) cites the ablation technique resulting in rejection of 9095% of the unprocessed ore mass but with a loss of only 510% of the uranium originally present. Another process under development (Becker et al., 2015) is for certain low-grade, surficial ores. It can reportedly increase the ore grade by a factor of 30 times or more without the use of chemicals, producing an inert waste and providing a leach feed suitable for acid leaching, although details of this second process were not yet released at that time. To date, neither has been undertaken at a commercial scale.
In tin beneficiation, the main new technology being adopted is the high-gravity concentrator, examples being the Kelsey jig, Falcon and Knelson concentrators, and the Mozley multigravity separator. Radical change in tin smelting and refining technology is not expected. In smelting, use of the TBRC and the Sirosmelt technologies will be more widely adopted, using fuming instead of reduction smelting for second-stage processing. Economies of scale are leading to the dominance of a few large smelters in countries such as China, Malaysia, and Bolivia. Refining technology will in general continue to rely on the same chemical principles, but will see greater adoption of automated technology such as the centrifuge for dross removal, and the vacuum process. Hydrometallurgical technologies may make an impact with developments in ion exchange, solvent extraction, and biooxidation and reduction.
Microbially induced mineral beneficiation involves three strategies, namely, selective bioleaching of the undesirable mineral from an ore or concentrate, selective flotation of the mineral, or selective dispersion/flocculation. Such microbially induced beneficiation will find applications in a number of areas such as:
Besides bioleaching using Acidithiobacillus bacteria and bioremediation using SRB, many mining organisms which inhabit ore deposits find applications in mineral beneficiation such as microbially induced flotation and flocculation. Acidithiobacillus spp can be used also to bring about microbially induced flotation and flocculation of minerals. Heterotrophic bacteria such as Paenibacillus polymyxa and Bacillus subtilis, yeasts such as Saccharomyces cerevisiae, and SRB such as D. desulfuricans can be used to bring about surface chemical changes on minerals, Principles and examples of microbially induced mineral beneficiation processes are illustrated in Chapter 10, Microbially Induced Mineral Beneficiation. Experimental protocols for such applications are illustrated below: 
Fully grown bacterial culture is centrifuged at 10,000g for 15min at 5C. The supernatant is decanted and filtered through sterile Millipore (0.2m) filter paper to remove all insoluble materials and any remaining bacterial cells.
CFE of a fully grown culture contains different bioreagents such as proteins, polysaccharides along with trace amount of other constituents. Proteins and polysaccharides can be isolated and used as flotation and flocculation reagents.
A suitable volume of a fully grown culture is initially centrifuged, and the supernatant filtered through a sterile millipore (0.2m) filter paper. Analytical reagent grade, extra pure, and fine powdered ammonium sulfate is added slowly to a saturation level of 65% with constant shaking at 4C. The solution is allowed to stay under refrigeration for 12h at 4C. The precipitated protein is dissolved in a minimum volume of 0.1M Tris hydrochloride buffer of pH 7 and dialyzed against the same buffer for over 18h at 4C. The precipitate formed during dialysis is removed through centrifugation and disposed. The clear supernatant is lyophilized, and the resultant solids weighed, and kept at 4C for further use.
An actively grown bacterial culture is harvested by centrifugation and dissolved in lysis buffer (10mM Tris-HCl (pH 8.0), 0.1M NaCl, 1mM EDTA (pH 8.0), 5% (v/v) Triton X-l00). The suspension is sonicated and centrifuged at 10,000g for 5min. The supernatant is subjected to 65% ammonium sulfate precipitation, the solution kept in a refrigerator for 1012h, and centrifuged. The precipitated protein is dialyzed with trisbuffer at neutral pH range and then precipitated with acetone. This protein precipitate is dissolved in trisbuffer and SDS-PAGE is carried out.
A suitable volume of fully grown batch culture is centrifuged to remove bacterial cells. The supernatant containing the extracellular polysaccharide (ECP) is filtered and lyophilized. The dehydrated sample is dissolved in 10mL of distilled water and cooled to <10C. Twenty milliliter of double-distilled ethanol is added to selectively precipitate ECP and purified. It is stored in a refrigerator for 8h at 4C. The precipitate is washed with double-distilled water. The ethanol precipitation procedure is repeated two to three times further, to purify the polysaccharide, and the solution dialyzed with double-distilled water. After dialysis, ECP is stored at low temperature (4C). The concentration of ECP is determined by the phenolsulfuric acid method .
