US2994581A - Method of processing pegmatite - Google Patents

Description

Aug. 1, 1961 A. M. THOMSEN 2,994,581

METHOD OF PROCESSING PEGMATITE Filed June 4. 1959 By mfpite mm all:

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flecaw bv ef IN V EN TOR- I xt s (arbanator i i United States Patent 2,994,581 METHOD OF PROCESSING PEGMATITE Alfred M. Thomsen, 265 Buckingham Way, Apt. 402, San Francisco, Calif. Filed June 4, 1959, Ser. No. 818,221 1 Claim. (Cl. 2318) In the instant application I have described the way in which I make from a pegmatite a silica skeleton similar to the one derived from a substituted magnesium silicate, but with the double advantage that the commercially valuable materials simultaneously extracted from the pegmatite pay all expenses, and that both the internal area and the porosity of said skeleton is subject to much variation depending upon the technique employed. Contrariwise, when operating upon a natural product we are compelled to accept what nature gives, we may decrease but we cannot enhance the inherent characteristics.

While pegmatites are exceedingly variable as to minor constituents they are composed principally of some type of feldspar. In the drawing I have represented a preferred version of my process, and the diversity between my procedure and that of others is best understood after I have explained said drawing with comments upon strategic points. I will, therefore, commence by expounding what I have represented in said drawing.

The pegmatite, being a hard, granitoid rock, generally coarsely crystalline, is first ground to a 4060 mesh product and passed over a Wilfley table or its equivalent, to obtain as a concentrate most of the heavy minerals of the pegrnatite. Such minerals are magnetite, ilrnenite, rutile, monazite, tantalite, microlite, euxenite, Wolframite, zircon, garnet, staurolite, uraninite, and a host of others. Individually, each ingredient is present in minute amounts but, collectively, this concentrate can represent considerable economic advantage. It is impossible to specify the actual amount of this concentrate but it will rarely be below 5% or above of the weight of the pegmatite. Just how this concentrate is treated is not represented on the drawing but it will be described later on after the procedure on the feldspar fraction is better understood.

The tailings from the Wilfiey are next passed through a set of conventional flotation cells, such as are universally used in the wet milling of feldspars. The object is to separate as much barren quartz as possible. Of course, such separation is only partial. The quartz is not entirely barren, but it will be more profitable to discard it than to treat it. Some feldspar is also still attached to it and other minerals as well. Only actual large scale work will determine the economic limit in this quartz separation. On most pegmatites it is rather small at best, say from 10% to 20% of the original pegmatite. Optimum results will only be available as a bookkeeping factor in over-all mill work.

In most pegmatites the more important minerals are those containing beryl and lithium. Being light, they are not removed in the concentrate of heavy minerals but remain with the feldspars. The most important single factor in flotation is to see that they are not rejected with the quartz so frequent analysis is needed as a safeguard. However, as I have called this phase of milling entirely conventional no further description is called for.

The accepted fraction of the pegmatite, approximately 80% of the total, is essentially a crude type of feldspar containing small amounts of beryllium and lithium minerals as well as that portion of the heavy minerals not separated on the Wilfley. This product is substantially unattacked by any acid, save hydrofluoric, so recourse is had to a smelting step. I find that if said product be Patented Aug. 1, 1961 commingled with enough alkali metal sulphide to roughly double the natural amount present, then the resultant smelt is readily attacked. I, therefore, add enough potassium and/or sodium sulphate to furnish the alkali and enough carbon for reduction to sulphide and fuse the mixture in an electric furnace. A previous pelletizing is an advantage though not essential.

When such a product as this smelt is attacked by an acid a hydrated silica remains as the final residue, but this is not a formless blob of silica-gel, but has a definite structure depending upon the removal of the soluble molecules of alkali metal and of aluminum compounds. Contrariwise, if a large excess of alkali be added to the fusion step then a type of silica-gel is produced in the acid treatment step. Such an excess was specified in Ser. No. 724,252. Another distinction is found in the solubility in water of the smelt from the two methods of approach. As prepared herein, said smelt is essentially insoluble in water. With the large excess formerly specified much alumina and silica would go into water solution, particularly if the extraction were made at super-atmospheric temperatures and pressures.

