This page isdevoted to the subject From theComminution Theory to Practice by selection of the correct Process Equipment by taking you step by step through some of the variables encountered in the specific part of Comminution calledgrinding and how each of these affect your operations.Should it be possible to reduce all of these variables to a simple mathematical formula the selection of a grinding mill would, of course, be simple. Many approaches to this have been made but to date a foolproof formula, both mathematically and practically applicable, has not been devised. We must, therefore, take each variable into consideration on its own merits and then correlate such ideas into a single selection. To do this a broad experience and understanding of the complete subject of grinding is essential. This is a part of the problem of your engineers and our own consulting staff. General points have been discussed briefly on wet or dry grinding, and fineness of grind. Two main categories of grinding equipment, namely rod mills and ball mills, have also been mentioned.Whether grinding is to be performed wet or dry, or in a ball mill or rod mill, a choice must be made between open or closed circuit. Other factors which require thought are mill size, speed of mill rotation, moisture content, retention time, circulating load, type and sizes of grinding media, mill pulp level, mill shape, power, and relation between diameter and length. These all influence operating results and are evaluated and incorporated in the selection and design of the Mill.
Mills are essentially cylindrical in shape and this design has been selected for very definite reasons.Mill capacity is a function of the mill volume and the load of grinding media. Therefore to obtain a mill of greatest capacity for any given space, pure logic dictates a mill having the greatest volume. While a square section would provide the greatest volume, smooth continuity of operation and uniformity of media action must also be considered and thus a true circle is the only practical answer. Should the diameter vary from one end to another there is but one thing which occurs reduced volume, or in other words, reduced capacity.
The cylinder simplifies mill construction, resulting in a minimum amount of maintenance and reflecting in less downtime. Power wise, cylindrical mills provide the most economical piece of equipment for grinding work. Floor space for any mill is proportional to the diameter of the mill and its length. Therefore, floor space is kept at a minimum. A mill, keeping uniform diameter throughout its full length obtains maximum volume for a given floor space.
The relationship of mill diameter to length is of considerable importance. Rod mills should have a length greater than the diameter to avoid entanglement of rods. The construction of ball mills is different in that the diameter may be larger, equal to, or smaller than the length.
The selection of mill length is dependent upon the size of feed, size of product and type of grinding circuit selected. Considerations given a short mill are the reduced floor space, shorter retention time producing less fines in the discharge product, and the possibility of producing a slight amount of tramp oversize particles. Corresponding conditions to be expected from a longer mill are greater floor space requirements, higher capacity (closely proportional to mill length), greater retention time thereby producing a finer mill discharge product and a greater amount of extreme fines, less tramp oversize in the product. Since most mill variables act as a function of the mill length, this consideration is relatively simple. On pages 10 and 11 considerable discussion is provided on the subject of mill diameter.
The method of operating a grinding mill may be classified into two methods, open circuit or closed circuit. In open circuit grinding feed enters one end of the mill at a predetermined rate so as to make the desired finished product during a single pass through that mill. In other words there is no size classification made on the discharge product. One important application is on ores containing damp and clay-like material which causes difficulty in fine crushing. This problem is generally solved by wet grinding in a rod mill or in this case it may be called wet fine crushing.
In closed circuit grinding the feed enters one end of the mill and is discharged from the other end into some type of classifier. This classifier is to limit maximum particle size removed from the mill circuit. The oversize material is returned to the grinding mill for additional size reduction. Such material returned to the mill is defined as the circulating load. Classifying equipment may consist of vibrating screens on coarse separations for wet or dry grinding. For wet grinding in the finer size ranges wet classifiers and/or cyclones are employed, generally to make a size separation from 20 mesh down to 325 mesh. Under dry grinding conditions air classifiers are used to make the size classification.
1. Simplicity of mill layout. 2. May be used where classifying is not practical. 3. May be used where control of product size is not important. 4. The use of rod mills will produce an ideal fine feed for ball mills. 5. May be used where classifier dilutuion would be objectionable.
1. Provides a close control of finished product size. 2. Mill capacity is greatly increased. 3. Power requirements per ton of finished material are lower. 4. Less overgrinding or production of extreme fines.
Generally speaking circulating loads for rod mill operation will be less than 200%. In most cases it will more closely approach 100% to 120%. In ball mill operations the circulating load will vary between 300% and 1000% depending upon the grind required and type of material. As an average it will approach 350%.
Proper speed, or most efficient speed, at which mills are to operate depends upon the action desired by the grinding media, the amount of media, its size and shape, percentage of solids in each mill, and shape of liners. In the following discussion we refer to critical speed applying to ball mills and peripheral speed referring to rod mills. Reference graphs giving these speeds for various mill diameters will be found below.
Critical speed may be considered as the speed at which an infinite particle will continue its travel around the periphery of the mill, thus becoming part of a flywheel action. Grinding balls actually will not centrifuge at this theoretical critical speed since they are larger than an infinite particle and also because of slippage.The following table illustrates the action of a normal ball charge at various percentages of critical speed.
Generally speaking ball mills operate within the range of 50% to 90% of critical speed. The average is found to be approximately 75%. Pebble Mills have been found to operate more efficiently at higher speeds than ball mills. When reaching the higher percentages of critical speed caution must be used and consideration given the action of the scoop feeder.
When considering rod mills, peripheral speeds only should be considered. In the case of ball milling, with a free moving grinding medium, ball paths obtained are based on critical speeds. In a rod mill with a comparatively rigid grinding medium, a certain cascading and roll of rods are obtained, which does not resemble the action of loosely projected ball paths. Therefore to simulate similar rod actions in mills of various diameters it is necessary to operate between 60% and 98% of critical speed. Therefore, Critical Speed is misleading if used in conjunction with rod mills. It has been found that low pulp level rod mills show increases in efficiency as peripheral speeds are increased from 300 per minute to the present practical maximum of around 500 per minute.
Theoretically critical speed is the point at which centrifugal and gravity forces acting on an infinite particle traveling on the shell liner offset each other or become equal. The formula used in calculating critical speed is shown on the graph below.
Often grinding capacity and power are used hand in hand since power is an index to the potentialities of any grinding mill. The grind achieved as in direct relation to the power applied in rotating a mill. This rotation transmits energy input to the grinding media and energy is consumed in reducing particle sizes. When any particle is split, producing two or more smaller particles, the total surface area of the smaller particles will be greater than the surface area of the initial size. Therefore surface area often is used to express the amount of grinding work which is performed.
There are two methods of looking at power. First and easily understood is the reference to connected horsepower, or the actual consumed horsepower required to drive the mill. The second is basing power on the amount of work done. We prefer to express this as kilowatt hours per ton of material ground. The following formula containing three factors may be found useful in calculating power consumed per ton of material ground. Wherever two of the factors are available, the third may easily be solved.
There are several variables in mill horsepower the most important has to do with mill diameter. Several of these variables also reflect similarly on capacity. There have been various statements made as to how power and capacity vary with mill diameter each using a figure of the diameter raised to some power such as D3 D2.65 D2.6 and D2.5. For your convenience we have listed on page 11 a table giving these various diameters raised to the appropriate figures. We have found in the low pulp level mills that the capacity varies closely as the diameter cubed. The mill power varies closely as the diameter to the 2.5 power. With overflow type mills, or high pulp level mills, the theoretical exponents more closely approach the 2.6 or 2.65 power. The difference lies in the waste of energy when transmitted through a cushioning deep quantity of pulp.
Power required in relationship to mill length is a straight line function or direct proportion within limits. In other words each foot of mill will require a definite amount of power. Capacity of a mill also varies in the same manner.
You are operating an 8 feet diameterBallMill consuming 245 HP and grinding 500 tons per day to 65 mesh. What will be the capacity of a 5 feet grinding Mill? From the table below, the 8 feet diameter mill cubed is 512; the 5 feet diameter mill cubed is 125; the 8 feet mill diameter to the 2.5 power is 181; and the 5 feet milldiameter to the 2.5 power is 55.91. Such diameters are inside new liners.
Power consumption is also a reflection of the amount of media carried within the mill. The maximum power requirements for any mill will be when it is 45% to 50% filled with grinding media. Above or below this power drops off. Similarly mill capacity will behave the same way. Within limits the effect of adding or decreasing grinding media will be proportional to that weight.
The above refers to wet grinding. Under dry grinding conditions it has been found that the power will be between 60% and 90% that of a wet grinding mill. Wet grinding capacity will be 1 to 2 times that of dry grinding.
