The control of a milling operation is a problem in imponderables: from the moment that the ore drops into the mill scoop the process becomes continuous, and continuity ceases only when the products finally come to rest at the concentrate bins and on the tailing dams. Material in process often cannot be weighed without a disturbance of continuity; consequently, mill control must depend upon the sampling of material in flux. From these samples the essential information is derived by means of analyses for metal content, particle size distribution, and content of water or other ingredient in the ore pulp.
The following formulas were developed during a long association not only with design and construction, but also with the operation of ore dressing plants. These formulas are herein the hope that they would prove of value to others in the ore dressing industry.
Pulp densities indicate by means of a tabulation the percentages of solids (or liquid-to-solid ratio) in a sample of pulp. This figure is valuable in two waysdirectly, because for each unit process and operation in milling the optimum pulp density must be established and maintained, and indirectly, because certain important tonnage calculations are based on pulp density.
As used in these formulas the specific gravity of the ore is obtained simply by weighing a liter of mill pulp, then drying and weighing the ore. With these two weights formula (2) may be used to obtain K, and then formula (1) to convert to S, the specific gravity. A volumetric flask of one liter capacity provides the necessary accuracy. In laboratory work the ore should be ground wet to make a suitable pulp. This method does not give the true specific gravity of the ore, but an apparent specific gravity which is more suitable for the intended purposes.
A mechanical classifier often receives its feed from a ball mill and produces (1) finished material which overflows to the next operation and (2) sand which returns to the mill for further size-reduction. The term circulating load is defined as the tonnage of sand that returns to the ball mill, and the circulating load ratio is the ratio of circulating load to the tonnage of original feed to the ball mill. Since the feedto the classifier, the overflow of the classifier, and the sand usually are associated with different proportions of water to solid, the calculation of circulating load ratio can be based on a pulp density formula.
Example: A mill in closed circuit with a classifier receives 300 dry tons of crude ore per day, and the percentages of solid are respectively 25, 50, and 84% in the classifier overflow, feed to classifier, and sand, equivalent to L: S ratios of 3.0, 1.0, and 0.190. Then the circulating load ratio equals
A more accurate basis for calculation of tonnage in a grinding circuit is the screen analysis. Samples of the mill discharge, return sand, and the classifier overflow are screen sized, and the cumulative percentages are calculated on several meshes. Let:
The efficiency of a classifier, also determined by means of screen analyses, has been defined as the ratio, expressed as percentage, of the weight of classified material in the overflow to the weight of classifiable material in the feed. Overflow having the same sizing test as the feed is not considered classified material. Let:
When no other method is available an approximation of the tonnage in a pulp stream or in a batch of pulp can be quickly obtained by one of these methods. In the dilution method water is added to astream of pulp at a known rate, or to a batch of pulp in known quantity, and the specific gravity of the pulp ascertained before and after dilution.
In both cases Dx and D2 are dilutions (tons of water per ton of ore) before and after addition of water. These are found from the specific gravities of the pulp, by formulas (4) and (6) or directly by the use of the tabulation on these of Pulp Density Tables.
The Pulp Density Tables were compiled to eliminate the many complicated calculations which were required when using other pulp density tables. The total tank volume required for each twenty-four hour period of treatment is obtained in one computation. The table gives a figure, in cubic feet, which includes the volume of a ton of solids plus the necessary volume of water to make a pulp of the particular specific gravity desired. Multiply this figure by the number of dry tons of feed per twenty-four hours. Then simply adjust this figure to the required treatment time, such as 16, 30, 36, 72 hours.
In the chemical method a strong solution of known concentration of common salt, zinc sulphate, or other easily measured chemical is added to the flowing pulp at a known rate, or to a batch of pulp in known quantity. The degree of dilution of this standard solution by pulp water is ascertained by chemical analysis of solution from a filtered sample, and the tonnage of ore is then calculated from the percentage solid. This method is impractical for most purposes, but occasionally an exceptional circumstance makes its employment advantageous. It has also been suggested as a rapid and accurate method of determining concentrate moistures, but in this application the expense is prohibitive, since ordinary chemicals of reasonable cost are found to react quickly with the concentrate itself.
With the above chart the per cent solids or specific gravity of a pulp can be determined for ores where gravities do not coincide with those in the Pulp Density Tables.This chart can also be used for determining the specific gravity of solids, specific gravity of pulps, orthe per cent solids in pulp if any two of the three are known.
These are used to compute the production of concentrate in a mill or in a particular circuit. The formulas are based on assays of samples, and the results of the calculations are generally accurate as accurate as the sampling, assaying, and crude ore (or other) tonnage on which they depend.
The simplest case is that in which two products only, viz., concentrate and tailing, are made from a given feed. If F, C, and T are tonnages of feed r on-centrate, and tailing respectively; f, c, and t are the assays of the important metal; K, the ratio of concentration (tons of feed to make one ton of concentrate); and R, the recovery of the assayed metal; then
When a feed containing, say, metal 1 and metal z, is divided into three products, e.g., a concentrate rich in metal 1, another concentrate rich in metal z, and a tailing reasonably low in both l and z, several formulas in terms of assays of these two metals and tonnage of feed can be used to obtain the ratio of concentration, the weights of the three products, and the recoveries of 1 and z in their concentrates. For simplification in the following notation, we shall consider a lead-zinc ore from whicha lead concentrate and a zinc concentrate are produced:
The advantages of using the three-product formulas (20-25) instead of the two-product formulas (14-19), are four-fold(a) simplicity, (b) fewer samples involved, (c) intermediate tailing does not have to be kept free of circulating material, (d) greater accuracy if application is fully understood.