Referring now to the drawing, 1 have shown this smelt as quenched in water and by this sudden cooling the molecular arrangement is still further modified, large crystal formation being made impossible. Fine grinding to say between mesh and 200 mesh is the next step as the particles remain largely the same size after acid treatment. Obviously, a 200 mesh product will contain much material passing a 325 mesh test. No actual sizing is specified but manifestly the reaction time with the finer material is much less. The ground material is then acid treated. No actual acid is used but instead ammonium bisulph-ate is substituted. The attack of this material is quite a lot slower than the use of sulphuric acid but with enough time it is equally thorough. Previously I specified such use of a mineral acid before I learned of the superiority of ammonium bi-sulphate. Apparently the structure of the silica skeleton is infiuenced by the strength of the acid by which the attack was made. Said bi-sulphate may be used in fused condition, being liquid at approximately C., or diluted with water as determined by the type of silica desired. The amount of said ammonium bi-sulphate employed must, of course, be in excess of the stoichiometric amount required to remove the resident bases as water-soluble sulphates in the subsequent leaching step.

Having indicated the importance of the technique employed in producing this silicious residual, I will continue with a description of what happens to the solution obtained thereby. At a later place I shall describe more fully the impregnation step, or at least my preferred Version, when I also describe the procedure for the separated heavy minerals. In the drawing I show this solution as commingled with ammonium carbonate for the precipitation of alumina, iron, etc. and further treatment of the purified solution. An intermediate step, not shown on the drawing may be introduced prior to this carbonate precipitation.

Inasmuch as said solution is composed essentially of the sulphates of potassium, sodium, aluminum, iron, etc. it is obvious that it is prime material for the conventional removal of iron by crystallization as an iron alum. Obviously, such a step would result in a purer subsequent precipitation of the resident alumina, which can thus be made even iron-free. Such an innovation would be at the discretion of the operator, it being as already stated quite conventional.

The next stage is the precipitation of beryllium as the hydroxide by the addition of ammonia, and separation of said precipitate. Obviously, any resident ammonium carbonate in the solution should first be removed by boiling, though such a step is not indicated. After removal of beryllium, lithium is next precipitated as the phosphate and removed as before indicated in the case of aluminum and beryllium. This is only ordinary, text book chemistry, but it should be noted that I have confined myself solely to ammonium compounds, where in my previous applications I specified alkali-metal compounds. The residual solution is, therefore, now a solution of ammonium sulphate with potash and soda. I have shown evaporation and crystallization as the means for removing potash and soda salts, leaving a solution of ammonium sulphate.

I have next indicated a re-cycling of such (K-Na) salts as required in the electric furnace step, the remainder being marketed, of course, after crystallization has removed soda from potash to the extent desired. Parenthetically, it may be stated that this separation is also well known. Some double salt will be produced and that is obviously suited for the recycling.

The ammonium sulphate solution, the final residual, is next fully dehydrated and fused to yield ammonium bisulphate with simultaneous evolution of one-half of the resident ammonia as a gaseous product. In the drawing I have shown all such material re-cycled; the bi-sulphate to the reactor, the ammonia used in part as such in the second precipitator, the remainder as being carbonated before its use in the first precipitator for the removal of alumina and iron. My process is, therefore, independent of both acid and alkali save for minor losses in operation. This is not merely of economic importance, but it has a much greater value as conserving all those minute amounts of rare metals that occur in all pegmatites. It is a fact that all the reactions used herein are incomplete in practical work though they may be considered analytically applicable in the laboratory. This is of no moment in a closed circuit as that which is lost on the first pass only increases the amount present in the circuit until the steps described remove enough to balance the amounts steadily introduced by new raw material. Minor obvious steps have been eliminated from the drawing. Thus, the phosphoric acid present in the recovered lithium phosphate will ultimately be recovered in processing the phosphate, and such phosphate will once more appear as ammonium phosphate for the initial, phosphate precipitation. In other words, all reagents used herein are cyclically returned to the circuit.

I will now describe the treatment of the heavy minerals. This material is mixed with about its own weight of alkali metal sulphates re-cycled of course, and with enough reducing carbon just as described under the former elec tric furnace treatment, except for the much larger addition of the alkali sulphate. Fusion is as before and Water quenching, but no subsequent grinding is needed. Quenching in water results in the formation of a black slurry consisting of alkali compounds in solution and a mass of insoluble sulphides. Separation is made by filtering or otherwise, and the black mud treated with a solution of ammonium bi-sulphate as before. In this manner I make a separation between such metals as titanium which hydrolyses or which has an insoluble sulphate, and the soluble group, chief of which in amount is iron. Much hydrogen sulphide is evolved, and as the alkali solution just described contains much caustic alkali it is prime material for the recovery of this hydrogen sulphide. After use as an absorbent it is de-hydrated and re-cycled to the furnace step for heavy minerals. In this manner, as the re-cycled material contains much sulphide it will require less carbon for reduction and less current for smelting. Having thus obtained the rare elements of the heavy minerals in a concentrated form and, with few exceptions, in solution further separation becomes conventional. The step on which the economics depends 4 is that of an inexpensive method of concentration as a part of an integrated process.