A Laboratory Ball Millis used for grinding in laboratory flotation test work, wet grinding is necessary in several stages in order to approximate the actual grinding conditions of a ball mill and classifier in plant operation. With this small ball millit is possible to grind successfully in several stages without dilution, because the large feed opening and discharge hole (which retains the balls while the mill is being discharged) permits quick and thorough draining with the use of a minimum of wash water. Used in conjunction with a (Closed) Ball Mill, and a laboratory batch classifier saves time.
The (Open) Batch Mill makes possible simulation of grinding in a ball mill-classifier circuit by grinding a fixed time, then hand classifying with a sieve and a bucket and returning the sand to the mill for further grinding. Pulp volume can be kept at a minimum by control of wash water since the ball load does not dump out. Thus fine or coarse grinding is accomplished with ease of manipulation. The mill is used considerably as a regrind unit for further size reduction of flotation concentrates, table concentrates, middlings, and other mill products. Air may be introduced to the pulp, if desired, through theopen end of the cast steel drum.The (Open) Laboratory Ball Mill can beused for the amalgamation of table and flotation concentrates, by using one 4 ball during the grinding action.A few slight changes provide for mounting an Abbe Jar onthe driving shaft opposite the drum. This affords anexcellent combination laboratory grinding unit.
Laboratory Ball-Rod Mill is for grinding large quantities of ore in batch or continuous work. A grate discharge allows quick emptying for batch work; and a trunnion discharge, for continuous work. This flexible unit can also be used as an amalgamator.
The welded steel base can be provided in lengths suitable for supporting one, two, or three drums. One drum constitutes a ball mill; two, a rod mill; and three, a tube mill. The constructionpermits easy and rapid addition or removal of drums forconversion from one type to another, the trunnion being moved along the channels to accommodate the various lengths.The grinding drums are made with thick walls, thus eliminating the necessity of liners and the possibility of salting samples.This grinding mill has many commercial applications, since special acid resisting drums and heads can be furnished. Hard iron or alloy steel replaceable liners are easily inserted where wear is an item as in continuous use.
For continuous pilot test plants utilizing No. 7 or No. 8 Sub-A Flotation Machines, ball mills with larger grinding capacities are required. The 30 Convertible Ball Mill is ideal for this application. Capacity of this mill with single section is 3 to 5 tons per 24 hours while the double section mill, illustrated above, has a capacity of 6 to 9 tons per 24 hours.
The Buckboard and Muller ball mills are an extremely useful addition to most ore dressing or industrial laboratories. It accomplishes the quick reduction of small quantities of ore or other crushable materials to a fine powder. The unit consists of a chilled iron buckboard grinding surface, two sides of which are rimmed, and the desired type and weight of muller. Thebuckboard grinding surface is planed smooth and standard mullers have rounded crushing faces. Mullers may be purchased separately from either standard types shown or special from the three sizes of each type listed in the table and classified by rounded or flattened crushing faces. Rounded face mullers as listed have hickory axe type handles and flattened face mullers are equipped with hickory pick type handles.
MILL, Ball, Braun-WelschA laboratory ball mill particularly suitable for the metallurgical laboratory for flotation, cyanidation or amalgamation tests, but useful for any type of fine grinding. Will grind either wet or dry.
The ball mill body, or shell, is of gray cast iron, 12 inches inside diameter, 7 inches between the two machined driving ribs. Openings in the centers of heads allow free access of air during the grind, duplicating commercial scale plant conditions. The driving power is peripheral, applied frictionally through the supporting rollers. For loading, discharging and cleaning, the mill body is raised from the rollers by a rack and pinion device. This allows it to be turned, bringing its axis into the vertical position shown above. Removing the upper head by loosening five lug screws makes the inside completely visible and accessible. Pulp can be discharged through the lower opening, the balls being retained by a spider. Discharging and cleaning are rapid and thorough. All possibility of salting is eliminated, provided advantage is taken of the facilities for thorough cleaning.
The capacity is from 10 grams up to 12 pounds. Preliminary crushing to minus 10 mesh is recommended. The time for reducing an average sample to minus 65 mesh is about 15 minutes. Iron balls are used. The usual ball charge is 22 pounds, assorted sizes, 1, 1 and 2 inch being used. Steel rods 1 inch diameter may be used with good results. The mill operates at 50 R. P. M. and is driven by a 1/6 H. P. 110 volt, belt-connected motor. Mounted on steel table; floor space required, 30 x 30 inches. Height over all, 51 inches. Net weight, including balls, 270 pounds. Gross shipping weight, approximately 350 pounds.
A Ball Mill of Laboratory size, for grinding wet or dry material. Specially recommended for experimental flotation work where the oil or reagent is mixed with the samples, and gives results in harmony with those obtained in regular practice. Can be used for any type of fine grinding.
The maximum capacity is 25 pounds of ore. An assortment of iron balls inches, 1 inch and 1 inch sizes are included. The sample should be ground 20 mesh or finer (preferably with a Braun type U A Pulverizer) before being placed in the ball mill.
Designed for laboratory needs with ready access for charging and discharging the sample. The oval opening measures approximately 10 x 4 inches, and is large enough for all practical purposes. The receiving pan is in two parts, the upper portion being fitted with a screen to catch the iron balls while the ground materials fall through into the lower pan.
Ball Mill Dimensions: Cylinder, 12 x 12 inches ; floor space 27 x 13 x 22 inches ; pulley, 12 x 2 inches. Speed recommended, 40-50 RPM. Shipping weight, 300 pounds.Iron Balls, ExtraFor use with above.Receiving Pan, Extra.Source
The Abbe Pebble Mill is particularly adapted to pulverizing or mixing either dry or wet materials. This unit is of the batch or intermittent type. The cylinder is approximately half filled with flint pebbles, porcelain balls, or metal balls; the material is put into the cylinder; a tight cover is fastened securely, to seal the mill hermetically; and the cylinder is then revolved until the fineness required is obtained. After that the tight cover is replaced by a grate discharge cover and the cylinder is revolved until the material is discharged, the grate retaining the pebbles or balls. For dry grinding it is customary to enclose the cylinder to prevent the spread of undesirable dustand also to preserve all of sample to insure accuracy of testingprocedures.
In wet grinding, the same directions are followed except there is no casing required; and instead of replacing the tight cover with a grate cover, a special wet discharge cover, or the Abbe patented discharge valve, is used for emptying the mill.
The Abbe Laboratory Jar Mill pulverizes materials by friction and the fall of pebbles or balls contained within a rotating jar. In operation the jar is filled almost to the center plane with pebbles, and enough crushed material is added to fill the interstices between the pebbles and bring the charge to about 3/5 of the capacity of the jar. Usually a coarse screen is used to separate the pebbles from the material after grinding.
This unit is particularly adapted to pulverizing and mixing dry or wet materials and is manufactured in many sizes and styles for various capacities and conditions. Number of jars handled varies with mill size.
Jars are manufactured in many sizes and are of the best material. The three kinds available are: (1) Porcelain Jars carefully molded and fired to obtain the proper degree of vitrification so as to give acid and wear resisting qualities for grinding and agitation. (2) Metal Jarsmade of the metal most resistant to the action of a grinding charge, such as Monel, stainless steel, cast steel, and bronze. (3) Pyrex Jarsare transparent and enable observation during grinding or agitating action.
The Jar type Laboratory Ball Mill is ideal for use in pulverizing, mixing of dry and wet materials, and agitation of all types of pulps. Two large bottles or as many as six smallbottles can be used at the same time, thus providing a very flexible type of jar mill. When used for agitation, large ammonia bottles filled with pulp can be agitated continuously for any length of time.
The Laboratory Pebble Mill is a simple, efficient laboratory batch grinding unit. This machine is also an effective laboratory bottle type agitator for use on all types of pulps especially cyanide pulp. Two large bottles or jars or as many as six small jars may be used at the same time. This provides a wide range of capacity for batch laboratory grinding, and, at thesame time, affords a very flexible unit for batch agitation.The speed reducer and chain drive mechanism is positiveand the entire unit is mounted on a steel base. Idler rolleris adjustable to various bottle sizes. Data on jar sizes and types given under Laboratory Agitator, (Bottle Type).
Pulverizing is effected in these ball mills by the friction and fall of balls or pebbles contained within a rotating porcelain or glass jar. In general practice, the jar is filled to a little below the centre plane with pebbles; well crushed material then being added to fill the interstices and bring the charge to about three-fifths of the total space in the jar. After the operation, the pebbles and ground material are usually separated by means of a coarse grid.
Either hard or soft material may be ground, but it should not be moist to the extent that it will pack in the cylinder. If sufficient liquid is present, this method is superior to others for fine pulverization. Equipment is supplied complete with porcelain jar and flint pebbles.