In further regard to (d) the three-product formulas have certain limitations. Of the three products involved, two must be concentrates of different metals. Consider the following examples (same as foregoing, with silver assays added):
In this example the formula will give reliable results when lead and zinc assays or silver and zinc assays, but not if silver and lead assays, are used, the reason being that there is no concentration of lead or silver in the second concentrate. Nor is the formula dependable in a milling operation, for example, which yields only a table lead concentratecontaining silver, lead, and zinc, and a flotation concentrate only slightly different in grade, for in this case there is no metal which has been rejected in one product and concentrated in a second. This is not to suggest that the formulas will not give reliable results in such cases, but that the results are not dependablein certain cases one or more tonnages may come out with negative sign, or a recovery may exceed 100%.
To estimate the number of cells required for a flotation operation in which: WTons of solids per 24 hours. RRatio by weight: solution/solids. LSpecific gravity, solution. SSpecific gravity, solids. NNumber of cells required. TContact time in minutes. CVolume of each cell in cu. ft.
Original feed may be applied at the ball mill or the classifier. TTons of original feed. XCirculation factor. A% of minus designated size in feed. B% of minus designated size in overflow. C% of minus designated size in sands. Circulating load = XT. Where X = B-A/A-C Classifier efficiency: 100 x B (A-C)/A (B-C)
Original feed may be applied at theball mill or the primary classifier. TTons of original feed. XPrimary circulation factor. YSecondary circulation factor. A% of minus designated size in feed. B% of minus designated size in primary overflow. C% of minus designated size in primary sands. D% of minus designated size in secondary overflow. E% of minus designated size in secondary sands. Primary Circulating Load = XT. Where X = (B-A) (D-E)/(A-C) (B-E) Primary Classifier Efficiency: 100 xB (A C)/A (B C) Secondary Circulating Load = YT. Where Y = (D-B)/(B-E) Secondary Classifier Efficiency: 100 xD (B-E)/B (D E) Total Circulating Load (X + Y) T.
Lbs. per ton = ml per min x sp gr liquid x % strength/31.7 x tons per 24 hrs.(26) Solid reagents: Lbs. per ton = g per min/31.7 x tons per 24 hrs.(27) Example: 400 ton daily rate, 200 ml per min of 5% xanthate solution Lbs. per ton = 200 x 1 x 5/31.7 x 400 = .079
Generally speaking, the purpose of ore concentration is to increase the value of an ore by recovering most of its valuable contents in one or more concentrated products. The simplest case may be represented by a low grade copper ore which in its natural state could not be economically shipped or smelted. The treatment of such an ore by flotation or some other process of concentration has this purpose: to concentrate the copper into as small a bulk as possible without losing too much of the copper in doing so. Thus there are two important factors. (1) the degree of concentration and (2) the recovery ofcopper.
The question arises: Which of these results is the most desirable, disregarding for the moment the difference in cost of obtaining them? With only the information given above the problem is indeterminate. A number of factors must first be taken into consideration, a few of them being the facilities and cost of transportation and smelting, the price of copper, the grade of the crude ore, and the nature of the contract between seller and buyer of the concentrate.
The problem of comparing test data is further complicated when the ore in question contains more than one valuable metal, and further still when a separation is also made (production of two or more concentrates entirely different in nature). An example of the last is a lead-copper-zinc ore containing also gold and silver, from which are to be produced. (1) a lead concentrate, (2) a copper concentrate, and (3) a zinc concentrate. It can be readily appreciated that an accurate comparison of several tests on an ore of this nature would involve a large number of factors, and thatmathematical formulas to solve such problems would be unwieldy and useless if they included all of these factors.
The value of the products actually made in the laboratory test or in the mill is calculated simply by liquidating the concentrates according to the smelter schedules which apply, using current metal prices, deduction, freight expense, etc., and reducing these figures to value per ton of crude ore by means of the ratios of concentration.
The value of the ore by perfect concentration iscalculated by setting up perfect concentrates, liquidating these according to the same smelter schedulesand with the same metal prices, and reducing theresults to the value per ton of crude ore. A simple example follows:
The value per ton of crude ore is then $10 for lead concentrate and $8.50 for zinc, or a total of $18.50 per ton of crude ore. By perfect concentration, assuming the lead to be as galena and the zinc as sphalerite:
The perfect grade of concentrate is one which contains 100% desired mineral. By referring to the tables Minerals and Their Characteristics (pages 332-339) it is seen that the perfect grade of a copper concentrate will be 63.3% when the copper is in the form of bornite, 79.8% when in the mineral chalcocite, and 34.6% when in the mineral chalcopyrite.
A common association is that of chalcopyrite and galena. In concentrating an ore containing these minerals it is usually desirable to recover the lead and the copper in one concentrate, the perfect grade of which would be 100% galena plus chalcopyrite. If L is the lead assay of the crude ore, and C the copper assay, it is easily shown that the ratio of concentration of perfect concentration is:
% Pb in perfect concentrate = K perfect x L.(30) % Cu in perfect concentrate = K perfect x C..(31) or, directly by the following formula: % Pb in perfect concentrate = 86.58R/R + 2.5.(32) where R represents the ratio:% Pb in crude ore/% Cu in crude ore Formula (32) is very convenient for milling calculations on ores of this type.
by (29) K perfect = 100/5.775+2.887 = 11.545 and % Pb in perfect concentrate = 11.545 x 5 = 57.7% and % Cu in perfect concentrate = 11.545 x 1 = 11.54% or, directly by (32), % Pb = 86.58 x 5/5 + 2.5 = 57.7%
Occasionally the calculation of the grade of perfect concentrate is unnecessary because the smelter may prefer a certain maximum grade. For example, a perfect copper concentrate for an ore containing copper only as chalcocite would run 79.8% copper, but if the smelter is best equipped to handle a 36% copper concentrate, then for milling purposes 36% copper may be considered the perfect grade.