The spectrophotometer is switched on and the wavelength adjusted to 280nm. Absorbance is calibrated to zero with buffer and the absorbance of the protein solution measured. The wavelength is adjusted to 260nm and absorbance calibrated to zero with buffer. The absorbance of the protein solution is measured.
One vial with 5mg of BSA is taken and 1mL of distilled water added (5mg/mL), 0.2mL is pipetted out into an Eppendorf tube and 0.8mL of distilled water added (1mg/mL) (Working standard). Standards can be prepared by taking 20, 40, 60, 80, 100L of BSA working standard in test tubes and made to 200L with distilled water. Two mL of Bradford reagent is added to all including test solutions and thoroughly mixed. After 10min, readings are measured at 595nm.
4 % Phenol: Conc. H2SO4, stock standard solution: 100mg of glucose was dissolved in 100mL of double-distilled water (100g/0.1mL). Working standard solution: 10mL of stock solution was made to 100mL with double-distilled water (10g/0.1mL).
Standards are prepared by taking 20, 40, 60, 80, 100L of glucose working standard in test tubes and made to 1mL with double-distilled water. To all the tubes including unknowns, 2mL of 4% phenol and 5mL of concentrated sulfuric acid are added and mixed thoroughly. Readings are taken at 490nm.
Coal is a sedimentary rock that occurs in seams bounded by layers of rock. The generation of waste is unavoidable during coal extraction and beneficiation. Mine waste or spoils are materials that are moved from its in situ location during the mining process but are not processed to obtain the final product. Dealing with mine waste is a major part of surface mining methods, where all of the rock above the coal seam (overburden and sometimes interburden) must be removed to expose the coal seam. This is done in a systematic fashion of digging pits with most of the overburden waste being cast from above the coal to be extracted into an adjacent pit from which the coal has already been extracted. Minimizing the handling of overburden waste is one of the keys to economic success in surface coal mining. Various techniques, such as cast blasting, are used to achieve this objective.
If underground mining methods are used, the amount of out-of-seam material handled is much less than in surface mining, and minimizing that amount has multiple economic benefits as discussed in Chapter 11. When large amounts of out-of-seam material have to be removed for underground infrastructure such as ventilation overcasts and undercasts and conveyor belt transfer points, it can be gobbed or left underground in untraveled mine openings. However, most of the out-of-seam material extracted in underground mines is mixed with the coal and constitutes part of the run-of-mine (ROM) or raw coal product.
Because modern mechanized mining equipment does not distinguish between the coal seam and layers of rock that encapsulate it and because complete or full extraction of a mineable coal seam is generally the objective of any coal-mining operation, there will always be some level of out-of-seam dilution in the ROM product. In most cases, out-of-seam material extracted with the coal must be separated from the coal before shipment to satisfy customer quality requirements. This is accomplished with coal preparation plants that generate a clean coal product and a waste material referred to as coal refuse. Coal preparation plants utilize various mineral processing technologies that, with few exceptions, are slurry-based and involve the use of substantial quantities of water . The efficiency of these processing systems depends on the size of material being treated. Hence, raw coal must be classified into different size fractions leading to two coal refuse products on the output side: (1) coarse coal processing waste (CCPW) and (2) fine coal processing waste (FCPW). Generally, CCPW is material larger than 150m (100 mesh) in size . CCPW includes reject streams from jigs, heavy media vessels, and heavy media cyclones. FCPW includes reject streams from spirals, flotation columns and cells, desliming cyclones, and effluent streams of filter presses, screenbowl centrifuges, and other dewatering equipment. All FCPW streams are typically concentrated in a thickener whose output is a waste slurry.
Most extraction and beneficiation wastes from coal mining (i.e., mine spoils and coal refuse) are categorized as special wastes that are exempted from regulation by hazardous waste rules and laws (e.g., Subtitle C of the US Resource Conservation and Recovery Act). However, coal utilization generates another type of waste known as coal combustion residuals (CCRs), which are regulated to some degree (e.g., Subtitle D of the US Resource Conservation and Recovery Act). CCRs are categorized into four groups based on physical and/or chemical forms that derive from the combustion method and the emission control system used. A brief description of each group follows :
Bottom ash is a coarse, angular, gritty material with similar chemical composition to fly ash. It is too large to be carried up by the smokestack, so it collects in the bottom of the coal furnace. It comprises 12% of all CCRs.