A complete laboratory service that includes preliminary examination, batch or pilot-grinding test, open or closed circuit grinding, both wet and/or dryprovides important data for determining accurate mill size, for determining circulating loads, sizing accessories, grindability indexes and power requirements. Preliminary tests are made at no cost to you. Send 100 lb. sample of material prepaid.
Under the foregoing conditions experiments were designed to obtain data on (1) the relation of time of grind to mill output in g.p.m., and (2) the relation of mill output in g.p.m. to size of finished product. The relation of time of grind to mill output was studied by two procedures: (1) By a batch continuous grind technic and (2) by a batch cycle grind technic. The batch cycle grind technic is described in Part II of this Paper, Batch Closed-Circuit Grinding.
A short period of grind was selected, which far purposes of discussion can be assigned a value of x minutes. A number of individual 1,750-gram charges were weighed out. One of them was ground for x minutes, one for 2x minutes, one for 3x minutes, and so on up to 8x minutes.
The g.p.m. output of the mill for any x-minute time increment is readily calculated by simple arithmetic. If Wx is the weight of finished product resulting in the first x-minute grind, the g.p.m. output is Wx % x. If W2x is the weight of finished product produced in 2x minutes time, the mill output for the second x-minute increment is W2x Wx/x g.p.m., and so on.
It is obvious that the maximum percentage (by weight) of unfinished product (feed) is present in the mill when x is zero, and that although the mill load remains constant, the weight of unfinished feed starts diminishing at the initial turn of the mill. For this reason it was thought that in batch grinding, which may be assumed to compare in a limited way to open-circuit grinding, the efficiency of the mill possibly should be a maximum at the very initial revolution of the mill. This proved not necessarily to be the case, as may be seen by reference to the experimental data given.
The many missing links in the science of ball-mill grinding as revealed in mailing the study here presented and as brought out in the subsequent analysis of the experimental data, led to the studies by the author and H. E. Lee which are presented elsewhere under the title of Ball Mill Studies, Parts I, II, III, etc., and to Part II of this paper.
The condensed experimental data of this research and calculations based on these data are given in Table 1. The time of grind is given in vertical column 1. The sieve analysis of the product of each grind test is given, showing total grams, weight per cent, weight per cent cumulative, grams per minute, grams per minute cumulative, total surface, and total surface cumulative of each sieve size. There is also a horizontal column for each time, increment of grind, giving grams per minute cumulative for each three-minute grind increment throughout the entire 24-minute time, rangethat is, the g.p.m. output for the 0 to 3 minute period, for the 3 to 6 minute period, for the 6 to 9 minute period, and so on for each period.
Figure 1 shows the finished product-per cent of total mill charge plotted vertically against time of grind horizontally for each of the finished sizes considerednamely, 65, 100, 150, 200, and 250 mesh. These lines may be termed cumulative rate curves.
A set of tests similar to those previously described for quartz was made with Morning mine ore. The data are presented in Table 2. This ore is an aggregate largely of quartzite, siderite, sphalerite, and galena. These, data are shown graphically in Figure 4, in which finished product, g.p.m. is plotted vertically against time of grind increment horizontally. These curves indicate that at all sizes the highest rate of production of finished product is for the initial two-minute grind increment, except in the cases of the finished minus-48 mesh and minus-65 mesh products. In the absence of surface data, such as were calculated for quartz, it can not be known if the maximum finished product points for the minus-48 and minus-65 mesh sizeswhich occur at the second two-minute grind incrementsare due to increases in unfinished sand surface in the mill or to some other element.
The significant fact to note from this set of experiments is that finished product is made most rapidly at the very beginning of the grind. This suggests that in the practical closed-circuit plant the circulating load should be large in order to keep a high percentage of unfinished sand in the mill.
For quartz grinding to, say, finished minus-100 mesh sand, a 100 per cent circulating load should give approximate maximum efficiency. For Morning ore the circulating load should be high for maximum efficiency. The charge should not remain in the mill longer than two minutes, or possibly one minute. For this (two min.) length of grind for, say, finished minus-100 mesh product, 20.6 per cent of the charge is reduced to finished size. Therefore, for high efficiency, a circulating load of not less than 500 per cent is required. This corresponds closely to practice now in vogue in the Morning Mill, Mullan, Idaho, where this material is being milled. In designing a plant to treat a new ore, a test of this kind should be of much value in selecting the type of ball mill and classifier to be used. These experiments led to the batch closed-circuit experiments, which are to be described in Part II of this paper.
If the capacity of the mill used in these experiments be rated at one unit of finished minus-65 mesh product in a unit of time, its capacity to make finished product at any finer mesh may be calculated from Table I for quartz and from Table 2 for Morning ore. See Table 3.
These data show that the mill has four times as much capacity when grinding to finished minus-65 mesh as when grinding to finished minus-250 mesh. It has approximately three times as much capacity to produce finished minus-150 mesh Morning ore as finished minus-150 mesh quartz. These figures all are based on the output of the mill for the initial grind periodthree minutes for quartz and two minutes for Morning ore.
To learn if a greater mill output could be obtained than in any of the previous tests, an experiment was designed in which the period of grind was one minute. Quartz was used, and all other experimental conditions were the same as in previous experiments. At the end of each one-minute grind the charge was removed from the mill and the weight, in grams, of finished minus-100 mesh sand was determined. A quantity of new ore equal to the determined weight of finished product with the original charge was returned to the mill. The grind was conducted for another one-minute period. This process was repeated seven times, and each time enough new water was added to keep the water-ore ratio at 30:70 by weight. At seven new ore additions the mill became choked, and further additions could not be made.
The object in the plan of this experiment was to keep in the mill a nearly constant weight of unfinished feed, without removing the finished product. This procedure should, of course, result in high unfinished sand surface in the mill. The experimental data of this run are compiled in Table 4.
The output of finished minus-100 mesh sand produced the first minute is 175 grams; the output per minute for the first two minutes is 154 grams; for the first three minutes, 127 grams; for the first four minutes, 143 grams; for the first five minutes, 108 grams; for the first six minutes, 119 grams; and for the full seven minutes, 106 grams. On the basis of these figures the output is a maximum for the initial one-minute grind during, which only original feed is in the mill. Considered on the basis of output for each of the time increments, calculation shows that the output of the mill is greatest during the fourth minute grind increment, when 195 grams were produced. This result may be due to experimental error, and it is believed to be, for during the next minute (the fifth) increment the output is negative and for the sixth-minute increment the output is 166 grams. That the mill output was less for the third and fifth minute increments than for the fourth-minute increment is almost proof of experimental error, and the conclusion may safely be drawn that nothing is to be gained by overfeeding a ball mill, and that choking of the mill with finished product cuts down the mill efficiency.
A detailed comparison is made of the capability of population balance models to predict steady-state product size distributions of a pilot-scale ball mill. The mill was operated at 60% solids with feed rates of crystalline limestone ranging from 90 to 450 kg/hr (200 to 1000 lb/hr). Two types of lumped parameter models are compared: a linear model in which size reduction parameters are independent of size consist and a nonlinear model in which these parameters are dependent on size consist. The nonlinear model is based on an empirical correlation between rate of breakage and size consist in the mill. Results indicate that the nonlinear model gives the most accurate predictions of product size distributions, however, at the cost of significantly more complex computations.
The milling of talc, as is the case with many non-metallic minerals, until recently, has not received adequate technical consideration, for the talc industry has become of importance only within the last decade. At first, talc was used only in the massive form, for foot warmers, griddles, and so on, and milling methods were unnecessary. As the demand for ground talc increased, producers adopted the machinery used in the milling of flour; in fact, many talc mills were rebuilt flour mills. Improvement has been slow, but today several types of grinding and separating machinery are in use; many of them, however, are still inadequate and inefficient.
To determine the best methods of milling talc, it is necessary to understand the essential properties a talc must possess to fit it for a particular use. For toilet purposes, whiteness, freedom from lime and grit, fineness of grain, and good slip are essential. As a paper filler or coating, the talc must be white, uniform, and fine grained, have a good slip, and be free from grit and iron; freedom from lime is a disputed point. Fibrous talc is supposed to be superior to massive on account of the interlocking of grains in the paper, thus increasing its strength; this point, though, is in doubt. In general, talc for the paper industry need not be of as high quality or as fine grained as that used for toilet purposes.
Talc is usually bought by sample, for which reason it is difficult to state the essential properties for use in paint, rubber, roofing, and so on. The adoption of standard screen tests and standard grades by producers is necessary before an accurate basis for manufacturing or selling standards can be established. Off-color talcs might be utilized, as in Germany and Austria, by the establishment of standard grades of colored talcs.