Similarly, in a zinc ore containing marmatite, in which it is known that the maximum possible grade of zinc concentrate is 54% zinc, there would be no point in calculating economic recovery on the basis of a 67% zinc concentrate (pure sphalerite). For example, the following assays of two zinc concentrates show the first to be predominantly sphalerite, the second marmatite:
The sulphur assays show that in the first case all of the iron is present as pyrite, and consequently the zinc mineral is an exceptionally pure sphalerite. This concentrate is therefore very low grade, from the milling point of view, running only 77.6% of perfect grade.On the other hand, the low sulphur assay of concentrate B shows this to be a marmatite, for 10% iron occurs in the form of FeS and only 2.5% iron as pyrite. The zinc mineral in this case contains 55.8% zinc, 10.7% iron, and 33.5% sulphur, and clearly is an intermediate marmatite. From the milling point of view cencentrate B is high grade, running 93% of perfect grade, equivalent to a 62% zinc concentrate on a pure sphalerite.
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Haver & Boecker Niagara is a leading provider in screening, pelletizing and primary crushing systems. The companys mission is to deliver the best of these technologies to customers in the aggregates, mining, minerals, chemical, cement and food industries. With deep roots and years of experience in these industries, Haver & Boecker Niagara uses its innovative and shared technologies to effectively meet the needs of customers around the world.
Haver & Boecker Niagara is a leading provider in screening, pelletizing and primary crushing systems. The companys mission is to deliver the best of these technologies to customers in the mining, aggregates, minerals, cement, building materials, fertilizer and salt. With deep roots and years of experience in these industries, Haver & Boecker Niagara uses its innovative and shared technologies to effectively meet the needs of customers around the world.
Casein is the major protein in cows milk, and comprises about 80 % of the total protein content of which the rest, some 20 %, are the whey or serum proteins. Casein is the basic component of ordinary cheese. In the cheese-making process, casein is precipitated by the action of rennet enzymes, and a coagulum is formed consisting of casein, whey proteins, fat, lactose and the minerals of the milk.Commercial casein is made from skim milk by one of two general methods precipitation by acid or coagulation by rennet. As much of the fat, whey proteins, lactose and minerals as possible must be removed by multistage washing in water, as they reduce the quality of the casein as well as its keeping quality. Dried, properly produced casein has a relatively good keeping quality and is used mainly in the food and chemical industries.
In order to produce high-quality casein, the raw material, skim milk, must be of good quality. If bacteria have had time to act on the protein in the milk as a result of a change in acidity, this will affect the colour and consistency of the casein, which will acquire a greyish colour and a smoother consistency. Excessive heating of the milk before precipitation will not only cause assorted interactions among the lactose, casein and whey protein constituents, but it will also give the casein a yellow or at worst a brownish colour. In order to produce casein of good bacteriological quality, without high heat treatment of the skim milk, the pasteurization plant may also contain a microfiltration (MF) plant. To satisfy the high demands on the quality of casein intended for use in the food industry, not only must the production line be carefully planned right from the reception of the milk, but the treatment and handling of the raw material prior to this stage must also be carefully controlled.
Skim milk, normally pasteurized at 72 C for 15 20 seconds, is used for the production of rennet casein, as well as other types of casein. Small amounts of fat are detrimental to the quality. It is therefore important that the milk has been separated efficiently. Figure 20.1 shows the various stages of rennet casein production. Renneting takes place with the help of the enzyme chymosin in the rennet. The milk is heated for a short period of time and then cooled to about 30 C. Then the rennet is added. A gel forms after 15 20 minutes. It is cut and the coagulum is stirred while being heated to about 60 C. The high temperature is needed to deactivate the enzyme. Cooking time is around 30 minutes.
The whey is drained off when the final temperature has been reached, and the remaining casein, while in the vat, is washed with water to remove whey proteins, lactose and salt. Washing takes place in two or three stages at a temperature between 45 and 60 C. After the water has been drained off, the casein is further dewatered in sieves or separators. It is then dried with hot air until the water content is 12 %, and finally ground to a powder. The drying temperature depends on the method used. In a two-stage drying process, the temperature is 50 55 C in the first stage and about 65 C in the second. Rennet casein should be white or slightly yellow. A darker colour is a sign of inferior quality and may be caused by too high a lactose content.
Rennet casein was originally produced in batches in special casein tanks, but nowadays continuous processes are also used. In a continuous plant, drainage of whey takes place before the casein passes through two or three washing tanks with agitators. Dewheying is normally done in a decanter centrifuge to reduce consumption of wash water. The casein is dewatered between washing stages, either on inclined static strainers or in decanters. After leaving the washing stages, the water/casein mixture goes through another decanter to discharge as much water as possible before final drying. In large scale production, coagulation of the casein is still done batchwise with a calculated number of casein vats emptied in sequence to feed the continuous de-wheying and washing plant.Washing takes place in countercurrent, which uses water more economically than concurrent washing. The latter system uses one litre of water per litre of skim milk, whereas only about 0.3 0.4 litre of water per litre of skim milk is needed in countercurrent washing. The number of washing stages is dependent on the requirements on the product. Two stages is the minimum. Fresh water is supplied in the last stage only. After washing, the casein is dewatered in a decanter to a DM content of 45 40 %. After drying, for example in a vibration dryer, the casein is ground to a particle size corresponding to 40, 60, or 80 mesh and packed in sacks. (Mesh = number of screen lines per inch; 40 mesh thus corresponds to 0.64 mm.)