Boiler slag is molten bottom ash that forms into pellets in the bottom of slag tap and cyclone type furnaces. It has a smooth glassy appearance after it is cooled with water. Boiler slag comprises 4% of all CCRs.
Flue gas desulfurization (FGD) material is residue from the sulfur dioxide emission scrubbing process. It can be a wet sludge consisting of calcium sulfite or calcium sulfate, or it can be a dry powdery material that is a mixture of sulfites and sulfates. FGD material comprises 24% of all CCRs.
A simple process of beneficiation has been selected, which will be low in capital cost. As the scheme is a simple one, the cost of operation and maintenance will be minimal. The process technology is so chosen that it should be able to meet the quality parameters laid down by consumers. The flow scheme is briefly described here:
The scheme of beneficiation indicated here is a simple and effective technique that does not take into consideration either small coal or fines. This simple scheme may be applicable both for consumption in the power sector and the cement industry. However, depending upon the raw coal characteristics and needs of the consumer, total washing may be needed, as in the case of coking coal (Fig. 9.6).
Commercial coal cleaning or beneficiation facilities are physical cleaning techniques to reduce the mineral matter and pyretic sulfur content. As a result, the product coal has a higher energy density and less variability (compared with feedstock coal) so that power plant efficiency and reliability are improved. A side benefit to these processes is that emissions of sulfur dioxide and other pollutants including mercury can be reduced. The efficiency of this removal depends on the cleaning process used, the type of coal, and the contaminant content of coal. Basic physical coal cleaning techniques have been commercial for over 50 years. The cleaning of coal takes place in water, in a dense medium, or in a dry medium.
Physical cleaning processes are based on either the specific gravity or the surface property differences between the coal and its impurities. Jigs, concentration tables, hydrocyclones, and froth flotation cells are common devices used in current physical coal cleaning facilities.
The removal efficiency ranged from 0% to 60% with 21% as average reduction. This efficiency is highly dependent on the type of coal and chloride content of the coal. Concerning other fuels, the cleaning of crude oil occurs mostly through the residue desulfurization (RDS). However, the content of Hg in crude oil is usually very low, and RDS is an inefficient method to even lower this content.
coal beneficiation - an overview | sciencedirect topics
Coal preparation, or beneficiation, is a series of operations that remove mineral matter (i.e., ash) from coal. Preparation relies on different mechanical operations (not discussed in detail here) to perform the separation, such as size reduction, size classification, cleaning, dewatering and drying, waste disposal, and pollution control. Coal preparation processes, which are physical processes, are designed mainly to provide ash removal, energy enhancement, and product standardization . Sulfur reduction is achieved because the ash material removed contains pyritic sulfur. Coal cleaning is used for moderate sulfur dioxide emissions control, as physical coal cleaning is not effective in removing organically bound sulfur. Chemical coal cleaning processes are being developed to remove the organic sulfur; however, these are not used on a commercial scale. An added benefit of coal cleaning is that several trace elements, including antimony, arsenic, cobalt, mercury, and selenium, are generally associated with pyritic sulfur in raw coal and they, too, are reduced through the cleaning process. As the inert material is removed, the volatile matter content, fixed carbon content, and heating value increase, thereby producing a higher quality coal. The moisture content, a result of residual water from the cleaning process, can also increase, which lowers the heating value, but this reduction is usually minimal and has little impact on coal quality. Coal cleaning does add additional cost to the coal price; however, among the several benefits of reducing the ash content are lower sulfur content; less ash to be disposed of; lower transportation costs, as more carbon and less ash is transported (coal cleaning is usually done at the mine and not the power plant); and increases in power plant peaking capacity, rated capacity, and availability . Developing circumstances are making coal cleaning more economical and a potential sulfur control technology and include :