The machinery to be used depends on whether the talc is fibrous, foliated, or massive; hard or soft; of uniform or variable grade; pure or impure. In some cases, it is possible to change the mining practice or to utilize different sections of the deposit in order to vary some of these factors, but in many cases these factors are fixed, so the milling must be designed to suit the conditions.
Crude talc from the mine, in many cases, is too moist and sticky to permit of efficient milling without first being dried. Sometimes it is dried in heaps exposed to the air; sometimes on a steam-heated, iron, drying floor; and sometimes in rotary, direct- or indirect- heat dryers fired with coal, coke, or steam. If direct-heat dryers are used, coke must be the fuel so that the color of the talc may not be impaired by smoke and soot. Where drying is necessary, rotary dryers are preferable.
The machinery for the primary reduction of talc rock does not differ essentially in various parts of the country. Crushing to about 1 in. is usually done in jaw crushers of the Blake type, although for large production gyratory crushers might be used to advantage. Rolls or rotary crushers reduce these lumps to about in. Between the primary and the secondary crushers there should be a rotary screen, which will remove material fine enough to pass the discharge opening of the secondary crusher.
The next step is accomplished by a variety of machines. The principal types used are the Raymond roller mill, the Fuller-Lehigh mill, various types of vertical and horizontal emery and burr mills, pulverizers of the swing-hammer and other types, so-called disintegrators, the Hardinge conical mill, and other continuous and intermittent types of ball and tube mills. These machines are followed by inclined vibrating screens, rotary screens, silk-bolting reels, or various systems of air separation; or the product may be bagged direct. Where screening or air separation is used, the product is conveyed to bins and then bagged. The product is carried from one stage of reduction to the next by bucket elevators, and belt or screw conveyors.
In the selection of all crushing, grinding, separation, and handling machinery care should be taken not only to provide sufficient capacity but some reserve; also that the capacities of the various machines are in proper proportion to one another. In one mill, the secondary crusher was too small for the rest of the plant, so that the mill capacity was lowered to the capacity of that crusher. Another mistake is to gage the capacity of a mill by the rated capacity of each machine running full time; it is impossible to maintain the maximum rated capacity for all machines simultaneously for any length of time.
The coarse and secondary crushing units are usually fairly efficient, but the full value of such machines is often not utilized. It is a fundamental principle that material already crushed fine enough should be removed before passing through the next reduction unit. Feed to a crusher making a 1-in. product should be freed from all material finer than 1 in. by screening. If allowed to go through the crusher, it not only reduces the capacity of the crusher for doing useful work but it tends to cushion the crushing effect on the coarser material. When belt conveyors are used to feed crushers, a magnetic pulley at the discharge end is advisable. This removes all tramp iron, and not only protects the machines from injury but keeps out fine iron which may contaminate the finished product.
Opinions have differed widely as to the most efficient type of fine grinding and separating machinery, but certain fundamental principles should be recognized. First, the feed to a machine grinding to 200 mesh should not be over in., and to 3/8 in. or finer is to be preferred. The very fine grinding of coarse material in one stage is not efficient, no matter what type of machine is used. Second, continuous operation is to be sought instead of intermittent, as in the case of certain types of pebble mills now in use.
Screening or bolting of material finer than 100 mesh is usually expensive, slow, and inefficient, though good results have occasionally been obtained, under careful competent supervision. Where dry grinding is used, some system of air separation is the cheapest and most efficient method of grading very finely ground talc. It has not been proved that fibrous talc from New York State, which contains fibers, flakes, and rounded grains, can be successfully separated by air, for, without fairly close sizing, large thin flakes settle in air at the same rate as small rounded grains. But here the problem may be solved by selective mining, close sizing before air separation, improvement in the design of air separation machinery, or the use of more than one separator in a closed circuit with the fine-grinding machinery.
The efficiency of ball or pebble mills for the fine grinding of talc is in dispute, but certain facts seem to be established. Pebble mills of the intermittent type, or dump cylinders, are not economical; they are slow and have a small capacity. The finished product is not uniform and its grade is dependent on the opinion of the man in charge, as it is not mechanically controlled. As the grinding progresses, the inefficiency increases: the most active grinding is done in the first hour, for the fine talc gradually cushions the impact of the pebbles and little grinding is done at the end. In one mill 5000 lb. of pebbles are used to one ton of talc in a mill revolving for 5 to 7 hr. at 22 r.p.m. In roller mills, which are continuous in operation, equipped with air separation, when the talc is ground sufficiently fine it is automatically removed, so that the full crushing effect is continuously applied to the coarser material. Such mills may have a capacity of about 5 tons per hour when producing a 200-mesh product.
Tube or pebble mills with continuous feed and discharge may be necessary in the milling of fibrous talc or talc mixed with tremolite; they may also be economical with certain other talcs, but the use of three or four large tube mills in tandem with no screening or separation between the mills nor for the sizing of the finished product is neither efficient nor economical. A good product is made by one talc producer by the use of short tube mills, but each mill is followed by bolting reels, which return the oversize for regrinding. In a plant being erected, a Hardinge mill and a large tube mill will be used with an air separation between them and another at the end of the tube mill; the coarse material from each air separator will be returned for regrinding. One plant practising close sizing between mills consumes 61 hp. per ton per hour; a plant that neither sizes nor separates between mills consumes over 375 hp. per ton per hour. Though there is some difference in hardness between the talcs ground in these cases, it is not enough to account for this difference in power consumption.
One argument given for the use of tube mills is that no iron produced by attrition of grinding surfaces can contaminate the product of a pebble mill lined with siliceous brick or pebbles. But iron can enter the pebble mill from the primary and secondary crushers. Furthermore, grinding machines in the most modern plants manufacturing the highest grades of toilet powders have all-iron grinding surfaces. It is stated that 0.5 per cent, of iron is not injurious in talc used for filling paper; that is, in milling a talc naturally free from iron, 10 lb. of metallic iron per ton of talc may be added without injury. In a mill producing 5 tons per hour, this would amount to 50 lb. of iron per hour, an amount of attrition that no mill could possibly produce and still be of practical use.
Where a large capacity at uniform fineness is desired, a vertical roller mill is probably the most efficient for soft talcs. Such a machine, with automatic feed, in which the product is continuously removed by a current of air introduced from outside, is used in some of the best designed plants. For smaller capacities of fine material, swing-hammer and similar types of pulverizers and disintegrators and vertical emery mills equipped with air separation are used. Some types of these machines are suitable only for coarse material, 40 to 100 mesh, but when equipped with air separation, these mills may be used to produce a certain percentage of 200-mesh product. The tailings from the air separator must be carefully reground or sent to screens making a coarse product. The tailings from all mills can be continuously returned for regrinding, but as there is a demand for coarser material, they are often sized and sold to the roofing industries. The best type of screen for this purpose is the flat, inclined, vibrating screen.
The inefficiency of horizontal burr mills, compared with vertical emery mills, for the grinding of most talc, seems to have been definitely established, as they have been discarded by nearly all producers.
The best type of hammer mill, pulverizer, disintegrator or emery mill for the grinding of a talc from a new deposit can be determined by a careful comparison of the physical properties of this talc with those of other talcs now successfully ground; or by actual tests made by companies selling grinding machinery. But whatever machines are used, the product from the last grinding machine should be carefully sized either by bolting or by air separation. No grinding machine can be relied on to produce continuously a fine uniform product without screening or air separation. Furthermore, the final, sized product should be tested frequently by hand screening. Probably nothing has injured the talc industry so much as the marketing of non-uniform, improperly sized product.
With the Raymond, the Sturtevant, or similar air separation systems, a small amount of almost impalpable dust is produced. This is usually of very high grade but it cannot be obtained in large enough quantities to warrant its exploitation. No machine now on the market can produce by air separation such fine material in large quantities at a cost that will permit its sale at a reasonable price, for with increasingly fine grinding, the machine capacity is rapidly reduced, proportionately increasing the cost per ton of product. But one plant has found that a somewhat larger production of this grade may be obtained at only a slightly increased cost, by using a secondary system of dust collectors in conjunction with the primary air separation system. If the demand for very finely ground material, that is through 300 mesh, increases it may be necessary to use wet grinding and water classification.
Most of the older plants are steam driven, but there is a tendency to install electric drives in new mills and in the remodeling of old mills. Individual motor drives for each important machine have many advantages; in one mill, each motor has its own voltmeter, ammeter, and circuit breaker. This permits an accurate record to be kept of the performance of each machine and is valuable in checking efficiency and in estimating the cost of producing different grades of talc. A recording wattmeter would also be valuable.