The milk is acidified to the isoelectric point of casein, which is thought to be pH 4.6, but it is shifted by the presence of neutral salts in solution and may be anywhere within a range extending from pH 4.0 to pH 4.8. The isoelectric point is the stage where the hydronium ion concentration neutralizes the negatively charged casein micelles, resulting in precipitation (coagulation) of the casein complex. Such acidification can be carried out biologically or by the addition of a mineral acid, e.g. hydrochloric acid (HCl) or sulphuric acid (H2SO4).
Lactic acid casein is produced by microbiological acidulation. The milk is pasteurized and cooled to 27 23 C. A mesophilic, non-gas-producing starter is then added. Acidulation to the required pH takes about 15 hours. If the acidulation process is too rapid, it can result in problems such as uneven quality and reduced casein yield. Large tanks are usually used, because it can take such a long time to empty the tank, that the degree of acidity may vary. When the required acidity has been reached, the milk is stirred and heated to 50 55 C in a plate heat exchanger. After a short hold, the continued treatment washing and drying is practically the same as for rennet casein.
The milk is heated to the required temperature, approx. 32 C. Mineral acid is then added to bring the pH of the milk to 4.3 4.6. Following the pH check, the milk is heated to 40 45 C in a plate heat exchanger and held for about two minutes, when smooth aggregates of casein are formed. To remove as much of the whey as possible before washing starts, the whey/casein mixture is passed through a decanter. This way, less water is needed for washing. Figure 20.2 shows a flow chart for a process line for the manufacture of acid casein. As can be seen, the plant downstream of acidification is almost identical to the one used for production of rennet casein. Before leaving the plant, the whey and wash water can be separated and the casein sludge is collected in a tank. When mixed with a lye solution, the casein dissolves and is then remixed with the skim milk intended for casein production.After dewatering, the acid casein is ground and packed in sacks. The technique for production of acid casein developed by Pillet, France, should also be mentioned. After pre-heating to 32 C, the skim milk is acidified and introduced into a coagulation unit (Figure 20.3). Coagulation is completed after heating to about 45 C by direct steam injection. Dewheying in a decanter is followed by countercurrent washing in one or two specially designed washing towers (Figure 20.4). Before being dried in a vibro-fluidized unit, the casein is dewatered in a decanter.
Co-precipitate contains practically all the protein fractions of milk. Following the addition of small quantities of calcium chloride or acid to the skim milk, the mixture is heated to 85 95 C and held at that temperature for a period of 1 20 minutes to allow interaction between the caseins and the whey proteins. Precipitation of the proteins from the heated milk is then effected by controlled addition of either calcium chloride solution (to produce high-calcium co-precipitate) or diluted acid (to produce medium-calcium or low-calcium co-precipitate, depending upon the amount of acid added and the pH of the resulting whey). The curd is subsequently washed and either dried to produce granular, insoluble co-precipitates or dissolved in alkali as described for the methods for the manufacture of caseinates to produce soluble or dispersible co-precipitates.
Caseinate may be defined as a chemical compound of casein and light metals, e.g. monovalent sodium (Na+) or divalent calcium (Ca++). Caseinates can be produced from freshly precipitated ("wet") acid casein curd or from dry acid casein by reaction with any of several diluted solutions of alkali as outlined in Figure 20.5.
Zoom Fig. 20.5 Basic steps involved in the manufacture of spray or roller-dried caseinates from acid casein curd or dry acid casein. Alkali may be sodium hydroxide, potassium hydroxide, calcium hydroxide, or ammonia.
Basic steps involved in the manufacture of spray or roller-dried caseinates from acid casein curd or dry acid casein. Alkali may be sodium hydroxide, potassium hydroxide, calcium hydroxide, or ammonia.
The most commonly used alkali in the production of sodium caseinate is sodium hydroxide (NaOH) solution, with a strength of 2.5 M or 10 %. The quantity of NaOH required is generally 1.7 2.2 % by weight of the casein solids in order to reach a final pH, generally about 6.7. Other alkalis, such as sodium bicarbonate or sodium phosphates, may be used, but the amounts required and their cost are both greater than those of NaOH. They are therefore generally used only for specific purposes, such as in the manufacture of citrated caseinates. The very high viscosity of sodium caseinate solutions of moderate concentration limits their solids content for spray drying to about 20 %. Regarding the processing procedures, it should be mentioned that the dissolving time is directly related to the particle size and that particle size reduction prior to addition of sodium hydroxide rather than afterwards produces a more rapid reaction. Consequently, the curd is passed through a colloid mill prior to addition of alkali. After the final casein wash, the curd may be dewatered to about 45% solids and then remixed with water (to 25 30% solids) before entering the colloid mill. The temperature of the emerging slurry should be below 45 C since it has been observed that milled curd can re-agglomerate at higher temperatures. Generally the slurry is collected in a jacketed tank provided with an effective agitator and also integrated in a circulation system with a high capacity pump. The addition of diluted alkali must be carefully controlled with the aim of reaching a final pH of about 6.7. Preferably, the alkali is dosed into the recirculation line just upstream of the pump. Once the alkali has been added to the slurry, it is important to raise the temperature as quickly as possible to 60 75 C, to reduce the viscosity. The dissolving time for sodium caseinate prepared in batches is usually 30 60 min. For efficient atomization, the sodium caseinate solution must have a constant viscosity when it is fed to the spray drier. It is common practice to minimize the viscosity by pre-heating the solution to 90 95 C just prior to spray drying.
The preparation of calcium caseinate follows the same general lines as for sodium caseinate, with a couple of important exceptions. Calcium caseinate solutions are liable to be destabilized by heating, especially at pH values below 6. It has been found that during the dissolving process, the reaction between acid casein curd and calcium hydroxide proceeds at a much slower rate than between curd and sodium hydroxide. To increase the rate of reaction between casein and calcium hydroxide, the casein may first be dissolved completely in ammonia. Calcium hydroxide in sucrose solution is then added, and the calcium caseinate solution is dried on rollers. Most of the ammonia evaporates during this process.