The silica previously separated is now to be impregnated with the suitable catalyst. conventionally this may be done by an alternate soaking in a solution of the metallic base alternating with immersion in ammonia, all being done in such a diluted form that the recovery of anything from the resultant solution becomes impossible. I prefer to operate in a very different fashion.

First I dry, but do not de-hydrate the separated silica. Such material is an excellent adsorbent for ammonia gas. I may improve this adsorption by a decrease in temperature or an increase in pressure. I then release this ammonia-impregnated silica beneath the surface of a solution of the desired metallic salt, but here again I prefer to make such solution by dissolving a hydroxide or a carbonate in a solution of ammonium bi-sulphate, which, of course, becomes neutralized in the process. Inasmuch as the ammonia adsorbed on the silica coats the surface with a uniform film, it follows that the precipitated hydroxide is spread uniformly over the entire internal area. Washing and calcining completes the process, and as the silica is not gelatinous it can be mechanically handled so that a strong solution of ammonium sulphate is produced, said material being then recycled to fresh hydroxide and/or carbonate, after heating to reform the bisulphate exactly as previously described. Not only is this form of impregnation superior but it is held by many experts that in the conventionally made catalyst only a relatively small amount of the catalyst present is truly active. In my case tests indicate that the customary amounts, say 13% to 25% alumina on a silica base, can be cut to less than one-half with no loss in activity.

In thus describing my drawing with many additional comments I have completely described what I regard as new in the process I have developed. That which is truly conventional I have left out with the certainty that any operator who is familiar with the themes involved will have no difficulty in following my instructions and will obtain similar results to those I have obtained. Naturally, I do not confine myself slavishly to a socalled preferred version as it is self evident that much latitude is possible in other directions than in the specific situations that I have described.

Having thus fully described my process, I claim:

The method of processing a pegmatite consisting essentially of a potash feldspar which always contains aluminum, free quartz, and heavy minerals which virtually always contain iron, with subordinate amounts of lithium and beryllium minerals, which comprises; grinding said pegmatite with water to form a thin slurry; passing said slurry over gravity concentrating devices so as to remove as a concentrate most of the heavy minerals liberated by said grinding; passing the tailings from said gravity concentration over flotation devices so as to remove at least a portion of the quartz liberated in said grinding; commingling the residual feldspar and other complex silicates with a suificient amount of re-cycled alkali metal sulphate selected from the group consisting of sodium sulphate, potassium sulphate and mixtures thereof obtained at a later step, to approximately double the sodium and potassium content of said feldspar and with sufiicient carbon to reduce said sulphate to sulphide and fusing the mixture with electrically generated heat to form a liquid slag of complex silicates; quenching said fluid slag in agitated water in which said silicates are substantially insoluble and fine-grinding the resultant product; commingling said ground slag with ammonium bi-sulphate in suflicient amount to remove substantially all resident bases when leached with water thus leaving a residue of hydrated silica with a high internal area; separating said silica and commingling the resultant solution with sufiicient carbonated ammonia to precipitate resident iron and aluminum as carbonate and hydroxide, and removing said precipitate; commingling the solution thus obtained with ammonia to precipitate resident beryllium as its hydroxide, and removing said hydroxide; commingling the resultant solution with ammonium phosphate to precipitate resident lithium as its phosphate, and removing said precipitate; concentrating the residual solution from said phosphate removal and removing alkali metal sulphate selected from the group consisting of sodium sulphate, potassium sulphate and mixtures thereof therefrom by crystallization; recycling a sufiicient amount of said sulphates to the fusion step to supply the additional sodium and potassium required for said fusion; dehydrating the ultimate ammonium sulphate solution remaining after separation of the sodium and potassium sulphates and heating same to convert resident ammonium sulphate into bi-sulphate with simultaneous evolution of ammonia gas; re-cycling the ammonium bi-sulphate thus formed by commingling it with fresh 6 ground slag to remove all bases therefrom, leaving a silicious residue after leaching; and re-cycling the ammonia liberated in the ammonium sulphate decomposition as the ammonium component of the various precipitating compounds employed herein.

References Cited in the file of this patent UNITED STATES PATENTS 1,062,278 Hart May 20, 1913 2,716,591 Thomsen Aug. 30, 1955 2,886,512 Winyall May 12, 1959 FOREIGN PATENTS 22,557 Great Britain May 22, 1913 of 1912 502,987 Great Britain Mar. 29, 1939

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