A point in mill design often overlooked is that of providing ample bin capacity for crude talc, crushed rock, and for several grades of finished products. Most mills have adequate capacity for one of these, but rarely for all three. Bins are not expensive but they are of great value in insuring steady production when the mine is temporarily closed or when sections of the mill are shut down for a few hours. Among the objections to the use of bins for finished products are: that if through accident, improperly ground material is made a large quantity of good talc in the bin is contaminated; and that certain talcs when finely ground are very sticky and will not flow freely from a bin. The first objection may be removed by using a bin divided into compartments and frequently checking the product by screen tests. The use of several compartments is advantageous, in any circumstances, in order that several grades of product may be made. Frequent screen testing is likewise desirable in order that a high quality may be maintained; in several mills screen tests are made every hour. The second objection may be removed by heating the inside of the bins with exhaust steam, or by installing some system of mechanical stirring, or agitation. Finely ground material if dry and warm will often flow freely, when it will not flow at all if cold and slightly moist.
There are no definite specifications for the selling of talc and I have not found a consumer who could tell exactly what he wanted. He tries a sample and if that is satisfactory he uses that talc. In fact, the tests are so poor that consumers have refused one kind of talc and accepted a talc made in just the same way, saying that it was satisfactory.
The first thing needed is to devise specifications. Certain consumers demand a talc with a slip, but no one can define or measure slip. There are no specifications whatever for those tests. So, first, the tests must be devised. The idea that fibrous talc must be used in the paper, paint, and other industries was probably originally fostered by some producers of that material. As a matter of fact, the fibrous talc is not so well fitted for paper as the granular talc, for unless it is carefully prepared the fibers are apt to pierce the paper and make a bad spot. Less fibrous than granular talc is now used in paper. Some paint men object to fibrous talc because the fibers, if long, are apt to upend themselves in a film of paint, and be brushed off, leaving an opening. The talc men have not taught the consumers the uses of the various varieties of talc.
The grinding test referred to has not been completed. All the screen tests have been made. The silica determinations have not been made, but from microscopic examinations of material remaining on the screens it seems probable that a method can be devised whereby probably 80 per cent, of the silica can be removed. The talc in question was foliated and when ground broke up into fine, thin scales, whereas the quartz was in little, rounded grains. I believe it is possible to grind that talc in either a Raymond roller mill, by which the coarser particles can be removed continuously from the inner cone, or by some form of a hammer mill equipped to throw out the heavier particles. I think that by one of these methods most of the grit can be eliminated.
If the talc is granular, I doubt if this could be done because the specific gravities of talc and silica or quartz are about the same, and when ground probably the grains would be about the same size. It would be difficult to separate grit or silica from talc in the case of a granular talc. Wet grinding and water flotation have not been practiced at all. Some experiments have been made, but no definite results have been obtained. I think that eventually the finer grades of talc, possibly the 300 mesh and finer, will have to be made by this method. I understand, however, that by air separation they are able to make a product of 350 mesh; but that would not apply to all talcs.
The results of experiments on a 40 cm 40 cm grate-discharge ball mill have been analysed for variation of mill hold-up weight of solids with solids feed rate, weight percent solids, mean feed particle size, material specific gravity and work-index. It is shown that mill hold-up weight is independent of the material specific gravity and mean feed particle size, and it varies linearly with solids feed rate and weight percent solids, at least over the range of practical interest. The variations in the transport behaviour of different materials have been attributed to the differences in the size distribution of the mill hold-up solids. It is shown that work-index can be used as the material characteristic for the development of an empirical correlation. Variation of mean residence time of solids with solids feed rate and weight-percent solids is also discussed.
The present is to discuss some of the current silver-treatment plants and also reviews briefly some of the older practices in important silver-mining areas since closed down.The greater part of the worlds production of silver is derived from the refining of the base metals, particularly lead ores, and complex ores of lead, copper, antimony, and zinc. Most of these ores are concentrated by flotation methods, and the concentrates smelted.
The previous pageshave been devoted to the treatment of gold and silver ores in which the recovery of silver, because of the relatively small amount present, is not ordinarily of economic importance. There are, however, certain mining areas where the recovery of the high silver values is or has been the principal metallurgical problem.
The ores of the Cobalt area were remarkable for their high content of silver and for the complex assemblage of minerals found in the veins and enclosing rock. Of the silver-bearing minerals, native silver was of outstanding importance, as fully 97 percent of the values occurred in this form. It was found in masses ranging from large slabs to the finest, filmy leaf. Other minerals included cobalt and nickel in the form of arsenides, sulphides, antimonides, and various combinations of these, associated and oftenintimately mixed with a number of base-metal compounds.
A variety of methods were used for treating high-grade ore and concentrates, including the amalgamation and cyanide process, the hypochlorite-cyanidation process, the sulphuric acid-cyanidation process, and chloridizing roasting, while the lower grade material was treated by a combination of gravity concentration and flotation or cyanidation.
The ore was oxidized and siliceous, the principal constituent of the gangue being quartz with some calcite. The silver minerals contained in the ore were principally argentite and cerargyrite, the former predominating. The lead minerals, all of which were argentiferous, were chiefly cerussite and galena, with occasionally a little anglesite. The gold was free, but most of the ore contained merely a trace.
The milling scheme included tabling at 10 mesh followed by regrinding to 80 to 90 per cent minus 200 mesh and cyanidation by Pachuca agitation. The solutions were maintained at 2.8 to 3.2 lb. per ton NaCN, and the reagent consumption was 6 lb. lime, 2.5 lb. NaCN, and 0.25 lb. zinc dust per ton of ore. Table concentrates averaged about 426 oz. silver per ton and carried 52.5 per cent lead. The zinc precipitate analyzed 20,000 oz. silver and 5.60 oz. gold per ton and carried 1.02 per cent zinc and 24.1 per cent lead. Both products were shipped to Carteret, N. J.
The property was purchased by the Mexican government from the United States Smelting, Refining and Mining Co. in September of 1948. For several years past the tonnage and grade of ore has been dropping and now stands at about 100,000 tons per month, assaying 300 grams silverand 3 grams gold per ton. The higher ratio gold than previously is due to the discovery some 5 years ago of a new vein carrying about 10 grams gold per ton.
Abstracted from: Oxygen as an Aid in the Dissolution of Silver by Cyanide, R.I. 3064, U.S.B. of M. and Flotation of Silver Minerals, R.I. 3436, U.S.B. of M. Ag dissolves according to the reaction:
Where argentine is intimately mixed with pyrite, sphalerite, and gangue, roasting for 1 hr. up to 460C was necessary for a +75 per cent extraction. Time is the important factor for argentite when pure. If temperature of roast exceeds 600C an insoluble silver silicate is formed.
In the case of polybasite, silver may be partially replaced by copper and antimony partially replaced by arsenic. Not all samples of tetrahedrite are as refractory as the one tested. Some yield up to 83 per cent extraction.
Two-stage grinding in cyanide solution is practiced. For primary grinding, 8- by 6-ft. Marcy grate mills and 6- by 12-ft. trunnion Traylor ball mills are used in closed circuit with 6- by 22-ft. Dorr classifiers, which are the only type operated in Pachuca. Nearly 80 per cent of the feed to these mills is coarser than 3 mesh and up to 1 in. The classifier overflow is 64 per cent solids. For secondary grinding, 6- by 10-ft. Traylor mills of the trunnion type and 5- by 10-ft. trunnion mills of local make are used in closed circuit with an 8- by 22- and a 6- by 22-ft. classifier, respectively. The classifier overflow contains 20 per cent solids; a sieve test of the final product shows on 48 mesh 2.20 per cent; on 65, 7.90; on 100, 9.17; on 150, 13.13; on 200, 8.09; and through 200, 59.51 per cent.
Ten Dorr thickeners, 48 by 15 ft., yield a pulp of 45 per cent solids. Between 1934 and 1945 the ore became more difficult to settle and underflows dropped to as low as 30 per cent solids. The trouble was largely overcome by removing about 1000 tons per day of plant solution and replacing with fresh cyanide solution. The solution removed is plant barren solution, which is first passed through the regeneration plant to recover its cyanide, silver, and gold content and then discarded with the tails.
Eighteen Pachuca tanks, 15 by 60 ft., and 32 flat tanks, 20 and 24 by 30 ft., do the agitating. The latter is a tank equipped with a Dorr-thickener mechanism and air jets. Air at 35 lb. pressure is used in the Pachucas and at 18 lb. in the flat tanks. Agitation proceeds, for 73 and 70 hr., respectively.