Magnesium caseinate has been briefly mentioned in the literature. Compounds of casein with aluminium have been prepared for medical use or for use as an emulsifier in meat products. Heavy metal derivatives of casein which have been used principally for therapeutic purposes include those containing silver, mercury, iron and bismuth. Iron and copper caseinates have also been prepared by ion exchange for use in infant and dietic products.
It is possible to produce sodium caseinate from casein in the presence of a limited amount of water by using extrusion techniques. Some European companies dealing with extrusion cooking Werner & Pfleiderer GmbH (Germany), Clextral (France) and a few others report good results from production of sodium caseinate by extrusion cooking. Most of the published information gives dry casein as the starting material. Water and alkali are added to form a mixture for extrusion. The casein/water mixture may have a moisture content of 10 30%. The extrusion technique used in the production of caseinates is likely to become highly competitive with the traditional batch technique. Furthermore, extrusion processing has also been tested in production of acid casein from skim milk powder. J Fichtali and F R van der Vort have run trials in a pilot plant at the MacDonald College of McGill University, Quebec, Canada. They summarize the results of their trials (1990) as follows: "Our initial work on the production of an acid curd from SMP (skim milk powder) by extrusion processing indicated that significantly more effort had to go into developing the process to produce a quality product. The United States, Canada and the European Economic Community have at times experienced a chronic oversupply of milk, of which substantial amounts are converted into skim milk powder. By modifying the extrusion process conditions, studying high solids coagulation and optimizing the coagulation and washing steps, acid casein of an acceptable quality can be produced by extrusion. This process is continuous, controllable, uses high solid SMP and may reduce labour and floor space requirements relative to conventional processes. This material can serve as a feed for further conversion by extrusion to sodium caseinate, which will be discussed in a succeeding paper."
Rennet casein is a product different from acid casein. In industry, it is used principally in the production of artificial substances in the plastics category. Casein polymerized with formalin is known as galalith, and synthetic fibres of casein are known as lanital. In spite of the large supply of various plastics which compete directly with galalith, there is still some demand for casein for galalith production. Small quantities of rennet casein are also used as a raw material for processed cheese. Rennet casein is insoluble in water.
Acid casein dominates the world markets. It is used in the chemical industry as an additive in paper manufacture for the glazing of paper of fine quality. For paper industry applications, it is particularly important that the casein is free from fat and contains no particles of foreign or burnt matter that might make spots on the paper. To obtain extremely low fat content in skim milk, it should be passed through a microfiltration plant (MF) in combination with pasteurization. Each industry has its own strict quality specifications. The paint and cosmetic industries are also large users of casein.
A casein application of growing importance is its use as a raw material for the manufacture of sodium caseinate. The casein is easily dissolved in a diluted alkali, and the liquid is then spray-dried to a powder. This powder is much more soluble than casein and is being increasingly used by the food industry. It is often used as an emulsifier in cured meats and is found in a number of new products, such as milk and cream substitutes. As sodium caseinate is highly viscous when dissolved, the maximum obtainable concentration is 20% at 55 60 C.
For certain applications, calcium caseinate may be chosen instead of sodium caseinate, one reason being the wish to reduce the sodium content of the product to a minimum. The viscosity of calcium caseinate is somewhat lower than that of sodium caseinate at the same concentration.
This product can also be dissolved in alkali and spray dried, and has much the same field of application as caseinate, however, in the production of calcium co-precipitate, it is possible to adapt the process for the purpose of regulating colour, solubility, and ash content in closer conformity to the users requirements. One of the most important advantages of casein and caseinate from a nutritional point of view is the relatively high content of the essential amino acid lysine. Moreover, tests have shown that the lysine keeps much longer, thanks to the absence of lactose in the environment. This suggests that milk proteins can be more conveniently stored in the form of casein and caseinate than, for example, as dried milk powder. Casein produced for industrial use must satisfy long-established demands for chemical purity. The new trend shows that casein and precipitate are intermediate products which find their way into a host of food products and must therefore satisfy strict demands with respect to bacteriological as well as chemical purity. Process lines must be designed and constructed so that they ensure hygienic manufacturing conditions. As casein is much more of a seasonal product than many other dairy products, the possibility of running the production line in multiple shifts without an undue demand for manual labour must be provided. Water consumption must also be kept within reasonable limits. Therefore, in these circumstances, it is of interest to be able to plan continuous production lines, e.g. incorporating centrifugal machines for dewatering the casein and recovery of casein losses from the whey and wash water.
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[Introduction]: Proportion of antimony ore is far greater than proportion of gangue, so it will be separated by using the method gravity separation. This method has many features, high efficiency, energy saving, and environment protection, which can make the low-grade ore enrichment advance. After gravity separation, the antimony will be purified by floatation. So the processing method of Xinhai is gravity separation-flotation process.
The ore after hand sorting will go through coarse crushing and fine crushing, during which the size of mineral should be under 30mm, and screening classification have three size fractions, 8-30mm, 2-8mm, 0-2mm.
Then the three kind of minerals will go respectively into Xinhai AM30 Jig, Xinhai LTA1010/2 Jig, and Xinhai Sawtooth Wave Jig to gravity separation. The mixted concentrate from the last stage will be go into the gravity concentrate district.
Then after the process of gravity, there will be a process of flotation, the tailings from the preview gravity separation will be grinded, separated, stirred, and separated. And the flotation will apply the process of one roughing, three cleanings, and two scavenging. Then high grade antimony fine powder will be produced. The tailings from the flotation will be reelected by shaking table to recycle the fine antimony particles, which we can get high grade antimony concentrate and tailings.