Coal preparation, or beneficiation, is a series of operations that remove mineral matter (i.e., ash) from coal. Preparation relies on different mechanical operations, which will not be discussed in detail, to perform the separation, such as size reduction, size classification, cleaning, dewatering and drying, waste disposal, and pollution control. Coal preparation processes, which are physical processes, are designed mainly to provide ash removal, energy enhancement, and product standardization (Elliot, 1989). Sulfur reduction is achieved because the ash material removed contains pyritic sulfur. Coal cleaning is used for moderate sulfur dioxide emissions control as physical coal cleaning is not effective in removing organically-bound sulfur. Chemical coal cleaning processes are being developed to remove the organic sulfur, but these are not used on a commercial scale. An added benefit of coal cleaning is that several trace elements, including antimony, arsenic, cobalt, mercury, and selenium, are generally associated with pyritic sulfur in raw coal, and they too are reduced through the cleaning process. As the inert material is removed, the volatile matter content, fixed carbon content, and heating value increase, thereby producing a higher-quality coal. The moisture content, from residual water from the cleaning process, can also increase; this lowers the heating value, but it is usually minimal so as to have little impact on coal quality. Coal cleaning does add additional cost to the coal price; however, there are several benefits to reducing the ash content which includes lower sulfur content, less ash to be disposed, lower transportation costs since more carbon and less ash is transported (since coal cleaning is usually done at the mine and not the power plant), and increases in power plant peaking capacity, rated capacity, and availability (Harrison, 2003). Developing circumstances are making coal cleaning more economical and a potential sulfur control technology, and they include the following (Elliot, 1989):
Coal beneficiation, or coal preparation as it is also termed, refers to the processes through which inorganic impurities are separated from raw mined coal, thereby providing improved combustion characteristics to the fuel produced. The separation processes used are primarily based on exploiting the physical differences between the organic (i.e., coal) and inorganic (i.e., ash) components. Given the low unit value of coal, it is imperative for these separation processes to be both efficient and cost effective. The most commonly used processes are jig washing, density separation, sizing, and froth flotation. Typical configurations divide the run of mine coal into size fractions and utilize different separation processes for each size fraction (Luttrell, Barbee, & Stanley, 2003).
Density separation exploits the differences in density between the organic and inorganic components found in mined coal. As previously described, coal typically is comprised of an assemblage of macerals and inorganic material. Macerals containing primarily organic matter generally have a density of <1.4g/cm3, and as the amount of ash associated with the macerals increases, the density of the particles also increases, because the primary composition of ash associated with coal is essentially the weathered products of quartz (density 2.65g/cm3). Thus particles in the density range of 1.61.8g/cm3 have a higher ash content. Pyrite (FeS2), another commonly associated mineral, has a much higher density of 5.0g/cm3. Given the difference in density between the desired material (coal) and undesired material (ash and pyrite), density separation can be an efficient approach for producing low-ash coal, provided the high-ash content particles are liberated from the low-ash particles.
Density separation processes employed in coal preparation are typically performed in a medium suspension of fine ground (45m) magnetite (Fe3O4) dispersed in water. Magnetite is added to the suspension to maintain the desired medium density. For example, if the medium density is maintained at a density of 1.45g/cm3, all particles with lower density will float to the top of a separation vessel while the higher density particles sink. The float- and- sink products are separately removed and washed on an appropriately sized screen. Magnetite particles are recovered from washwater with magnetic separators and recycled back into the process. Dense medium separation of coarse particles (>50mm) is typically accomplished in vessels, while intermediate-size particles (501mm) are treated in cyclones. The operating principles of dense medium cyclones are essentially the same as those of conventional cyclone sizing processes; however, with dense medium cyclones, the fluid density can be increased to the desired separation density by the addition of magnetite. Jig washing employs similar separation principles, but rather than adjusting the medium density, particles are separated in a water medium that is pulsated pneumatically or hydraulically. The pulsation of the jigging motion stratifies particles based on density. Lighter particles migrate to the top of the particle bed, and denser particles migrate to the bottom, thus producing a separation based on particle density. The choice between using jigging or dense medium separation is generally made depending on the amount of near-gravity material, or the amount of material within 0.1 specific gravity units of the desired separation specific gravity. With 07% of the feed near gravity, almost any separation process will work effectively, though jigs are commonly employed under these conditions. With 7%10% near-gravity material, jigs operate with decreased efficiency, and so dense medium separation processes are appropriate. With >10% near gravity material, dense medium separation processes have application, but the process needs to be more closely controlled. With >25% near-gravity material, dense medium separation is very difficult, but can still find application in limited situations (Wills, 2006).