Aero-brand cyanide is dissolved in barren solution to make a strong solution, and this is added to the agitators to bring the strength to 0.17 per cent NaCN. Litharge is added in the dissolving tank to eliminate soluble sulphides. Cyanide consumption, excluding regeneration, amounts to 1.62 kilograms perton ore. Lime consumption is 9.0 kilograms.
Butters tanks, each with 187 leaves, 67 by 117 in., do the filtering. Each tank averages 11 cycles of 128 min. each day, and each cycle is divided into 26 min. for caking, 38 min. for barren wash, 15 min. for water wash to mill, and 20 min. for water wash to regeneration, the remaining 29 min. being required for filling transfers, discharging, etc. A vacuum of 18 in. is maintained. Average cake is 7/8 in. thick.
Solution from the filters is clarified in 12 Sweetland presses, which can handle 2 tons per day per square foot of surface. They are discharged twice and cleaned once each day, and leaves are acid-treated every 10 days.
The Merrill-Crowe system of zinc-dust precipitation is used. Centrifugal pumps force the solution through the presses. Zinc consumption is 170 grams per ton of ore. The dried precipitate assays 83 per cent silver and 0.46 per cent gold, also 0.25 per cent selenium and some other metals.
Precipitate is melted to bullion in the usual manner, granulated borax and bottle glass being used, in an oil-fired reverberatory furnace of 15 tons capacity. The temperature is raised to 1050C., and slag is skimmed off. Air is then blown in, and the slag is skimmed for 60 hr. Then the metal is tapped into a continuous anode-casting machine. Anodes weigh 10 kilograms, and a furnace charge makes 2000 of them in 5 hr. of casting. The bullion is increased in fineness from 950 to 993; copper is the principal impurity remaining.
The anodes are next parted in 200 Thum-type electrolytic cells, and the resultant silver is 999 plus fine. The gold mud is reduced to anodes, which are treated in Wohlwill cells, giving gold 999.8 fine.
The current extraction (1948) now averages 85 per cent of the silver and 90 per cent of the gold contained in the ore. Cyanide Regeneration. The plant for the regeneration of 3800 tons of cyanide solution per day is describedearlier.
In Tonopah there are five mills: the Belmont, Extension, MacNamara, Montana, and West End, while at Millers, 12 miles north, are the Belmont and Tonopah mills, ore being shipped to these at a cost of 70 cents per ton. In nearly every case gyratory crushers are used for breaking ore as it comes from the mines, the procedure being to crush first in a large crusher up to the No. 7 type K Gates size and pass through revolving trommels, the oversize being again reduced in No. 3 size gyratories, the final product for the stamps being about 1 in. Sorting is done at the Belmont and MacNamara mills, at the former on a pan conveyor from which 15 per cent is rejected, and at the latter on a 30-in. rubber belt from which 6 per cent is sorted out. From the crushing department, the ore is taken to mill bins by 20-in. belt conveyors, or bucket elevators, and distributed by the usual automatic devices.
There is no amalgamation at Tonopah, nor is it necessary on this class of ore. Crushing is done in weak and warm (from 50 to 80F.) cyanide solutions, so the ore is in contact with solution from the stamps to filtration. This is necessary as well as the heating, which, although somewhat expensive, quickens the solution and accelerates the dissolving action. Solutions are usually heated to about 95 and in one case to 120 by live steam introduced in the agitators.
The practice of using hot solutions is briefly as follows: At the new Belmont mill the temperature at the stamps is from 60 to 70F., and at the Pachuca agitators exhaust steam from the mill air compressor is fed in, increasing it from 90 to 100. In the M. and S. Press of Jan. 27, 1912, A. H. Jones, metallurgist at this plant, gave some valuable data on this subject. On an ore carrying 0.05 oz. gold and 18.2 oz. silver per ton, 60 hr. agitation with both 60 and 90 solutions, the tailing averaged 0. 0175 and 3.45 and 0.0125 and 1.90 oz., respectively. Tests on 48 and 69 hr. at similar temperatures gave as marked results. Besides the effect on extraction, the hot solutions flowing through the mill kept the whole place at a good working temperature. At the Montana-Tonopah, ore is crushed in 50 to 60 solution, which is increased to 110 at the Hendryx agitators by live steam. It is found also that the heat aids settling. There is a marked decrease in extraction without hot solutions.
Tonopah ores carry as much as 3 per cent pyrite, but concentration is not always employed, it being done only at the Belmont, Montana, Tonopah, and West End. It would seem that, if the grade of the ore and percentage of mineral are not too high, tables are not necessary, and this varies from time to time in the various plants. At any rate, a very close saving is not attempted. The Extension Company dispensed with their Deister tables, selling them to the West End. The Belmont, Montana, and Tonopah use Wilfley tables. Concentrate is collected, steam dried in large trays, sacked, and shipped to smelters. Freight and treatment cost nearly $70 per ton.
All-sliming is the standard method, with the exception of the Tonopah mill at Millers, where three products are made: concentrate, sand, and slime. At this plant reduction is by stamps and Chilean and Huntington mills, while at Tonopah the procedure is as follows: The pulp from the stamps is fed into Dorr duplex classifiers making 12 strokes per minute, from which slime overflows and coarse material is fed into tube mills by means of a special feeder. Discharge from these is elevated to the Dorr classifiers, where a further classification takes place, followed by further grinding in the tube mill, and so on.
Various types of thickeners or dewaterers are in use, the practice being to allow the clear solution to overflow and decant off as much as possible for battery storage. When it gets too high in gold content, it is decanted to the tank for precipitation. As at many other mining centers there is quite a difference of opinion regarding the efficiency of agitators, the Trent being used at the MacNamara, Montana, and West End; the Hendryx at the Montana; Pachuca tanks at the new Belmont mill; and ordinary mechanical agitators and air lifts at the Belmont and Tonopah at Millers,these being in series at the Belmont plant. Centrifugal pumps and air at about 20-lb. pressure are used for the Trent system, and better results are obtained if pulp is drawn off near the top of a full vat and pumped through the arms as usual. Agitation proceeds for upward of 48 hr. At the new Belmont mill, slime is first agitated in six Pachuca tanks, and from these it is elevated to Dorr thickeners by an air lift, prior to going to another set of six Pachucas, making a total of 48 hr. agitation, the idea being to get rid of as much valuable solution as possible before sending slime to the filter plant. Cyanide and lead acetate are added to the agitators, the former being from 2 to 5 lb. solution, while regular addition of the acetate is found necessary at all mills. Lime is usually slaked and added to the tube-mill feed. Consumption of chemicals at the Extension is as follows:
Agitated slime is drawn off to stock tanks, which serve the purpose of storage from agitators and excess from filter plants. The latter have little of special note about them, being of the ordinary stationary leaf type which has been described so often in technical papers.
Zinc-dust precipitation is used at the new Belmont and Montana mills, and zinc shavings at the Belmont, Extension, MacNamara, Tonopah, and West End. Methods of dealing with precipitate vary somewhat. At the new Belmont precipitate is dried, mixed with 5 per cent borax, and smelted in double-compartment, oil-fired Rockwell furnaces lined with carborundum, kaolin, and water glass. At the Extension it is dried, fluxed, and smelted in oil-fired Steele-Harvey tilting furnaces which contain a No. 250 graphite crucible, while at the Tonopah mill the fine zinc- shaving precipitate is incompletely dried, mixed with crude borax which swells up through the mass, and then smelted in six coke-fired tilting furnaces. Crucibles last from 90 to 130 hr. and are turned once. Tonopah bullion will average 950 fine in silver and a trifle over 10 in gold and is sampled by being bored at opposite corners of top and bottom bars. The bullion is shipped by freight like any other merchandise.
The Montana Tonopah closed its 500-ton mill in 1923, and thereafter no operating mill in the district employed concentration. The average gold extraction in this district was 94 per cent, and the average silver extraction 92 per cent.
The Sunshine mill operated by the Sunshine Mining Company is situated 6 miles from Kellogg, Coeur dAlene district, Idaho. The mine is the largest silver producer in the United States. In 1937 its output was 12,147,719 oz. silver and 2,784,289 lb. copper. In 1947, on a curtailed basis due to labor shortage, it produced 5,034,160 oz. silver, 1,249,555 lb. copper, and 5,881,796 lb. lead. The 1200-ton mill employs the straight flotation flow sheet shown in Fig. 96, the concentrate being shipped to a lead smelter. In 1947 the ore averaged 44.5 oz. silver (associated with galena and tetrahedrite, Cu8Sb2S7), 0.55 per cent copper, and 2.61 percent lead. The recoveries were 98.53 per cent of the silver and 98.04 per cent of the lead. Milling costs were 95 cents per ton.