Mineral processing, mineral beneficiation, or upgradation involves handling three primary types of ROM material, which have been blasted, fragmented, and brought out from an insitu position. These materials can be used directly or by simple or complex processing and even by applying extractive metallurgy like hydrometallurgical or pyrometallurgical methods. The categories are:
The journey from ROM ore to concentrate and finallymetal travels through many operations of liberation, separation, concentration, and extraction before it reaches the end users. These activities have been diagrammatically summarized in Figs.13.53 and 13.54. Apanoramic view of State of the Art zinc and lead smelting is depicted in Fig. 13.55.
Figure13.54. A complete flow diagram, including crushing, grinding, density media separation, froth flotation, and pyrometallurgical and hydrometallurgical process route to achieve the highest purity of metals. PGE, platinum-group elements.
Figure13.55. Panoramic view of hydro-metallurgical smelter of Hindustan Zinc Limited at Rajpura-Dariba, Rajasthan, India. The smelter has an annual production capacity of 210,000 t zinc and 100,000 t lead metal, and 160 MW captive power plant.
Mineral processing or mineral beneficiation or upgradation involves handling of three primary types of ROM ore material which has been blasted, fragmented and brought out from in situ position. These materials can be used directly or by simple or complex processing and even applying extractive metallurgy like hydrometallurgical or pyrometallurgical methods. The categories are as follows:
The journey from ROM ore to concentrate and ultimately to metal has been conceptualized. The various unit operations used for liberation, separation, concentration and extraction have been discussed in the previous pages of this chapter. The activities and the typical sequence of operations in the process plant have been diagrammatically summarized in Fig. 12.54.
Mining and mineral processing industries have been the key focus of research in many countries due to its increasing sustainability concerns that affect global warming and climate change. This chapter analysed and summarised the significant research outputs published on the environmental impact assessment of mining and mineral processing industries through life cycle assessment (LCA). This chapter presents valuable insights in identifying the gaps, where should the focus be in the mining and mineral processing industries for a sustainable future.
The review results reveal the assessment indicators in human health and ecosystems are key factors that are mostly missing in the previous studies which are crucial for people or community living nearby mining area. This chapter identifies the research gaps to the existing literature that can form the base for future research direction in the field of LCA and sustainable energy integration in mining and mineral processing industries.
Mineral processing operations involve a number of process variables that change randomly with uncertain frequencies. The control strategies developed with the use of PID controllers have been found to be inadequate especially in non-linear systems and systems with large lag times. The present development to solve these problems fall under two categories:
The self tuning control algorithm has been developed and applied on crusher circuits and flotation circuits [22-24] where PID controllers seem to be less effective due to immeasurable change in parameters like the hardness of the ore and wear in crusher linings. STC is applicable to non-linear time varying systems. It however permits the inclusion of feed forward compensation when a disturbance can be measured at different times. The STC control system is therefore attractive. The basis of the system is:
The disadvantage of the set up is that it is not very stable and therefore in the control model a balance has to be selected between stability and performance. A control law is adopted. It includes a cost function CF, and penalty on control action. The control law has been defined as:
A block diagram showing the self tuning set-up is illustrated in Fig. 18.27. The disadvantage of STC controllers is that they are less stable and therefore in its application a balance has to be derived between stability and performance.
The empirical model predicts the process output for a certain predicted time. The error is not fixed as in a PID system, but extends over a time period and minimized. The concept is therefore time based and known as an extended horizontal control system. The algorithm is known as Multivariable, Optimal Control Action or MOCCA . The MOCCA system can be considered as an improvement on the level concept described earlier. It is based on the fact that the prediction of output equals the sum of the future actions plus past control action. It is developed around a step response under steady state conditions by combining:
To derive the model, Sripada and Fisher  considered a steady state condition. Also for a single input-single output system (SISO), the predicted output for horizon 1 to P is obtained in N number of step responses. The future and past control actions were written as:
The predicted horizon P, is the number of predicted outputs that the control objective has been optimized The control horizon H is the number of future control actions which minimize the cost function against the predicted horizon.
Optimization of the control system is achieved from performance criteria including any constraints. It is necessary to know the set point and predicted output trajectories for future control effort. The errors and control efforts have to be minimized. For the error trajectory the square of the difference of set point trajectory and the predicted output trajectory is taken. Taking these into consideration Vien et al  describes the cost function, Cf, in terms of minimizing the error trajectory plus control effort. Taking the weighted least square performance, the cost function Cf is given as:
Based on the process model, the control block calculates the predictions for future control actions, the supervisory block generates the desired set point trajectory. The feedback loop with filter and disturbance predictor corrects incongruity between the model and unaccounted, therefore unmeasured, disturbances. It also reduces the noise levels. The predictor in the feed back control loop intimates the future effects of disturbances. Combination of the feed back corrections and the predictions from the model provide the necessary estimate of output.
MOCCA has been found to be far superior to the conventional PID or PI controllers and is being increasingly used. It is particularly useful where long time delays are involved. Its advantage is that it uses discrete step response data and can be used to model processes with unusual dynamic behaviour. Its added advantage over the PID system of control is that it rises faster and has no overshoot. This system has been used successfully in control of grinding circuits.
Mining and mineral processing generates large volumes of waste, including waste rock, mill tailings, and mineral refinery wastes. The oxidation of sulfide minerals in the materials can result in the release of acidic water containing high concentrations of dissolved metals. Recent studies have determined the mechanisms of abiotic sulfide-mineral oxidation. Within mine wastes, the oxidation of sulfide minerals is catalyzed by microorganisms. Molecular tools have been developed and applied to determine the activity and role of these organisms in sulfide-mineral-bearing systems. Novel tools have been developed for assessing the toxicity of mine-waste effluent. Dissolved constituents released by sulfide oxidation may be attenuated through the precipitation of secondary minerals, including metal sulfate, oxyhydroxide, and basic sulfate minerals. Geochemical models have been developed to provide improved predictions of the magnitude and duration of environmental concerns. Novel techniques have been developed to prevent and remediate environmental problems associated with these materials.