Size separation processes are the simplest to implement. These processes exploit distinct difference in sizes between coal and ash particles. If, for example, the coal to be processed is coarse while the ash is fine, then efficient separation can be achieved by a simple screening at the appropriate size. The same is true for the converse (i.e., coarse ash and fine coal). As this approach is so simple, it is used wherever possible; however, it is dependent upon the size distribution of the coal and ash particles. When particles are too small to screen efficiently, the size difference between coal and ash particles is exploited using classifying cyclones.
For fine particles (<150m), dense medium separation and sizing do not produce efficient separations. These particles are separated by flotation, which exploits differences in particle hydrophobicity. Most bituminous and higher-rank coals have some natural hydrophobic properties, while ash particles are hydrophilic. Coal hydrophobicity can be increased by selective adsorption of small quantities (100200g/tonne) of nonpolar collectors, such as diesel or fuel oil. The coal/ash suspension (1015% solids w/w) is agitated in a tank or cell, and air bubbles are introduced at the bottom of the cell. Surface-active agents, such as short-chain alcohols, are typically added to increase bubble surface area by reducing surface tension at the air/liquid interface, thus producing copious amounts of small air bubbles. Hydrophobic coal particles adsorb onto the rising air bubble and are transported to the top of the cell, where they coalesce and form a stable froth layer. The froth layer overflows the cell or is removed by mechanical scrapers while ash particles remain in suspension and are withdrawn from the cell. Flotation cells used in coal preparation are either mechanically agitated or column flotation cells with no agitator.
Economics of coal beneficiation using oil agglomeration approach is very sensitive to the quality and price of oil used. As it was shown in (23), the cost of oil comprised about 31 percent of the total product cost of the beneficiation plant or $14 per ton of coal processed. Total process capital and fixed costs comprised about 8 percent and cost of electricity for coal grinding about 2 percent of the total product cost. (No. 2 fuel oil at a rate of 10 percent on the dry-ash-free-coal weight and oil price of $200/t were considered. Energy consumption for coal grinding of 30 kWhr/t of feed at c2.75/kWh and coal cost of $26/t were assumed).
It is obvious from above that at the oil and coal prices presented, the oil agglomeration approach considered for coal beneficiation is uneconomical, unless oil consumption is drastically reduced, for example, by economic recovery of oil from beneficiated coal, or oil cost may be mostly written off as in the case of coal beneficiation integrated with a process where oil would be utilized together with the coal. Also, reduction of coal cost would substantially improve the economics.
On the other hand, as it was discussed earlier, direct liquefaction (hydrogenation) of a coal with reduced ash content may substantially increase liquid product yields. Preliminary calculations have shown, that for Canadian conditions, the expected overall liquid product cost of an integrated direct coal liquefaction plant producing 25,000 bbl/day of syncrude would be in the $4050/bbl range for lignite at a cost of $10-15/t having ash content of about 8-10 percent on dry coal weight.
The hard coal beneficiation process in mechanical preparation plants generates coarse, small or fines rejects and coal tailings slurries. The tailings are the finest grain size, with the majority below ~0.25mm, whereby material sized below 0.035mm makes up to 60% share in the slurry composition. Depending on the quality parameters (ash and sulphur content, calorific value, etc.), such slurries can be transferred as an ingredient to energy mixtures, or are dumped in earth settlers of individual mines. Most slurries to date have been collected in settlers, as there were no customers interested in buying them at the time they were produced. Dumped slurries were therefore treated as waste from coal preparation processes. Most of this waste is actually a potentially viable energy source. For this reason, in recent years, the interest in combustion options has increased as other fossil energy sources have increased in delivered cost. There is also interest in using coal tailings in construction products and engineering projects.
Some coal tailings are transferred to preparation plants for recovery of coal contained in the waste. Currently about 9% of generated waste is utilized in this way. The residue after the recovery of coal is re-dumped or used, for example in hydraulic backfilling or the building construction materials industry. Energy generation from coal tailings is covered in more detail in the sub-section below.
Coal tailings are quite commonly used in the manufacture of construction products for the building industry as an essential raw material for obtaining slate aggregate, i.e., a lightweight building construction aggregate used in the manufacture of lightweight concrete, as well as an essential raw material or component for the production of various building construction elements, such as bricks or roofing tiles. Currently, only about 0.5% of generated waste is utilized in this way. The waste is also added to the charge in the production of cement, in order to adjust the main module of cement clinkers. Coal tailings may also be useful for the production of refractory materials, but only if they have a high content of Al2O3.