The flotation reagents used are 0.13 lb. per ton butyl xanthate and 0.75 lb. per ton frother. The frother is a mixture of 1 part Barrett No. 4 with 3 parts methyl amyl alcohol.Average concentrate analysis for 1947 is shown in Table 93.
The New York and Honduras Rosario Mining Company operates two mills in Honduras, the Rosario and Mochito mills, and the El Dorado mill in El Salvador. The following information in regard to the latest practice at these mills was supplied to the author through the courtesy of the president of the company, W. A. Prendergast.
Rosario Mill (Type IIa). The mill treats 550 tons daily of an ore carrying 13.25 oz. per ton silver, 0.071 oz. per ton gold, 0.5 per cent zinc, 0.5 per cent lead, and 2.0 per cent manganese. Primary crushing is carried out in two gyratory crushers making a 2-in. product, 350 tons of which is crushed in twenty 1800-lb. stamps and 200tons in a 6- by 5-ft. Allis-Chalmers ball mill charged with 5-in. alloy-steel balls. This mill is in closed circuit with a 5- by 25-ft. 6-in. DSFXM Dorr classifier overflowing a 35-mesh product. The stamp milling is carried out in cyanide solution (3 lb. KCN per ton of solution). The product passing the -in. battery screens is dewatered in two 6- by 20-ft. Dorr DSC classifiers, the overflow going to thickeners and the underflow to two 5- by 9-ft. ball mills in closed circuit with two 6- by 18-ft. DSC Dorr classifiers, also overflowing a 35-mesh product. A rationed ball charge of 70 per cent 3-in. and 30 per cent 4-in. moly-chrome alloy balls is used.
The minus 35-mesh product from both classifiers flows to an 8- by 6- by 20-ft. DSF classifier which is close-circuited with two 5- by 9-ft. ball mills using a 2-in. ball charge. The final pulp runs 26 per cent plus 150 mesh. The total steel consumption for crushing and grinding is 1.86 lb. per ton of ore milled.
Four 35 by 10-ft. Dorr thickeners and one 35- by 15-ft. Dorr balanced-type tray thickener produce pulp underflows, by means of direct-connected4- in. Dorrco diaphragm pumps, of 40 per cent solids. This thickened pulp is agitated for 83 hr. in batches in eighteen 15- by 45-ft. Pachuca tanksand in three 35- by 10-ft. Dorr mechanical agitators. The air pressure in the Pachucas is 35 lb. per sq. in., and about 95 cu. ft. per min. is used. The cyanide is added to the Pachucas to maintain a strength of 4.6 lb. KCN per ton of solution, and lime is held at 0.8 lb. per ton of solution. The cyanide consumption is 2.956 lb. KCN per ton of ore, and the lime consumption is 15.21 lb. of crude lime of 8.15 lb. CaO per ton of ore.
Filtration is done in three Merrill center washing slime presses with one hundred 3-in. by 4-ft. by 6-ft. frames, the plates covered with 8-oz. sail canvas, which has a life of 1100 charges or 79 days. The press cycle consists of charging with pulp for 10 min., a barren solution wash of 28 min., a water wash of 34 min. under 55-lb. pressure, and sluicing of the presses for 20 min. with water at 75-lb. pressure on the nozzles. For the washes at the presses 825 tons of barren solution and 1,025 tons of water are used; 2,425 tons of water are used for sluicing the presses. The dissolved- values loss in the tailings are 7 cents in silver and 3 cents in gold.
The precious metals are precipitated from the solution by means of zinc duct of which 0.4735 lb. per ton of ore milled or 0.03929 lb. per fine oz. of bullion is consumed. The pregnant solution averages about $2.88 per ton, and 2000 tons is precipitated per 24 hr. The effluent carries a trace of the metals.
Thirty-five per cent of the silver and 71 per cent of the gold are dissolved in the grinding circuit, 55 per cent of the silver and 24 per cent of the gold during agitation and 0.8 per cent of the silver and 0.7 per cent of the gold in the filters.
In 1948 the Mills-Crowe cyanide recovery process regenerated 174,066 lb. KCN from the barren solution. This enabled the carrying of a high cyanide strength in the agitators, a longer water wash on the Merrill filter presses, and a low mechanical loss in the tailings from the filters.
This mill, which is located at Mochito, near Lake Yojoa, treats 100 tons per day of a high-grade silver ore carrying about 39 oz. silver per ton. It has a relatively high manganese content (4 per cent) mostly in the form of pyrolustite. The silver occurs with lead and zinc sulphides.
The ore is delivered from the mine to the mill, a distance of about a mile, by means of Diesel trucks. The ore is crushed to a 1-in. product through a No. 3 gyratory crusher. Primary grinding is done in a 6 by 5 Allis-Chalmers ball mill in closed circuit with a Denver mineral jig and a 4-ft. by 18-ft. 4-in. Dorr DSFH classifier, and the overflow carries about 10 per cent plus 150 mesh in the pulp. The pulp from the secondary classifier is thickened in a Denver 38- by 10-ft. thickener, the underflow going to eight Massco Fahrenwald flotation cells, five of which are used as roughers, two as cleaners, and one as recleaner. Pine oil, Aerofloat 31, reagents 404 and 301 are used. The pH is maintained at about 7.6. Flotation tailings are thickened in a 38- by 10-ft. thickener to about 40 per cent solids, the underflow being elevated by Oliver slurry pumps to two Denver disk filters, 6-ft. diameter and five leaves each.
The cake from the filters, which carries about 20 per cent moisture, is repulped in barren cyanide solution from the precipitation plant and is then aerated in an 8-ft. Denver agitator, where the lime emulsion is added. Cyanide is then added to the pulp as it flows to agitation. The pulp is agitated 63 hr. in six 12- by 36-ft. Pachuca tanks, the cyanide being- maintained at about 6 lb. KCN per ton of solution and the lime at 0.8 lb.
The pulp from the Pachucas then flows to four 38- by 10-ft. countercurrent washing thickeners. Two 8- by 10-ft. Oliver filters are now beinginstalled for filtration of the pulp from washer 4 in order to lower the mechanical loss in cyanide, which is high because of the high cyanide strength required during agitation.
The washing is done in cyanide solution by means of a 4- by 19-ft. washing trommel with a 1-in. punched-plate screen. The undersize is dewatered in a Stearns-Roger dewatering drag 4 by 26 ft., the overflow of which, containing about 11 per cent of the tonnage, is pumped direct to the primary thickener. The trommel oversize is discharged onto a 24-in. picking belt where waste is hand-picked and the ore delivered to a No. 50 Kue-Ken crusher set to crush at 1 in. The crusher product plus the dewatered drag sands are delivered to the fine ore bin at the grinding plant by means of a 16-in. by 280-ft., 15-deg. inclined conveyor.
Primary grinding is done by means of a 6- by 5-ft. Marcy ball mill in closed circuit with a 3- by 15-ft. Wemeo screw classifier. The secondary grinding is carried out in a 4- by 10-ft. ball mill charged with 1.5-in. balls and in closed circuit with a 6-ft. by 21-ft. 4-in. model F Dorr classifier, the overflow being allminus 150-mesh product and flowing directly to the primary thickener, which is 38 ft. 7 in. in diameter by 10 ft. deep. The underflow is maintained at 40 to 45 per cent solids and is agitated in three 21-ft. 6-in. by 16-ft. mechanical Wemco agitators with an air pressure of about 12 lb. on the air lift. These are in series and flow into four 38-ft.7- in. by 10-ft. washing countercurrent thickeners, all underflows being maintained at 40 to 45 per cent solids.
Owing to the colloidal slimes present in the ore, it has been found necessary to add to the washing solution 0.015 lb. caustic starch and a lime emulsion before washing the ore in order to get flocculation of the slimes. About 10 lb. of lime is consumed in washing the ore.
Oxidized silver ores containing the higher oxides of manganese are generally refractory to metallurgical treatment. Manganese fouls mercury if amalgamation of the gold content is attempted. A refractory compound of manganese and silver is formed, probably a manganite, which is insoluble in cyanide solution and other common solvents for silver.