In any mineral processing operation, the term benefits of scale is used to denote the significant economic advantages can be obtained by having larger production volumes and using larger ships. Larger tonnage operations operate with fewer man-hours per ton, while capital costs for larger machines are less than the multiples of their relative production capacities. In order to compete on world markets, category 1 producers must consider the benefits of scale. For example, in the kaolin industry during the 1970s, a 100,000 tons/year operation was considered to be a reasonable commercial operation. For the current developments in Brazil, a minimum plant size of 300,000 tons per year is being quoted.
For category 2, the annual tonnage requirement is governed by market size rather than benefits of scale. Annual productions from such processing operations typically fall between 10,000 and 100,000 tons per year. The sizes of category 3 operations are typically governed by other factors such as market size or accessible market share.
Mining and mineral-processing industries producing lithium minerals, metals, and salts contribute to the lithium burden in the environment. The processing of lithium-containing minerals such as spodumene, in general, comprises crushing, wet grinding in a ball mill, sizing, gravity concentration, and flotation using a fatty acid (oleic acid) as the collector. The major lithium mineral in lithium ore is spodumene, which is considered insoluble in water and dilute acids. However, a small amount of dissolution may occur during processing of the ore especially in the grinding and flotation stages where some dilute (0.01M) sulfuric acid is used (see Table 6). Tailings are discharged to storage areas, and the decanted water is usually recovered for reuse. Lithium concentrations in tailing dams increase gradually. The dissolved lithium found in the tailing dams of lithium mineral beneficiation plants could be as high as 15mgl1. The repeated use of tailing waters without any treatment further increases the dissolved lithium levels in these waters.
Some of the lithium minerals are more soluble than the others. Manufacturing of lithium chemicals could contribute to the lithium burden in the environment. Most of the lithium chemicals are often more soluble than lithium minerals, and therefore, the risk to the environment could be higher than the risk introduced by the lithium minerals (see Table 5).
Mining and mineral processing can cause arsenic contamination of the atmosphere (in the form of airborne dust), sediment, soil, and water. The contamination can be long-lasting and remain in the environment long after the activities have ceased (Camm et al., 2003). Recent estimates suggest that there are approximately 11 million tonnes of arsenic associated with copper and lead reserves globally (USGS, 2005). In developing mines containing significant amounts of arsenic, careful consideration is now given to treatment of wastes and effluents to ensure compliance with legislation on permitted levels of arsenic that can be emitted to the environment. Such legislation is becoming increasingly stringent. Arsenic contamination from former mining activities has been identified in many areas of the world including the United States (Plumlee et al., 1999; Welch et al., 1999, 1988, 2000), Canada, Thailand, Korea, Ghana, Greece, Austria, Poland, and the United Kingdom (Smedley and Kinniburgh, 2002). Groundwater in some of these areas has been found with arsenic concentrations as high as 48000gl1. Elevated arsenic concentrations have been reported in soils of various mining regions around the world (Kreidie et al., 2011). Some mining areas have AMD with such low pH values that the iron released by oxidation of the iron sulfide minerals remains in solution and therefore does not scavenge arsenic. Well-documented cases of arsenic contamination in the United States include the Fairbanks gold-mining district of Alaska (Welch et al., 1988; Wilson and Hawkins, 1978), the Coeur d'Alene PbZnAg mining area of Idaho (Mok and Wai, 1990), the Leviathan Mine (S), California (Webster et al., 1994), Mother Lode (Au), California (Savage et al., 2000), Summitville (Au), Colorado (Pendleton et al., 1995), Kelly Creek Valley (Au), Nevada (Grimes et al., 1995), Clark Fork river (Cu), Montana (Welch et al., 2000), Lake Oahe (Au), South Dakota (Ficklin and Callender, 1989), and Richmond Mine (Fe, Ag, Au, Cu, Zn), Iron Mountain, California (Nordstrom et al., 2000).
Phytotoxic effects attributed to high concentrations of arsenic have also been reported around the Mina Turmalina copper mine in the Andes, northeast of Chiclayo, Peru (Bech et al., 1997). The main ore minerals involved are chalcopyrite, arsenopyrite, and pyrite. Arsenic-contaminated groundwater in the Zimapan Valley, Mexico, has also been attributed to interaction with AgPbZn, carbonate-hosted mineralization (Armienta et al., 1997). Arsenopyrite, scorodite, and tennantite were identified as probable source minerals in this area. Increased concentrations of arsenic have been found as a result of arsenopyrite occurring naturally in CambroOrdovician lode gold deposits in Nova Scotia, Canada. Tailings and stream sediment samples show high concentrations of arsenic (39ppm), and dissolved arsenic concentrations in surface waters and tailing pore waters indicate that the tailings continue to release significant quantities of arsenic. Biological sampling demonstrated that both arsenic and mercury have bioaccumulated to various degrees in terrestrial and marine biota, including eels, clams, and mussels (Parsons et al., 2006).
Data for 34 mining localities of different metallogenic types in different climatic settings were reviewed by Williams (2001). He proposed that arsenopyrite is the principal source of arsenic released in such environments and concluded that in situ oxidation generally resulted in the formation of poorly soluble scorodite, which limited the mobility and ecotoxicity of arsenic. The Ron Phibun tin-mining district of Thailand is an exception (Williams et al., 1996). In this area, arsenopyrite oxidation products were suggested to have formed in the alluvial placer gravels during the mining phase. Following cessation of mining activity and pumping, groundwater rebound caused dissolution of the oxidation products. The role of scorodite in the immobilization of arsenic from mine workings has been questioned by Roussel et al. (2000), who point out that the solubility of this mineral exceeds drinking water standards irrespective of pH.