Attempts have been made to recover metal concentrates from coal tailings, including aluminium, iron, titanium, germanium and gallium. Fine coal waste can also, after mixing with a compound fertilizer and peat, be used for biological reclamation and restoration of the fertility of devastated land, or reclamation of soil.
Flotation tailings wastes, a specific type of tailings, have not yet found an industrial application due to a number of factors including significant thixotropy, high humidity and difficulties in transport. However, such wastes can be used as a material for filling abandoned workings in mines or to seal the surface stockpiles. Post-flotation wastes from beneficiation of coking coals with calorific value more than 5 000kJ/kg can be used as fuel for the production of building construction ceramics, and after further beneficiation as an additive to energy fuel.
As no commercial coal beneficiation is perfectly efficient, some indices are required to measure the efficiency of the process. The best way of indicating the efficiency of a density separation device is the distribution of the partition curve (Fig. 7.4), which was first proposed by Tromp (1937). This curve depends on the equipment used, the relative density or cut-off point and the size range of the feed coal. Various simpler measures of efficiency have been defined but none are as accurate in predicting the performance of a density separation device as the Tromp curve. It denotes the probability of a particle reporting with the floats to its specific gravity. The distribution numbers are marked on the vertical axis against the various specific gravity fractions shown on the horizontal axis. Thus, if the vertical axis has value x, then the corresponding value of the horizontal axis is denoted as dx. The partition density is denoted by d50 the distribution number is 50, in this condition the particle will have an equal chance of floating or sinking. The Tromp curve is nothing but an error curve, the steeper the curve, the most efficient is the separation. To measure the inclination of the curve, Terra introduced Ecart Probable (Ep) which is defined as,
When, Ep = 0, the curve becomes a straight vertical line at the specific gravity of separation the efficiency of separation will be 100%. The Ep value does not consider the tails of the Tromp curve that are above the distribution number 75 and below the distribution number 25. The larger tails in the Tromp curve result in lower yield at the desired ash.
The efficiency at any relative density is defined as (Sarkar and Das, 1978) the recovery % of clean coal (ash % of raw coal ash % of clean coal) divided by the recovery % of float coal (ash % of raw coal ash % of float coal).
It may be noted that the most widely accepted measure of the efficiency with which a cleaning device separates coal from impurities is referred to as probable error (Ep) and Ep/dp which is sometimes called the generalised probable error (Gottfried and Jacobsen, 1977).
Currently, the wet coal beneficiation process is the predominant method for coal upgrading. The wet beneficiation processes include heavy media separation, cyclone (water only), froth flotation, and spiral separation [23,24]. The use of these technologies depends on the particle size of the feed and the quality of the product required. The quality of product and the recovery from the wet method is generally better than those obtained from the dry beneficiation method . Slimes and acidic water generated from the wet process require tailings ponds. Dewatering of the washed coal may cause leaching out of pollutants, which in turn can cause ground water pollution if not managed properly. Wet cleaning is mostly used for metallurgical coals, whereas there is a general trend to use dry beneficiation for thermal coals.
As noted, one of the fundamental reasons for coal beneficiation is the reduction of ash yield and deleterious minerals and elements with an inorganic affinity. The partitioning of major, minor, and trace elements depends on the degree of liberation of the minerals, their inorganic versus organic association, and the specific gravity of the separation. Mineral matter occurring as discrete bands and lenses within the coal can often be removed easily, but that disseminated within the coal matrix or within the organic compounds of the macerals will be more difficult to remove by simple density separation and may require extensive (and expensive) grinding to beneficiate. In low rank coals, dissolved salts or inorganic elements incorporated within the organic compounds of the macerals are common.
An overview of analytical methods used to determine inorganics in coal is given by Huggins (2002), and mineral matter in coal is presented in Chapter 2 of this book. A common method for determining whether a mineral or element will partition during beneficiation is through analysis of the float/sink fractions for different size fractions (Querol et al., 2001). As stated in Huggins (2002), the higher the organic affinity, the more the element reports to light-specific gravity fractions, and hence, the more it is associated with the organic fraction of the coal. One would assume that these lighter fractions would be dominated by vitrain, but that is not always the case. Various studies (Zubovic, 1966; Gluskoter et al., 1977; Cavallaro et al., 1978; Fiene et al., 1978; and Kuhn et al., 1980) suggest that the organic affinity of many elements varies significantly from coal to coal. More direct methods of analyzing maceral separates or scanning electron microscopy will assist in characterizing this variability for specific macerals and minerals.