The Caron Process (U. S. Patent 1,232,216, Aug. 3, 1917), described by G. H. Clevenger and M. H. Caron in Bul. 226, U.S.B. of M., is based on the following principle: When oxidized ores containing a refractory compound of manganese and silver are heated in a reducing atmosphere, the higher manganese oxides are reduced to manganous oxides, and if cooled so as to prevent reoxidation, the refractory compound is rendered amenable to cyanidation. Refractory compounds of silver also can be so treated.
Manganese-silver ores occur generally in acid-eruptive rocks, chiefly rhyolite and dacite flows of later Tertiary age. Potassium-aluminum silicate is a vein material, and the vein quartz replaces the calcite. The manganese oxide is generally of secondary origin and is formed by atmospheric agencies. For the foregoing reason manganiferous ore from near the surface may be refractory but from depth may be amenable to treatment. Wad, a hydrous manganese manganate, is common in the zone of oxidation.
Various treatments of the raw ore have proved unsuitableconcentration (including flotation), magnetic separation, chloridizing, roasting, volatilization, sulphuric acid, and heating with organic matter. The Ag to Mn ratio persists in all sieve sizes from plus 20 to minus 200 mesh.
Laboratory tests were made in the United States and in Sumatra, followed by plant-scale runs in the latter country and a 50-ton plant at Pachuca, Mexico. Direct cyanidation of raw ore containing 2 to 10 per cent MnO2 gave 50 per cent extraction of the silver, but ore with 25 per cent MnO2 gave only 25 per cent extraction. The Caron process, on the other hand, extracted 92 per cent of the gold and 90 per cent of the silver. The pilot plant in Mexico successfully treated ore containing 2.8 to 13 per cent MnO2 and 12 to 20 oz. Ag. The Clevenger kiln (U. S. Patent 1,379,083, May 24, 1921) was fired with producer gas with the following analysis: CO, 15 per cent; CH4, 5.5 per cent; H2, 4.6 per cent; CO2, 6 per cent, the remainder being nitrogen.
Caron described the application of the process at Tambang Sawah, Sumatra, Dutch East Indies, a translation appearing in the M.J. (London) Feb. 19, 1927. The ore, mainly quartz, carried 18 per cent manganese dioxide, 30 oz. silver, and 8 dwt. gold. After being crushed to 1 in., it was heated 4 hr. in a rotary kiln. It remained 1 hr. in the reducing section of the lain in producer gas at 600C. The MnO2 was reduced to 2 per cent. The ore was then ground and cyanided, yielding 87 per cent of the silver and 97 per cent of the gold, as compared with 25 per cent by raw treatment. Chemical consumption was 2.2 lb. cyanide, 5.5 lb. lime, and 1.3 lb. zinc per ton ore. A very complete bibliography covering the treatment of manganese-silverores is given in Bul. 226, U.S.B. of M., 1925, by Clevenger and Caron.
McClusky Process. Manganese is found in varying percentages in the ores at Fresnillo, Mexico; although, fortunately, the average content is not enough to necessitate special treatment, ores from some parts of the mine contain sufficient manganese to affect seriously the extraction of the silver. To improve the extraction on this relatively small quantity of refractory ore, S. P. McClusky, formerly metallurgist with the Fresnillo company, developed a modified method of what has become known as the sulphur dioxide process for manganiferous silver ores. This is described by W. E. Crawford in Trans. 112, A.I.M.E., 1934, as follows:
The content of silver in the ore and the gain in extraction by sulphur dioxide treatment, plus the cost of the process, are the criteria by which the applicability of the process to manganiferous silver ores may be judged. At Fresnillo argentite is the predominant silver mineral. It is associated with pyrite and manganese minerals.
In practice, the ore is ground in a weak, spent solution of 0.008 per cent KCN so that 30 or 35 per cent passes through a 200-mesh sieve and 1 per cent is coarser than 10 mesh. The pulp then flows to a 4-in. Wilfley centrifugal pump, which delivers the pulp to the top of the first and second of the sulphur dioxide treatment towers, which are sealed, airtight chambers, three in number. These towers are of wood, 3 by 3 ft. in cross section and 17 ft. high, with wooden baffles lined with white- iron plates. The purpose of the baffle plates is to disperse the pulp as it falls through the tower, so that it may come into intimate contact with the ascending current of SO2 gas. The flow of the pulp and SO2 gas is countercurrent, the pulp is constantly enriched in acidity and the gas mixture progressively depleted of SO2 from unit to unit. The final result is that the gas exhausting to the atmosphere contains slightly more than 1 per cent SO2 indicating a total absorption of 85 per cent of the available SO2. The gas used in this process and in the cyanide-regeneration plant at Fresnillo is produced by roasting the pyrite from flotation in seven-hearth Herreshoff furnaces.
From the SO2 absorption towers the pulp passes through five conditioner tanks arranged in series. An emulsion of lime is added to the fourth tank for the purpose of precipitating the dissolved manganous and ferrous compounds, as manganous and ferrous hydrates. The low-pressure air in this and the fifth tank assists in oxidizingthe manganous and ferrous compounds to manganic and ferric compounds. After passing through the last conditioner tank the pulp is returned to the mill for regrinding in a 6- by 14-ft. Traylor ball mill in closed circuit with a Dorr bowl duplex classifier. The overflow of this classifier, which averages 60 per cent minus 200-mesh material, joins the feed of the plant treating the regular silver ore.
The gain in extraction accomplished by the sulphur dioxide treatment varies considerably with different ores, but it appears to be in direct proportion to the amount of manganese dissolved by the gas, approximately 7 grams silver for every 0.1 per cent dissolved manganese. An increased recovery by this treatment, of 25 grams silver per ton, represented a substantial economic advantage when silver was quoted at around 30 cents (United States currency) per ounce.
The laboratory pilot test, carried out daily in conjunction with the plant treatment, often showed as much as 35 grams additional recovery of silver. Mixing of the SO2 treated slimes with the general mill slimes made it difficult to check the actual additional recovery in the plant.
An interesting point is noted in connection with tests for the oxygen content of solution in the pulp leaving the final treatment tank of this unit. This solution is entirely devoid of free oxygen; moreover, it required several hours of vigorous agitation with air to satisfy the oxygen-consuming requirement and to render it susceptible to the absorption of free oxygen. In view of this, it is quite possible that a separate cyanide circuit for these treated slimes would be a distinct advantage, especially if it were so designed that several hours of agitation and aeration could be given prior to the addition of cyanide.
There are several approaches to the problem of silver recovery from manganiferous ores. These are listed with the knowledge that particular ores may be economically amenable to some unmentioned processes and that not all attempted processes are known to the author.
In all the above methods where cyanidation is used, product recovery is frequently accomplished by zinc dust precipitation with subsequent fire refining. In cases of gold-bearing ores, concentration of values from low-grade cyanide solutions by adsorption on activated charcoal with subsequent electrowinning of a dore bullion is a useful method but in the case of silver ores running several or more ounces per ton, the sheer bulk of the necessary carbon points to the direct zinc dust precipitation method as preferable. An alternate precipitation method involves the use of NaHS or Na2S as suggested by Reno USBM to precipitate a silver sulfide product. Use of this method has been held back by poor settling and filtration rates for the silver sulfide precipitate and the insidious toxicity of possible hydrogen sulfide gas, to say nothing of hydrogen cyanide.
The recovery methods employing sulfur dioxide to dissolve the manganese offer the possibility of a manganese recovery step. This may be economic in those ores containing substantial manganese values. Parenthetically, it is felt that even very low values of manganese may be sufficient to render the silver difficult to dissolve with cyanide if the manganese and silver are syngenetic.
Fines production kinetics in batch grinding analyzed using the PBM framework.Well-known functional forms used for describing material breakage characteristics.Conditions for different patterns of variation with grinding time established.Variation of Bond Work Index with the classifying (closing) screen explained.Correct approach to determination of the breakage distribution function discussed.
The rate of production of fine material in the batch mode of grinding operation forms the basis for determination of the grindability parameter of the Bond approach and the breakage distribution function of the population balance model (PBM) approach to the mill scale-up design. For a given set of mill operating conditions, the rate of production of fines is determined by the breakage characteristics and production history of the material being ground. Another important aspect is the variation in the rate of production of fines with grinding time. With a view to developing a clear understanding of these aspects, a detailed analysis of variations in the rate of production of fines was carried out using the PBM framework and two well-known functional forms for the specific breakage rate and breakage distribution parameters. In this paper, it has been shown how the results of this analysis can be used for: (i) obtaining more accurate estimates of the breakage distribution parameters by performing just one short-duration batch grinding experiment, and (ii) explaining variation in the Bond Work index with the product size in terms of the exponent of particle size in the expression for the specific breakage rate function: kj=Axj.Get in Touch with Mechanic