A wide variety of minerals processing routes are used for REE deposits (Jordens et al., 2013; Krishnamurthy and Gupta, 2015). For many REE ores, processing techniques for the minerals are unproven on a commercial scale and processing is a major challenge that needs to be considered early in exploration. Physical concentration using density, magnetic and electrostatic properties are normally the most cost-effective. Monazite and xenotime, if reasonably well liberated and coarse-grained, are amenable to physical separation from mineral sands and some carbonatite ores. Finer grained phosphates, and most fluorcarbonates, require more complex and expensive processing via flotation, and/or acid leaching. Eudialyte can be concentrated by physical beneficiation but is difficult to dissolve, although techniques to solve this problem are now available at laboratory and pilot scale.
With yearsaccumulation of experience in R&D, the HCH ultra-fine grinding mill is new ultra fine pulverizing equipment designed by HongCheng. This mill is widely used to grind any non-metallic minerals with Mohs hardness below 7 and moisture below 6%, such astalc,calcite, calciumcarbonate, dolomite,bentonite,kaolin, graphite, carbon black etc.. Product fineness can be adjusted within a range from 325 mesh to 2500mesh and its disposable fineness can reach D97 5um. HCH ultra-fine grinding mill is especially suitable for ultra fine grinding. After a long period of market application practice and user authentication, the device HC1395 model was certified by the China Association of calcium carbonate for energy-saving equipment in China's calcium carbonate ultra-fine processing industry. HCH1395 is the biggest ultra fine circle-roll grinding mill in China.
Technological process: The pre-grinding raw ore material will be crushed into particles10mm and transport to the feeding hopper by the elevator, then feed into the grinding chamber by the feeder. The grinding rollers equip on the rotary table rotates around the centre shaft. There is flexible gap between the roller and ring. The rollers rotate outward by the centrifugal function to compress the fixed ring. The rollers also self- rotates around the roller pin. When material passed through the gap between the ring and roller, the material will be smashed by the rotating rollers. Four layers of rollers. Material will be grinded 1st time when passing the 1st layer of roller and ring. Then be grinded second, third and fourth time when loop through each layer rollers. Thus the materials were grinded sufficient and obtain much fine powder. The powder fallen down onto the bottom table by gravity will goes up to the classifier for separation by the airflow from blower. The qualified fineness passing from the classifier will be collected by the pulse bag collector as final product, while unqualified fallen down for regrinding till passing through. The powder goes with air flow into the pulse bag filter and collects by the discharging valve.The wind path is in circulation and the airflow is in negative pressure. There will be no dust escape, so the equipment can ensure a no dust operation in workshop.
HCH Ultra-fineGrinding Mill is widely used to grind any non-metallic minerals with Mohs hardness below 7 and moisture below 6%, such as talc, calcite, calcium carbonate, dolomite, bentonite, kaolin, graphite, carbon black etc.. This kind of mill is especially suitable for ultra finegrinding. The fineness can be adjusted from 0.045mm(325 mesh) to 0.005mm(2500 mesh), whose range is much wider than that of traditional RaymondMill.
Higher production capacity and lower power consumption:Non metallic mineral particles feed which feedingsize is less than 10mm, canbeone-time processedas < 10 m powder (97% passing). The particlesize less than 3um accounted for about 40%,which contributes to a largerspecific surface area.It has the advantages of low cost,high efficiency, andgoodproduct fineness.
Wide fineness and flexible adjustment: Turbine classifier (patent no.: ZL201030143470.6). The fineness can be adjusted flexibly from 0.04mm (400 mesh) to 0.005mm (2500 mesh). Products with various fineness can meet the market needs and improve your competitiveness.
Environmental protection: The pulse collecting system will remove 99.9% of the dust, ensuring dust-free operation environment. The pulse dust collection system is Hongcheng special invent comply for the environment protection requirements.
Thorough CleaningThe pulse dust collection system is adopted with pulse-jet type of cleaning. By utilize the compression air to shoot clean each filtering bag. High and complete dust cleaning. Prevent bags from powder stocking.
Simply complete the form below, click submit, you will get the price list and a Hongcheng representative will contact you within one business day. Please also feel free to contact us by email or phone. ( * Denotes a required field).
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The term talc refers both to the pure mineral and a wide variety of soft, talc-containing rocks that are mined and utilized for a variety of applications. Talc forms mica-like flakes. Talc is the softest mineral on the Mohs hardness scale at 1 and can be easily cut and crushed. Talc has perfect cleavage in one direction. This means that it breaks into thin sheets. As a result, it feels greasy to the touch (which is why talc is used as a lubricant).
Most talc is mined today by conventional open-pit, drill-and-blast, shovel-and-truck techniques. The major difference from conventional technology is that blasting is minimized to reduce breakage of soft talc ore.
Most talc in the United States is produced from an open pit mine where the rock is drilled, blasted, and partially crushed in the mining operation. The highest grade ores are produced by selective mining and sorting operations.
Great care is taken during the mining process to avoid contaminating the talc with other rock materials. These other materials can have an adverse effect on the color of the product. Contamination can introduce hard particles that cause problems in applications where talc is being used because of its softness or lubricating properties.
Partially crushed rock is taken from the mine to a mill, where it is further reduced in particle size. Impurities are sometimes removed by froth flotation or mechanical processing. The mills produce crushed or finely ground talc that meets customer requirements for particle size, brightness, composition, and other properties.
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