Mitchell and McCabe (1937), Helfinstine et al. (1971, 1974), Cavallaro et al. (1976), and, more recently, Mastalerz and Padgett (1999) studied the ash and sulfur partitioning of (generally) high-S Pennsylvanian Illinois Basin coals. Because of the fine nature of much of the pyrite and an organic association of about half of the total S, the S in the clean product was generally above 2%.
Finkelman (1994b) discussed the associations of the hazardous trace elements. His work was based both on his own research (Finkelman, 1981) and on comprehensive works by others (e.g., Gluskoter et al., 1977; Raask, 1985b; Eskenazy, 1989; and Swaine, 1990). Akers and Dospoy (1994) demonstrated the magnitude of element reduction through a number of coal beneficiation schemes and DeVito et al. (1994) examined the trends in a large collection of coal company data from the Illinois Basin and the Northern Appalachians. Summaries of the associations of elements and the estimated ease of removal by conventional coal cleaning are shown in Table 3.1. As shown in the table, because of the varying modes of occurrence and the fine mineral associations, removal of trace elements by coal beneficiation can be quite inefficient. Further studies of the association of trace elements in coals have been conducted by Senior et al. (2000a) and Palmer et al. (2004) and their partitioning by size and gravity separation by W. Wang et al. (2006). Specific studies (and reviews of other studies) have been conducted for As (Kolker et al., 2000b; Yudovich and Ketris, 2005a), Hg (Yudovich and Ketris, 2005b,c; Brownfield et al., 2005; Wang M. et al., 2006), and Se (Yudovich and Ketris, 2006a). It should also be noted that trace elements are not uniformly distributed in minerals, such as As in pyrite (Ruppert et al., 2005) and Hg in pyrite (Hower and Robertson, 2003, citing unpublished work from 2000 by same authors).
Source: Fuel Processing Technology 39, M. S. DeVito, L. W. Rosendale, and V. B. Conrad, Comparison of trace element contents of raw and clean commercial coals, 87106, copyright 1994, with permission from Elsevier.
Not all element concentrations will be reduced by beneficiation. Organic sulfur is an obvious example of an element that will not easily be eliminated in beneficiation. Hower et al. (1998) noted an increase in total S from run-of-mine to clean Eastern Kentucky coals, since organic sulfur will go with the product rather than high density reject coal. Similarly, chlorine (associations reviewed by Spears, 2005; Yudovich and Ketris, 2006b) is generally associated with the organic fraction; therefore, removal of the diluent mineral matter increases relative Cl concentration.
Coal oxidation has an important influence on some relevant properties in relation with the coal beneficiation and utilization. For example its influence on plastic properties, which can be completely destroyed by the effect of air oxidation is well known . The reactivity of chars and cokes produced can be substantially modified by preoxidation of coal , and is strongly influenced by the conditions and the extent of oxidation and by coal rank. In this work a study was made in order to contribute to a better understanding of the effect of coal preoxidation on the reactivity of the chars produced. Textural properties of the chars and the gasified materials were determined an a study of their relation with kinetic parameters was carried out.
development and calibration of a dynamic flotation circuit model - sciencedirect
A dynamic simulator is developed for a mini-pilot scale beneficiation plant.The effects of ore type variation and pH are included in the simulator.The flotation circuit model is identified with experimental data.Quantitative validation shows good performance.The simulator can be used, for example, in process design, monitoring and as a decision support system.
Monitoring of mineral beneficiation processes is difficult due to lack of reliable measurements and hazardous environment. Therefore robust models describing steady-state and dynamic behaviour of the processes are needed when aiming for improved monitoring and control. Specific characteristics of models used in mineral processes are that they require spatial mineralogical information of raw material. In this study, a dynamic simulator combining ore characteristic physical data, process operating parameters and mineralogical properties is developed for the Oulu Mining School (OMS) mini-pilot scale mineral beneficiation plant. The mini-pilot process, theoretical part of the model and the model development are described. Open and closed flotation circuit experiments were carried out in mini-pilot research environment for model identification. Experimental and simulated results of ore type variation and pH change are presented. Based on the results the effects of the aforementioned factors on flotation performance are predicted.